prompt,ground_truth
"Swift 5: How to add Marker on GoogleMap on tap/touch with coordinates?
*I was learning GoogleMap-SDK for iOS, in the process i did searched for the above question and found satisfactory answers but not the one that's practically useful.*


E.G. : [Swift 3 google map add markers on touch](https://stackoverflow.com/questions/43877300/swift-3-google-map-add-markers-on-touch/43879138)


It adds Marker but does not have Place Name or Place Details and without that Marker is not that useful and for that i have to search other answers to get Address from the coordinates.


So , here i have combined both answers to save time of fellow developers and make it Marker more practical.


","**For Swift 5.0+**



> 
> First, make sure you have added `GMSMapViewDelegate` delegate to your ViewController Class
> 
> 
> 


We do not need `UILongPressGestureRecognizer` or `UITapGestureRecognizer` for this, `GMSMapViewDelegate` provides convenient default method for this.



```
///This default function fetches the coordinates on long-press on `GoogleMapView`
func mapView(_ mapView: GMSMapView, didLongPressAt coordinate: CLLocationCoordinate2D) {
     
     //Creating Marker
     let marker = GMSMarker(position: coordinate)
    
     let decoder = CLGeocoder()

     //This method is used to get location details from coordinates
     decoder.reverseGeocodeLocation(CLLocation(latitude: coordinate.latitude, longitude: coordinate.longitude)) { placemarks, err in
        if let placeMark = placemarks?.first {

            let placeName = placeMark.name ?? placeMark.subThoroughfare ?? placeMark.thoroughfare!   ///Title of Marker
            //Formatting for Marker Snippet/Subtitle       
            var address : String! = """"
            if let subLocality = placeMark.subLocality ?? placeMark.name {
                address.append(subLocality)
                address.append("", "")
            }
            if let city = placeMark.locality ?? placeMark.subAdministrativeArea {
                address.append(city)
                address.append("", "")
            }
            if let state = placeMark.administrativeArea, let country = placeMark.country {
                address.append(state)
                address.append("", "")
                address.append(country)
            }

            // Adding Marker Details
            marker.title = placeName
            marker.snippet = address
            marker.appearAnimation = .pop
            marker.map = mapView
        }
    }
}

```

***HOPE IT HELPS !!!***


"
"C++ Using the function of the class two class above
I have the current setup:



```
class Interface1
{
  public:
    virtual ~Interface1() {}
    virtual void DoSomething() = 0;
};

class Interface2 : public virtual Interface1
{
  public:
    virtual ~Interface2() {}
    virtual void DoSomething() override = 0;
    virtual void DoSomethingElse() = 0;
};

class MyClass1 : public Interface1
{
  public:
    MyClass1();
    void DoSomething() override;
};

class MyClass2 : public Interface2
{
  public:
    MyClass2();
    void DoSomething() override;
    void DoSomethingElse() override;
};

int main()
{
    std::map<std::string, boost::any> items;
    items.insert(make_pair(""item1"", shared_ptr<Interface1>(new MyClass1())));
    items.insert(make_pair(""item2"", shared_ptr<Interface2>(new MyClass2())));

    auto object = items.at(""item2"");
    auto item = boost::any_cast<shared_ptr<Interface1>>(object);
    item->DoSomething();

    return 0;
}

```

When I run this code, nothing happens. `MyClass2` doesn't appear to be calling `DoSomething()`, which is what I would like. How can I make the call to `Interface1::DoSomething()` actually call `Interface2::DoSomething()`? I would think it would be possible because they all inherit from each other, but I can't seem to make it work.


The reason I want this is because I have some functions which will only work with classes inherited from `Interface2`, but some functions need to support classes derived from either `Interface1` and `Interface2`. Once boost::any takes over I loose which type it originally was, but it shouldn't be a problem if I could use the setup described above, so even if my original class was derived from `Interface2`, it could call the same function in `Interface1` and get the same result.


Is there a way of doing what I want with the current setup?


EDIT:
Sorry, the `void` in front of the constructors where my bad, but that is not the issue.


","Why do you need the `boost::any`?


If you need to determine the difference between `Interface1` and `Interface2`, and you have a `std::shared_pointer` stored in your map, then just store a `std::shared_pointer<Interface1>` and use `std::dynamic_pointer_cast<Interface2>` to determine whether you have an `Interface1` or an `Interface2`


**Example:**



```
#include <map>
#include <memory>
#include <iostream>

class Interface1
{
  public:
    virtual ~Interface1() = default;
    virtual void DoSomething() = 0;
};

class Interface2 : public Interface1
{
  public:
    virtual ~Interface2() = default;
    virtual void DoSomethingElse() = 0;
};

class MyClass1 : public Interface1
{
  public:
    MyClass1() {}
    void DoSomething()     override { std::cout << ""\t\t"" << __PRETTY_FUNCTION__ << '\n'; }
};

class MyClass2 : public Interface2
{
  public:
    MyClass2() {}
    void DoSomething()     override { std::cout << ""\t\t"" << __PRETTY_FUNCTION__ << '\n'; }
    void DoSomethingElse() override { std::cout << ""\t\t"" << __PRETTY_FUNCTION__ << '\n'; }
};

int main()
{
    std::map<std::string, std::shared_ptr<Interface1>> items;

    items.emplace(""item1"", std::make_shared<MyClass1>());
    items.emplace(""item2"", std::make_shared<MyClass2>());

    auto check = [&items](const std::string& name)
        {
            auto object = items.at(name);
            auto item = std::dynamic_pointer_cast<Interface2>(object);
            if (item)
            {
                std::cout << name << "" is an Interface2\n"";
                item->DoSomething();
                item->DoSomethingElse();
            }
            else
            {
                std::cout << name << "" is an Interface1\n"";
                object->DoSomething();
            }
        };

    check(""item1"");
    check(""item2"");

    return 0;
}

```

**Output:**



```
item1 is an Interface1
        virtual void MyClass1::DoSomething()
item2 is an Interface2
        virtual void MyClass2::DoSomething()
        virtual void MyClass2::DoSomethingElse()

```

**Some final notes:**


- I also question the need for virtual inheritance between `Interface2` and `Interface1`
- I don't believe you need to override `DoSomething` in `Interface2` - it's already there by publically inheriting from `Interface1`


`virtual void DoSomething() override = 0;` is unnecessary


"
"Julia: Macro threads and parallel
as we know, Julia supports parallelism and this is something rooted in the language which is very good.


I recently saw that Julia supports threads but it seems to me to be experimental. I noticed that in the case of using the `Threads.@Threads` macro there is no need for Shared Arrays which is perhaps a computational advantage since no copies of the objects are performed. I also saw that there is the advantage of not declaring all functions with `@everywhere`.


Can anyone tell me the advantage of using the `@parallel` macro instead of the `@threads` macro?


Below are two simple examples of using non-synchronized macros for parallelism.


**Using the @threads macro**



```
addprocs(Sys.CPU_CORES)

function f1(b)
   b+1 
end

function f2(c)
   f1(c)
end

result = Vector(10)

@time Threads.@threads for i = 1:10
  result[i] = f2(i)
end 

```

*0.015273 seconds (6.42 k allocations: 340.874 KiB)*


**Using the @parallel macro**



```
addprocs(Sys.CPU_CORES)

@everywhere function f1(b)
   b+1 
end

@everywhere function f2(c)
   f1(c)
end

result = SharedArray{Float64}(10)
@time @parallel for i = 1:10
    result[i] = f2(i)
end

```

*0.060588 seconds (68.66 k allocations: 3.625 MiB)*


It seems to me that for Monte Carlo simulations where loops are mathematically independent and there is a need for a lot of computational performance the use of the `@threads` macro is more convenient. What do you think the advantages and disadvantages of using each of the macros?


Best regards.


","Here is my experience:


### Threads


Pros:


- shared memory
- low cost of spawning Julia with many threads


Cons:


- constrained to a single machine
- number of threads must be specified at Julia start
- possible problems with false sharing (<https://en.wikipedia.org/wiki/False_sharing>)
- often you have to use locking or atomic operations for the program to work correctly; in particular many functions in Julia are not threadsafe so you have to be careful using them
- not guaranteed to stay in the current form past Julia 1.0


### Processess


Pros:


- better scaling (you can spawn them e.g. on a cluster of multiple machines)
- you can add processes while Julia is running


Cons:


- low efficiency when you have to pass a lot of data between processes
- slower to start
- you have to explicitly share code and data to/between workers


### Summary


Processes are much easier to work with and scale better. In most situations they give you enough performance. If you have large data transfers between parallel jobs threads will be better but are much more delicate to correctly use and tune.


"
"Numpy concatenate is slow: any alternative approach?
I am running the following code:



```
for i in range(1000)
    My_Array=numpy.concatenate((My_Array,New_Rows[i]), axis=0)

```

The above code is slow. Is there any faster approach? 


","This is basically what is happening in all algorithms based on arrays.


Each time you change the size of the array, it needs to be resized and every element needs to be copied. This is happening here too. (some implementations reserve some empty slots; e.g. doubling space of internal memory with each growing).


- If you got your data at np.array creation-time, just add these all at once (memory will allocated only once then!)
- If not, collect them with something like a linked list (allowing O(1) appending-operations). Then read it in your np.array at once (again only one memory allocation).


This is not much of a numpy-specific topic, but much more about data-strucures.


**Edit:** as this quite vague answer got some upvotes, i feel the need to make clear that my linked-list approach is one possible example. As indicated in the comment, python's lists are more array-like (and definitely not linked-lists). But the core-fact is: list.append() in python is [fast](https://wiki.python.org/moin/TimeComplexity) (amortized: O(1)) while that's not true for numpy-arrays! There is also a small part about the internals in the [docs](https://docs.python.org/2/faq/design.html#how-are-lists-implemented):



> 
> 
> > 
> > How are lists implemented?
> > 
> > 
> > **Python’s lists are really variable-length arrays, not Lisp-style linked lists**. The **implementation uses a contiguous array of references to other objects**, and keeps a pointer to this array and the array’s length in a list head structure.
> > 
> > 
> > This makes indexing a list a[i] an operation whose cost is independent of the size of the list or the value of the index.
> > 
> > 
> > When items are appended or inserted, the array of references is resized. **Some cleverness is applied to improve the performance of appending items repeatedly**; when the array must be grown, some extra space is allocated so the next few times don’t require an actual resize.
> > 
> > 
> > 
> 
> 
> 


(bold annotations by me)


"
"Collatz Sequence
I'm a beginner and I only know limited amounts such as `if`, `while`, `do while`. So I'm here to check if I'm coding with best practise and the most effective methods to my current knowledge.



```
import java.util.Scanner;


public class SixtyTwo {

    public static void main(String[] args) {

            Scanner keyboard = new Scanner(System.in);

            System.out.print(""Starting Number: "");
            int n = keyboard.nextInt();
            int counter = 0;
            int stepsTaken = 0;
            int largestNumber = 0;
            System.out.println();

            while ( n != 1 ){
                    if ( ( n & 1 ) == 0 ) {
                            System.out.print( (n = ( n / 2 )) + "" "" );
                            stepsTaken++;
                            counter++;
                    }       else {
                            System.out.print( (n = ( n * 3 ) + 1) + "" "" );
                            stepsTaken++;
                            counter++;
                    }

                    if ( n > largestNumber ){
                            largestNumber = n;
                    }

                    if (counter == 9){
                            counter = 0;
                            System.out.print(""\n"");
                    }
            }

            System.out.println();
            System.out.println(""\nTerminated after "" + stepsTaken + "" steps."");
            System.out.println(""The largest value was "" + largestNumber + ""."");
    }
}

```

","That is some overall quite good code you have there, but I have a couple of comments.


- `Scanner` should be closed when you are done with the input. Simply call `keyboard.close();` when you have acquired the input that you need.
- Your class is called `SixtyTwo` but I have no clue what that has to do with anything. A better name would be `Collatz`
- `if ( ( n & 1 ) == 0 )` although it is a nice way of checking if a number is even, it is for Java programmers more readable to use the `%` (modulo) operator. I would use `if (n % 2 == 0)`.
- `System.out.print( (n = ( n / 2 )) + "" "" );` ...now this I'm not a big fan of. Sure, it works to both modify a value and outputting it on the same line, but I would separate the assignment to it's own line. Assign on one line, output on another, makes things much clearer to read.
- `stepsTaken++;` and `counter++;` don't need to be inside the if-else blocks, put them outside them as they are always done no matter if your condition is true or false.
- `if (counter == 9)` can instead be `if (stepsTaken % 9 == 0)`, which would remove the need for the `counter` variable entirely.


"
"Make a backup of the installed packages on Ubuntu like Linux Mint
On the Linux mint you have a tool that list and save all your installed packages in a simple way. There is any way to do the same on Ubuntu/Xubuntu/Lubuntu/...?


","**EDIT:** This is no longer a free application.


### Aptik


After seeing the various answers here (and not disagreeing with any of them) it strikes me that you asked for simplicity. In my comment, I linked to an application called Aptik and I'm going to show you why I think this meets your criteria best.


Aptik is simple to install and trivially easy to use. It is also a handy dandy GUI (Graphical User Interface) that has easy to use buttons and requires absolutely no real advanced knowledge. If you can click on a button with your mouse then you can install Aptik.


If you'd like to *try* read more about Aptik then you can click [here](http://www.teejeetech.in/) and visit their home page. I'm not actually sure what good that will do. Their home page doesn't seem to have a whole lot of information.


-
**It is Simple and it Does Work!**


Not only does it do what you asked, it does a bit more. I think a screen shot should be fairly self-explanatory.


[![Aptik in Action](https://i.stack.imgur.com/nyPQR.png)](https://i.stack.imgur.com/nyPQR.png)


### How to Install Aptik


1. Open your terminal by pressing `CTRL`+`ALT`+`T`.
2. Copy and Paste this to the terminal:



```
 sudo apt-add-repository ppa:teejee2008/ppa

```
3. Press `ENTER`
4. Enter your password. Nothing will show on your screen, the cursor will not move, no asterisks will appear - this is normal behaviour.
5. Copy and paste this to your terminal:



```
 sudo apt-get update 

```
6. Press `ENTER`.
7. Copy and paste this to your terminal:



```
 sudo apt-get install aptik  

```
8. Press `ENTER`.
9. Allow the application to install and follow any on-screen prompts.


### Running Aptik


If your computer is anything like mine then Aptik should magically appear in the Start menu. In my case, it appears under System Tools. Your case is probably not like mine. Open the launcher and search for `aptik`. If, for some peculiar reason, Aptik is not available then it can be launched from the terminal (use above commands) by running `sudo aptik-launcher`.


If, for some even stranger reason, you want to go ahead and run the application from the terminal, entirely, you can do that too but you're on your own. For the sake of completeness, this is a list of commands.



```
Aptik v1.6.4 by Tony George (teejee2008@gmail.com)

Syntax: aptik [options]

Options:

  --list-available      List available packages
  --list-installed      List installed packages
  --list-top            List top-level installed packages
  --list-{manual|extra} List top-level packages installed by user
  --list-default        List default packages for linux distribution
  --list-ppa            List PPAs
  --list-themes         List themes in /usr/share/themes
  --list-icons          List icon themes in /usr/share/icons
  --backup-ppa          Backup list of PPAs
  --backup-packages     Backup list of manual and installed packages
  --backup-cache        Backup downloaded packages from APT cache
  --backup-themes       Backup themes from /usr/share/themes
  --backup-icons        Backup icons from /usr/share/icons
  --restore-ppa         Restore PPAs from file 'ppa.list'
  --restore-packages    Restore packages from file 'packages.list'
  --restore-cache       Restore downloaded packages to APT cache
  --restore-themes      Restore themes to /usr/share/themes
  --restore-icons       Restore icons to /usr/share/icons
  --take-ownership      Take ownership of files in your home directory
 --backup-dir          Backup directory (defaults to current directory)
  --[show-]desc         Show package description if available
  --yes                 Assume Yes for all prompts
  --h[elp]              Show all options

```

### Closure


Regardless of which method you choose, this has been included to ensure that there's a nice, simple, GUI way to do this. There are many ways to accomplish things in Linux and it is always nice to have options.


"
"How to pause MySQL before taking an LVM/ZFS snapshot?
How can I instruct MySQL to complete all ""in-progress"" transactions, but to delay starting new ones (without kicking clients off) until I have taken a ZFS or LVM snapshot (which takes less than a second).


e.g.


1. pause MySQL, waiting for ""in-progress"" transactions to complete
2. sync to disk
3. take ZFS/LVM snapshot
4. resume MySQL
The point of this is to get a consistent snapshot for backup purposes. Step 2 takes a fraction of a second.


Step one should not cause client errors, just a very short pause until step 4 is reached.


Are there MySQL commands which can do 1 and 4? What are they?


","A hacky way would be, to wait for the transactions to finish:


`mysql> FLUSH LOCAL TABLES;
Query OK, 0 rows affected (11.31 sec)`


and then getting a read lock:


`mysql> FLUSH TABLES WITH READ LOCK;
Query OK, 0 rows affected (22.55 sec)`


Now all queries are blocked (ie. they wait for the lock to get released) until your session ends. Mind you - you still have to wait until all transactions are finished. Depending on your workload this can take a while (UPDATE-ing a few million rows...).


You can code this in your favourite script language.


But seriously - why not use [Xtrabackup](https://www.percona.com/doc/percona-xtrabackup/2.3/index.html ""Xtrabackup"") ? It does take care of a consistent snapshot of mysql for you and you can dump it on the filesystem, and zfs/lvm snapshot it.


"
"Randomly remove 30% of values in numpy array
I have a 2D numpy array which contains my values (some of them can be NaN). I want to remove the 30% of the non-NaN values and replace them with the mean of the array. How can I do so? What I tried so far:



```
def spar_removal(array, mean_value, sparseness):
    array1 = deepcopy(array)
    array2 = array1
    spar_size = int(round(array2.shape[0]*array2.shape[1]*sparseness))
    for i in range (0, spar_size):
        index = np.random.choice(np.where(array2 != mean_value)[1])
        array2[0, index] = mean_value
    return array2

```

But this is just picking the same row of my array. How can I remove from all over the array? It seems that choice works only for one dimension. I guess what I want is to calculate the `(x, y)` pairs that I will replace its value with `mean_value`.


","There's likely a better way, but consider:



```
import numpy as np

x = np.array([[1,2,3,4],
              [1,2,3,4],
              [np.NaN, np.NaN, np.NaN, np.NaN],
              [1,2,3,4]])

# Get a vector of 1-d indexed indexes of non NaN elements
indices = np.where(np.isfinite(x).ravel())[0]

# Shuffle the indices, select the first 30% (rounded down with int())
to_replace = np.random.permutation(indices)[:int(indices.size * 0.3)]

# Replace those indices with the mean (ignoring NaNs)
x[np.unravel_index(to_replace, x.shape)] = np.nanmean(x)

print(x)

```

**Example Output**



```

[[ 2.5  2.   2.5  4. ]
 [ 1.   2.   3.   4. ]
 [ nan  nan  nan  nan]
 [ 2.5  2.   3.   4. ]]

```

NaNs will never change and floor(0.3 \* number of non-NaN elements) will be set to the mean (the mean ignoring NaNs).


"
"How to programmatically replace an HyperLinkField in a ASP.NET GridView
I have an ASP.NET Web Forms application. In my application I have a **GridView** that works smoothly. I have several text fields and the last one is a `<asp:hyperlinkfield>`.


Now I would like to programmatically change the field by placing a **simple link** instead of the `hyperlinkfield` if a specific condition is fulfilled. Therefore I catch the `onRowDataBound` event:



```
Sub myGridView_RowDataBound(ByVal sender As Object, ByVal e As GridViewRowEventArgs) Handles myGridView.RowDataBound

    If (condition) Then
           Dim link = New HyperLink()
           link.Text = ""login""
           link.NavigateUrl = ""login.aspx""
           e.Row.Cells(3).Controls.Add(link)
        End If
    End If
End Sub

```

where **n** is the cell where the `hyperlinkfield` is placed. With this code it just adds to the `hyperlinkfield` the new `link`. How can I replace it?


PS: The code is in VB6 but I am a C# programmer, answers with both languages are accepted


","Remove the control you want to replace from the collection before adding the new one:



```
protected void TestGridView_RowDataBound(object sender, GridViewRowEventArgs e)
{
  if (e.Row.RowType == DataControlRowType.DataRow)
  {
    HyperLink newHyperLink = new HyperLink();
    newHyperLink.Text = ""New"";
    e.Row.Cells[3].Controls.RemoveAt(0);
    e.Row.Cells[3].Controls.Add(newHyperLink);
  }
}

```

But I agree with the others, just change the existing link's properties:



```
protected void TestGridView_RowDataBound(object sender, GridViewRowEventArgs e)
{
  if (e.Row.RowType == DataControlRowType.DataRow)
  {
    HyperLink link = e.Row.Cells[0].Controls[0] as HyperLink;
    if (link != null)
    {
      link.Text = ""New"";
      link.NavigateUrl = ""New"";
    }
  }
}

```

"
"How can I install MinGW-w64 and MSYS2?
I am trying to build some open source library. I need a package management system to easily download the dependencies. At first I am using [MinGW](https://en.wikipedia.org/wiki/MinGW) and [MSYS](https://en.wikipedia.org/wiki/MinGW#History). But the included packages are limited. Someone told me to use [Mingw-w64](https://en.wikipedia.org/wiki/Mingw-w64) and [MSYS2](https://en.wikipedia.org/wiki/Mingw-w64#MSYS2).


I downloaded the `mingw-w64-install` from [here](http://mingw-w64.yaxm.org/doku.php/download). When running, it reports the following error. How can I fix it?


![Enter image description here](https://i.stack.imgur.com/F0IoK.png)


And by the way, from the Mingw-w64 download page, I see a lot of download links. Even [Cygwin](https://en.wikipedia.org/wiki/Cygwin) is listed. How are Cygwin and Mingw-w64 related?


![Enter image description here](https://i.stack.imgur.com/uzlMg.png)


My current understanding is, in the time of MinGW and MSYS, MSYS is just a nice addon to MinGW, while in Mingw-w64 + MSYS2, MSYS2 is stand-alone and Mingw-w64 is just a set of libraries it can work with. Just like Cygwin can download many different packages.


","Unfortunately, the MinGW-w64 installer you used sometimes has this issue. I myself am not sure about why this happens (I think it has something to do with Sourceforge URL redirection or whatever that the installer currently can't handle properly enough).


Anyways, if you're already planning on using MSYS2, there's no need for that installer.


1. Download MSYS2 from [this page](https://msys2.github.io/).
2. After the install completes, click on the `MSYS2 UCRT64` in the Start menu (or `C:\msys64\ucrt64.exe`).


If done correctly, the terminal prompt will say `UCRT64` in magenta letters, not `MSYS`.
3. Update MSYS2 using `pacman -Syuu`. If it closes itself during the update, restart it and repeat the same command to finish the update.


You should routinely update your installation.
4. Install the toolchain: (i.e. the compiler and some extra tools)



```
pacman -S mingw-w64-ucrt-x86_64-toolchain

```
5. Install any libraries/tools you may need. You can search the repositories by doing



```
pacman -Ss name_of_something_i_want_to_install

```

e.g.



```
pacman -Ss gsl

```

and install using



```
pacman -S package_name_of_something_i_want_to_install

```

e.g.



```
pacman -S mingw-w64-ucrt-x86_64-gsl

```

and from then on the GSL library will be automatically found by your compiler!


Make sure any compilers and libraries you install have this package prefix: `mingw-w64-ucrt-x86_64-`. Only use unprefixed packages for misc command-line utilities (such as `grep`, `sed`, `make`, etc), unless you know what you're doing.
6. Verify that the compiler is working by doing



```
gcc --version

```


If you want to use the toolchains (with installed libraries) outside of the MSYS2 environment, all you need to do is add `C:/msys64/ucrt64/bin` to your `PATH`.


MSYS2 provides [several compiler flavors](https://stackoverflow.com/q/76552264/2752075), UCRT64 being one of them. It should be a reasonable default.


"
"Stochastic volatility model - why not identified?
For the following model


$y\_t = \beta e^{h\_t/2} \epsilon\_t$


$h\_{t+1} = \mu + \phi(h\_t - \mu) + \sigma\_n \eta\_t$


[this Kim et al (1998)](http://apps.olin.wustl.edu/faculty/chib/papers/KimShephardChib98.pdf) paper writes that 



> 
> For identifiability reasons either $\beta$ must be set to one or $\mu$ to zero
> 
> 
> 


Why is that? (to me this is not immediately obvious)


","Intuitively, it's because both $\beta$ and $\mu$ are responsible for the average *scale* of the volatility.


Say you define a new parameter $\mu' = \mu + \Delta$. Then, it follows that:


$$h\_{t+1} + \Delta = \mu' + \phi((h\_t + \Delta) - \mu') + \sigma\_\eta \eta\_t$$


Substituting in the other equation:


$$y\_t = \beta e^{-\Delta/2} e^{\frac{h\_t + \Delta}{2}} \varepsilon\_t$$


If you then define $\beta' = \beta e^{-\Delta/2}$ and $h\_t'=h\_t + \Delta$, it follows that:


$$y\_t = \beta' e^{h'\_t/2} \varepsilon\_t$$
$$h'\_{t+1} = \mu' + \phi(h'\_t - \mu') + \sigma\_\eta \eta\_t$$


This is the same model form that you started with, so you will get the same likelihood no matter the value of $\Delta$, if you adjust both parameters in this way by the same value of $\Delta$; this is the identifiability issue that was alluded to. 


In theory it doesn't really matter which parameter you get rid of, but in the paper you linked, the specific application (MCMC) typically works better when parameters are less correlated in the posterior. In this case, it means you should get rid of $\beta$. This is demonstrated empirically in the paper.


"
"Django: How to call management custom command execution from admin interface?
Referring to, [executing management commands from code](https://stackoverflow.com/questions/907506/how-can-i-call-a-custom-django-manage-py-command-directly-from-a-test-driver), 


Is their a way to call this command execution code from the django admin interface? 


I've a custom command to do some periodic update of database which has been [scheduled as cron](https://stackoverflow.com/questions/573618/django-set-up-a-scheduled-job). The cron is working fine. I need to update the database manually from the admin interface whenever needed.


","**update:** you can run any management command simply by calling the function `call_command('compilemessages')` from anywhere within your python code. Note: this obviously is a *blocking* process from the caller's perspective. With the ajax example below you can have a form of non-blocking/asynchronous user experience. Depending on the backend implementation you can take this isolation level further.


Example:



```
from django.core.management import call_command
call_command('compilemessages')

```

If the task is bound to the object currently viewed in the admin, a nice way might be to implement an extra view called by an ajax script when clicking a button. The extra view could optionally be wrapped as a celery task, e.g.


**models.py**



```
class Foo(models.Model):
    # fields...

    def my_task_init(self):
        return mark_safe(""<img class='loading' src='/static/img/loading.gif' alt='loading' style='display:none;' /><a data-identifier='task_%i' class='task'><img src='/static/img/process.png' style='cursor:pointer;' /></a>"") % self.id
    my_task_init.allow_tags = True
    my_task_init.short_description = _(u""Execute Task"")

```

**admin.py**



```
class FooAdmin(admin.ModelAdmin):
    list_display = ['other_field', 'my_task_init']

    class Media:
        js = (
            'https://ajax.googleapis.com/ajax/libs/jquery/1.7.2/jquery.js',
            '/static/js/admin_tasks.js',
        )

    def get_urls(self):
        urls = super(FooAdmin, self).get_urls()
        extra_urls = patterns('',
            (r'^my-task/$', self.admin_site.admin_view(self.parse_view))
        )
        return extra_urls + urls

    # optionally decorated by celery
    def task_view(self, request):
        if not request.is_ajax():
            raise Http404
        task_id = request.GET.get('task_id')
        # your logic
        return HttpResponse('Success')

```

**admin\_tasks.js**



```
$(document).ready(function (){
   $('.task').click(function(){
       var image = $(this).find('img'),
           loading = $(this).parent().find('.loading'),
           task_id = $(this).data('identifier').replace('task_', '');
       $.ajax({
           type: ""GET"",
           data: ({'task_id': task_id}),
           url: ""/admin/app/model/my-task/"",
           beforeSend: function() {
               image.hide();
               loading.show();
           },
           statusCode: {
               200: function() {
                   loading.hide();
                   image.attr('src', '/static/img/success.png');
                   image.show();
               },
               404: function() {
                   loading.hide();
                   image.attr('src', '/static/img/error.png');
                   image.show();
               },
               500: function() {
                   loading.hide();
                   image.attr('src', '/static/img/error.png');
                   image.show();
               }
           }
       });
   });
});

```

If you're trying to initiate an unbound task you could just override a template element or add some html.


"
"Nginx: Permission denied to Gunicorn socket on CentOS 7
I'm working in a Django project deployment. I'm working in a CentOS 7 server provided ma EC2 (AWS). I have tried to fix this bug by many ways but I cant understand what am I missing.


I'm using ningx and gunicorn to deploy my project. I have created my `/etc/systemd/system/myproject.service`file with the following content:



```
[Unit]
Description=gunicorn daemon
After=network.target

[Service]
User=centos
Group=nginx
WorkingDirectory=/home/centos/myproject_app
ExecStart=/home/centos/myproject_app/django_env/bin/gunicorn --workers 3 --bind unix:/home/centos/myproject_app/django.sock app.wsgi:application
[Install]
WantedBy=multi-user.target

```

When I run `sudo systemctl restart myproject.service`and `sudo systemctl enable myproject.service`, the `django.sock` file is correctly generated into `/home/centos/myproject_app/`.


I have created my `nginx` conf flie in the folder /etc/nginx/sites-available/ with the following content:



```
server {
    listen       80;
    server_name  my_ip;
    charset      utf-8;

    client_max_body_size       10m;
    client_body_buffer_size    128k;

    # serve static files
    location /static/ {
        alias /home/centos/myproject_app/app/static/;
    }

    location / {
        include proxy_params;
        proxy_pass http://unix:/home/centos/myproject_app/django.sock;
    }
}

```

After, I restart `nginx` with the following command:



```
sudo systemctl restart nginx

```

If I run the command `sudo nginx -t`, the reponse is:



```
nginx: configuration file /etc/nginx/nginx.conf test is successful

```

When I visit my\_ip in a web browser, I'm getting a 502 bad gateway response.


If I check the nginx error log, I see the following message:



```
1 connect() to unix:/home/centos/myproject_app/django.sock failed (13: Permission denied) while connecting to upstream

```

I really have tried a lot of solutions changing the sock file permissions. But I cant understand how to fix it. How can I fix this permissions bug?... Thank you so much


","If all the permissions under the `myproject_app` folder are correct, and `centos` user or `nginx` group have access to the files, I would say it looks like a Security Enhanced Linux (SELinux) issue.


I had a similar problem, but with RHEL 7. I managed to solve it by executing the following command:



```
sudo semanage permissive -a httpd_t

```

It's related to the security policies of SELinux, you have to add the `httpd_t` to the list of permissive domains.


This post from the NGINX blog may be helpful: [NGINX: SELinux Changes when Upgrading to RHEL 6.6 / CentOS 6.6](https://www.nginx.com/blog/nginx-se-linux-changes-upgrading-rhel-6-6/)


Motivated by a similar issue, I wrote a tutorial a while ago on [How to Deploy a Django Application on RHEL 7](https://simpleisbetterthancomplex.com/tutorial/2017/05/23/how-to-deploy-a-django-application-on-rhel.html). It should be very similar for CentOS 7.


"
"Change colour navbar header Ionic 2
I have this problem... My colour is white right now, my code is like this:



```
<ion-header >
  <ion-navbar>
    <ion-title>
      HELLO
    </ion-title>
  </ion-navbar>
</ion-header>

```

Change color with this opcion is easy (primary, secondary, danger, light, dark)



```
<ion-header >
      <ion-navbar danger>
        <ion-title>
          HELLO
        </ion-title>
      </ion-navbar>
 </ion-header>

```

but my problem is when I want to use custom colors. 
Somebody know how can I resolve it?
Thanks inadvance.


Best regards.


","There're two ways of doing this, based on if you want to change the color only in a single page, or if you want to change it in all the pages from your app:


### 1) Change it in a single page/view


Just like you can see [here](http://ionicframework.com/docs/v2/theming/theming-your-app/)



> 
> To change the theme, just tweak the `$colors` map in your
>  `src/theme/variables.scss` file:
> 
> 
> 



```
$colors: (
  // ...
  newcolor:    #55acee

)

```

And then use it in the view



```
 <ion-header>
      <ion-navbar color=""newcolor"">
        <ion-title>
          HELLO
        </ion-title>
      </ion-navbar>
 </ion-header>

```

### 2) Change it in all the pages/views


In this case, you'd need to add the following in your `variables.scss` file to override Ionic's defaults:



```
$toolbar-ios-background: #55acee;
$toolbar-md-background: #55acee;
$toolbar-wp-background: #55acee;

```



---


### Edit



> 
> Hi, how can I add gradient in app/theme/app.variables.scss?
> 
> 
> 


You could add the colors that you're going to use in the `src/theme/variables.scss`:



```
$header-first-color: #AAAAAA;
$header-last-color: #000000;

```

And then set a rule to use it (in your `app.scss` file if you want to apply it to every page, or in the `page-name.scss` file if you want to apply it to a single page):



```
ion-header {
    .toolbar-background {
        background: linear-gradient(135deg, $header-first-color 0%, $header-last-color 100%);
    }
}

```

"
"Can I use a library that uses guice to bind contact and implementations in applications without issues?
I have a play application and want to take a common operation out from the application and make it as a library in order to use in other play applications.


This proposing library has a contract(interface) and several implementations of top of that.
I have 2 questions;


1. What if I use guice to bind contact and implementations with named injection(using annotations) and call them with the appropriate annotation in the play application without instantiating library classes in the apply application?
2. Can we use such libraries in other java applications without any issue? Ex: Use a library which used Guice in a Spring application


","You are confusing dependency injection (DI) with the more specific idea of using a DI framework to supply those injected dependencies. remember that it's perfectly possible to use pure DI to inject dependencies without a framework. So the answer to the question, ""*Is it good to use dependency injection in a java library?*"", is ""yes"". As a default state, you should use DI everywhere, only not using it when there's a compelling reason not to.


But the body of your question then asks about tying your library to a specific DI framework. Unless you are sure that all of your users will be using that framework, then that is likely a bad idea. You couple people to a specific framework that they may not wish to use. Further, they may already be using another framework and they tend not to be compatible with each other. As a result, you risk driving folk away from your library as it's too difficult to use.


So the answer to the question, ""*should I couple my library to guice (or another framework)*"" is likely to be ""no"" for most scenarios. 


"
"Codeigniter Cart - saving data in database - how to approach?
I need help with handling orders and cart in my web application.


I decided to use Cart library built in Codeigniter 2.


I've seen some tutorials about that Cart library and i know how to use it, but i dont know:


- when shall I create/save that order in database? when user is adding items to cart, or shall i keep all order data in session and ""move"" it do database when user accepts order?
- shall I store session id or something other in database?


I tried to see how cart/order functionality is implemented in PrestaShop, but it looks too complicated for me, im amateur PHP programmer.


","The ideal and good way to use cart is keep it in session, codeigniter's cart class is doing the same thing, and when user gives order use this data put this order in database and do other stuff like payment gateway, shipping. 
 If you want to use user to keep his order in next session, like if user add some product in cart and he quits before giving order and he is a registered user, then you can save his cart every time in database, so that if he gone away without putting order you can show him his orders next time when he logged in.


You can store users cart data in database using `$this->cart->contents();` method of cart class. 
use like this 



```
$cartContentString = serialize($this->cart->contents());

```

you will get a string of cart content, you can save this string in database and later use it like 



```
$cartArray = unserialize($cartContentString);

```

"
"why in an 'if' statement 'then' has to be in the next line in bash?
`if` is followed by `then` in bash but I don't understand why `then` cannot be used in the same line like `if [...] then` it has to be used in the next line. Does that remove some ambiguity from the code? or bash is designed like that? what is the underlying reason for it?


I tried to write `if` and `then` in the same line but it gave the error below:



```
./test: line 6: syntax error near unexpected token \`fi'
./test: line 6: \`fi'

```

the code is:



```
#!/bin/bash
if [ $1 -gt 0 ] then
echo ""$1 is positive"" 
fi

```

","It has to be preceded by a separator of some description, not *necessarily* on the next line(a). In other words, to achieve what you want, you can simply use:



```
if [[ $1 -gt 0 ]] ; then
    echo ""$1 is positive""
fi

```

As an aside, for one-liners like that, I tend to prefer:



```
[[ $1 -gt 0 ]] && echo ""$1 is positive""

```

But that's simply because I prefer to see as much code on screen as possible. It's really just a style thing which you can freely ignore.




---


(a) The reason for this can be found in the `Bash` manpage (my emphasis):



> 
> RESERVED WORDS: Reserved words are words that have a special meaning to the shell. The following words are recognized as reserved when unquoted and either ***the first word of a simple command*** (see SHELL GRAMMAR below) or the third word of a `case` or `for` command:
> 
> 
> `! case coproc do done elif else esac fi for function if in select then until while { } time [[ ]]`
> 
> 
> 


Note that, though that section states it's the ""first word of a simple command"", the manpage seems to contradict itself in the referenced `SHELL GRAMMAR` section:



> 
> A simple command is a sequence of optional variable assignments followed by blank-separated words and redirections, and terminated by a control operator. ***The first word specifies the command to be executed,*** and is passed as argument zero.
> 
> 
> 


So, whether you consider it part of the next command or a separator of some sort is arguable. What is *not* arguable is that it needs a separator of some sort (newline or semicolon, for example) before the `then` keyword.


The manpage doesn't go into *why* it was designed that way but it's *probably* to make the parsing of commands a little simpler.


"
"How to create a CLI in Python that can be installed with PIP?
As the title suggests, I'm trying to make a python script accessible from the command line. I've found libraries like [click](https://pypi.org/project/click/) and [argv](https://pypi.org/project/argv/) that make it easy to access arguments passed from the command line, but the user still has to run the script through Python. 


Instead of



```
python /location/to/myscript.py

```

I want to be able to just do 



```
myscript

```

from any directory


From what I understand, I can achieve this on my computer by editing my PATH variables. However, I would like to be able to simply do:



```
pip install myscript

```

and then access the script by typing `myscript` from anywhere. Is there some special code I would put in the `setup.py`?


","You can do this with `setuptools`


an example of a nice `setup.py` (say your package requires pandas and numpy):



```
import setuptools
setuptools.setup(
    name='myscript',
    version='1.0',
    scripts=['./scripts/myscript'],
    author='Me',
    description='This runs my script which is great.',
    packages=['lib.myscript'],
    install_requires=[
        'setuptools',
        'pandas >= 0.22.0',
        'numpy >= 1.16.0'
    ],
    python_requires='>=3.5'
)

```

Your directory should be setup as follows:



```
[dkennetz package]$ ls
lib scripts setup.py

```

inside lib would be:



```
[dkennetz package]$ ls lib
myscript

```

inside of `myscript` would be:



```
[dkennetz package]$ ls lib/myscript
__main__.py
__init__.py
helper_module1.py
helper_module2.py

```

main would be used to call your function and do whatever you want to do.


inside scripts would be:



```
[dkennetz package]$ ls scripts
myscript

```

and the contents of `myscript` would be:



```
#!/usr/bin/env bash

if [[ ! $@ ]]; then
    python3 -m myscript -h
else
    python3 -m myscript $@
fi

```

then to run you do: `python setup.py install`


which will install your program and all of the dependencies you included in `install_requires=[]` in your setup.py and install `myscript` as a command-line module:



```
[dkennetz ~]$ myscript

```

"
"Redis - Default blocking VM

> 
> The blocking VM performance is better overall, as there is no time lost in
>  synchronization, spawning of threads, and resuming blocked
>  clients waiting for values. So if you are willing to accept an higher
>  latency from time to time, blocking VM can be a good pick. Especially
>  if swapping happens rarely and most of your often accessed data
>  happens to fit in your memory.
> 
> 
> 


This is default mode of Redis (and the only mode going forward I believe now VM is deprecated in 2.6), leaving the OS to handle paging (if/when required). I am correct in my understanding that it will take some time to get ""hot"" when booted/started. When working on a 1gb RAM node with a 16gb dataset, does Redis attempt to load it all into virtual memory at boot and thus 90%+ is immediately paged out, and only after some good amount of usages does the above statement hold true? 


","Redis VM was already deprecated in Redis 2.4, and has been removed in Redis 2.6. It is a dead end: don't use it.


I think you are confusing the blocking VM with OS paging. They are two different things.


OS paging is the default mode of Redis when Redis VM is not configured at all (whatever the blocking mode). The OS will swap Redis memory if it does not fit in physical memory. The event loop can be frozen at any time. When it happens, performance is abysmal because none of the Redis internal data structures is designed for this (no locality, no paging system).


Redis VM can be configured in non blocking mode (using I/O threads). When I/Os are done, the event loop is not blocked, and Redis is still responsive. However, when too many I/Os pile up, the I/O threads will be completely busy, and you end up with a responsive Redis, but unable to process any queries requiring I/Os.


Redis VM can also be configured in blocking mode. In this mode all I/Os are synchronously performed in the main event loop thread. So the event loop is frozen in case of I/O (for instance in case of a key miss). All clients are impacted. However, general performance (CPU consumption and latency) is better than with the non blocking mode because some threading scheduling/synchronization is saved.


In practice, the difference between OS paging and the Redis blocking VM is the granularity level. With Redis VM, the granularity is the key. With OS paging, well it is the page (a 4 KB block which can span on several unrelated keys).


In all 3 cases, the initial load of the dump file will be extremely slow and generate a peak of random I/Os on your system. As you pointed out, most objects will be loaded and then swapped out. The warm-up time will be significant.


Except if you have extreme locality in your data, or if you do not care at all about the latencies, using 1 GB RAM for a 16 GB dataset with the Redis VM is science-fiction IMO. 


There is a reason why the Redis VM was phased out. By design, it will never perform as well as a disk-based datastore (which can exploit file mapping or direct I/Os to avoid the double buffering, and use adapted data structures like B-trees).


Redis as an in-memory store is excellent. But if you need to store something which is bigger than RAM, don't use it. Other (disk-based) stores will all perform much better.


"
"What are namespaces for ? what about usages?
1. what is the purpose of namespaces ?
2. and, more important, should they be used as objects in java (things that have data and functions and that try to achieve encapsulation) ? is this idea to far fetched ? :)
3. or should they be used as packages in java ?
4. or should they be used more generally as a module system or something ?


","Given that you use the Clojure tag, I suppose that you'll be interested in a Clojure-specific answer:



> 
> what is the purpose of namespaces ?
> 
> 
> 


Clojure namespaces, Java packages, Haskell / Python / whatever modules... At a very high level, they're all different names for the same basic mechanism whose primary purpose is to prevent name clashes in non-trivial codebases. Of course, each solution has its own little twists and quirks which make sense in the context of a given language and would not make sense outside of it. The rest of this answer will deal with the twists and quirks specific to Clojure.


**A Clojure namespace groups Vars**, which are containers holding functions (most often), macro functions (functions used by the compiler to generate macroexpansions of appropriate forms, normally defined with `defmacro`; actually they are just regular Clojure functions, although there is some magic to the way in which they are registered with the compiler) and occasionally various ""global parameters"" (say, `clojure.core/*in*` for standard input), Atoms / Refs etc. The protocol facility introduced in Clojure 1.2 has the nice property that protocols are backed by Vars, as are the individual protocol functions; this is key to the way in which protocols present a solution to the expression problem (which is however probably out of the scope of this answer!).


It stands to reason that namespaces should group Vars which are somehow related. In general, creating a namespace is a quick & cheap operation, so it is perfectly fine (and indeed usual) to use a single namespace in early stages of development, then as independent chunks of functionality emerge, factor those out into their own namespaces, rinse & repeat... Only the things which are part of the public API need to be distributed between namespaces up front (or rather: prior to a stable release), since the fact that function such-and-such resides in namespace so-and-so is of course a part of the API.



> 
> and, more important, should they be used as objects in java (things that have data and functions and that try to achieve encapsulation) ? is this idea to far fetched ? :)
> 
> 
> 


Normally, the answer is no. You might get a picture not too far from the truth if you approach them as classes with lots of static methods, no instance methods, no public constructors and often no state (though occasionally there may be some ""class data members"" in the form of Vars holding Atoms / Refs); but arguably it may be more useful not to try to apply Java-ish metaphors to Clojure idioms and to approach a namespace as a group of functions etc. and not ""a class holding a group of functions"" or some such thing.


There is an important exception to this general rule: namespaces which include `:gen-class` in their `ns` form. These are meant precisely to implement a Java class which may later be instantiated, which might have instance methods and per-instance state etc. Note that `:gen-class` is an interop feature -- pure Clojure code should generally avoid it.



> 
> or should they be used as packages in java ?
> 
> 
> 


They serve some of the same purposes packages were designed to serve (as already mentioned above); the analogy, although it's certainly there, is not that useful, however, just because the things which packages group together (Java classes) are not at all like the things which Clojure namespaces group together (Clojure Vars), the various ""access levels"" (`private` / `package` / `public` in Java, `{:private true}` or not in Clojure) work very differently etc.


That being said, one has to remember that there is a certain correspondence between namespaces and packages / classes residing in particular packages. A namespace called `foo.bar`, when compiled, produces a class called `bar` in the package `foo`; this means, in particular, that namespace names should contain at least one dot, as so-called single-segment names apparently lead to classes being put in the ""default package"", leading to all sorts of weirdness. (E.g. I find it impossible to have VisualVM's profiler notice any functions defined in single-segment namespaces.)


Also, **`deftype` / `defrecord`-created types do not reside in namespaces**. A `(defrecord Foo [...] ...)` form in the file where namespace `foo.bar` is defined creates a class called `Foo` in the package `foo.bar`. To use the type `Foo` from another namespace, one would have to `:import` the class `Foo` from the `foo.bar` package -- `:use` / `:require` would not work, since they pull in Vars from namespaces, which records / types are not.


So, in this particular case, there is a certain correspondence between namespaces and packages which Clojure programmers who wish to take advantage of some of the newer language features need to be aware of. Some find that this gives an ""interop flavour"" to features which are not otherwise considered to belong in the realm of interop (`defrecord` / `deftype` / `defprotocol` are a good abstraction mechanism even if we forget about their role in achieving platform speed on the JVM) and it is certainly possible that in some future version of Clojure this flavour might be done away with, so that the namespace name / package name correspondence for `deftype` & Co. can be treated as an implementation detail.



> 
> or should they be used more generally as a module system or something ?
> 
> 
> 


They *are* a module system and this is indeed how they should be used.


"
"BackChannelLogoutUri with multi-tenant scenario
I am currently working with Identity server 4, where i am trying to enable [BackChannelLogoutUri](https://github.com/IdentityServer/IdentityServer4/tree/master/samples/Clients/src/MvcHybridBackChannel).


Each client has been given a BackChannelLogoutUri in the config of the client



```
BackChannelLogoutUri = ""http://localhost:44322/home/LogoutBackChannel"",

```

Each client application has registered the cookieEventHandler and LogoutSessionManager.



```
services.AddTransient<CookieEventHandler>();
        services.AddSingleton<LogoutSessionManager>();

        services.AddAuthentication(options =>
        {
            options.DefaultScheme = CookieAuthenticationDefaults.AuthenticationScheme;
            options.DefaultChallengeScheme = ""oidc"";
        })
            .AddCookie(options =>
            {
                options.ExpireTimeSpan = TimeSpan.FromMinutes(60);
                options.Cookie.Name = ""mvchybridbc"";

                options.EventsType = typeof(CookieEventHandler);
            })

```

My logout view on the identity server contains the Iframe



```
@if (Model.PostLogoutRedirectUri != null)
{
    <div>
        Click <a class=""PostLogoutRedirectUri"" href=""@Model.PostLogoutRedirectUri"">here</a> to return to the
        <span>@Model.ClientName</span> application.
    </div>
}
@if (Model.SignOutIframeUrl != null)
{
    <iframe width=""0"" height=""0"" class=""signout"" src=""@Model.SignOutIframeUrl""></iframe>
}

```

This is all well and good. But my problem is that the BackChannelLogoutUri is a single url. When hosted it will need to be passed some how from each tennent


- ""<http://one.mysite.com/app1/home/LogoutBackChannel>""
- ""<http://one.mysite.com/app2/home/LogoutBackChannel>""
- ""<http://two.mysite.com/app1/home/LogoutBackChannel>""
- ""<http://three.mysite.com/app3/home/LogoutBackChannel>""


We cant really have a client for each tenant and app. That would be a lot of clients. That and clients that are only users of tenant one would not need to be logged out of tenant two. 


I am not sure how to address this issue.


","I've implemented the backchannel logout without having to rely on iframes. What it basically does is, collect the necessary urls and then send the notifications.


I don't have tenants, so this will work for me. But you can adapt the code and add the logic for tenants, as commented in the code:



```
// Injected services:

//private readonly IUserSession _userSession;
//private readonly IClientStore _clientStore;
//private readonly IBackChannelLogoutService _backChannelClient;

private async Task LogoutUserAsync(string logoutId)
{
    if (User?.Identity.IsAuthenticated == true)
    {
        // delete local authentication cookie
        await HttpContext.SignOutAsync();

        // Get all clients from the user's session
        var clientIds = await _userSession.GetClientListAsync();
        if (clientIds.Any())
        {
            var backChannelClients = new List<BackChannelLogoutModel>();
            var sessionId = await _userSession.GetSessionIdAsync();
            var sub = User.Identity.GetSubjectId();

            foreach (var clientId in clientIds)
            {
                var client = await _clientStore.FindEnabledClientByIdAsync(clientId);
                // This should be valid in any case:
                if (client == null && !string.IsNullOrEmpty(client.BackChannelLogoutUri))
                    continue;

                // Insert here the logic to retrieve the tenant url for this client
                // and replace the uri:
                var tenantLogoutUri = client.BackChannelLogoutUri;

                backChannelClients.Add(new BackChannelLogoutModel
                {
                    ClientId = client.ClientId,
                    LogoutUri = tenantLogoutUri,
                    SubjectId = sub,
                    SessionId = sessionId,
                    SessionIdRequired = true
                });
            }

            try
            {
                await _backChannelClient.SendLogoutNotificationsAsync(backChannelClients);
            }
            catch (Exception ex)
            {
                // Log message
            }
        }

        // raise the logout event
        await _events.RaiseAsync(new UserLogoutSuccessEvent(User.GetSubjectId(), User.GetDisplayName()));
    }
}

```

"
"How does copy and paste for large files work
I am curious to know how computers execute ""copy"" and ""paste"" of large folders. I've read that copy and paste of text between different processes or same process is achieved by saving the content into RAM and then copying it from there to destined location. 


So, how does computer instructions flow while copying a folder of say 10 GB on a machine which has 2 GB RAM and 4 GB max virtual memory. Is file copy is different from text copy.


I think it's basic question but any links or insights appreciated.


","Clipboard doesn't have to hold entire file. When you copy a file (or files), only its path is put into the clipboard. It's also marked as a file - clipboard keeps track of its content's type, like plain text, formatted text, file, image, Word text etc. This is why you can't for example open up an image in Paint, press `Ctrl`+`C` and then paste it into a directory - because you have copied a picture, and directories hold files, not pictures.


When you paste a compatible content (i.e. file(s) and/or folder(s)) into a directory, some application will handle the copying/moving operation. By default it will be the `explorer` process (the same one that's responsible for displaying Start menu and all file explorer windows), but some apps may replace it.[1]


What happens now depends on what you're doing:


- If you're **moving a file to another directory on the same partition**, it won't be physically moved on the disk, only its path will be updated[2].
- If you're **moving a file to another partition**, it will be split into chunks of the same size[3] and those will be copied one by one, then the original file will be deleted. Too small chunks will slow down the process, too big chunks will consume more memory.
- Exactly the same will happen when you're **copying a file** (no matter if it's the same partition or not), except that the original file won't be deleted.
- **Writing to an external storage** (like USB drives) doesn't work exactly like that[4] and I'm not sure what exactly happens then. My guess is that it's not a continuous chunk-by-chunk process, but something else happens every few chunks (buffer-related?). If anybody knows something about this, then feel free to edit.




---


**Annotations:**


[1]. For example [TeraCopy](http://codesector.com/teracopy), which is a nice advanced copy window replacement.


[2]. Physical file structure on the hard disk doesn't resemble the directory structure - it's flat and all hierarchy information is stored in a separate part of the partition. How it's exactly done depends on filesystem (for example see [MFT](http://codesector.com/teracopy)). That information block holds all the information about file locations etc. So when you move a file inside one partition, there's no need to move it physically - only the path information has to be updated.


[3]. Not literally, nothing will be split on the hard disk. The program that handles the copying process will work like there are multiple separate chunks, but the original file will be untouched. It's purely virtual.


[4]. You can see it when using TeraCopy: the ""predicted progress"" doesn't work like on fixed drives, instead the ""real progress"" catches up with it, then ""predicted progress"" is expanded and so on. Explorer's default copy window has USB hiccups too.


"
"Creating unitary tests in Spring 3
I'm starting on testing applications in general and I want to create several tests to learn Mockito in Spring. I've been reading several information but I have some general doubts I'd like to ask.


1. I have seen come Mockito tests and they annotate the test of the class with: **@RunWith(MockitoJUnitRunner.class)** while in the Spring documentation it is used **@RunWith(SpringJUnit4ClassRunner.class)**. I don't know what's the difference between them and which one should I use for a Spring application where tests use Mockito.
2. As I haven't seen any real application that has test I'd like to know typical test that a developer would do. For example in a typical CRUD application for users (users can be created, updated...) can anyone a usual test that it would be done.


Thanks.


","
```
@RunWith(MockitoJUnitRunner.class)

```

With this declaration you are suppose to write a **unit test**. Unit tests are exercising a single class mocking all dependencies. Typically you will inject mocked dependencies declared like this in your test case:



```
@Mock
private YourDependency yourDependencyMock;

```



---



```
@RunWith(SpringJUnit4ClassRunner.class)

```

Spring runner is meant for **integration test** (*component test*?) In this type of tests you are exercising a whole bunch of classes, in other words you are testing a single class with real dependencies (testing a controller with real services, DAOs, in-memory database, etc.)


You should probably have both categories in your application. Althought it is advices to have more unit tests and only few smoke integration tests, but I often found myself more confident writing almost only integration tests.


As for your second question, you should have:


- **unit tests** for each class (controller, services, DAOs) separately with mocked all other classes
- **integration tests** for a whole single CRUD operation. For instance creating a user that exercises controller, service, DAO and in-memory database.


"
"Create or update a section in XML file with Augeas
I would like to update a section in an XML config file or add a new one if does not exist already using Augeas.


The XML file looks like this:



```
<?xml version=""1.0""?>
<security>
          <users>
            ...
            <user>
              <id>deployment</id>
              <firstName>Deployment</firstName>
              <lastName>User</lastName>
              <password>somepasshere</password>
              <status>active</status>
              <email>changeme1@yourcompany.com</email>
            </user>
          </users>
</security>

```

I would like to update the last name/first name/email if the ID exists already or add a new user section if it's a new ID.


In AugTool I use:



```
augtool> set /augeas/load/Xml/lens Xml.lns
augtool> set /augeas/load/Xml/incl /security.xml
augtool> load

```

I'm still learning Augeas, so this was my first try to get the node :



```
augtool> print /files/security.xml/security/users/user/*[ #text= 'deployment']

```

What would be the command to update or create a new section user in users ?


thank you!


","First, with recent Augeas versions (>= 1.0.0), you can use `--transform` to set up the transformation. Let's say the file is `./users.xml`:



```
$ augtool -r . --noautoload --transform ""Xml.lns incl /users.xml""
augtool> defnode user /files/users.xml/security/users/user[id/#text=""deployment""] # Create a new user entry if it doesn't exist yet, assign node to the ""user"" variable
augtool> set $user/id/#text ""deployment""         # Set id if node was just created
augtool> set $user/firstName/#text ""Deployment""  # Set firstName
augtool> set $user/lastName/#text ""User""         # Set lastName
augtool> set $user/email/#text ""changeme1@yourcompany.com""  # set email
...
augtool> save
Saved 1 file(s)

```

You can even turn this into a script, say `user.augtool`:



```
#!/usr/bin/augtool -sf
defnode user /files/users.xml/security/users/user[id/#text=""deployment""]
set $user/id/#text ""deployment""
set $user/firstName/#text ""Deployment""
set $user/lastName/#text ""User""
set $user/email/#text ""changeme1@yourcompany.com""

```

which you can then launch:



```
$ chmod +x user.augtool
$ ./user.augtool --transform ""Xml.lns incl /users.xml"" -r .
Saved 1 file(s)

```

"
"Regular expression quoting in Python
How should I declare a regular expression?



```
mergedData = re.sub(r'\$(.*?)\$', readFile, allData)

```

I'm kind of wondering why this worked. I thought that I need to use the `r''` to pass a regular expression.



```
mergedData = re.sub(""\$(.*?)\$"", readFile, allData)

```

What does `""\$""` result in in this case? Why? I would have thought `""$""`.


","
> 
> I thought that I need to user the r'' to pass a regular expression.
> 
> 
> 


`r` before a string literal indicates raw string, which means the usual escape sequences such as `\n` or `\r` are no longer treated as new line character or carriage return, but simply `\` followed by `n` or `r`. To specify a `\`, you only need `\` in raw string literal, while you need to double it up `\\` in normal string literal. This is why it is usually the case that raw string is used in specifying regular expression1. It reduces the confusion when reading the code. You would have to do escaping twice if you use normal string literal: once for the normal string literal escape and the second time for the escaping in regex.



> 
> What does `""\$""` result in this case? Why? I would have thought `""$""`
> 
> 
> 


In Python normal string literal, if `\` is not followed by an escape sequence, the `\` is preserved. Therefore `""\$""` results in `\` followed by `$`.


This behavior is slightly different from the way C/C++ or JavaScript handle similar situation: the `\` is considered escape for the next character, and only the next character remains. So `""\$""` in those languages will be interpreted as `$`.


**Footnote**


1: There is a small defect with the design of raw string in Python, though: [Why can't Python's raw string literals end with a single backslash?](https://stackoverflow.com/questions/647769/why-cant-pythons-raw-string-literals-end-with-a-single-backslash)


"
"Copy data from numpy to ctypes quickly
I have a ctypes object shared between a c++ and python process. The python process takes the input values from that object, runs them through Tensorflow, and I'm left with a numpy array as output. As these arrays are pretty big I'm wondering if there is a better approach to copying the data from the output of tensorflow back to the shared ctypes object so the c++ process can act on them. (speed is the issue, yes.)


Right now I'm copying each value one by one:



```
output = np.array([12, 13, 11, 10]) # in reality this is fairly large (the Tensorflow result)
for i, value in enumerate(output):
    data.pressure[i] = ctypes.c_double(value)

```

where data is the ctypes object shared in memory. (structured after [this example](https://github.com/teeks99/py_boost_shmem))


On the other hand, copying data from the ctypes object into numpy is easy, I'm wondering if there is something that does the opposite (from numpy to the ctypes array) Here is that easy code:



```
# Creating a numpy array from the ctypes array
input = np.reshape(data.velocity, (1, 29791))
# Tensorflow returns a numpy array
output = sess.run(final, feed_dict={Input: input}) 
# Now how do I get the data from output into data.pressure?

```

Edit: For reference, this is how the ctypes looks (python side)



```
class TransferData(ctypes.Structure):
    _fields_ = [
        ('statusReady', ctypes.c_bool),
        # ...
        ('velocity', ctypes.c_double * (31 * 31 * 31)),
        ('pressure', ctypes.c_double * (31 * 31 * 31))
    ]

```

","This shows how to copy a whole data block from numpy array to ctypes array:



```
import numpy as np
import ctypes


# Preparing example

class TransferData(ctypes.Structure):
    _fields_ = [
        ('statusReady', ctypes.c_bool),
        ('velocity', ctypes.c_double * 4),
        ('pressure', ctypes.c_double * 4)
    ]


data = TransferData()
output = np.array([12., 13., 11., 10.])


# Actual code

# Both values should be equal but there could be problems with alignment settings
assert ctypes.sizeof(data.pressure) == output.nbytes

ctypes.memmove(ctypes.byref(data.pressure), output.ctypes.data, output.nbytes)


print(list(data.pressure))

```

"
"I've been trying to reconfigure my SSH key
I keep getting the following error:



```
 git commit -sam ""blah blah blah""
fatal: either user.signingkey or gpg.ssh.defaultKeyCommand needs to be configured


```

I just updated Git completely trying to figure this out, so it's completely up to date, and then I successfully added all of my usual configurations to the new version now running in Git Bash. I've repeatedly gone to Github's SSH key generator and followed the directions one by one. AND AFTER ALL THAT, I'M STILL GETTING THIS ERROR.


I am unable to make commits ANYWHERE on my local machine (Git Bash, Terminal, GitKraken; I even broke down and tried Github Desktop), much less signed, annotated commits, as is my habit. I removed the expired keys from my Github account. I DON'T KNOW WHAT ELSE TO DO.


[![I'm about to lose my mind.](https://i.stack.imgur.com/10n5v.png)](https://i.stack.imgur.com/10n5v.png)


","You're mixing together the concepts of ssh keys (which are somewhat generalized and apply across all of ssh, and which you can use to authenticate yourself to GitHub) and Git's *signed commits* (and signed annotated tags). These are different, although they use related mechanisms.


In particular, to use an ssh key to *sign a Git commit*, you must:


- configure your ssh locally so that it can sign commits (this may or may not already be supported depending on your OpenSSH version);
- tell Git *how* to use your ssh to sign commits (this depends on your ssh version); and
- tell Git *to* use ssh to sign commits.


None of these three steps use or require anything on *GitHub*. But this is what is failing here: you have not set up `user.signingkey` or `gpg.ssh.defaultKeyCommand` in Git, which is where that second bullet point comes in. (You're already doing the third one, but Git doesn't know *how* to run your ssh yet!) You'll need to figure out how to get Git to invoke the right commands on your system (which will depend somewhat on your OS and OpenSSH version).


Once you have such signed commits, however, these digital signatures are useful only to you, not to anyone else, *unless* you have spread the key(s) involved in these digital signatures. This is where you get GitHub involved.


See [How do I sign git commits using my existing ssh key](https://stackoverflow.com/q/72844616/1256452) and particularly VonC's answer [here](https://stackoverflow.com/a/45120525/1256452) to [Why does git sign with GPG keys rather than using SSH keys?](https://stackoverflow.com/q/45119932/1256452), to see how to do the first part. See both VonC's and other answers, particularly [Jakuje's here](https://stackoverflow.com/a/45126508/1256452), for some cautions involving using ssh keys here.


I don't know any of the GitHub side details here, but VonC's answers have more.


In general, it's a lot easier to use GPG for signing commits and/or annotated tags.


"
"How to know/change current directory in Python shell?
I am using Python 3.2 on Windows 7. When I open the Python shell, how can I know what the current directory is? How can I change it to another directory (where my modules are)?


","You can use the `os` module.



```
>>> import os
>>> os.getcwd()
'/home/user'
>>> os.chdir(""/tmp/"")
>>> os.getcwd()
'/tmp'

```

But if it's about finding other modules: You can set an environment variable called `PYTHONPATH`, under Linux would be like



```
export PYTHONPATH=/path/to/my/library:$PYTHONPATH

```

Then, the interpreter searches also at this place for `import`ed modules. I guess the name would be the same under Windows, but don't know how to change.


**edit**


Under Windows:



```
set PYTHONPATH=%PYTHONPATH%;C:\My_python_lib

```

(taken from <http://docs.python.org/using/windows.html>)


**edit 2**


... and even better: use `virtualenv` and `virtualenv_wrapper`, this will allow you to create a development environment where you can add module paths as you like (`add2virtualenv`) without polluting your installation or ""normal"" working environment.


<http://virtualenvwrapper.readthedocs.org/en/latest/command_ref.html>


"
"How to copy DLL files into the same folder as the executable using CMake?
We use CMake for generating the Visual Studio files of our sources in our SVN. Now my tool requires some DLL files to be in the same folder as the executable. The DLL files are in a folder alongside the source.


How can I change my `CMakeLists.txt` such that the generated Visual Studio project will either have already the particular DLL files in the release/debug folders or will copy them upon compilation?


","I'd use `add_custom_command` to achieve this along with `cmake -E copy_if_different...`. For full info run



```
cmake --help-command add_custom_command
cmake -E

```

  



So in your case, if you have the following directory structure:



```
/CMakeLists.txt
/src
/libs/test.dll

```

and your CMake target to which the command applies is `MyTest`, then you could add the following to your CMakeLists.txt:



```
add_custom_command(TARGET MyTest POST_BUILD        # Adds a post-build event to MyTest
    COMMAND ${CMAKE_COMMAND} -E copy_if_different  # which executes ""cmake - E copy_if_different...""
        ""${PROJECT_SOURCE_DIR}/libs/test.dll""      # <--this is in-file
        $<TARGET_FILE_DIR:MyTest>)                 # <--this is out-file path

```

  



If you just want the entire contents of the `/libs/` directory copied, use `cmake -E copy_directory`:



```
add_custom_command(TARGET MyTest POST_BUILD
    COMMAND ${CMAKE_COMMAND} -E copy_directory
        ""${PROJECT_SOURCE_DIR}/libs""
        $<TARGET_FILE_DIR:MyTest>)

```

  



If you need to copy different dlls depending upon the configuration (Release, Debug, eg) then you could have these in subdirectories named with the corresponding configuration: `/libs/Release`, and `/libs/Debug`. You then need to inject the configuration type into the path to the dll in the `add_custom_command` call, like this:



```
add_custom_command(TARGET MyTest POST_BUILD
    COMMAND ${CMAKE_COMMAND} -E copy_directory
        ""${PROJECT_SOURCE_DIR}/libs/$<CONFIGURATION>""
        $<TARGET_FILE_DIR:MyTest>)

```

"
"is there any way to get samples under each leaf of a decision tree?
I have trained a decision tree using a dataset. Now I want to see which samples fall under which leaf of the tree. 


From here I want the red circled samples.


[![enter image description here](https://i.stack.imgur.com/DYhwf.png)](https://i.stack.imgur.com/DYhwf.png)


I am using Python's Sklearn's implementation of decision tree . 


","If you want only the leaf for each sample you can just use



```
clf.apply(iris.data)

```


> 
> array([ 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1,
>  1,
>  1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1,
>  1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 1, 5,
>  5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5,
>  5, 5, 14, 5, 5, 5, 5, 5, 5, 10, 5, 5, 5, 5, 5, 10, 5,
>  5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 5, 16, 16,
>  16, 16, 16, 16, 6, 16, 16, 16, 16, 16, 16, 16, 16, 16, 16, 16, 16,
>  8, 16, 16, 16, 16, 16, 16, 15, 16, 16, 11, 16, 16, 16, 8, 8, 16,
>  16, 16, 15, 16, 16, 16, 16, 16, 16, 16, 16, 16, 16, 16])
> 
> 
> 


If you want to get all samples for each node you could calculate all the decision paths with



```
dec_paths = clf.decision_path(iris.data)

```

Then loop over the decision paths, convert them to arrays with `toarray()` and check whether they belong to a node or not. Everything is stored in a `defaultdict` where the key is the node number and the values are the sample number.



```
for d, dec in enumerate(dec_paths):
    for i in range(clf.tree_.node_count):
        if dec.toarray()[0][i] == 1:
            samples[i].append(d)

```

**Complete code**



```
import sklearn.datasets
import sklearn.tree
import collections

clf = sklearn.tree.DecisionTreeClassifier(random_state=42)
iris = sklearn.datasets.load_iris()
clf = clf.fit(iris.data, iris.target)

samples = collections.defaultdict(list)
dec_paths = clf.decision_path(iris.data)

for d, dec in enumerate(dec_paths):
    for i in range(clf.tree_.node_count):
        if dec.toarray()[0][i] == 1:
            samples[i].append(d) 

```

**Output**



```
print(samples[13])

```


> 
> [70, 126, 138]
> 
> 
> 


"
"incomprehensible performance improvement with openmp even when num\_threads(1)
The following lines of code



```
int nrows = 4096;
int ncols = 4096;
size_t numel = nrows * ncols;
unsigned char *buff = (unsigned char *) malloc( numel );

unsigned char *pbuff = buff;
#pragma omp parallel for schedule(static), firstprivate(pbuff, nrows, ncols), num_threads(1)
for (int i=0; i<nrows; i++)
{
    for (int j=0; j<ncols; j++)
    {
        *pbuff += 1;
        pbuff++;
    }
}

```

take 11130 usecs to run on my i5-3230M when compiled with



```
g++ -o main main.cpp -std=c++0x -O3

```

That is, when the openmp pragmas are ignored.


On the other hand, it only takes 1496 usecs when compiled with



```
g++ -o main main.cpp -std=c++0x -O3 -fopenmp

```

This is more than 6 times faster, which is quite surprising taking into acount that it is run on a 2-core machine. In fact, I have also tested it with *num\_threads(1)* and the performance improvement is still quite important (more than 3 times faster).


Anybody can help me to understand this behaviour?


EDIT: following the suggestions, I provide the full piece of code:



```
#include <stdlib.h>
#include <iostream>

#include <chrono>
#include <cassert>


int nrows = 4096;
int ncols = 4096;
size_t numel = nrows * ncols;
unsigned char * buff;


void func()
{
    unsigned char *pbuff = buff;
    #pragma omp parallel for schedule(static), firstprivate(pbuff, nrows, ncols), num_threads(1)
    for (int i=0; i<nrows; i++)
    {
        for (int j=0; j<ncols; j++)
        {
            *pbuff += 1;
            pbuff++;
        }
    }
}


int main()
{
    // alloc & initializacion
    buff = (unsigned char *) malloc( numel );
    assert(buff != NULL);
    for(int k=0; k<numel; k++)
        buff[k] = 0;

    //
    std::chrono::high_resolution_clock::time_point begin;
    std::chrono::high_resolution_clock::time_point end;
    begin = std::chrono::high_resolution_clock::now();      
    //
    for(int k=0; k<100; k++)
        func();
    //
    end = std::chrono::high_resolution_clock::now();
    auto usec = std::chrono::duration_cast<std::chrono::microseconds>(end-begin).count();
    std::cout << ""func average running time: "" << usec/100 << "" usecs"" << std::endl;

    return 0;
}

```

","The answer, as it turns out, is that `firstprivate(pbuff, nrows, ncols)` effectively declares `pbuff`, `nrows` and `ncols` as local variables within the scope of the for loop. That in turn means the compiler can see `nrows` and `ncols` as constants - it cannot make the same assumption about global variables!


Consequently, with `-fopenmp`, you end up with the huge speedup because *you aren't accessing a global variable each iteration*. (Plus, with a constant `ncols` value, the compiler gets to do a bit of loop unrolling).


By changing



```
int nrows = 4096;
int ncols = 4096;

```

to



```
const int nrows = 4096;
const int ncols = 4096;

```

**or** by changing



```
for (int i=0; i<nrows; i++)
{
    for (int j=0; j<ncols; j++)
    {
        *pbuff += 1;
        pbuff++;
    }
}

```

to



```
int _nrows = nrows;
int _ncols = ncols;
for (int i=0; i<_nrows; i++)
{
    for (int j=0; j<_ncols; j++)
    {
        *pbuff += 1;
        pbuff++;
    }
}

```

the anomalous speedup vanishes - the non-OpenMP code is now just as fast as the OpenMP code.


The moral of the story? Avoid accessing mutable global variables inside performance-critical loops.


"
"Why is FastCGI fast?

> 
> FastCGI provides a way to improve the
>  performance of the thousands of Perl
>  applications that have been written
>  for the Web. -[Source](http://www.fastcgi.com/drupal/)
> 
> 
> 


and how does it do that?


","Mark R. Brown's [whitepaper on the subject](http://www.fastcgi.com/drupal/node/6?q=node/16#S3) claims that one of the primary benefits of FastCGI is that different requests can share a single cache, making caching practical:



> 
> Today's most widely deployed Web server APIs are based on a pool-of-processes server model. The Web server consists of a parent process and a pool of child processes. Processes do not share memory. An incoming request is assigned to an idle child at random. The child runs the request to completion before accepting a new request. A typical server has 32 child processes, a large server has 100 or 200.
> 
> 
> In-memory caching works very poorly in this server model because processes do not share memory and incoming requests are assigned to processes at random. For instance, to keep a frequently-used file available in memory the server must keep a file copy per child, which wastes memory. When the file is modified all the children need to be notified, which is complex (the APIs don't provide a way to do it).
> 
> 
> FastCGI is designed to allow effective in-memory caching. Requests are routed from any child process to a FastCGI application server. The FastCGI application process maintains an in-memory cache.
> 
> 
> 


"
"Set href in attribute directive in Angular
I'm trying to set up a Bootstrap tab strip with Angular 2. I have the tabs rendering in an `ngFor` but I'm getting template errors when I try to put the `#` infront of the `href` expression. So this template compiles but isn't what I want:



```
<ul class=""nav nav-tabs"" role=""tablist"">
  <li *ngFor=""let aType of resourceTypes; let i = index""
      [ngClass]=""{'active': i == 0}""
      role=""presentation"">
    <a [attr.href]=""aType.Name""
       [attr.aria-controls]=""aType.Name""
       role=""tab""
       data-toggle=""tab"">
      {{aType.Name}}
    </a>
  </li>
</ul>

```

What I want to do is `[attr.href]=""#aType.Name""`
but that blows up. What is the correct syntax to prepend the `#` in front of the expression in the attribute directive?


","There is no need to prefix with `#`


In this code



```
<ul class=""nav nav-tabs"" role=""tablist"">
  <li *ngFor=""let aType of resourceTypes; let i = index""
      [ngClass]=""{'active': i == 0}""
      role=""presentation"">
    <a [attr.href]=""aType.Name""
       [attr.aria-controls]=""aType.Name""
       role=""tab""
       data-toggle=""tab"">
      {{aType.Name}}
    </a>
  </li>

```

`aType` already refers to the `*ngFor` variable.


If you want to prefix a literal `#` you can use 



```
[attr.href]=""'#' + aType.Name"" 

```

There is also no need for `attr.` if the element actually has the property `href` which is the case with `<a>`.


"
"Does a wildcard inside double-quotes glob?
On a ""standard BASH"" does a wildcard inside double-quotes glob? For example:



```
$ touch abc
$ ls ""*abc*""

```

would that, or wouldn't that work on bash? 


I was told Ubuntu shipped with a bash variant that doesn't conform to POSIX or BASH. Is that true?


","Short answer: no


Long answer from man bash:


Enclosing characters in double quotes preserves the literal value of all characters within the quotes, with the exception of $, `, \


So:



```
~$ touch sas
~$ ls *a*
sas
~$ ls ""*a*""
ls: cannot access *a*: No such file or directory

```

Of course when talking about double quotes and asterisks there is another exception: parameter expansion (ie: ""$\*"") ...but that's another history 


For the second question: are you asking about dash? dash is a lightweight shell for scripts. In debian, and in ubuntu per extension (as is debian based), is aliased to sh; It's major advantage over bash is its speed, so it is used by default for system scripts; however bash is still available as the interactive shell for users by default



```
 ~$ ls -l /bin/sh 
lrwxrwxrwx 1 root root 4 2011-04-21 11:54 /bin/sh -> dash

```

"
"ionic3 multiple swiper sliders on samepage
I am using ionic 3 while placing 3 swiper sliders on same page i am unable to control their behavior separately i am bit new for angular js as well ts and ionic 


here is code



```
<ion-slides style=""height: 30%;"" pager>
  <ion-slide style=""background-color: green"" *ngFor=""let post of posts"">
    <h1>{{post.title}}</h1>
    <img [src]=""post.image"" />
  </ion-slide>
</ion-slides>


<ion-slides style=""height: 30%;"" pager>
  <ion-slide style=""background-color: green"" *ngFor=""let post of posts"">
    <h1>{{post.title}}</h1>
    <img [src]=""post.image"" />
  </ion-slide>
</ion-slides>

```

ts is 



```
 @ViewChild(Slides) slides: Slides;

  goToSlide() {
    this.slides.slideTo(2, 500);

  }

   ngAfterViewInit() {
    //this.slides.freeMode = true;
    this.slides .slidesPerView=3,
    this.slides.spaceBetween=5;
    //this.slides.loop=true;
  }

```

","Instead of `ViewChild`, you can use `ViewChildren` to get all the instances of the slider in that page ([Angular docs](https://angular.io/api/core/ViewChildren)). Please take a look at **[this working plunker](http://plnkr.co/edit/L1eFQjY0v9TxFXS6ga7e)**. The end result would be something like this:


[![Slider demo](https://i.stack.imgur.com/2Xfng.gif)](https://i.stack.imgur.com/2Xfng.gif)


Like you can see in the plunker, we get all the instances of the slider, and then we just get the target instance by using its index:



```
import { Component, ViewChildren, QueryList } from '@angular/core';
import { NavController, Content, Slides } from 'ionic-angular';

@Component({...})
export class HomePage {
  @ViewChildren(Slides) slides: QueryList<Slides>;

  constructor() {}

  public move(index: number): void {
    this.slides.toArray()[index].slideNext(500);
  }      

}

```

An then in the view:



```
<ion-header>
  <ion-navbar>
    <ion-title>Ionic Demo</ion-title>
  </ion-navbar>
</ion-header>

<ion-content padding>

  <ion-slides style=""height: 75px"" pager>
    <ion-slide><h1>Slide 1</h1></ion-slide>
    <ion-slide><h1>Slide 2</h1></ion-slide>
    <ion-slide><h1>Slide 3</h1></ion-slide>
  </ion-slides>

  <ion-slides style=""height: 75px"" pager>
    <ion-slide><h1>Slide 1</h1></ion-slide>
    <ion-slide><h1>Slide 2</h1></ion-slide>
    <ion-slide><h1>Slide 3</h1></ion-slide>
  </ion-slides>

  <ion-slides style=""height: 75px"" pager>
    <ion-slide><h1>Slide 1</h1></ion-slide>
    <ion-slide><h1>Slide 2</h1></ion-slide>
    <ion-slide><h1>Slide 3</h1></ion-slide>
  </ion-slides>

  <ion-row>
    <ion-col>
      <button (click)=""move(0)"" ion-button text-only>Move First</button>
    </ion-col>
    <ion-col>
      <button (click)=""move(1)"" ion-button text-only>Move Second</button>
    </ion-col>
    <ion-col>
      <button (click)=""move(2)"" ion-button text-only>Move Third</button>
    </ion-col>
  </ion-row>

</ion-content>

```

"
"How to hide a badge when the value is empty or zero
I would like my PrimeFaces [`p:badge`](https://www.primefaces.org/showcase/ui/misc/badge.xhtml) not to be visible when the value is zero or empty.
The following will display a badge with the value `0`:



```
<p:badge value=""0"">
  <p:avatar label=""Badge""/>
</p:badge>

```

And if I use an empty value an empty badge will be displayed:



```
<p:badge value="""">
  <p:avatar label=""Badge""/>
</p:badge>

```

Using the `rendered` attribute is no option as it will also hide the contents of the badge (the `p:avatar` in this case):



```
<p:badge rendered=""#{bean.value ne '0'}"">
  <p:avatar label=""Badge""/>
</p:badge>

```

","## PrimeFaces 10


Have a look at the HTML that is rendered when an empty value is used:



```
<div id=""..."" class=""ui-overlay-badge"">
  <span class=""ui-badge ui-widget""></span>
  ...
</div>

```

You can simply hide empty elements using the `:empty` CSS selector. So you can hide empty badges using CSS like:



```
.ui-badge:empty { display: none; }

```

This CSS rule requires you to have an empty value in order to hide the badge.
If you want to hide the badge if the value is `0`, you could use a ternary expression to map `0` to empty:



```
<p:badge value=""#{bean.value eq '0' ? '' : bean.value}"">
  <p:avatar label=""Badge""/>
</p:badge>

```

See also


- [How do I override default PrimeFaces CSS with custom styles?](https://stackoverflow.com/questions/8768317/how-do-i-override-default-primefaces-css-with-custom-styles)


## PrimeFaces 11+


In PrimeFaces 11 the `visible` attribute is [introduced](https://github.com/primefaces/primefaces/pull/7677). It can be used like:



```
<p:badge value=""#{bean.value}"" visible=""#{not empty bean.value}"">
  <p:avatar label=""Badge""/>
</p:badge>

```

or



```
<p:badge value=""#{bean.value}"" visible=""#{bean.value ne '0'}"">
  <p:avatar label=""Badge""/>
</p:badge>

```

"
"How do I set the lifetime of a return value as the lifetime of the variable I move into it?
I am trying to teach myself some rust, and have written something that looks like:



```
let args:Vec<String> = env::args().collect();
let parsed = parser::sys(args.as_slice());

```

...



```
pub fn sys<'a>(args:&'a [String]) -> Parsed<'a> {
  parsed(args)
}

```

where `parsed` is a function that parses and loads configs.


This works fine. Now I am trying to abstract away explicitly calling `env::args()` and hide it in the call to `sys` , so I write a new version of `sys`



```
pub fn sys<'a>() -> Parsed<'a> {
  let args:Vec<String> = env::args().collect();
  parsed(args.as_slice())
}

```

And this fails with: 



```
error: `args` does not live long enough
src/test.rs:66      parsed(args.as_slice())

```

I think the error is because there's no way for compiler to infer that I want the lifetime of this newly created struct to be the lifetime of the variable I want to move this into. Is this correct? How would I annotate the lifetime on this return value/fix this? 


","
> 
> I think the error is because there's no way for compiler to infer that I want the lifetime of this newly created struct to be the lifetime of the variable I want to move this into.
> 
> 
> 


Actually, no.


The error is because you are trying to create references into the variable `args` which will no longer be valid after you return from `sys` since `args` is a local variable and thus is dropped at the end of `sys`.


If you wish to use references, you could supply `sys` with a `&'a mut Vec<String>` (empty), fill it in `sys`, and return references to it:



```
pub fn sys<'a>(args: &'a mut Vec<String>) -> Parsed<'a> {
    *args = env::args().collect();
    parsed(args.as_slice())
}

```

which guarantees that `args` outlive the `sys` call. This will borrow `args` for the lifetime of the result.


Another solution is to do away with `'a` and simply have `Parsed` own its elements rather than have references to them; however without the definition of `Parsed` I cannot advise how best to do so.


"
"Reading Large Files in Java getting really slow
I am trying to read a large file ~200 MB (~300 million lines of text). I am using a relatively standard way of reading like: 



```
for (int i = 1; i <= numProcs; i++) {
    System.out.println(""Parsing "" + traceFilePrefix + i + "".prg"");
    System.out.println(""##########################"");

    try {
        String line;
        int instrType;
        String hexAddr;
        ArrayList<Trace> traces = new ArrayList<Trace>();

        int lineNum = 1;

        BufferedReader br = new BufferedReader(new FileReader(traceFilePrefix + numProcs + ""/"" + traceFilePrefix + i + "".PRG""));
        while ((line = br.readLine()) != null) {
            line = line.toUpperCase();
            // validate line format
            if (line.matches(regexLineFormat)) {
                StringTokenizer tokenizer = new StringTokenizer(line);
                instrType = Integer.parseInt(tokenizer.nextToken());
                hexAddr = tokenizer.nextToken();

                traces.add(new Trace(instrType, hexAddr));
            } else {
                // line is in invalid format
                System.err.println(""ERR: line "" + lineNum + "" is invalid \"""" + line + ""\"""");
            }
            lineNum++;

            if (lineNum % 1000 == 0) System.out.println(""line "" + lineNum); 
        }
        br.close();
        processorsTraces.add(i-1, traces);
    } catch (Exception e) {
        e.printStackTrace();
    }
}

```

I find that reading slows down at about line 20 million... in face it doesnt seem to progress at all ... how can I improve this? Whats the problem in this simple piece of code? 


A latest run seem to run out of heap space ... I guess I cant store data in an arraylist like that? 


This is supposed to be a simulator. It first parses trace files containing hexadecimal addresses of memory accesses. Then is supposed to run simulation of these data. This code snipplet only parses and stores the traces in an arraylist. 


","The issue is almost certainly in the structure of your `Trace` class, and it's memory efficiency. You should ensure that the `instrType` and `hexAddress` are stored as memory efficient structures. The instrType appears to be an `int`, which is good, but just make sure that it is declared as an `int` in the Trace class.


The more likely problem is the size of the hexAddress String. You may not realise it but Strings are notorious for 'leaking' memory. In this case, you have a `line` and you think you are just getting the hexString from it... but in reality, the hexString contains the entire line.... yeah, really. For example, look at the following code:



```
public class SToken {

    public static void main(String[] args) {
        StringTokenizer tokenizer = new StringTokenizer(""99 bottles of beer"");
        int instrType = Integer.parseInt(tokenizer.nextToken());
        String hexAddr = tokenizer.nextToken();
        System.out.println(instrType + hexAddr);
    }
}

```

Now, set a break-point in (I use eclipse) your IDE, and then run it, and you will see that hexAddr contains a char[] array for the entire line, and it has an offset of 3 and a count of 7.


Because of the way that String substring and other constructs work, they can consume huge amounts of memory for short strings... (in theory that memory is shared with other strings though). As a consequence, you are essentially storing the entire file in memory!!!!


At a minimum, you should change your code to:



```
hexAddr = new String(tokenizer.nextToken().toCharArray());

```

But even better would be:



```
long hexAddr = parseHexAddress(tokenizer.nextToken());

```



---


EDIT: Request to get help on how to debug the code:


To debug the code, you need to set a break-point in the method. Select the line `System.out.println(...);` and use the key-board short-cut Ctrl-Shift-B which will toggle a break-point, and you will get the 'dot' marker like the picture below:


![enter image description here](https://i.stack.imgur.com/EJx7t.jpg)


Then, with that break-point enabled, you need to debug the program:


Right-click on the 'main' word in the `public static void main (String[] ...)` and select 'Debug As -> Java Application'


This should prompt you to go to the 'Debug' perspective, and you will have a window with the 'Variables' displayed. One of the variables will be `hexAddr` and you can expand it to show its contents. Here's a screen-shot:


![enter image description here](https://i.stack.imgur.com/YU9ec.jpg)


From the variables section there are a few things of note. At the bottom, the 'value' variable is highlighted, and it's internal value is shown in the area below ( `[9, 9, , b, o, ......, r]`


You can see the `count` and `offset` variables for `hexAddr` too. This means that the `hexAddr` variable stores the char[] array for the entire line, but also stores an offset and count in that array which it uses for managing the subset of the chars that this String instance is interested in


"
"""Invalid Handle"" Create CGBitmapContext
I've got a problem with the CGBitmapcontext.
I get en error while creating the CGBitmapContext with the message ""invalid Handle"".


Here is my code:



```
var previewContext = new CGBitmapContext(null, (int)ExportedImage.Size.Width, (int)ExportedImage.Size.Height, 8, (int)ExportedImage.Size.Height * 4,                                                    CGColorSpace.CreateDeviceRGB(), CGImageAlphaInfo.PremultipliedFirst);

```

Thank you;


","That is because you are passing null to the first parameter. The CGBitmapContext is for drawing directly into a memory buffer. The first parameter in all the overloads of the constructor is (Apple docs):



> 
> *data* 
>  A pointer to the destination in memory where the drawing is to be rendered. The size of this memory block should be at least
>  (bytesPerRow\*height) bytes.
> 
> 
> 


In MonoTouch, we get two overloads that accept a byte[] for convenience. So you should use it like this:



```
int bytesPerRow = (int)ExportedImage.Size.Width * 4; // note that bytes per row should 
    //be based on width, not height.
byte[] ctxBuffer = new byte[bytesPerRow * (int)ExportedImage.Size.Height];
var previewContext = 
    new CGBitmapContext(ctxBuffer, (int)ExportedImage.Size.Width, 
    (int)ExportedImage.Size.Height, 8, bytesPerRow, colorSpace, bitmapFlags);

```

"
"How to disable all other date except from selected date in jquery datepicker
I have two datepickers named startdate and enddate. I want to disable all other dates except the selected date from startdate. for example if i selected 15/12/2016 from startdate datepicker then i want to disable all other dates in enddate datepicker except 15th day.


Demofiddle : <https://jsfiddle.net/d7vzxn8s/>


This is my code :



```
 <p>Start Date: <input type=""text"" id=""startdate""></p>
 <p>End Date: <input type=""text"" id=""enddate""></p>

 <script>
   $(""#startdate"").datepicker({
    onSelect: function (selected) {
        var dt = new Date(selected);
        dt.setDate(dt.getDate());
        $(""#enddate"").datepicker(""option"", ""minDate"", dt);
      }
    });
  $(""#enddate"").datepicker();

```

","Modified Fiddle : <https://jsfiddle.net/d7vzxn8s/2/>



```
var dateToday = new Date();
        var selectedDate;
        $(""#startdate"").datepicker({
            minDate: dateToday,
            onSelect: function (dateText, inst) {
                selectedDate = $(this).datepicker( ""getDate"" );
                var slctdDate = selectedDate.getDate()
               // alert(selectedDate);
                $(""#enddate"").datepicker({
                minDate: inst.day,
                beforeShowDay: function(date){
                    //Only allow fri, sat

                        return [date.getDate()==slctdDate];

                }
                });

            }
        });

        $(""#enddate"").datepicker(""option"");

```

"
"How do I measure how long a (linux C) function takes to execute?
I have a particular function (well, set of functions) that I want to start every 400ms. I'm not much of a C programmer, and so anything outside of the standard libraries is a mystery to me, as well as quite a bit within them.


My first thought is to use nanosleep to pause execution for 400ms in some sort of loop, but this of course doesn't take into account the execution time of the code I will be running. If I could measure it, and if it seemed fairly certain that it ran for the same approximate duration after 10 or 20 tests, I could then nanosleep() for the difference. This wouldn't be perfect, of course... but it might be close enough for a first try.


How do I measure the execution time of a C function? Or is there a better way to do this altogether, and what keywords do I need to be googling for?


","You should be able to use [settimer](http://linux.die.net/man/2/setitimer)



```
int setitimer(int which, const struct itimerval *value,
              struct itimerval *ovalue);

```

Just put the code that you want to execute every 400ms inside the SIGALRM handler. This way you don't need to account for the time that your code takes to run, which could potentially vary. I'm not sure what happens if the signal handler doesn't return before the next signal is generated.


An outline of what some of the code might look like is shown below.



```
void periodic_fuc(int signal_num)
{
  ...
  signam(SIGALRM, periodic_func);
}

int main(...)
{
  struct itimerval timerval = {0};

  signal(SIGALRM, periodic_func);      
  ...
  timerval.it_interval.tv_usec = 400000;
  timerval.it_value.tv_usec = 400000; // Wait 400ms for first trigger
  settimer(ITIMER_REAL, &timerval, NULL);

  while (!exit)
    sleep(1);
  return 0;
}

```

"
"Java which element in the ListIterator is considered previous?
I'm currently trying to learn how to implement my own ListIterators. I have most of it implemented and ready to go, except I'm confused by the previous() method. By standard convention, can I get an explanation of how previous() is usually interpreted.


i.e.:



```
             >cursor<
dog     cat    fish     bird     frog    snake

```

According to Oracles Java Platform 7 API:


E previous()


Returns the previous element in the list and moves the cursor position backwards. This method may be called repeatedly to iterate through the list backwards, or intermixed with calls to next() to go back and forth. (Note that alternating calls to next and previous will return the same element repeatedly.)


What I don't quite understand is if previous() is called, does it return 'fish' or 'cat'.


I understand it in 2 ways:


1) 'fish' was the object that you were previously at


2) 'cat' is the object at the index numerically previous to the index of 'fish'


If previous returns 'fish', then does remove() actually remove the same element regardless of the direction of traversal?


","It's easiest to think of the cursor as being *between* two elements. So at the start, it's *before* `dog`. Calling `next()` returns `dog`, and moves it to between `dog` and `cat`, etc. You've finished iterating when the cursor is *after* `snake`.


So `next()` always returns the value after the cursor, and moves the cursor to the position after that. `previous()` always returns the value *before* the cursor, and moves the cursor to the position before that.


EDIT: As noted by David Conrad in comments: `remove()` always removes the last value that was returned by either `next()` or `previous()`. That's probably the least obvious part, when it comes to the way I've explained it above... it's much easier to model that behaviour using the sort of ""on item"" cursor you've described.


"
"Bootstrap complex form layout for combined vertical/inline fields
I'm very new to Bootstrap but growing fast in my skill and also appreciation for it. I have a need to build a fairly complex page that will contain a lot of form elements, including multiple address fill-in components. I've attached here an image of the basic design that I am aiming for, and can't seem to achieve it with the basic form design elements available to me. I feel like my approach is getting ""out of control"" in the sense that I've got too many nesting levels and not enough understanding of what I'm actually doing. So rather than post the code I've got so far, can anyone give me a simplified example of how my design might be achieved?


![enter image description here](https://i.stack.imgur.com/PdJOy.png)


My dotted-line rectangles represent the fact that my page should have a two-column display, with most of the form elements off to the wider RH side. And then, within that column, my form has some additional levels of complexity. There are many additional form elements that I'm not even showing here, and that's my main reason for wanting to design it this way--because I need to save on vertical screen real estate as much as possible.


Thanks for your help.


","You should use <http://getbootstrap.com/css/#grid> for this.


A simple sample from me. Hope this will help you understand the concept.



```
<div class=""row"">
  <div class=""col-md-8"">
    <div class=""row"">
        <div class=""col-md-3"">
            Left side for labes
        </div>
        <div class=""col-md-9"">
            Right side for forms
            <div class=""row"">
                <div class=""col-md-12"">
                    longer forms
                </div>
            </div>
            <div class=""row"">
                <div class=""col-md-3"">
                    shorter forms
                </div>
                <div class=""col-md-3"">
                    shorter forms
                </div>
                <div class=""col-md-3"">
                    shorter forms
                </div>
                <div class=""col-md-3"">
                    shorter forms
                </div>
            </div>
        </div>
    </div>
  </div>
  <div class=""col-md-4"">
    Right panel with instructions
  </div>
</div>

```

"
"DEoptim appears to freeze in R
I have encountered a situation where `DEoptim` appears to freeze. I can't figure out why and was hoping somebody with more experience in C could take a look at it.


It is rather difficult to create a reproducible example, so I simply saved the entire environment 50 iterations before the DEoptim freezes. The file below, `'Envir650.Rdata'` can be found [here](https://drive.google.com/open?id=0B9O8SpdmUPxQV1ZOcHFQaEN1Umc). 



```
rm(list = ls())
library(DstarM)
library(DEoptim)

load('Envir650.Rdata') # load the environment

# Adjust one function
argsList$fun.density = DstarM::Voss.density 

argsList$control$trace = 1 # show intermediate output
argsList$control$parallelType = 0 # don't use parallel processing
.Random.seed = randomseed # set seed

out = do.call(DEoptim, argsList) # freezes at iteration 21 and crashes R.

```

Many thanks in advance!


EDIT: I hope the problem is now reproducible.


","The problem is in rtdists package, source file density.c, function integrate. The loop 



```
for(x = a+0.5*step; x < b; x += step) {
    result += step * F->f(x, F->data);
}

```

becomes infinite because `step` is too small. It is so small that `x+step==x` and `x` never reaches `b`. The code of `integrate` should be changed so that `step` is never smaller than `EPSILON`:



```
--- orig/rtdists/src/density.c  2016-07-15 10:28:56.000000000 +0200
+++ mine/rtdists/src/density.c  2016-08-29 17:41:53.831078335 +0200
@@ -72 +72 @@
-   double step = width / N;
+   double step = fmax(width / N, EPSILON);

```

With this change applied, your example finishes at iteration 51 without looping or crashing. I've
[notified](https://github.com/rtdists/rtdists/pull/3)
the rtdists authors; the fix is now in the [github version](https://github.com/rtdists/rtdists) of the package.


"
"How to Get the current location of my device

> 
> **Possible Duplicate:**  
> 
> [How do I get the current GPS location programmatically in Android?](https://stackoverflow.com/questions/1513485/how-do-i-get-the-current-gps-location-programmatically-in-android) 
> 
> 
> 


I have an application in which I want to get the current location.


also I want to get the latitude and longitude so I store that and return back that position.


how can do it...


","try following code segment,



```
import android.location.Location;
import android.location.LocationListener;
import android.os.Bundle;

public class LocListener implements LocationListener
{
    private static double lat =0.0;
    private static double lon = 0.0;
    private static double alt = 0.0; 
    private static double speed = 0.0;

    public static double getLat()
    {
        return lat;
    }

    public static double getLon() 
    {
        return lon;
    }

    public static double getAlt()
    {
        return alt;
    }

    public static double getSpeed()
    {
        return speed;
    }

    @Override
    public void onLocationChanged(Location location) 
    {
        lat = location.getLatitude();
        lon = location.getLongitude();
        alt = location.getAltitude();
        speed = location.getSpeed(); 
    }

    @Override
    public void onProviderDisabled(String provider) {}
    @Override
    public void onProviderEnabled(String provider) {}
    @Override
    public void onStatusChanged(String provider, int status, Bundle extras) {}
}

```

"
"How can I prevent the ""... has stopped working"" window on my Server 2008 R2 dev box?
I'm using a Windows Server 2008 x64 R2 machine as a development box. Amongst many other things I've got Visual Studio 2008 SP1 installed on it. When I'm working on a project, I sometimes need to use Debug > Start without Debugging (Ctrl + F5). If the program throws an exception, I get a new R2-style window appear. I'm pretty sure I didn't see this on my XP box - this is the first time I've developed directly on a server. Is there any way to avoid this - it's really beginning to bug me? 


E.g. my current project is accessing a WCF service - I'm using Ctrl+F5 to start a console program client. I run it and get the window. The title is the name of the project I've just started and it contains:



```
[insert-project-name-here] has stopped working

Windows can check online for a solution to the problem.

--> Check online for a solution and close the program

--> Close the program

--> Debug the program

V  View problem details.

```

Clicking on ""Close the program"" will actually close the window and the exception message appears in the console, which is what I want (but without the extra window-faffing).


How can I avoid this annoyance?


","
> 
> **Just-In-Time Debugging on Windows Vista and 2008 Server**
> 
> 
> By default on Windows Vista, the
>  just-in-time debugger is configured to
>  display a user-friendly dialog shown
>  below. Although you can click the
>  Debug the program button and start the
>  debugger successfully, your original
>  Visual Studio instance will hang in
>  the end of the debugging session.
> 
> 
> In order to avoid having to constantly
>  kill the Visual Studio process after
>  debugging, you will want to change
>  Just-In-Time debugger configuration to
>  work the same way it does on Windows
>  XP and Windows Server 2003. This
>  configuration is stored in registry as
>  DbgJITDebugLaunchSetting value in
>  HKEY\_LOCAL\_MACHINE\SOFTWARE\Microsoft.NETFramework
>  key. It’s default value on Windows
>  Vista is 0×10. Change it to 0×2, which
>  is the default value on Windows XP.
>  After making this change, you should
>  no longer see the User-Defined
>  Breakpoint dialog and will be able to
>  continue using Visual Studio after
>  debugging.
> 
> 
> 


from [this page](http://www.olegsych.com/2008/09/t4-tutorial-debugging-code-generation-files/) and scroll down to 'Just-In-Time Debugging on Windows Vista and 2008 Server'


Also note: if you are running x64 the key is going to be in the WOW section.


"
"Color values for Buttons (Normal state) in holo dark theme ICS
Need help regarding the color codes for buttons in HOLO DARK theme ICS. This [question](https://stackoverflow.com/questions/9947156/what-are-the-default-color-values-the-holo-theme-on-android-4-0) throws light into all the color codes for Holo theme except for the buttons in normal state. Thanks in advance. 
[basic buttons](http://developer.android.com/design/building-blocks/buttons.html#basic) holo theme ics.


","Ive got it now. If you want the default colors of Android ICS; you just have to go to your Android SDK and look after platforms\android-15\data\res\values and then colors. 
There you go:



```
 <!-- For holo theme -->
    <drawable name=""screen_background_holo_light"">#fff3f3f3</drawable>
    <drawable name=""screen_background_holo_dark"">#ff000000</drawable>
    <color name=""background_holo_dark"">#ff000000</color>
    <color name=""background_holo_light"">#fff3f3f3</color>
    <color name=""bright_foreground_holo_dark"">@android:color/background_holo_light</color>
    <color name=""bright_foreground_holo_light"">@android:color/background_holo_dark</color>
    <color name=""bright_foreground_disabled_holo_dark"">#ff4c4c4c</color>
    <color name=""bright_foreground_disabled_holo_light"">#ffb2b2b2</color>
    <color name=""bright_foreground_inverse_holo_dark"">@android:color/bright_foreground_holo_light</color>
    <color name=""bright_foreground_inverse_holo_light"">@android:color/bright_foreground_holo_dark</color>
    <color name=""dim_foreground_holo_dark"">#bebebe</color>
    <color name=""dim_foreground_disabled_holo_dark"">#80bebebe</color>
    <color name=""dim_foreground_inverse_holo_dark"">#323232</color>
    <color name=""dim_foreground_inverse_disabled_holo_dark"">#80323232</color>
    <color name=""hint_foreground_holo_dark"">#808080</color>
    <color name=""dim_foreground_holo_light"">#323232</color>
    <color name=""dim_foreground_disabled_holo_light"">#80323232</color>
    <color name=""dim_foreground_inverse_holo_light"">#bebebe</color>
    <color name=""dim_foreground_inverse_disabled_holo_light"">#80bebebe</color>
    <color name=""hint_foreground_holo_light"">#808080</color>
    <color name=""highlighted_text_holo_dark"">#6633b5e5</color>
    <color name=""highlighted_text_holo_light"">#6633b5e5</color>
    <color name=""link_text_holo_dark"">#5c5cff</color>
    <color name=""link_text_holo_light"">#0000ee</color>

```

This for the Background:



```
 <color name=""background_holo_dark"">#ff000000</color>
 <color name=""background_holo_light"">#fff3f3f3</color>

```

Probably you wont get the same colors if you look this up in Photoshop etc. because its set up with Alpha values.


"
"LMER model with uneven time points
Does it matter that my time series are taken at random points? All the examples I've seen have nice time series, like every day or whatever. I have what is otherwise a pretty simple model: one fixed treatment condition and then a random intercept and slope for each subject, but the observations are essentially random, like s1 has observations at 1,3,7, and 14 days, whereas s2 has them at, say day 5, 6, and 15 (not even the same number of observations!) I was thinking of either binning (early/late), or considering the missing days as missing data. Can someone advise, and/or point me to a relevant example? 


P.s., here's the model I'm using: 



```
lmer(KPS ~ Day + Tx + (Day | Tx) + (Day | SID), data = d)

```

","Short answer: this should be no problem. I'm pretty sure you'll have somewhat lower power than if you had a perfectly balanced design, but there is no fundamental difficulty.


Here's an example where I subsample the (complete/balanced) `sleepstudy` data set and show that it works fine (and the results don't change very much).


The only specific issue that I can think of is that if you are fitting a model with autoregressive structure (e.g. in `lme` with the `correlation` argument), you'll have to switch from an ""AR1"" specification to a ""continuous AR"" or ""Ornstein-Uhlenbeck"" or ""exponential correlation"" specification (in `lme` from `corAR1` to `corCAR`: in `glmmTMB` from `ar()` to `ou()`).


[This analysis of carbon exchange in tundra ecosystems](https://bbolker.github.io/mixedmodels-misc/ecostats_chap.html#tundra-carbon) shows an example of mixed models applied to an unevenly sampled time series.


## subsample data set



```
library(lme4)
set.seed(101)
table(sleepstudy$Subject)
ss <- do.call(""rbind"",
 lapply(split(sleepstudy,sleepstudy$Subject),
                     function(x) {
                         x[sort(sample(1:10,
                          size=rbinom(1,size=10,prob=0.7),
                                       replace=FALSE)),]
                     }))
lapply(split(ss,ss$Subject), function(x) x$Days)

```

## fit unbalanced/subsampled data set



```
lmer(Reaction~Days+(Days|Subject), data=ss)
## Linear mixed model fit by REML ['lmerMod']
## Formula: Reaction ~ Days + (Days | Subject)
##    Data: ss
## REML criterion at convergence: 1130.586
## Random effects:
##  Groups   Name        Std.Dev. Corr
##  Subject  (Intercept) 22.634
##           Days         6.303   0.15
##  Residual             21.628
## Number of obs: 119, groups:  Subject, 18
## Fixed Effects:
## (Intercept)         Days
##      250.30        10.44

```

## fit full data set



```
lmer(Reaction~Days+(Days|Subject), data=sleepstudy)

## Linear mixed model fit by REML ['lmerMod']
## Formula: Reaction ~ Days + (Days | Subject)
##    Data: sleepstudy
## REML criterion at convergence: 1743.628
## Random effects:
##  Groups   Name        Std.Dev. Corr
##  Subject  (Intercept) 24.737       
##           Days         5.923   0.07
##  Residual             25.592       
## Number of obs: 180, groups:  Subject, 18
## Fixed Effects:
## (Intercept)         Days  
##      251.41        10.47  

```

"
"Output to console from a Win32 GUI application on Windows 10
I tried this code to output to the console:



```
#include <Windows.h>

#include <stdio.h>
#include <io.h>
#include <fcntl.h>

int APIENTRY WinMain(HINSTANCE hInstance,
                     HINSTANCE hPrevInstance,
                     LPSTR     lpCmdLine,
                     int       nCmdShow)
{
    AllocConsole();

    HANDLE handle_out = GetStdHandle(STD_OUTPUT_HANDLE);
    int hCrt = _open_osfhandle((long) handle_out, _O_TEXT);
    FILE* hf_out = _fdopen(hCrt, ""w"");
    setvbuf(hf_out, NULL, _IONBF, 1);
    *stdout = *hf_out;

    HANDLE handle_in = GetStdHandle(STD_INPUT_HANDLE);
    hCrt = _open_osfhandle((long) handle_in, _O_TEXT);
    FILE* hf_in = _fdopen(hCrt, ""r"");
    setvbuf(hf_in, NULL, _IONBF, 128);
    *stdin = *hf_in;

    printf(""Hello!"");
}

```

the console opens but nothing is outputted to it. What's wrong with that code?


I tried all these suggestions:


<https://justcheckingonall.wordpress.com/2008/08/29/console-window-win32-app/>


<http://dslweb.nwnexus.com/~ast/dload/guicon.htm>


[How do I print to the debug output window in a Win32 app?](https://stackoverflow.com/questions/1333527/how-do-i-print-to-the-debug-output-window-in-a-win32-app)


but I couldn't get any output to the console created with AllocConsole() on Windows 10 in WinMain. 
Note: I actually didn't create any real Window.
Was something changed in Window 10 that prevents the above solutions from working or is there something that I might be missing (compiler flags or something)?


What do you think?


","Based on the accepted answer from ProXicT link with a few modifications. The following code works for std::cout. The other methods won't work on 64bit with Visual Studio 2015:



```
#include <iostream>
#include <cstdio>
#include <fstream>

#include <Windows.h>


// For debugging
#include <io.h>
#include <fcntl.h>


#define UNUSED(x) (void)(x)       // Unused param (C compatible - not applicable to expressions)

class outbuf : public std::streambuf {
public:
    outbuf() {
        setp(0, 0);
    }

    virtual int_type overflow(int_type c = traits_type::eof()) {
        return fputc(c, stdout) == EOF ? traits_type::eof() : c;
    }
};

int CALLBACK
WinMain (HINSTANCE hInstance,
         HINSTANCE /*hPrevInst*/, // Unused param (C++ only)
         LPSTR lpCmdLine,
         int (nShowCmd))
{
    UNUSED(hInstance);
//    UNUSED(hPrevInst);
    UNUSED(lpCmdLine);
    UNUSED(nShowCmd); // This param is used


    // create the console
    if (AllocConsole()) {
        FILE* pCout;
        freopen_s(&pCout, ""CONOUT$"", ""w"", stdout);
        SetConsoleTitle(L""Debug Console"");
        SetConsoleTextAttribute(GetStdHandle(STD_OUTPUT_HANDLE), FOREGROUND_GREEN | FOREGROUND_BLUE | FOREGROUND_RED);
    }

    // set std::cout to use my custom streambuf
    outbuf ob;
    std::streambuf *sb = std::cout.rdbuf(&ob);

    // do some work here
    printf(""Hello!\n"");

    std::cout << ""nShowCmd = "" << nShowCmd << std::endl;
    std::cout << ""Now making my first Windows window!"" << std::endl;


    // make sure to restore the original so we don't get a crash on close!
    std::cout.rdbuf(sb);

    MessageBoxW (NULL,
                 L""Hello World!"",
                 L""hello"",
                 MB_OK | MB_ICONINFORMATION);

    return 0;
}

```

EDIT:


The same stuff as above but with colors for completeness:



```
#include <iostream>
#include <cstdio>
#include <fstream>

#include <Windows.h>


// For debugging
#include <io.h>
#include <fcntl.h>


#define UNUSED(x) (void)(x)       // Unused param (C compatible - not applicable to expressions)

class outbuf : public std::streambuf {
public:
    outbuf() {
        setp(0, 0);
    }

    virtual int_type overflow(int_type c = traits_type::eof()) {
        return fputc(c, stdout) == EOF ? traits_type::eof() : c;
    }
};

int CALLBACK
WinMain (HINSTANCE hInstance,
         HINSTANCE /*hPrevInst*/, // Unused param (C++ only)
         LPSTR lpCmdLine,
         int (nShowCmd))
{
    UNUSED(hInstance);
//    UNUSED(hPrevInst);
    UNUSED(lpCmdLine);
    UNUSED(nShowCmd); // This param is used


    // create the console
    if (AllocConsole()) {
        FILE* pCout;
        freopen_s(&pCout, ""CONOUT$"", ""w"", stdout);
        SetConsoleTitle(L""Debug Console"");

    }

    // set std::cout to use my custom streambuf
    outbuf ob;
    std::streambuf *sb = std::cout.rdbuf(&ob);

    // do some work here

    printf(""Hello!\n"");

    HANDLE stdHandle = GetStdHandle(STD_OUTPUT_HANDLE);
    CONSOLE_SCREEN_BUFFER_INFO consoleInfo;
    WORD defaultConsoleTextAttributes;
    GetConsoleScreenBufferInfo(stdHandle, &consoleInfo);
    defaultConsoleTextAttributes = consoleInfo.wAttributes;
    WORD currentConsoleTextAttributes = FOREGROUND_GREEN | FOREGROUND_BLUE | FOREGROUND_RED;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""nShowCmd = "" << nShowCmd << std::endl;
    currentConsoleTextAttributes = FOREGROUND_GREEN;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_GREEN | FOREGROUND_INTENSITY;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_BLUE;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_BLUE | FOREGROUND_INTENSITY;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_RED;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_RED | FOREGROUND_INTENSITY;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_GREEN | FOREGROUND_BLUE;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_GREEN | FOREGROUND_BLUE | FOREGROUND_INTENSITY;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_GREEN | FOREGROUND_RED;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_GREEN | FOREGROUND_RED | FOREGROUND_INTENSITY;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_BLUE | FOREGROUND_RED;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;
    currentConsoleTextAttributes = FOREGROUND_BLUE | FOREGROUND_RED | FOREGROUND_INTENSITY;
    SetConsoleTextAttribute(stdHandle, currentConsoleTextAttributes);
    std::cout << ""Now making my first Windows window!"" << std::endl;



    // make sure to restore the original so we don't get a crash on close!
    std::cout.rdbuf(sb);

    Sleep(5000);
    ShowWindow(GetConsoleWindow(), SW_HIDE);
    FreeConsole();

    MessageBoxW (NULL,
                 L""Hello World!"",
                 L""hello"",
                 MB_OK | MB_ICONINFORMATION);

    return 0;
}

```

All that remains now is to make it into a complete logging class.


"
"How can I reroute in BeforeEnterObserver in the view before AppLayout is rendered in Vaadin 14 (Vaadin Flow)
So for example let's I have:



```
public class MainView extends AppLayout {
   public MainView() {
      User user = VaadinSession.getAttribute(""user"");
      if(user.isCheckSomething())
          Span span = new Span(""Hello "" + user.getFirstname());
   }
}

```

This will fail with a NPE if the user is not already logged in as getting the User from the session will be null. Now I could add a null check before the `if(user.isCheckSomething())` but ideally I would prefer not to have the `AppLayout` rendered at all if the user is not logged in. That is I'd rather fail in the view through `BeforeEnterObserver` in the `public void beforeEnter(BeforeEnterEvent event)` method however the `AppLayout` is called and created before the `beforeEnter(...)` method is called.


In other words how can I force the instantiation of `AppLayout` to be skipped entirely if the user isn't logged in through the view so that the `AppLayout` is never constructed.


","As a rule of thumb, it's really not reliable to work with the UI in the constructor in any way, since it's not initialized yet.


AppLayout has its own `afterNavigation` method which should be called after BeforeEnterEvent.



```
public class MainView extends AppLayout {
    @Override
    protected void afterNavigation() {
        super.afterNavigation();
        User user = VaadinSession.getAttribute(""user"");
        if(user.isCheckSomething())
            Span span = new Span(""Hello "" + user.getFirstname());
    }

```

So in beforeEnter method you will reroute to login, otherwise the after navigation from the parent layout will be fired.



```
@Override
public void beforeEnter(BeforeEnterEvent event) {
    if (!isAuthenticated()) {
        event.rerouteTo(LoginView.class);
    }
}

```

You can either implement `BeforeEnterObserver` in each of your views and check if user is authenticated over and over again... or you can do it once:



```
@SpringComponent
public class AuthenticationControl implements VaadinServiceInitListener {
     @Override
    public void serviceInit(ServiceInitEvent event) {
        event.getSource().addUIInitListener(uiEvent -> {
            final UI ui = uiEvent.getUI();
            ui.addBeforeEnterListener(/* do the check here, either with a method reference or create a separate listener class */);
        });
     }
}

```

"
"AutoExec Module not Initiating
My data storage form has a new module (AutoExec) which is meant to test the connection to a SQL server db upon opening the form. Only it is not firing when opened. I was wondering if there was something in my code causing this, this is my first Module so I am not familiar with the proper forms yet.



```
Public Sub AutoExec()
Dim cnn As ADODB.Connection
Dim localrst As New ADODB.Recordset
Dim remoterst As New ADODB.Recordset

Set cnn = New ADODB.Connection
cnn.Open ""Provider=SQLOLEDB; Data Source=DB\P003,49503; Initial Catalog=HRLearnDev;"" _
    & ""User Id=USERNAME; Password=PASSWORD;""

If cnn.State = adStateOpen Then

    MsgBox (""You have an established connection."")
  Else
    MsgBox (""Cannot connect to remote server. Data will be stored locally  to CDData Table until application is opened again."")

End If

cnn.Close
Dim rst As New ADODB.Recordset


End Sub

```

","When a database includes a macro named *AutoExec*, Access will run that macro at database startup.


In order for that to work, *AutoExec* must be an Access *macro object*. In your case, you have a VBA procedure named *AutoExec*. Since that is not a *macro object*, Access does not run the procedure automatically at database startup.


I suggest you create a VBA function ...



```
Public Function Startup()

```

Add the code from the body of your *AutoExec* procedure to the function.


Then create a new Access macro (an actual Access macro object --- on the ribbon in Access 2007, choose *Create->Macro*), use the macro `RunCode` action to run your new `Startup()` function. Name that macro object as *AutoExec*.


Here is a screenshot of my simple *AutoExec* example macro open in Design View. 


![AutoExec macro in design view](https://i.stack.imgur.com/AUso2.png)


"
"Apache POI XSSF header not created
As I experiment with xlsx creation, I'm stuck on creating headers. I'm able to create a file with rows and merged cells, but headers never seem to work. Here's what I have:



```
var WorkBook = CreateObject( 
    ""java"",
    ""org.apache.poi.xssf.usermodel.XSSFWorkbook""
).Init();

var Sheet = WorkBook.CreateSheet(
    JavaCast( ""string"", 'my sheetname' )
);

// create the default header if it doesn't exist
var header = sheet.getHeader(); // have also tried getEvenHeader() and getOddHeader()

header.setText('&LLeft Section');
// have also tried the following:
//header.setLeft('left header');
//header.setCenter('CENTER');
//header.setRight('right header');

// open the file stream
var FileOutputStream = CreateObject(
    ""java"",
    ""java.io.FileOutputStream""
).Init(
    JavaCast( ""string"", filename )
);

// Write the workbook data to the file stream.
WorkBook.Write( 
    FileOutputStream 
);

// Close the file output stream.
FileOutputStream.Close();

```

When I run this code, no errors are thrown. The file is created and can be opened without throwing any errors, but no headers appear. And like I said, if I create rows/cells instead of a header, those are created correctly. What am I missing?


EDIT:
As Leigh points out below, headers/footers have a different meaning in Excel than how I was thinking of them (as in PDFs). I got thrown off by the way adding a header in Excel shows it above the first row, and thought that adding one through POI would do the same thing.


","(Promoted from comments, in case the answer is helpful for the next guy)


Silly question, but how are you verifying the headers are not present? In Excel, [headers and footers should only be visible when printing (or in print preview mode)](https://support.office.com/en-us/article/Headers-and-footers-in-a-worksheet-CAE2A88C-64A7-42AB-96A4-28D2FC16AD31). 



> 
> ...Headers and footers are not displayed on the worksheet in Normal
>  view — they are displayed only in Page Layout view and on the printed
>  pages.
> 
> 
> 


FWIW, the code works fine for me under CF10 and 11, after I populated at least one cell (so there was something to print).


[**Runnable Example on trycf.com**](http://trycf.com/gist/011e3494c818f8c63c29b6e28dddfec3/acf?theme=monokai)



```
<cfscript>
    workBook = CreateObject( ""java"", ""org.apache.poi.xssf.usermodel.XSSFWorkbook"").Init();
    sheet = WorkBook.CreateSheet( JavaCast( ""string"", 'my sheetname' ) );
    header = sheet.getHeader(); // have also tried getEvenHeader() and getOddHeader()
    header.setText('&LLeft Section');
    // add some data so there is something to print
    sheet.createRow(0).createCell(0).setCellValue(""sample value"");
    // Using binary stream because trycf.com does not support files for security reasons
    baos = createObject(""java"", ""java.io.ByteArrayOutputStream"").init();
    // Write the workbook data to the binary stream
    workBook.write( baos );
    baos.close();
</cfscript>

<!--- CF10 lacks support for script version of cfcontent --->
<cfcontent type=""application/vnd.openxmlformats-officedocument.spreadsheetml.sheet""
    variable=""#baos.toByteArray()#"">

```

"
"What does (String[] arguments) mean?
I added the following code to a new class I created in Java:



```
public static void main(String[] arguments) {

```

I understand what public, static and void mean, but what does (`String[] arguments`) mean?


","Your `main()` method can take input parameters of type `String` if your program is run through a console like



```
java YourClass arg1 arg2

```

Now, within `main()` if you iterate the `String []` like



```
for (arg : arguments)
    System.out.println(arg);

```

it should print



```
arg1
arg2

```

***Demo :***  




```
public class AddTwoNumbers {

    public static void main(String[] args) {
      if(args.length == 2) {
        try {
            int a = Integer.parseInt(args[0]);
            int b = Integer.parseInt(args[1]);
            System.out.println(""a + b = "" + a + "" + "" + b + "" = ""+ (a + b));
        } catch (NumberFormatException e) {
            System.err.println(""Invalid Input: Please enter numbers."");
        }
      } else {
         System.err.println(""Missing Input: Please enter TWO numbers."");
      }
    }
}

```

You can run this on your console as



```
java AddTwoNumbers 2 3

```

and it should print



```
a + b = 2 + 3 = 5

```

"
"Unique classes in a nested list
I am looking for a straight forward way to get the unique classes of the first-level elements in a nested list. 


Reproducible example:



```
x = list(list(1,""A""), # Unique classes:  character, numeric
         4, # Unique classes: numeric
         list(1,2,list(4,5, list(6)))) # Unique classes: numeric

# Expected return: list(c('character', 'numeric'), c('numeric'), c('numeric'))

```

","This works



```
list_classes <- function(ls) {
  unique(sapply(ls, class))
}

lapply(x, list_classes)

#[[1]]
#[1] ""numeric""   ""character""

#[[2]]
#[1] ""numeric""

#[[3]]
#[1] ""numeric"" ""list""

```

**Edit:** Recursive version



```
list_classes <- function(ls) {

  # get classes of list
  cls <- sapply(ls, class)

  # which are lists themselves
  which_ls <- cls == ""list""
  lists <- ls[which_ls]

  # which are non lists
  non_lists <- cls[!which_ls]

  # recursively evaluate contents of list for their classes
  return(
    unique(
      c(non_lists, sapply(lists, list_classes), recursive = T)
    )
  )

}

lapply(x, list_classes)


#[[1]]
#[1] ""numeric""   ""character""

#[[2]]
#[1] ""numeric""

#[[3]]
#[1] ""numeric""

```

"
"Is there are nice way of logging function names when inside a lambda?
Our logging framework, like most logging frameworks, uses the `__FUNCTION__` preprocessor macro to insert the current function into log files so that our logging looks like:



```
L4  T11332 609661594 [PreProcessorFunctions::RegenerateLinkIDs] [ENTER]
L4  T11332 609661594 [PreProcessorFunctions::RegenerateLinkIDs] [EXIT]
L4  T11332 609661594 [ConfigMerger::ValidateConfigObject] [ENTER]
L3  T11332 609661594 [ConfigMerger::ValidateConfigObject] Configuration Exists: 1

```

As we've started to use C++11 more, I've noticed that labmdas produce accurate yet unhelpful `__FUNCTION__` output:



```
L4  T9604 609661594 [`anonymous-namespace'::<lambda1>::operator ()] Writing EMX config file: C:\windows\TEMP\CBE01448-32A2-493A-A9A1-2112F5709028\CA37BE5C-B398-4D61-980D-66B8E1E6D001\\EMXConfiguration.xml
L4  T11332 609661594 [`anonymous-namespace'::<lambda3>::operator ()] Writing Auditing config file: C:\windows\TEMP\CBE01448-32A2-493A-A9A1-2112F5709028\CA37BE5C-B398-4D61-980D-66B8E1E6D001\\Auditing.xml
L4  T11960 609661594 [`anonymous-namespace'::<lambda2>::operator ()] Writing UEM config file: C:\windows\TEMP\CBE01448-32A2-493A-A9A1-2112F5709028\CA37BE5C-B398-4D61-980D-66B8E1E6D001\\Configuration.xml

```

As you can see, all class scope has been lost and all we now know is that this logging statement came from an anonymous lambda. Does anyone have a good strategy for logging out the enclosing function? This would appear to be the most useful thing to log...


","I think the only workaround would be to expand `__FUNCTION__` (or `__func__` in C++11) in the enclosing function and capture that inside the lambda.



```
void f(){
  auto& _func_ = __func__;
  [&]{ std::cout << _func_; };
}

```

I would tag this as a ""quality of implementation"" issue, though. GCC, for example, provides the `__PRETTY_FUNCTION__` macro that expands much better for lambdas (and in general):



```
#include <iostream>

struct X{
  void f(){
    []{ std::cout << __PRETTY_FUNCTION__ << ""\n""; }();
  }
};

int main(){
  X().f();
}

```

Output: `X::f()::<lambda()>`


[Live example.](http://ideone.com/TF8cD9)


"
"Iterate a select statement in SQLite
I have prepared a select statement.



```
sqlite3_prepare(db, ""SELECT symbol,image FROM test WHERE template=? AND field=?"", -1, &stmt, 0);
sqlite3_bind_text16(stmt, 1, pParams->pszFile, -1, SQLITE_STATIC);
sqlite3_bind_text16(stmt, 2, pParams->pszFieldId, -1, SQLITE_STATIC);

```

How to go about iterating through the results?


","
```
int rc = sqlite3_prepare(db, ""SELECT symbol,image FROM test WHERE template=? AND field=?"", -1, &stmt, 0);
sqlite3_bind_text16(stmt, 1, pParams->pszFile, -1, SQLITE_STATIC);
sqlite3_bind_text16(stmt, 2, pParams->pszFieldId, -1, SQLITE_STATIC);  

 if(rc != SQLITE_OK)
 {
    fprintf(stderr, ""SQL error: %s\n"", sqlite3_errmsg(db));
    return;
  }
  rc = sqlite3_step(stmt);
  int ncols = sqlite3_column_count(stmt);

  while(rc == SQLITE_ROW)
  {

      for(i=0; i < ncols; i++)
      {
         fprintf(stderr, ""'%s' "", sqlite3_column_text(stmt, i));
      }

      fprintf(stderr, ""\n"");
      rc = sqlite3_step(stmt);
  }

  sqlite3_finalize(stmt);

```

"
"How to convert variant value into integer?
Below is my code:



```
var
  i : integer;
  ...
  ...
if not VarIsNull(TcxLookupComboBox(Sender).EditValue) then
begin
  i := Integer(TcxLookupComboBox(Sender).EditValue);
end;

```

I can use `VarToStr` to convert variant into string but there is no `VarToInt` in Delphi for this. So, I have converted it like this `Integer(TcxLookupComboBox(Sender).EditValue)`. Is this the right approach?


","Have a look at this: <http://docwiki.embarcadero.com/RADStudio/Rio/en/Variant_Types>


Specifically check the Variant Type Conversions section.


You should be able to assign is directly using implicit type casting. As in Delphi just handles it for you.


As an example:



```
var
  theVar: Variant;
  theInt: integer;
begin

  theVar := '123';
  theInt := theVar;
  showmessage(IntToStr(theint));
end;

```

This works without issue.


To ensure that your data is an integer and that it is safe to do at runtime (as in to make use that you didn't have a string value in the variant, which would result in a runtime error) then have a look at the Val function: <http://docwiki.embarcadero.com/Libraries/Rio/en/System.Val>


Hope this helps.


"
"Confusion about pooling layer, is it trainable or not?
I have read in many places such as Stanford's Convolutional neural networks course notes at [CS231n](http://cs231n.github.io/convolutional-networks/) (and also [here](http://andrew.gibiansky.com/blog/machine-learning/convolutional-neural-networks/), and [here](http://ufldl.stanford.edu/tutorial/supervised/ExerciseConvolutionalNeuralNetwork/) and [here](https://www.quora.com/How-do-I-derive-and-implement-the-gradients-of-subsampling-max-pooling-and-normalization-in-deep-learning)), that pooling layer does not have any trainable parameters!  

And yet today I was informed by someone that in some paper([here it is](http://arxiv.org/abs/1506.01195))
they say and I quote : 



> 
> S1 layer for sub sampling, contains six feature map, each feature map
>  contains 14 x 14 = 196 neurons. the sub sampling window is 2 x 2
>  matrix, sub sampling step size is 1, so the S1 layer contains 6 x 196
>  x (2 x 2 + 1) = 5880 connections. Every feature map in the S1 layer
>  contains a weights and bias, so a total of 12 parameters can be
>  trained in S1 layer .
> 
> 
> 


What is this ?  

Can anyone please enlighten me on this? 


","In the paper you read



> 
> a total of 12 parameters can be trained in S1 layer
> 
> 
> 


meant the number of output planes in the pooling layer, not the number of parameters in the weight matrix. Normally, what we train within a neural network model are the parameters in the weight matrix. We don't train parameters in input planes or output planes. So students who wrote the paper didn't express themselves clearly, which made you confused about what a pooling layer really is.


There are no trainable parameters in a max-pooling layer. In the **forward pass**, it pass maximum value within each rectangle to the next layer. In the **backward pass**, it propagate error in the next layer to the place where the max value is taken, because that's where the error comes from.




---


For example, in **forward pass**, you have a image rectangle:



```
1 2
3 4

```

and you would get:



```
4

```

in the next layer.


And in **backward pass**, you have error:



```
-0.1

```

then you propagate the error back to where you get it:



```
0 0 
0 -0.1

```

because the take the number 4 from that location in the forward pass.


"
"Running standard optimization passes on a LLVM module
Say I have a valid LLVM module:



```
std::unique_ptr<llvm::Module> module;

```

I want to run LLVM traditional optimization passes on it:



```
llvm::PassBuilder passBuilder;
llvm::ModulePassManager modulePassManager = passBuilder.buildPerModuleDefaultPipeline(llvm::PassBuilder::OptimizationLevel::O3);
llvm::ModuleAnalysisManager moduleAnalysisManager;
passBuilder.registerModuleAnalyses(moduleAnalysisManager);
modulePassManager.run(*module, moduleAnalysisManager);

```

Unfortunately, the call crashes and debugging shows that the `moduleAnalysisManager` has only the module passes, but
not the function ones that are wrapped with the proxy class.


How should I setup `modulePassManager` to handle all (module) passes of a specific level? I don't have individual functions, so I can't run the function passes just on them.


","The proper LLVM way is to create all `analysisManager`s and then link all of them together. Let's start by creating them:



```
llvm::PassBuilder passBuilder;
llvm::LoopAnalysisManager loopAnalysisManager(true); // true is just to output debug info
llvm::FunctionAnalysisManager functionAnalysisManager(true);
llvm::CGSCCAnalysisManager cGSCCAnalysisManager(true);
llvm::ModuleAnalysisManager moduleAnalysisManager(true);

```

Then we register each manager individually, and then we cross register them. This means that the number of managers here is fixed by design and if LLVM (7 at this time) changes the number of managers, this will need to be adapted:



```
passBuilder.registerModuleAnalyses(moduleAnalysisManager);
passBuilder.registerCGSCCAnalyses(cGSCCAnalysisManager);
passBuilder.registerFunctionAnalyses(functionAnalysisManager);
passBuilder.registerLoopAnalyses(loopAnalysisManager);
// This is the important line:
passBuilder.crossRegisterProxies(
    loopAnalysisManager, functionAnalysisManager, cGSCCAnalysisManager, moduleAnalysisManager);

```

Once the `passBuilder` is created, we can finally make the optimization passes for the module with a call to the `moduleAnalysisManager`.



```
llvm::ModulePassManager modulePassManager =
    passBuilder.buildPerModuleDefaultPipeline(llvm::PassBuilder::OptimizationLevel::O3);
modulePassManager.run(*module, moduleAnalysisManager);

```

This will run module level passes as well as all inner passes that LLVM can run on pieces of the module (function level, loop level...).


"
"How to specify default value for List<> parameter in a c# function?
I've got a C# method that takes a bunch of parameters, all of which have default values. One of the parameters is a List. I can't figure out how to specify that the List should default to empty. Here's what it looks like:



```
    public static void execute(
        String condition = ""Unnamed condition"",
        List<String> messages,
        Object actual    = null,
        Object expected  = null)

```

I can't quite figure out how to specify that messages should be empty by default. When I enter:



```
...
List<String> messages = new List<String> ()
...

```

it complains that ""default parameter value for 'messages' must be a compile-time constant"".


Any ideas?


","Because default parameter values must be compile-time expressions, the only acceptable default parameter value for reference types is `null`.


You can get around this with an overload, though:



```
 public static void execute(String condition = ""Unnamed condition"")
 {
     execute(condition, new List<String>(), null, null);
 }

```

Or constructing a list if the argument is `null`. If you must need a list and want to treat all `null` as an empty list, this would also handle if they called with `null` explicitly.



```
 public static void execute(String condition = ""Unnamed condition"",
        List<String> messages = null, Object actual = null,
        Object expected = null)
{
    // if you really want this to be empty if null, can check and assign.
    if (messages == null)
    {
        messages = new List<String>();
    }

    // your other logic
}

```

Or, if messages is only used in one place, you can use the null-coallescing operator to substitute an empty enumeration:



```
 public static void execute(String condition = ""Unnamed condition"",
        List<String> messages = null, Object actual = null,
        Object expected = null)
{
    // assuming you are using messages once for iteration or something...
    foreach(var msg in messages ?? Enumerable.Empty<String>())
    ...
}

```

Though obviously a simple if-guard can be more efficient. Really depends if you want to treat it as an empty enumerable or as an empty list or just bypass logic...


"
"calculating intersection point of quadratic bezier curve
This is definitely pushing the limits for my trig knowledge.


Is there a formula for calculating an intersection point between a quadratic bezier curve and a line?


**Example:**


in the image below, I have P1, P2, C (which is the control point) and X1, X2 (which for my particular calculation is just a straight line on the X axis.)


What I would like to be able to know is the X,Y position of T as well as the angle of the tangent at T. at the intersection point between the red curve and the black line.


![curve example](https://i.stack.imgur.com/2w9u0.png)


After doing a little research and finding [this](https://stackoverflow.com/questions/5634460/quadratic-bezier-curve-calculate-point ""info"") question, I know I can use:



```
t = 0.5; // given example value
x = (1 - t) * (1 - t) * p[0].x + 2 * (1 - t) * t * p[1].x + t * t * p[2].x;
y = (1 - t) * (1 - t) * p[0].y + 2 * (1 - t) * t * p[1].y + t * t * p[2].y;

```

to calculate my X,Y position at any given point along the curve. So using that I could just loop through a bunch of points along the curve, checking to see if any are on my intersecting X axis. And from there try to calculate my tangent angle. But that really doesn't seem like the best way to do it. Any math guru's out there know what the best way is? 


I'm thinking that perhaps it's a bit more complicated than I want it to be.


","![enter image description here](https://i.stack.imgur.com/5TO6R.png)


**Quadratic curve formula:**



```
y=ax^2+bx+c // where a,b,c are known

```

**Line formula:**



```
// note: this `B` is not the same as the `b` in the quadratic formula ;-)

y=m*x+B  // where m,B are known.

```

**The curve & line intersect where both equations are true for the same [x,y]:**


Here's annotated code and a Demo:





```
// canvas vars
var canvas=document.getElementById(""canvas"");
var ctx=canvas.getContext(""2d"");
var cw=canvas.width;
var ch=canvas.height;

// linear interpolation utility
var lerp=function(a,b,x){ return(a+x*(b-a)); };

// qCurve & line defs
var p1={x:125,y:200};
var p2={x:250,y:225};
var p3={x:275,y:100};
var a1={x:30,y:125};
var a2={x:300,y:175};

// calc the intersections
var points=calcQLintersects(p1,p2,p3,a1,a2);

// plot the curve, line & solution(s)
var textPoints='Intersections: ';
ctx.beginPath();
ctx.moveTo(p1.x,p1.y);
ctx.quadraticCurveTo(p2.x,p2.y,p3.x,p3.y);
ctx.moveTo(a1.x,a1.y);
ctx.lineTo(a2.x,a2.y);
ctx.stroke();
ctx.beginPath();
for(var i=0;i<points.length;i++){
  var p=points[i];
  ctx.moveTo(p.x,p.y);
  ctx.arc(p.x,p.y,4,0,Math.PI*2);
  ctx.closePath();
  textPoints+=' ['+parseInt(p.x)+','+parseInt(p.y)+']';
}
ctx.font='14px verdana';
ctx.fillText(textPoints,10,20);
ctx.fillStyle='red';
ctx.fill();

///////////////////////////////////////////////////

function calcQLintersects(p1, p2, p3, a1, a2) {
  var intersections=[];

  // inverse line normal
  var normal={
    x: a1.y-a2.y,
    y: a2.x-a1.x,
  }

  // Q-coefficients
  var c2={
    x: p1.x + p2.x*-2 + p3.x,
    y: p1.y + p2.y*-2 + p3.y
  }

  var c1={
    x: p1.x*-2 + p2.x*2,
    y: p1.y*-2 + p2.y*2,
  }

  var c0={
    x: p1.x,
    y: p1.y
  }

  // Transform to line 
  var coefficient=a1.x*a2.y-a2.x*a1.y;
  var a=normal.x*c2.x + normal.y*c2.y;
  var b=(normal.x*c1.x + normal.y*c1.y)/a;
  var c=(normal.x*c0.x + normal.y*c0.y + coefficient)/a;

  // solve the roots
  var roots=[];
  d=b*b-4*c;
  if(d>0){
    var e=Math.sqrt(d);
    roots.push((-b+Math.sqrt(d))/2);
    roots.push((-b-Math.sqrt(d))/2);
  }else if(d==0){
    roots.push(-b/2);
  }

  // calc the solution points
  for(var i=0;i<roots.length;i++){
    var minX=Math.min(a1.x,a2.x);
    var minY=Math.min(a1.y,a2.y);
    var maxX=Math.max(a1.x,a2.x);
    var maxY=Math.max(a1.y,a2.y);
    var t = roots[i];
    if (t>=0 && t<=1) {
      // possible point -- pending bounds check
      var point={
        x:lerp(lerp(p1.x,p2.x,t),lerp(p2.x,p3.x,t),t),
        y:lerp(lerp(p1.y,p2.y,t),lerp(p2.y,p3.y,t),t)
      }
      var x=point.x;
      var y=point.y;
      // bounds checks
      if(a1.x==a2.x && y>=minY && y<=maxY){  
        // vertical line
        intersections.push(point);
      }else if(a1.y==a2.y && x>=minX && x<=maxX){
        // horizontal line
        intersections.push(point);
      }else if(x>=minX && y>=minY && x<=maxX && y<=maxY){
        // line passed bounds check
        intersections.push(point);
      }
    }
  }
  return intersections;
}
```


```
body{ background-color: ivory; padding:10px; }
#canvas{border:1px solid red;}
```


```
<h4>Calculate intersections of QBez-Curve and Line</h4>
<canvas id=""canvas"" width=350 height=350></canvas>
```






"
"fetching data using rownum in oracle
I have a query in oracle to fetch data from table using `rownum` but i didn't get any data.  

My query is like this : 


`select * from table-name where rownum<5` 


is this is a wrong query to fetch data whose row number is less than 5.  

when i used query like :  

`select * from table-name where rownum<=4`  

than it will gives a result record. 


My question is what is wrong here with ?  

Is this is syntax error or anything else??..


","rownum is a pseudo column that counts rows in the result set after the where clause has been applied. 



```
  SELECT table_name
    FROM user_tables
    WHERE rownum > 2;
TABLE_NAME                     
------------------------------

0 rows selected

```

However, this query will always return zero rows, regardless of the number of rows in the table.


To explain this behaviour, we need to understand how Oracle processes ROWNUM. When assigning ROWNUM to a row, Oracle starts at 1 and only increments the value when a row is selected; that is, when all conditions in the WHERE clause are met. Since our condition requires that ROWNUM is greater than 2, no rows are selected and ROWNUM is never incremented beyond 1.


<http://blog.lishman.com/2008/03/rownum.html>


[another stackoverflow link](https://stackoverflow.com/questions/834702/sql-query-with-row-number-order-by-and-where-clause) 


**Edited** 


this paragraph i find on [oracle](http://download.oracle.com/docs/cd/B14117_01/server.101/b10759/pseudocolumns008.htm) website which is much better 


Conditions testing for ROWNUM values greater than a positive integer are always false. For example, this query returns no rows:



```
SELECT * FROM employees
    WHERE ROWNUM > 1;

```

The first row fetched is assigned a ROWNUM of 1 and makes the condition false. The second row to be fetched is now the first row and is also assigned a ROWNUM of 1 and makes the condition false. All rows subsequently fail to satisfy the condition, so no rows are returned.


You can also use ROWNUM to assign unique values to each row of a table, as in this example:


"
"Select first element that has certain property if there are any duplicates otherwise select first element
For example I have a list of objects, with those properties



```
{ ID: ""1"", STATUS: ""INA"" }
{ ID: ""1"", STATUS: ""ACT"" } 
{ ID: ""2"", STATUS: ""INA"" } 
{ ID: ""2"", STATUS: ""BLO"" }

```

And now I want to group them by ID, to reduce duplicates, but when the duplicates are found I want to check if any of the statuses is ACT, if it is ACT i want to select this record, but if there's no ACT status I want to select first occurrence of the duplicate.


In example above I would want to select 



```
{ ID: ""1"", STATUS: ""ACT"" } //since it has ACT
{ ID: ""2"", STATUS: ""INA"" } //since it is first duplicate found

```

I know that the first step would be



```
var NoDup = from l in list.AsEnumerable()
                  group l by l.ID into c

```

But I don't know what to do next


","
```
var q = list.GroupBy(x => x.ID)
            .Select(g => g.OrderByDescending(x => x.STATUS == ""ACT"").First());

```

So first group-by the ID, then order the groups by the `bool` that is returned from the comparison `x.STATUS == ""ACT""`. True is ""higher"" than false, that's why i used `OrderByDescending`. `First` ensures that i only get the first record of every duplicate-group. Since `OrderBy...` has a stable sort the original order is maintained even if there is no ACT-status.




---


Another similar approach which might be little bit more efficient is:



```
var q = list.GroupBy(x => x.ID)
    .Select(g => g.Where(x => x.STATUS == ""ACT"").DefaultIfEmpty(g.First()).First());

```

This could be more efficient if the duplicae-groups are very large because the whole group doesn't need to be ordered by the `bool` if the first was already `STATUS == ""ACT""`. 


"
"How to process multiple connections simultaneously with HttpListener?
In the application that I build, there is a need for webserver that can serve, simultaneously, multiple clients.  

For that I use the `HttpListener` object. with its `Async` methods\events `BeginGetContext` and `EndGetContext`.  

In the delegated method, there is a call for the listener to start listening again, and it works.. mostly.


The code provided is a mix of code that i found here and there, and a delay, to simulate a data processing bottleneck.


The problem is, it starts to manage the next connection only AFTER the last one was served.. no use for me.



```
public class HtServer {


    public void startServer(){
        HttpListener HL = new HttpListener();
        HL.Prefixes.Add(""http://127.0.0.1:800/"");
        HL.Start();
        IAsyncResult HLC = HL.BeginGetContext(new AsyncCallback(clientConnection),HL);   
    }

    public void clientConnection(IAsyncResult res){
        HttpListener listener = (HttpListener)res.AsyncState;
        HttpListenerContext context = listener.EndGetContext(res);
        HttpListenerRequest request = context.Request;
        // Obtain a response object.
        HttpListenerResponse response = context.Response;
        // Construct a response. 
        // add a delay to simulate data process
        String before_wait = String.Format(""{0}"", DateTime.Now);
        Thread.Sleep(4000);
        String after_wait = String.Format(""{0}"", DateTime.Now);
        string responseString = ""<HTML><BODY> BW: "" + before_wait + ""<br />AW:"" + after_wait + ""</BODY></HTML>"";
        byte[] buffer = System.Text.Encoding.UTF8.GetBytes(responseString);
        // Get a response stream and write the response to it.
        response.ContentLength64 = buffer.Length;
        System.IO.Stream output = response.OutputStream;
        // You must close the output stream.
        output.Write(buffer, 0, buffer.Length);
        output.Close();
        listener.BeginGetContext(new AsyncCallback(clientConnection), listener);
    }
}

```

## edit



```
    private static void OnContext(IAsyncResult ar)
    {
        var ctx = _listener.EndGetContext(ar);
        _listener.BeginGetContext(OnContext, null);

        Console.WriteLine(DateTime.UtcNow.ToString(""HH:mm:ss.fff"") + "" Handling request"");

        var buf = Encoding.ASCII.GetBytes(""Hello world"");
        ctx.Response.ContentType = ""text/plain"";

        // prevent thread from exiting.
        Thread.Sleep(3000);
        // moved these lines here.. to simulate process delay
        ctx.Response.OutputStream.Write(buf, 0, buf.Length);
        ctx.Response.OutputStream.Close();
        Console.WriteLine(DateTime.UtcNow.ToString(""HH:mm:ss.fff"") + "" completed"");
    }

```

the output is
![enter image description here](https://i.stack.imgur.com/zY7mK.png)


","Well. That's because you start to fetch the next context after you have processed the first. Don't do that. Get the next context directly:



```
public void clientConnection(IAsyncResult res){
    HttpListener listener = (HttpListener)res.AsyncState;
    HttpListenerContext context = listener.EndGetContext(res);

    //tell listener to get the next context directly.
    listener.BeginGetContext(clientConnection, listener);

    HttpListenerRequest request = context.Request;
    // Obtain a response object.
    HttpListenerResponse response = context.Response;
    // Construct a response. 
    // add a delay to simulate data process
    String before_wait = String.Format(""{0}"", DateTime.Now);
    Thread.Sleep(4000);
    String after_wait = String.Format(""{0}"", DateTime.Now);
    string responseString = ""<HTML><BODY> BW: "" + before_wait + ""<br />AW:"" + after_wait + ""</BODY></HTML>"";
    byte[] buffer = System.Text.Encoding.UTF8.GetBytes(responseString);
    // Get a response stream and write the response to it.
    response.ContentLength64 = buffer.Length;
    System.IO.Stream output = response.OutputStream;
    // You must close the output stream.
    output.Write(buffer, 0, buffer.Length);
    output.Close();
}

```

Here is my sample code that proves that it work (updated per request of the OP):



```
class Program
{
    private static HttpListener _listener;

    static void Main(string[] args)
    {
        _listener = new HttpListener();
        _listener.Prefixes.Add(""http://localhost/asynctest/"");
        _listener.Start();
        _listener.BeginGetContext(OnContext, null);

        Console.ReadLine();
    }

    private static void OnContext(IAsyncResult ar)
    {
        var ctx = _listener.EndGetContext(ar);
        _listener.BeginGetContext(OnContext, null);

        Console.WriteLine(DateTime.UtcNow.ToString(""HH:mm:ss.fff"") + "" Handling request"");

        var buf = Encoding.ASCII.GetBytes(""Hello world"");
        ctx.Response.ContentType = ""text/plain"";

        // simulate work
        Thread.Sleep(10000);

        ctx.Response.OutputStream.Write(buf, 0, buf.Length);
        ctx.Response.OutputStream.Close();


        Console.WriteLine(DateTime.UtcNow.ToString(""HH:mm:ss.fff"") + "" completed"");
    }
}

```

Generates:


![enter image description here](https://i.stack.imgur.com/9HRvJ.png)


Both requests starts to get processed directly. 


**Why the above code works**


HTTP have something called pipelining. It means that all requests that are received over the same connection must get their responses in the same order. However, the built in HttpListener doesn't seem to support pipelining, instead it's completes the response for the first request before taking care of the second. It's therefore important that you make sure that every request is sent over a new connection. 


The easiest way to do that is to use different browsers when trying out the code. I did that, and as you see both my requests are handled at the same time.


"
"Programatically hiding and showing individual tabs in React Native Router Flux Tabbar
I have a tabbar in my app using React Native Router Flux. There are a couple use cases where hiding or showing specific tabs based on the current user would be very helpful. The main ones I have run into are:


- AB testing new tabs to specific users
- Showing a special admin tab to certain users with certain privileges


The react-native-router-flux library does not support any options to do this from what I can see. How can I achieve this functionality?


","The default tabbar component in react-native-router-flux is just the component from the `react-navigation-tabs` library. You can import this component directly into your code, customize as needed, and then pass it to `react-native-router-flux` through the `tabBarComponent` prop ([documented here](https://github.com/aksonov/react-native-router-flux/blob/master/docs/API.md#tabs-tabs-or-scene-tabs)).


I created a new component, which you should be able to copy directly and just change the logic for actually hiding the tabs based on your state:



```
import React from 'react'
import { BottomTabBar } from 'react-navigation-tabs'
import { View, TouchableWithoutFeedback } from 'react-native'
import { connect } from 'react-redux'

const HiddenView = () => <View style={{ display: 'none' }} />
const TouchableWithoutFeedbackWrapper = ({
  onPress,
  onLongPress,
  testID,
  accessibilityLabel,
  ...props
}) => (
  <TouchableWithoutFeedback
    onPress={onPress}
    onLongPress={onLongPress}
    testID={testID}
    hitSlop={{
      left: 15,
      right: 15,
      top: 5,
      bottom: 5,
    }}
    accessibilityLabel={accessibilityLabel}
  >
    <View {...props} />
  </TouchableWithoutFeedback>
)
const TabBarComponent = props => (
  <BottomTabBar
    {...props}
    getButtonComponent={({ route }) => {
      if (
        (route.key === 'newTab' && !props.showNewTab) ||
        (route.key === 'oldTab' && props.hideOldTab)
      ) {
        return HiddenView
      }
      return TouchableWithoutFeedbackWrapper
    }}
  />
)

export default connect(
  state => ({ /* state that you need */ }),
  {},
)(TabBarComponent)

```

And then simply imported and used that in my Tabs component:



```
<Tabs
  key=""main""
  tabBarComponent={TabBarComponent} // the component defined above
  ...

```

**Detailed look at where these things are getting passed to**


Looking at [the line of the source](https://github.com/aksonov/react-native-router-flux/blob/b37a24b780d75879512903b5cf7ef297f5b3721a/src/navigationStore.js#L833) of react-native-router-flux, it is using `createBottomTabNavigator` from the `react-navigation` library, and passing no component if you do not pass a custom tabBarComponent. The `createBottomTabNavigator` method in `react-navigation` comes [from this line of the library](https://github.com/react-navigation/react-navigation/blob/4502061a1081704b34db82aa742a4ba8f7f04e4a/src/react-navigation.js#L113), and is actually defined in `react-navigation-tabs`. Now, we can [here](https://github.com/react-navigation/react-navigation-tabs/blob/master/src/navigators/createBottomTabNavigator.js#L48) see in `react-navigation-tabs` that if no tabBarComponent has been passed, it simply uses BottomTabBar, also defined in `react-navigation-tabs`. This `BottomTabBar`, in turn, [takes a custom tab button renderer through props](https://github.com/react-navigation/react-navigation-tabs/blob/master/src/views/BottomTabBar.js#L232), called `getButtonComponent`.


"
"Hyperparameter Tuning of Tensorflow Model
I've used Scikit-learn's GridSearchCV before to optimize the hyperparameters of my models, but just wondering if a similar tool exists to optimize hyperparameters for Tensorflow (for instance **number of epochs, learning rate, sliding window size etc.**)


And if not, how can I implement a snippet that effectively runs all different combinations?


","Another viable (and documented) option for grid search with Tensorflow is [Ray Tune](https://docs.ray.io/en/latest/tune.html). It's a scalable framework for hyperparameter tuning, specifically for deep learning/reinforcement learning.


You can try out [a fast tutorial here](https://github.com/ray-project/tutorial/tree/master/tune_exercises).


It also takes care of Tensorboard logging and efficient search algorithms (ie, [`HyperOpt`](http://hyperopt.github.io/hyperopt) integration and [HyperBand](https://people.eecs.berkeley.edu/%7Ekjamieson/hyperband.html)) in about 10 lines of Python.



```
from ray import tune

def train_tf_model(config):  
    for i in range(num_epochs):
        accuracy = train_one_epoch(model)
        tune.report(acc=accuracy)

tune.run(train_tf_model,
         config={
            ""alpha"": tune.grid_search([0.2, 0.4, 0.6]),
            ""beta"": tune.grid_search([1, 2]),
         })

```

(Disclaimer: I contribute actively to this project!)


"
"Kendo UI grid with MVC fails to use Template when ajax turned on
I am using KendoUI with MVC.NET to create a grid. The data source for the grid is an AJAX call back to my controller.


The TEMPLATE property on bound columns should allow me to format the columns with RAZOR templating.


This works until I ""turn"" on ajax. The code below illustrates the issue:



```
@using Kendo.Mvc.UI;
@model IEnumerable<Test>
@{
    ViewBag.Title = ""Test - Index"";
}

<h2>Index</h2>

@{
        //public class Test {
        //    public string A;
        //    public string B;
        //}

        var data = new List<Test>() { new Test() { A=""a1"", B=""b1""}, new Test() { A=""a2"",B=""b2""}};

        // This one fails to honor TEMPLATE
        Html.Kendo().Grid(data).Name(""test"").Columns(columns => {
                                                              columns.Bound(""A"").Template(@<div>@item.A TEMPLATE APPLIED</div>);
                                                              columns.Bound(""B"");
                                                          }).DataSource(ds=>ds.Ajax()).Render(); // THIS BREAKS IT!

        // This one honors template               
        Html.Kendo().Grid(data).Name(""test"").Columns(columns => {
                                                              columns.Bound(""A"").Template(@<div>@item.A TEMPLATE APPLIED</div>);
                                                              columns.Bound(""B"");
                                                          }).Render();


    }

```

Two grids are displayed in this sample. The second grid does razor templating ok. The first fails to show the templatized data. The first grid just shows the original data.


NOTE: I've simplified this code as much as possible to illustrate the problem. Normally, after calling the Ajax() method, I also call other methods to setup the callback.


","As stated in the [documentation](http://docs.kendoui.com/getting-started/using-kendo-with/aspnet-mvc/helpers/grid/faq#how-do-i-customize-the-way-a-property-is-displayed-in-a-grid-bound-column?), for Ajax datasource you should use `ClientTemplate` instead of `Template`. Be careful, the syntax is slightly different :



```
Html.Kendo().Grid(data).Name(""test"").Columns(columns => {
    columns.Bound(""A"").ClientTemplate(""<div>#=A# TEMPLATE APPLIED</div>"");
    columns.Bound(""B"").ClientTemplate(""<div>#=A# AND #=B# JUST FOR KICKS</div>"");;
}).DataSource(ds=>ds.Ajax()).Render();

```

"
"Lock movement with PanResponder in react native
Using react native [PanResponder](https://facebook.github.io/react-native/docs/panresponder.html), how can I block the movement when the screen touch coordinates are outside of a certain range of values?


For example, how can I prevent users from moving a component below a certain y position on the screen?


The PanResponder uses the [Gesture Responder System](https://facebook.github.io/react-native/docs/gesture-responder-system.html).


I am carefully reading the documentation, but I cannot find the answer.


Any help is greatly appreciated.


","Looking at the PanResponder documentation page you provided (<https://facebook.github.io/react-native/docs/panresponder.html>), I think you could modify the example to meet your needs.


The function responsible for taking action in response to a gesture is a property of PanResponder called onPanResponderMove, in the example code from the documentation this looks like:



```
_handlePanResponderMove: function(e: Object, gestureState: Object) {
    this._circleStyles.style.left = this._previousLeft + gestureState.dx;
    this._circleStyles.style.top = this._previousTop + gestureState.dy;
    this._updateNativeStyles();
},

```

Where the gestureState object looks like this:



```
stateID - ID of the gestureState- persisted as long as there at least one touch on screen
moveX - the latest screen coordinates of the recently-moved touch
moveY - the latest screen coordinates of the recently-moved touch
x0 - the screen coordinates of the responder grant
y0 - the screen coordinates of the responder grant
dx - accumulated distance of the gesture since the touch started
dy - accumulated distance of the gesture since the touch started
vx - current velocity of the gesture
vy - current velocity of the gesture
numberActiveTouches - Number of touches currently on screen

```

You might add a conditional check in \_handlePanResponderMove to determine if the gestureState.y0 is below some threshold, and only apply a change if so


"
"turn off axis, keep ticks
I am plotting an image using plt.imshow() using the seaborn add on....



```
from astropy.io import fits
import numpy as np
import matplotlib.pyplot as plt
import seaborn as sns

sns.set_style(style='white') #turn off grid

def plotter():
    mapy = np.zeros((100,100))
    pf = 2.8 #my physical pixel size
    areaX = mapy.shape[0]/2*pf # half of the area!
    areaY = mapy.shape[1]/2*pf # half of the area!
    cmapz ='Reds'
    fig = imshow(mapy,interpolation='spline16',origin='lower',cmap=cmapz,extent=[-areaX*pf,areaX*pf,-areaY*pf,areaY*pf])
    plt.grid(False)
    plt.show()

```

I now want to get rid off the ugly axis around my plot, but **KEEP** the ticks in place as they refer to the distance.  

Everything I found and tried, e.g. `plt.axis('off')`, `fig.axes.get_yaxis().set_visible(False)`, `plt.tick_params(top='off', bottom='on', left='off', right='off', labelleft='on', labelbottom='on')` etc. turns off both the axis **AND** the ticks. 


","You may decide not to use seaborn and turn the axes spines invisible:



```
for d in [""left"", ""top"", ""bottom"", ""right""]:
    plt.gca().spines[d].set_visible(False)

```




```
import numpy as np
import matplotlib.pyplot as plt

mapy = np.random.rand(100,100)
pf = 2.8
areaX = mapy.shape[0]/2*pf # half of the area!
areaY = mapy.shape[1]/2*pf # half of the area!

fig = plt.imshow(mapy,interpolation='spline16',origin='lower',
                 cmap='Reds',extent=[-areaX*pf,areaX*pf,-areaY*pf,areaY*pf])
plt.grid(False)

for d in [""left"", ""top"", ""bottom"", ""right""]:
    plt.gca().spines[d].set_visible(False)

plt.show()
```






[![enter image description here](https://i.stack.imgur.com/AGUgj.png)](https://i.stack.imgur.com/AGUgj.png)


The same can be done, using rcParams, 



```
s = {""axes.spines.left""   : False,
    ""axes.spines.bottom"" : False,
    ""axes.spines.top""    : False,
    ""axes.spines.right""  : False}
plt.rcParams.update(s)

```




```
import numpy as np
import matplotlib.pyplot as plt
s = {""axes.spines.left""   : False,
""axes.spines.bottom"" : False,
""axes.spines.top""    : False,
""axes.spines.right""  : False}
plt.rcParams.update(s)

mapy = np.random.rand(100,100)
pf = 2.8
areaX = mapy.shape[0]/2*pf # half of the area!
areaY = mapy.shape[1]/2*pf # half of the area!

fig = plt.imshow(mapy,interpolation='spline16',origin='lower',
             cmap='Reds',extent=[-areaX*pf,areaX*pf,-areaY*pf,areaY*pf])
plt.grid(False)

plt.show()
```






Alternatively you can set the axes edgecolor to transparent. 



```
plt.rcParams[""axes.edgecolor""]=(1,1,1,0)

```




```
import numpy as np
import matplotlib.pyplot as plt
plt.rcParams[""axes.edgecolor""]=(1,1,1,0)

mapy = np.random.rand(100,100)
pf = 2.8
areaX = mapy.shape[0]/2*pf # half of the area!
areaY = mapy.shape[1]/2*pf # half of the area!

fig = plt.imshow(mapy,interpolation='spline16',origin='lower',
             cmap='Reds',extent=[-areaX*pf,areaX*pf,-areaY*pf,areaY*pf])
plt.grid(False)

plt.show()
```






Or, if you want to use seaborn (and it's white-style), additionally reactivate the ticks using 



```
plt.tick_params(axis=""both"", which=""major"", length=5)

```




```
import numpy as np
import matplotlib.pyplot as plt
import seaborn as sns

sns.set_style(style='white') #turn off grid

mapy = np.random.rand(100,100)
pf = 2.8
areaX = mapy.shape[0]/2*pf # half of the area!
areaY = mapy.shape[1]/2*pf # half of the area!

fig = plt.imshow(mapy,interpolation='spline16',origin='lower',
                 cmap='Reds',extent=[-areaX*pf,areaX*pf,-areaY*pf,areaY*pf])
plt.grid(False)

for d in [""left"", ""top"", ""bottom"", ""right""]:
    plt.gca().spines[d].set_visible(False)

plt.tick_params(axis=""both"", which=""major"", length=5)
plt.show()
```






[![enter image description here](https://i.stack.imgur.com/VdylG.png)](https://i.stack.imgur.com/VdylG.png)


As @mwaskom points out in the comments, Seaborn also offers `sns.despine()` to get rid of the spines, which you would then call like



```
sns.despine(left=True, top=True, bottom=True, right=True)

```

Note the double negation (**de**spine True means not to have spines).





```
import numpy as np
import matplotlib.pyplot as plt
import seaborn as sns

sns.set_style(style='white') #turn off grid

mapy = np.random.rand(100,100)
pf = 2.8
areaX = mapy.shape[0]/2*pf # half of the area!
areaY = mapy.shape[1]/2*pf # half of the area!

fig = plt.imshow(mapy,interpolation='spline16',origin='lower',
             cmap='Reds',extent=[-areaX*pf,areaX*pf,-areaY*pf,areaY*pf])
plt.grid(False)
sns.despine(left=True, top=True, bottom=True, right=True)

plt.tick_params(axis=""both"", which=""major"", length=5)
plt.show()
```






"
"How to send a message to workstation issuing Capistrano task?
If I execute `cap deploy` from a workstation how can I specify custom feedback to be displayed in the workstation's console as the recipe runs? 


It's confusing to me because Capistrano runs its 'tasks' on the deployment machines (or servers) so that `run 'echo ""this is my message""'` would just appear in the server's console. 


","Can't you just use puts/print? 


# Shell commands


Anyway, if you need to run commands in the workstation issuing the task, you should use the backtick operator. Even Capistrano documentation recommends it. See the Caveats section, at the bottom of this [page](https://github.com/capistrano/capistrano/wiki/2.x-DSL-Action-Invocation-Run ""Caveats"").


Quoting:



> 
> The run helper may not be appropriate for running commands which do not need to be executed remotely. In such instances, it may be more appropriate to utilize the backtick (```) instead.
> 
> 
> 


The backtick operator executes system commands and returns the output to you. For example, if you want to execute `echo ""this is my message""`, you would do: ``echo ""this is my message""``


That would call the echo ""system command"", and return the output to **you**: `""this is my message\n""`. But it wouldn't print it to the console, which is what you want.


To print it, you should use: `puts `echo ""this is my message""``  

But that's redundant, you might as well call puts or print directly.  

See [this link](http://www.devdaily.com/blog/post/ruby/how-execute-external-shell-command-expansion ""this link"") for more information about executing shell commands in Ruby.


# Puts and print


Puts and print prints a message to the console that is running the Ruby code. That seems like what you're trying to achieve. In your case, you should just do:



```
puts ""this is my message""
# or, depending on what you want
print ""this is my message""

```

`puts` is like `print`, but it adds a new line to the end of the string. This:



```
puts ""my message""

```

Is the same as this: 



```
print ""my message\n""

```

For some documentation, follow these links: 


- [print](http://ruby-doc.org/core/classes/Kernel.html#M005952)
- [puts](http://ruby-doc.org/core/classes/Kernel.html#M005954)


## Anyway


Just use `puts ""this is my message""` and be happy.


"
"php exec: does not return output
I have this problem: 
On a ISS web server, windows 7 x64 professional, zend server installed. Running this command under php:



```
exec('dir',$output, $err);

```

`$output is empty, $err=1.` So exec is not returing the output, it seems has some errors.
Php `disable_functions` is empty, php is not on safe mode, is standard, I check all option.
It seems to be a general errors, even search on google do not give results.


Please write every one his experince and eventual solutions or workarounds.


","There are a few posts to the relevant sections of the PHP Manual such as [this one](http://www.php.net/manual/en/function.exec.php#85930):



> 
> I was having trouble using the PHP exec command to execute any batch
>  file. Executing other commands (i.e., ""dir"") works fine). But if I
>  executed a batch file, I receieved no output from the exec command. 
> 
> 
> The server setup I have consists of Windows Server 2003 server running
>  IIS6 and PHP 5.2.3. On this server, I have:
> 
> 
> 1. Granted execute permissions to the Internet User on c:\windows\system32\cmd.exe.
> 2. Granted Everyone->Full Control to the directory in which the batch file is written.
> 3. Granted Everyone->Full Control on the entire c:\cygwin\bin directory and its contents.
> 4. Granted the Internet User ""log on as batch"" permissions.
> 5. Specified the full path to each file being executed.
> 6. Tested these scripts running from the command line on the server and they work just fine.
> 7. Ensured that %systemroot%\system32 is in the system path.
> 
> 
> It turns out that even with all of the above in place on the server, I
>  had to specify the full path to cmd.exe in the exec call.
> 
> 
> When I used the call: `$output = exec(""c:\\windows\\system32\\cmd.exe`
> `/c $batchFileToRun"");`
> 
> 
> then everything worked fine. In my situation, `$batchFileToRun` was the
>  actual system path to the batch file (i.e., the result of a call to
>  realpath()).
> 
> 
> 


There are a few more on both the [`exec`](http://php.net/manual/en/function.exec.php) and [`shell_exec`](http://php.net/manual/en/function.shell-exec.php) manual pages. Perhaps following through them will get it up and working for you.


"
"Why does the positional parameter syntax `${18}` need braces?
I'm currently learning the command line and I'm learning Positional Parameters.
And I saw that the Positional Parameters end at nine, To use more than $9, I must put them in curly braces, like: ${18}. But what is the story of the nine Positional Parameters and not allowed to put more of them in shape:$18?


","The list of positional parameters can be as long as required and as current resource limits allows. This means that there may be well over 9 elements in the list. As you have already noticed, elements 10 and later may be accessed by adding braces around the number, as in `${12}`.


The reason the braces are needed to access elements beyond element 9 in most POSIX-like shells is because the POSIX standard says so.



> 
> A positional parameter is a parameter denoted by the decimal value represented by one or more digits, other than the single digit 0. The digits denoting the positional parameters shall always be interpreted as a decimal value, even if there is a leading zero. **When a positional parameter with more than one digit is specified, the application shall enclose the digits in braces.**
> 
> 
> 


([From here](https://pubs.opengroup.org/onlinepubs/9699919799/utilities/V3_chap02.html#tag_18_05_01), ""Shell Command Language: Positional Parameters"")


The word ""application"" in the text above means ""your script"".



> 
> The simplest form for parameter expansion is:
> 
> 
> `${parameter}`
> 
> 
> The value, if any, of `parameter` shall be substituted.
> 
> 
> The parameter name or symbol can be enclosed in braces, which are optional **except for positional parameters with more than one digit** or when parameter is a name and is followed by a character that could be interpreted as part of the name.
> 
> 
> 


([From here](https://pubs.opengroup.org/onlinepubs/9699919799/utilities/V3_chap02.html#tag_18_06_02), ""Shell Command Language: Parameter Expansion"")


Note also that addressing any other than the first few positional parameters *directly* is quite a rare thing to do. It's common to access `$1` and `$2`, but `$3` and the others see very little use. It's more common to iterate over the list, as in



```
for argument do
    # use ""$argument"" here
done

```

... or to use the whole list at once with `""$@""`.


Also note that some shells may choose to ignore the standard. Here's the `zsh` shell, which does not try to be strictly POSIX compliant:



```
$ set -- a b c d e f g h i j k l m n o p
$ print $13
m
$ print $16
p

```

... unless it's emulating `sh`:



```
$ emulate sh
$ print $13
a3

```

The `a3` string comes from outputting `$1` followed by `3`.


"
"how to list rooms on socket.io nodejs server
after progress on the question [how to create socket.io multicast groups](https://stackoverflow.com/questions/6616922/how-to-create-socket-io-multicast-groups/6624604#6624604), I found making rooms a great way to do what I needed.


However, it would be great to know about all the rooms, without an extra data-structure.


Is it possible to get a list of all rooms on the server from the server socket?


","The short answer:



```
io.sockets.adapter.rooms

```

I analysed `io`:


I got the following output:



```
{ server: 
   { stack: 
      [ [Object],
        [Object],
        [Object],
        [Object],
        [Object],
        [Object] ],
     connections: 3,
     allowHalfOpen: true,
     watcher: { host: [Circular], callback: [Function] },
     _events: 
      { request: [Function],
        connection: [Function: connectionListener],
        listening: [Object],
        upgrade: [Object] },
     httpAllowHalfOpen: false,
     cache: {},
     settings: { home: '/', env: 'development', hints: true },
     redirects: {},
     isCallbacks: {},
     _locals: { settings: [Object], app: [Circular] },
     dynamicViewHelpers: {},
     errorHandlers: [],
     route: '/',
     routes: 
      { app: [Circular],
        routes: [Object],
        params: {},
        _params: [],
        middleware: [Function] },
     router: [Getter],
     __usedRouter: true,
     type: 'tcp4',
     fd: 7 },
  namespaces: 
   { '': 
      { manager: [Circular],
        name: '',
        sockets: [Object],
        auth: false,
        flags: [Object],
        _events: [Object] } },
  sockets: 
   { manager: [Circular],
     name: '',
     sockets: { '210837319844898486': [Object] },
     auth: false,
     flags: { endpoint: '', exceptions: [] },
     _events: { connection: [Function] } },
  settings: 
   { origins: '*:*',
     log: true,
     store: 
      { options: undefined,
        clients: [Object],
        manager: [Circular] },
     logger: { colors: true, level: 1 },
     heartbeats: true,
     resource: '/socket.io',
     transports: 
      [ 'websocket',
        'htmlfile',
        'xhr-polling',
        'jsonp-polling' ],
     authorization: [Function],
     'log level': 1,
     'close timeout': 25,
     'heartbeat timeout': 15,
     'heartbeat interval': 20,
     'polling duration': 20,
     'flash policy server': true,
     'flash policy port': 843,
     'destroy upgrade': true,
     'browser client': true,
     'browser client minification': false,
     'browser client etag': false,
     'browser client handler': false,
     'client store expiration': 15 },
  handshaken: 
   { '210837319844898486': 
      { headers: [Object],
        address: [Object],
        time: 'Mon Jul 18 2011 00:53:27 GMT+0200 (CEST)',
        xdomain: false,
        secure: undefined,
        PHPSESSID: '7qo6cht3q0rskhfes4eesb2d05' } },
  connected: { '210837319844898486': true },
  open: { '210837319844898486': true },
  closed: {},
  closedA: [],
  rooms: 
   { '': [ '210837319844898486' ],
     '/public-alfred': [ '210837319844898486' ] },
  roomClients: { '210837319844898486': [ '': true, '/public-alfred': true ] },
  oldListeners: [ [Function] ],
  _events: 
   { 'set:origins': [Function],
     'set:flash policy port': [Function],
     'set:transports': [Function] } }

```

after joining room ""public-alfred"" from a single client `io.sockets.adapter.rooms` contained:



```
 { '': [ '210837319844898486' ],
         '/public-alfred': [ '210837319844898486' ] }

```

"
"How to repeat prompt to user in a shell script?
In a set of if/elif/else/fi statements, I have made 'else' present the user with an error message, but I also want it to take the user back to the question which was asked before the if/else statements so that they can try to answer it again.


**How do I take the user back to a previous line of code? Or, if this is not possible, is there another way to do this?**


","I think the easiest way would be to wrap the prompting code into a function, and then drop it into an `until` loop.


Since all you need really is to call the function until it succeeds, you can put the noop command ""`:`"" in the until loop.


Something like this:



```
#!/bin/bash

getgender() {
  read -p ""What is the gender of the user? (male/female): "" gender
  case ""$gender"" in
    m|M)
      grouptoaddto=""boys""
      return 0
      ;;
    f|F)
      grouptoaddto=""girls""
      return 0
      ;;
    *)
      printf %s\\n ""Please enter 'M' or 'F'""
      return 1
      ;;
  esac
}

until getgender; do : ; done
sudo usermod -a -G ""$grouptoaddto"" ""$username""

```



---


The point here is the function called with `until`, so it is repeatedly called until it succeeds. The case switch within the function is just an example.




---


Simpler example, without using a function:



```
while [ -z ""$groupname"" ]; do
  read -p ""What gender is the user?"" answer
  case ""$answer"" in
    [MmBb]|[Mm]ale|[Bb]oy) groupname=""boys"" ;;
    [FfGg]|[Ff]emale|[Gg]irl) groupname=""girls"" ;;
    *) echo ""Please choose male/female (or boy/girl)"" ;;
  esac
done
sudo usermod -a -G ""$groupname"" ""$username""

```

In this last example, I'm using the `-z` switch to the `[` (test) command, to continue the loop as long as the ""groupname"" variable has zero length.


The keynote is the use of `while` or `until`.


To translate this last example into human readable pseudocode:



```
While groupname is empty,

  ask user for gender.

  If he answers with one letter ""m"" or ""B"",
    or the word ""Male"" or ""boy"",
    set the groupname as ""boys"".

  If she answers with one letter ""F"" or ""g"",
    or the word ""female"" or ""Girl"",
    set the groupname as ""girls"".

  If he/she answers anything else, complain.

(And then repeat, since groupname is still empty.)

Once you have groupname populated,
  add the user to that group.

```



---


Yet another example, without the `groupname` variable:



```
while true; do
  read -p ""What gender is the user?"" answer
  case ""$answer"" in
    [MmBb]|[Mm]ale|[Bb]oy)
      sudo usermod -a -G boys ""$username""
      break
      ;;
    [FfGg]|[Ff]emale|[Gg]irl)
      sudo usermod -a -G girls ""$username""
      break
      ;;
    *) echo ""Please choose male/female (or boy/girl)"" ;;
  esac
done

```

"
"OpenCv Core.line draws white when color is expected
In an `OpenCv4Android` environment, when I create a `Mat` image and use `Core.line()` to draw on the image, it always shows a white instead of the color I specify. 


[![white square instead of green square](https://i.stack.imgur.com/wGYRe.jpg)](https://i.stack.imgur.com/wGYRe.jpg)


I have seen [a question related to gray scale](https://stackoverflow.com/questions/23406368/scalar0-255-0-does-not-give-green-colored-line-in-opencv ""gray scale SO question""), but the image I have has not been converted to gray.



```
public class DrawingTest extends AppCompatActivity {
    public static final Scalar GREEN = new Scalar(0,255,0);
    private RelativeLayout mLayout;
    private ImageView imageView;

    static {
        System.loadLibrary(""opencv_java"");
    }

    @Override
    protected void onCreate(Bundle savedInstanceState) {
        super.onCreate(savedInstanceState);
        setContentView(R.layout.activity_drawing_test);

        mLayout = (RelativeLayout) findViewById(R.id.activity_drawing_test);
        mLayout.setDrawingCacheEnabled(true);

        imageView = (ImageView) this.findViewById(imageView_dt);

        //test.jpg is in the drawable-nodpi folder, is an normal color jpg image.
        int drawableResourceId = getResources().getIdentifier(""test"", ""drawable"", getPackageName());
        Bitmap bitmap = BitmapFactory.decodeResource(getResources(), drawableResourceId);
        //Mat matImage = new Mat(); // Also white
        Mat matImage = new Mat(bitmap.getHeight(), bitmap.getWidth(), CV_8UC4);
        Utils.bitmapToMat(bitmap, matImage);


        // Attempt to draw a GREEN box, but it comes out white
        Core.line(matImage, new Point(new double[]{100,100}), new Point(new double[]{100, 200}), GREEN,4);
        Core.line(matImage, new Point(new double[]{100,200}), new Point(new double[]{200, 200}), GREEN,4);
        Core.line(matImage, new Point(new double[]{200,200}), new Point(new double[]{200, 100}), GREEN,4);
        Core.line(matImage, new Point(new double[]{200,100}), new Point(new double[]{100, 100}), GREEN,4);

        Bitmap bitmapToDisplay = Bitmap.createBitmap(matImage.cols(), matImage.rows(), Bitmap.Config.ARGB_8888);
        Utils.matToBitmap(matImage, bitmapToDisplay);
        imageView.setImageBitmap(bitmapToDisplay);
    }
}

```

","The problem is with the color you have initialized



```
public static final Scalar GREEN = new Scalar(0,255,0); 

```

as per this statement 



```
Mat matImage = new Mat(bitmap.getHeight(), bitmap.getWidth(), CV_8UC4);`

```

you are creating a 4-channel Mat, but initializing the `GREEN` Scalar with 3 components only, hence the 4th component, which defines the color of fourth channel of your line, which is set of default value `0` in your case.


So, what you are perceiving as white color is transparent in reality. You may fix this either by creating `matImage` with `CV_8UC3` or changing your `GREEN` scalar to be `public static final Scalar GREEN = new Scalar(0,255,0, 255);`


"
"NSScrollView in Cocoa
I am trying to get an NSSrollView with an NSTextField in it to work, however the scrollbars do not seem to respond to anything that I am coding.


I am declaring the NSScrollView as an IBOutlet, add it as a property and then synthesizing it. However, it's to no avail.


The scrollbars do appear when I resize, however they serve no function at the moment.


I have tried to use the apple documentation, but again, no joy there.


Any help? Thanks!


","As You was saying *""For my purposes, any container which scrolls will do the trick""* You can use `NSTextView` in an `NSScrollView`. And for setting text You need to use ***setString*** instead *setStringValue*.


Example how to set text in `NSTextView`:


**.h**



```
IBOutlet NSTextView *textView; 

```

**.m**



```
-(void)awakeFromNib {
    NSString *string = @""my text"";
    [textView setString:string];
}

```

  

Don't forget IBOutlet `NSTextView` not `NSScrollView`.


`NSTextView` is in `NSScrollView`:


![enter image description here](https://i.stack.imgur.com/7B15Y.jpg)


"
"Fetch value from multiple drop down with same name row wise
I have a table like below-


[![enter image description here](https://i.stack.imgur.com/30U4l.jpg)](https://i.stack.imgur.com/30U4l.jpg)


the code of this table in my view is as following-



```
 <table class=""table table-bordered"">
   <tr>
   <td>Course Name</td>
   <td>Class</td>
   <td>Subject Name</td>
   <td>Month</td>
   <td>View</td>
   </tr>
   <?php if(!empty($data)) { $i=1; foreach($data as $key => $value) { ?>
      <tr>
        <td><?php echo $value->course_name; ?></td>
        <td><?php echo $value->class_name; ?></td>
        <td><?php echo $value->subjectName; ?></td>
        <td><select class=""form-control"" id=""month"" data-parsley-required=""true"" data-parsley-error-message=""Please Select Class"">
             <option value="""" hidden=""hidden"">Select Month</option>
             <option value=""1"">January</option>
             <option value=""2"">Febuary</option>
             <option value=""3"">March</option>
             <option value=""4"">April</option>
             <option value=""5"">May</option>
             <option value=""6"">June</option>
             <option value=""7"">July</option>
             <option value=""8"">August</option>
             <option value=""9"">September</option>
             <option value=""10"">October</option>
             <option value=""11"">November</option>
             <option value=""12"">December</option>
            </select>
          </td>
          <td>
           <button id=""<?php echo $i;?>"" class=""form-control"" data-eid=""<?php echo $this->session->userdata('e_id');?>""                                                data-subjectid=""<?php echo $value->subjectId;?>"" data-courseid=""<?php echo $value->course_id;?>"" data-classid=""<?php echo $value->class_id;?>"">
     View</button></td>
 </tr>
  <?php $i++; } } ?>
  </table>

```

//javascript



```
<script type=""text/javascript"">
$('button').on('click',function(){
   var eid = $(this).data('eid');
   var subId = $(this).data('subjectid');
   var courseId = $(this).data('courseid');
   var classId = $(this).data('classid');
   var month = $('#month').val();
   if(month == '')
    alert(""Select Month"");
   else { alert(""okay""); }       
</script>

```

Whenever I select `month` and click the `view` button the page is redirected correctly, but when I select `month` in the second row and click the `view` button again it gives alert `select month`.


I don't know how to differentiate all the dropdowns based on a name.


","The issue is because you're repeating the same `id` attributes in the elements you create in the loop. These need to be unique within the DOM. 


To fix the issue you need to change the `id` to `class`. Then you can use jQuery's DOM traversal methods to find the month `select` which is in the same row as the button which was clicked. To do that use `closest()` to get the parent `tr`, then `find()`:



```
<select class=""form-control month"" data-parsley-required=""true"" ... >

```


```
$('button').on('click', function() {
  var eid = $(this).data('eid');
  var subId = $(this).data('subjectid');
  var courseId = $(this).data('courseid');
  var classId = $(this).data('classid');

  var month = $(this).closest('tr').find('.month').val();
  if (month == '')
    alert(""Select Month"");
  else 
    alert(""okay"");  
});

```

"
"How do I hide virtual keyboard for select element in Win8 JavaScript app?
I'm creating a Windows 8 application in JavaScript. I have these html elements in the markup:



```
<input id=""myField1"" type=""number""></input>
<select id=""myField2"">
    <option value=""one"">One</option>
    <option value=""two"">Two</option>
    <option value=""three"">Three</option>
</select>

```

When I tap `myField1` on the device without physical keyboard, virtual keyboard appears. After that if I tap on the `myField2`, the keyboard remains in place. But I don't want to have typeahead in this particular `<select>` field. Is there a way to hide the virtual keyboard for it, enabling it only for `<input>` fields?


","As per the touch keyboard documentation, touch keyboard show and hide behavior is designed to be consistent across apps - keeping in focus the user. That is why show and hide methods are not available on the [Windows.UI.ViewManagement.InputPane class](http://msdn.microsoft.com/en-us/library/windows/apps/windows.ui.viewmanagement.inputpane.aspx).


This [touch keyboard whitepaper](http://msdn.microsoft.com/en-us/library/windows/apps/hh465404.aspx) describes the details, and calls out why the behavior to persist the keyboard while navigating/tabbing between input elements has been designed this way.



> 
> There is also a set of controls that can receive focus during a text
>  entry flow but that aren't editable. Rather than needlessly churn the
>  UI and potentially disorient the user in the middle of a flow, the
>  touch keyboard remains in view on controls because the user is likely
>  go back and forth between the controls and text entry with the touch
>  keyboard. Therefore, if the keyboard is already showing and focus
>  lands on a control of one of the types in the following list, the
>  keyboard will not hide itself. However, if the keyboard was not
>  already showing, it will not show itself unless the tap lands on an
>  editable region.
> 
> 
> 


In summary, as per the guidelines - apps should stay with the default touch keyboard. Think hard before trying change the default behavior. 


If you hit a scenario where it is really looking out of place to have the keyboard, the whitepaper also gives clues for a workaround (not intended). One can put the focus programmatically temporarily on a non-editable element (say button) and the keyboard will hide itself. But the select control will miss the tap event because of programmatic shift of focus, and will require another tap to open its drop-down. May be with more effort, this workaround can be improved. But again, this is a workaround and should be used judiciously.



```
_initializeEventHandlers: function initializeEventHandlers()
{
    myField2.addEventListener('focus', this._onfocus.bind(this));
},
_onfocus: function onfocus(event)
{
    console.log('on focus invoked');
    if (this._movingFocusInProgress)
    {
        this._movingFocusInProgress = false;
        return;
    }

    this._movingFocusInProgress = true;
    b1.focus();
    WinJS.Promise.timeout(200).then(function ()
    {
        myField2.focus();
    });
},

```

HTML will need to have this zero size button.



```
<input id=""myField1"" type=""number"" />
<select id=""myField2"" role=""banner"">
    <option value=""one"">One</option>
    <option value=""two"">Two</option>
    <option value=""three"">Three</option>
</select>
<button id=""b1"" style=""visibility:visible; border-width:0;
    border-color:transparent;outline-color:transparent; min-width:0; min-height:0""></button>

```

"
"Android 12 ForegroundServiceStartNotAllowedException while in Foreground?
While it seems there are a few questions on here regarding Android 12 service restrictions, I have yet to see one specific to this use case.


Background: I am starting a foreground service to handle a very long running background media player (Exoplayer). Due to the setup of the app, I cannot use the built in notification manager from exoplayer. However I am seeing an odd behavior from Crashlytics. I am getting `ForegroundServiceStartNotAllowedException` While the app is undoubtedly in the foreground. Per the logs, it is easy to see the user is navigating the app within a second of the `startForeground` call.


I am also listening to



```
override fun onStateChanged(source: LifecycleOwner, event: Lifecycle.Event)

```

to ensure the app is in a foreground state.


[![logs](https://i.stack.imgur.com/cuxXK.png)](https://i.stack.imgur.com/cuxXK.png)


[![crash](https://i.stack.imgur.com/yqhGu.png)](https://i.stack.imgur.com/yqhGu.png)


I am beginning to run out of ideas as to what would cause this. As you can see by the timestamps, less than 1 second had gone by. The app has no use cases where the service could even accidentally be started from the background.


Thanks


Edit:


I am also declaring my service type:



```
        <service
            android:name="".service.SoundPlayerService""
            android:enabled=""true""
            android:exported=""false""
            android:foregroundServiceType=""mediaPlayback""
            android:stopWithTask=""false"">
            <intent-filter>
                <action android:name=""android.media.browse.MediaBrowserService"" />
            </intent-filter>
        </service>

```


```
        if (!audioFocusTakenBackground && Application.instance.isAppInForeground) {
            if (Build.VERSION.SDK_INT >= Build.VERSION_CODES.Q) {
                startForeground(
                    ONGOING_NOTIFICATION_ID,
                    builder.build(),
                    ServiceInfo.FOREGROUND_SERVICE_TYPE_MEDIA_PLAYBACK
                )
            } else {
                startForeground(
                    ONGOING_NOTIFICATION_ID,
                    builder.build()
                )
            }
        }

```

Edit: Device State tab suggested by @Arlind


[![Device States](https://i.stack.imgur.com/vayHm.png)](https://i.stack.imgur.com/vayHm.png)


","One thing that's looking promising (I won't mark anything as answered until I get more data in) is by adding this bit to the notification builder:



```
if (Build.VERSION.SDK_INT >= Build.VERSION_CODES.S) {
    foregroundServiceBehavior = Notification.FOREGROUND_SERVICE_IMMEDIATE
}

```

Then to prevent crashes but still be notified of an issue I wrapped the `startForeground` in a try/catch that posts to crashlytics so i know there's an issue:



```
        if (!audioFocusTakenBackground && Application.instance.isAppInForeground && !isForegroundActive) {
            isForegroundActive = if (Build.VERSION.SDK_INT >= Build.VERSION_CODES.Q) {
                try {
                    startForeground(ONGOING_NOTIFICATION_ID, builder.build(), ServiceInfo.FOREGROUND_SERVICE_TYPE_MEDIA_PLAYBACK)
                    true
                } catch (e: Exception) {
                    Timber.e(""Unable to start foreground: ${e.message}"")
                    FirebaseCrashlytics.getInstance().recordException(e)
                    false
                }
            } else {
                try {
                    startForeground(ONGOING_NOTIFICATION_ID, builder.build())
                    true
                } catch (e: Exception) {
                    Timber.e(""Unable to start foreground: ${e.message}"")
                    FirebaseCrashlytics.getInstance().recordException(e)
                    false
                }
            }
        }

```

Note that in the area where `Build.VERSION.SDK_INT >= Build.VERSION_CODES.Q`, it adds that the notification is a `FOREGROUND_SERVICE_TYPE_MEDIA_PLAYBACK`. According to the google docs here: <https://developer.android.com/guide/components/foreground-services#notification-immediate> this should be enough to initially show the notification. I'm curious if some implementations don't use or listen to that.


I will update this when I get more data, but for now my crashes have subsided and I am not seeing any logging from the catch blocks in crashlytics. Hopefully someone else can use this and verify my findings.


EDIT:


After a few days my crashes from this error have gone down to 0 and no noticeable logs from the catch's have been observed. Finally calling this one closed


"
"Using Elasticsearch.Net/NEST to search parent documents based on child attributes, where parent/child documents are stored separately
I would like to use Elasticsearch.Net/NEST to search related documents. For example, I have:



```
Person:

id name address_id
-- ---- ----------
1  John 1
2  Mary 2

Address:

id city
-- ------
1  Boston
2  Berlin

```

I'd like to store the Person and Address documents separately, and do queries where I return Person documents based on Address fields. For example, return all documents for people living in Boston. I've seen some examples in the Elaticsearch documentation using mapping and parent/child directives, but nothing for Elasticsearch.Net/NEST. Any code samples or pointers would be greatly appreciated...


","Here's a small snippet where the address is the parent


EDIT:
Create the index:



```
var indicesOperationResponse = _client.CreateIndex(ci => ci.Index(""test"")
            .AddMapping<Address>(m => m.MapFromAttributes())
            .AddMapping<Person>(m => m.MapFromAttributes().SetParent<Address>()));

```

Index documents:



```
var bulkResponse = _client.Bulk(b => b
            .Index<Address>(bd => bd.Object(new Address { Name = ""Tel Aviv"", Id = 1 }).Index(""test""))
            .Index<Person>(bd => bd.Index(""test"").Object(new Person {Id = 5, Address = 1, Name = ""Me""}).Parent(1)));

```

And search by parent



```
var searchResponse = _client.Search<Person>(s => s
        .Query(q=>q.MatchAll())
        .Filter(q => q
            .HasParent<Address>(c => c
                .Query(cq => cq.Match(m=>m.OnField(t => t.Name).Query(""Tel Aviv""))))));

```

"
"What is the difference in approach to create DispatchQueue Swift3
I am a rookie in Swift, and there is such misunderstandings 


what is the difference how to create dispatch queue


sample 1



```
let backgroundQueue = DispatchQueue(label: ""com.app.queue"",
                                qos: .background, 
                                target: nil)

backgroundQueue.async {
print(""Dispatched to background queue"")
}

```

sample 2



```
let backgroundQueue = DispatchQueue.global()
backgroundQueue.async {
print(""Dispatched to background queue"")
}

```

as far as I understand this two approaches do the same 


or for example this approach 



```
DispatchQueue.global(qos: .userInitiated).async {
print(""user initiated task"")
}

```

what does it mean? 


","The queue you create in your first example is your own custom serial queue. As the somewhat dated, yet still relevant, [Concurrency Programming Guide](https://developer.apple.com/library/content/documentation/General/Conceptual/ConcurrencyProgrammingGuide/OperationQueues/OperationQueues.html#//apple_ref/doc/uid/TP40008091-CH102-SW2) says:



> 
> Serial queues (also known as *private dispatch queues*) execute one task at a time in the order in which they are added to the queue. The currently executing task runs on a distinct thread (which can vary from task to task) that is managed by the dispatch queue. Serial queues are often used to synchronize access to a specific resource.
> 
> 
> You can create as many serial queues as you need, and each queue operates concurrently with respect to all other queues. In other words, if you create four serial queues, each queue executes only one task at a time but up to four tasks could still execute concurrently, one from each queue.
> 
> 
> 


Whereas your latter examples are using simply retrieving system-provided global queues which are concurrent:



> 
> Concurrent queues (also known as a type of *global dispatch queue*) execute one or more tasks concurrently, but tasks are still started in the order in which they were added to the queue. The currently executing tasks run on distinct threads that are managed by the dispatch queue. The exact number of tasks executing at any given point is variable and depends on system conditions.
> 
> 
> 


Now, you can nowadays create your own custom concurrent queue, but a global queue is simply a concurrent queue that was created for us. 


So, what does this mean to you?


- If you dispatch blocks to serial queue (your first example), only one block can run at any time. This makes it really useful for synchronizing memory access in multi-threaded apps, but can be used in any environment where you need a background queue, but you want dispatched blocks to be run serially (i.e. sequentially) with respect to other blocks on that queue.
- The global queues that you are using in your latter examples are concurrent queues. This means that if you dispatch four separate tasks to this global queue, those blocks may run concurrently with respect to each other). This is ideal where you really want not only background execution, but don't care if these blocks also run at the same time as other dispatched blocks, too.
- In your latter examples, where you're accessing a global queue, recognize that because those are system-provided, you have some modest limitations on your interaction with these queues, namely:


	- You cannot suspend global queues;
	- You cannot use barriers on global queues;But, with that notwithstanding, if you are just looking for an simple way of dispatching blocks to run in the background (and you don't care if those dispatched blocks run at the same time as each other), then global queues are incredibly simple and efficient way to do that.




---


By the way, the difference between your second example (for which I assume you intended `let backgroundQueue = DispatchQueue.global()`) and the third example, is merely that in your third example, you assigned the explicit quality of service (qos), whereas in your second example, you're using the default qos. FWIW, it's generally advisable to specify a qos, so that the OS can prioritize threads contending for limited system resources accordingly.


"
"Web Workers - do they create actual threads?
I have always thought that web workers create separate threads, but today I ran into the spec on w3c website. Below is a citation about web workers:



> 
> This allows for thread-like operation with message-passing as the
>  coordination mechanism.
> 
> 
> 


The question is - if it is **thread-like**, not actual thread what is an advantage(performance wise) of using this technology?


Any help will be appreciated!


","Yes, web workers create actual threads (or processes, the spec is flexible on this). According to the [Web Workers](https://www.w3.org/TR/workers/) specification, when a worker is created the first step is:



> 
> 1. Create a separate parallel execution environment (i.e. a separate thread or process or equivalent construct), and run the rest of these steps in that context.
> 
> 
> For the purposes of timing APIs, this is the official moment of creation of the worker.
> 
> 
> (W3C Web Workers specification section 4.4)
> 
> 
> 


So it is explicitly specified that code running in Web Workers run in actual threads or processes.


Although it is possible to implement Workers without threads (note the ""equivalent construct"" language) for use on systems that don't support threads, all browser implementations implement Web Workers as threads.


"
"Python - decorator - trying to access the parent class of a method
This doesn't work:



```
def register_method(name=None):
    def decorator(method):
        # The next line assumes the decorated method is bound (which of course it isn't at this point)
        cls = method.im_class
        cls.my_attr = 'FOO BAR'
        def wrapper(*args, **kwargs):
            method(*args, **kwargs)
        return wrapper
    return decorator

```

Decorators are like the movie Inception; the more levels in you go, the more confusing they are. I'm trying to access the class that defines a method (at definition time) so that I can set an attribute (or alter an attribute) of the class.


Version 2 also doesn't work:



```
def register_method(name=None):
    def decorator(method):
        # The next line assumes the decorated method is bound (of course it isn't bound at this point).
        cls = method.__class__  # I don't really understand this.
        cls.my_attr = 'FOO BAR'
        def wrapper(*args, **kwargs):
            method(*args, **kwargs)
        return wrapper
    return decorator

```

The point of putting my broken code above when I already know why it's broken is that it conveys what I'm trying to do.


","I don't think you can do what you want to do with a decorator (quick edit: with a decorator of the method, anyway). The decorator gets called when the method gets constructed, which is *before* the class is constructed. The reason your code isn't working is because the class doesn't exist when the decorator is called.


jldupont's comment is the way to go: if you want to set an attribute of the *class*, you should either decorate the class or use a metaclass.


EDIT: okay, having seen your comment, I can think of a two-part solution that might work for you. Use a decorator of the method to set an attribute of the *method*, and then use a metaclass to search for methods with that attribute and set the appropriate attribute of the *class*:



```
def TaggingDecorator(method):
  ""Decorate the method with an attribute to let the metaclass know it's there.""
  method.my_attr = 'FOO BAR'
  return method # No need for a wrapper, we haven't changed
                # what method actually does; your mileage may vary

class TaggingMetaclass(type):
  ""Metaclass to check for tags from TaggingDecorator and add them to the class.""
  def __new__(cls, name, bases, dct):
    # Check for tagged members
    has_tag = False
    for member in dct.itervalues():
      if hasattr(member, 'my_attr'):
        has_tag = True
        break
    if has_tag:
      # Set the class attribute
      dct['my_attr'] = 'FOO BAR'
    # Now let 'type' actually allocate the class object and go on with life
    return type.__new__(cls, name, bases, dct)

```

That's it. Use as follows:



```
class Foo(object):
  __metaclass__ = TaggingMetaclass
  pass

class Baz(Foo):
  ""It's enough for a base class to have the right metaclass""
  @TaggingDecorator
  def Bar(self):
    pass

>> Baz.my_attr
'FOO BAR'

```

Honestly, though? Use the `supported_methods = [...]` approach. Metaclasses are cool, but people who have to maintain your code after you will probably hate you.


"
"Prevent short comments from wrapping in R Markdown/knitr output
Is it possible to get a code chunk `tidy`ed up, while leaving the comments alone?


Say I wanna put ASCII art in some comments in a function, I could just set `tidy = F` but then the rest of the code chunk is ""messy"" as in the following output: 



```
myfun=function(a,b){
  ##     ^_^
  ##    {o,o}
  ##    |)__)
  ##-----m-m-----
  c=sum(a,b)
  return(c) 
}

```

If I set `tidy = T` then the short comments all wrap together and I get the following as output:



```
myfun = function(a, b) {
    ## ^_^ {o,o} |)__) -----m-m-----
    c = sum(a, b)
    return(c)
}

```

What I would like to see is the following as output:



```
myfun = function(a, b) {
    ##      ^_^
    ##     {o,o} 
    ##     |)__)
    ## -----m-m-----
    c = sum(a, b)
    return(c)
}

```

","As suggested in the comments, I'll answer my own question. 


If one looks at Yihui's [documentation for `formatR`](https://github.com/yihui/formatR/wiki), one might notice that roxygen comments (which look like this: `#'`) will not be wrapped in any case.


So using the code chunk



```
```{r, tidy = T}
myfun=function(a,b){
  #'     ^_^
  #'    {o,o}
  #'    |)__)
  #'-----m-m-----
  c=sum(a,b)
  return(c)
}
```

```

will give me the desired output:



```
myfun = function(a, b) {
    #'     ^_^
    #'    {o,o}
    #'    |)__)
    #'-----m-m-----
    c = sum(a, b)
    return(c)
}

```

"
"Create a private network inside my Windows 7 machine
My Main Rig is a Windows 7 Professional laptop. On this machine is two Virtual Machines, both of which are also Windows 7 Professional. I need these VMs to be able to see each other but I also need them to access to the internet.


Currently they are all connected to the public internet at the hotel I am staying at but I am unable to see the other machines because this is a public network.


What I need to do is create a private network amongst my physical machine and the two virtual machines that reside within it. How can I do that?


I found how to make it think the public network a private network, but not how to create a private network that exists within the machine while still seeing the outside world.


My Google-fu hasn't revealed much - maybe I am using the wrong terms.


**UPDATE**: I added an additional NIC to both VM's and set them to local. I then started the VMs and tried to create a new network. However I can now see other machines on the hotel network and presumably they can see me (not desired). Can anyone give a detailed explanation of how to set this up or a link to a site that does? I can't be the only person who wants a private internal network on their PC?!


**UPDATE2**: [Found this](https://web.archive.org/web/20141226213852/http://blogs.msdn.com:80/b/smguest/archive/2005/02/10/370861.aspx) that describes what I want to do. Will try tomorrow to see if it works - seems I was almost there.


**UPDATE3**: Seems it didn't work. Added bounty which will be awarded for anyone who can do a step by step guide for getting two virtual machines to communicate. Extra cudos for using MySQL & Apache on the VMs!


","To have the VMs on a separate network that is private to your PC, use the Microsoft Loopback Adapter. The Microsoft Loopback adapter is a testing tool for a virtual network environment where network access is not available. It sets up a virtual network adapter that is private to your computer.


To install, see this : [How to install a Loopback Adapter in Windows 7](https://web.archive.org/web/20140716182022/http://www.windowsreference.com:80/windows-7/how-to-install-a-loopback-adapter-in-windows-7/).


After it is installed, from within Virtual PC :


1. Select your virtual machine
2. Go to the Settings
3. Click the ""Networking"" setting
4. For your network adapter, select ""Microsoft Loopback Adapter""


Now, you should be able to see your other running Virtual PC instances that also share the same loopback adapter, while this virtual network is inaccessible from outside your computer.


If you need to use Internet from within the VMs, see this article :  

[Configuring NAT via using the Microsoft Loopback Adapter and Internet Connection Sharing](https://web.archive.org/web/20151231160533/http://blogs.msdn.com/b/virtual_pc_guy/archive/2005/10/04/477195.aspx).


"
"Jquery | Function inside same function. It bad practice?
I am using this function for easiness, as I am going to use fadeTo a lot:



```
function fade_to(div, speed, opacity, after_fade) {
    $(div).fadeTo(speed, opacity, after_fade);
}

```

Then I am calling the same function for after\_fade parameter:



```
fade_to('#div', 3000, 1, function() { fade_to('#another_div', 3000, 1)});

```

Is that a bad thing to do? Will I have speed/smoothness issues?
Is it better to just use jQuery's default fadeTo function?


Thanks!


","There is no gain to be made with your method. Plus you are using the jQuery fadeTo function. There is nothing wrong with what you did, just no gain. You could save work with such a technique if for example you had less arguments in your custom function:



```
function fade_to(div, after_fade) {
    $(div).fadeTo(3000, 1, after_fade);
}

fade_to('#div', function(){ fade_to('#another_div', $.noop); });

```

This would actually save you work by preventing you from having to enter speed and opacity arguments. You could also curry it like this



```
function Fade_to(speed, opacity){
     return function(div, callback){
           $(div).fadeTo(speed, opacity, callback);
     }
}

```

Then you could make argument saving functions on the fly like



```
var fade_to_foo = Fade_to(3000, 1);
fade_to_foo('#div', function(){ fade_to_foo('#another_div'); });

```

Otherwise there is no reason not to just write it the jQuery way



```
$('#div').fadeTo(3000, 1, function(){ $('#another_div').fadeTo(3000, 1); });

```

"
"VS For Mac error generating android build - NDK
I installed Visual Studio for Mac and around to build the application for Android it gave me this error. I installed NDK and now this error appears to me.


Do you know how to solve this?



> 
> /Library/Frameworks/Mono.framework/External/xbuild/Xamarin/Android/Xamarin.Android.Common.targets(2873,3): error XA5101: C compiler for target Arm was not found. Tried paths: ""/Users/UserName/Library/Developer/Xamarin/android-sdk-macosx/ndk-bundle/toolchains/arm-linux-androideabi-4.9/prebuilt/darwin-x86\_64/bin/arm-linux-androideabi-gcc; /Users/UserName/Library/Developer/Xamarin/android-sdk-macosx/ndk-bundle/prebuilt/darwin-x86\_64/bin/gcc"" [/Users/UserName/Projetos/ProjectName/app/src/android/ProjectNameApp.Droid/ProjectName.Droid.csproj]
> 
> 
> 


An error occurred when executing task 'Release-Android'.
Teardown
Executing custom teardown action...
Environment: QA
Configuration: Release
Target: Release-Android
Error: One or more errors occurred. (MSBuild: Process returned an error (exit code 1).)
MSBuild: Process returned an error (exit code 1).


I have already seen several solutions but I can not solve it, I have already installed NDK, I have already validated the folder where it is installed, I have already inserted AndroidNdkDirectory in Droid.csproj and so far none of this has resulted.


[![I have already installed NDK](https://i.stack.imgur.com/VHS4f.png)](https://i.stack.imgur.com/VHS4f.png)
[![I have already validated the folder where it is installed](https://i.stack.imgur.com/ZnAT7.png)](https://i.stack.imgur.com/ZnAT7.png)
[![I have already validated the folder where it is installed](https://i.stack.imgur.com/eUfqf.png)](https://i.stack.imgur.com/eUfqf.png)
[![I have already inserted AndroidNdkDirectory in Droid.csproj](https://i.stack.imgur.com/l2jqn.png)](https://i.stack.imgur.com/l2jqn.png)


","There are two potential ways to resolve this issue:


- You can switch the Android NDK version to r17 or lower. One way to do that is to un-check **NDK** from the Android SDK manager (the **Tools** tab from your screenshot) to uninstall version r19, then select the gear icon from the lower right of the window, pick the **Microsoft (Recommended)** repository, and check **NDK** version r17 (which should now appear in the list) to install that version.
- Or, if you are working on code where you can use a preview version of Xamarin.Android, you can update to the current preview Xamarin.Android version 9.3.0.19 by switching to the **Preview** [updater channel](https://learn.microsoft.com/visualstudio/mac/update) in Visual Studio for Mac.


**Additional info**


The current Xamarin.Android SDK version 9.2.3.0 is not yet compatible with Android NDK versions r18 or higher, where [GCC has been removed](https://github.com/android-ndk/ndk/wiki/Changelog-r18).


Compatibility with those newer NDK versions (and in particular, compatibility with the change from `gcc` to `clang`) is [included in the preview Xamarin.Android SDK version 9.3](https://learn.microsoft.com/xamarin/android/release-notes/9/9.3#android-ndk-r19-compatibility).


"
"DirectX Shader Resource View in Shaders
I'm a bit confused right now and just want to ask you all to help me get a few ideas clarfied. 


In a HLSL shader (compute shader for example) I can declare a StructuredBuffer sb, right? Do I HAVE to bind it to a register, such as : register(t0)? 


From the application side, I can call CSSetShaderResources(...). The first argument (the StartSlot), does it have anything to do with the little number behind the ""t"" in the register declaration? 


If I set the StartSlot as 0 (for example), and I set the second argument as 2. Am I telling the API that I'll binding two Shader Resource Views, one in register(t0) and another in register(t1)?


If I declare Texture2D tex[10] : register(t0) I can set it by calling CSSetShaderResources(0, 10, ...). Does this mean that registers(t0~t9) are all used up?


Sorry for such a ""rapid fire"" of questions, but I'm just really confused and some of my tests seem to give confliting results...


Any help would be appreciated.


","So let's reply in order:


Yes you can of course declare a StructuredBuffer in a compute shader (actually you can declare it for any type of shader).


If you don't use effect framework (techniques), you need to declare a register, so the shader will know where to read from your data (using effect framework it just does it under the hood, but you still can explicitely declare).


CSSetShaderResources tells to bind N resources from a start slot, so your description of using 0,2 is correct.


For array of textures, I had to run PIX to check it out, but it's indeed the way you said.



```
Texture2D tex[10] : register(t0); 

```

Will mean that each texture index will be allocated a slot starting from the register you specified, so you need to call CSSetShaderResources(0,10,srvarray) to set them.


"
"Odoo 10 Developer mode and developer mode with assets
I'm new on Odoo v10. I'm using the debug mode with URL. But on the settings menu there is links to activate developer mode. I want to ask, 
What is the difference between ""Activate the developer mode"" and ""Activate developer mode ( with asset)""? 


","**Debug mode (most of the time used by Administrator)**
With this mode you can unable some advance features/options in odoo. so if you open odoo with debug mode you will see some extra menus, extra fields to more advance configuration.


**Debug with assets mode (most of the time used by developers)**
It enables all features from normal debug plus some extra. Odoo internally use assets bundles to improve performance which internal merge+minifies all static assets like css and JavaScript. So debug with assets will use actual css and javascript files instead of minifying/merging css and js files. This mode will reduce performance but it will give you proper error message in console if you are doing some JavaScript changes in odoo


**Debug with assets mode (tests assets)**
According to [this](https://www.odoo.com/documentation/user/14.0/general/developer_mode/activate.html#:%7E:text=Activate%20the%20developer%20mode%20%28with%20tests%20assets%29), starting from Odoo 13 (also available in Odoo 14), we have a new debug mode. This mode is mostly for developers & testers.


You can use this [**chrome plugin**](https://chrome.google.com/webstore/detail/odoo-debug/hmdmhilocobgohohpdpolmibjklfgkbi) to quick switch between all those modes.


"
"How to use LL (low level) drivers in CubeMX STM32?
I'm creating a blank project for a STM32F103 microcontroller using STM32CubeMX. Using HAL drivers (the default), I got a blinky example working pretty quickly, but I'd like to try out LL (low level) drivers instead of HAL. 


When I go to Project > Settings > Advanced Settings in CubeMX, I see the peripherals listed, and there is a pulldown next to each one which only has one option, HAL. According to the docs this is where I could select LL or HAL for each peripheral. 


How do I enable LL drivers in STM32CubeMX?


Details: I did a clean install of STM32CubeMX (4.23.0) on Ubuntu (16.04) and created a new project for STM32F103C8 microcontroller (this is on a ""Blue Pill"" board). I'm using SW4STM32 IDE, everything more or less worked out of the box except I had to select STLink V2 instead of V2.1 in a config file.


","Currently (STM32CubeMX v4.23.0) the LL drivers are only generated for L1, L4, F2, F4, F7 series. See the chapter 6.2 STM32Cube code generation using Low Layer drivers
For STM32L1, STM32L4, STM32L4+, STM32F2, STM32F4 and STM32F7 Series, 
STM32CubeMX allows generating peripheral initialization code based either on the 
peripheral HAL driver or on the peripheral Low Layer (LL) driver.


<http://www.st.com/content/ccc/resource/technical/document/user_manual/10/c5/1a/43/3a/70/43/7d/DM00104712.pdf/files/DM00104712.pdf/jcr:content/translations/en.DM00104712.pdf>


"
"Possible to defer loading of jQuery?
Let's face it, jQuery/jQuery-ui is a heavy download.


Google recommends [deferred loading of JavaScript](http://code.google.com/speed/page-speed/docs/payload.html#DeferLoadingJS) to speed up initial rendering. My page uses jQuery to set up some tabs which are placed low on the page (mostly out of initial view) and I'd like to defer jQuery until AFTER the page has rendered.


Google's deferral code adds a tag to the DOM after the page loads by hooking into the body onLoad event:



```
<script type=""text/javascript"">

 // Add a script element as a child of the body
 function downloadJSAtOnload() {
 var element = document.createElement(""script"");
 element.src = ""deferredfunctions.js"";
 document.body.appendChild(element);
 }

 // Check for browser support of event handling capability
 if (window.addEventListener)
 window.addEventListener(""load"", downloadJSAtOnload, false);
 else if (window.attachEvent)
 window.attachEvent(""onload"", downloadJSAtOnload);
 else window.onload = downloadJSAtOnload;

</script>

```

I'd like to defer loading of jQuery this way, but when I tried it my jQuery code failed to find jQuery (not completely unexpected on my part):



```
$(document).ready(function() {
    $(""#tabs"").tabs();
});

```

So, it seems I need to find a way to defer execution of my jQuery code until jQuery is loaded. How do I detect that the added tag has finished loading and parsing?


As a corollary, it appears that [asynchronous loading](http://code.google.com/speed/page-speed/docs/rtt.html#PreferAsyncResources) may also contain an answer.


Any thoughts?


","Try this, which is something I edited a while ago from the jQuerify bookmarklet. I use it frequently to load jQuery and execute stuff after it's loaded. You can of course replace the url there with your own url to your customized jquery. 



```
(function() {
      function getScript(url,success){
        var script=document.createElement('script');
        script.src=url;
        var head=document.getElementsByTagName('head')[0],
            done=false;
        script.onload=script.onreadystatechange = function(){
          if ( !done && (!this.readyState || this.readyState == 'loaded' || this.readyState == 'complete') ) {
            done=true;
            success();
            script.onload = script.onreadystatechange = null;
            head.removeChild(script);
          }
        };
        head.appendChild(script);
      }
        getScript('http://ajax.googleapis.com/ajax/libs/jquery/1.5.2/jquery.min.js',function(){
            // YOUR CODE GOES HERE AND IS EXECUTED AFTER JQUERY LOADS
        });
    })();

```

I would really combine jQuery and jQuery-UI into one file and use a url to it. If you REALLY wanted to load them separately, just chain the getScripts:



```
getScript('http://myurltojquery.js',function(){
        getScript('http://myurltojqueryUI.js',function(){
              //your tab code here
        })
});

```

"
"SwiftUI ButtonStyle - how to check if button is disabled or enabled?
To style a button in SwiftUI, according to my understanding, you extend `ButtonStyle` and implement `func makeBody(configuration: Self.Configuration) -> some View` within which you start applying modifications to the `configuration.label` which is a reference to the `Button` view.


Now, besides the `label` field, `ButtonStyle.Configuration` has a boolean field for `isPressed`, but that's seems to be all.


How do I check if the button is enabled or disabled?


For example, I want to draw a border around the button, and I want the border to be blue if the button is enabled and gray if disabled.


My first guess is something like this:



```
    func makeBody(configuration: Self.Configuration) -> some View {
        configuration.label
            .overlay(
                RoundedRectangle(cornerRadius: 4).stroke(configuration.isEnabled ? Color.blue : Color.gray, lineWidth: 1).padding(8)
            )
    }

```

But there's no `isEnabled` field.


The Apple-supplied `PlainButtonStyle` obviously has access to this information, as the doc comment in the header file that declares it says 'The style may apply a visual effect to indicate the pressed, focused, or enabled state of the button.'



```
/// A `Button` style that does not style or decorate its content while idle.
///
/// The style may apply a visual effect to indicate the pressed, focused,
/// or enabled state of the button.
@available(iOS 13.0, OSX 10.15, tvOS 13.0, watchOS 6.0, *)
public struct PlainButtonStyle : PrimitiveButtonStyle {
.....

```

Is there a way to access this information?


## EDIT:


Before you suggest closing this question, please read this:


There's a similar question asked a while ago. That question does not specify any context.


Here I have a specific context: creating a generic reusable button style.


The answer provided there does not suffice. I can't expect people to pass the disabled state to the style constructor. That's too much duplication of effort.


Can you imagine if people have to write code this way:


`Button() { .... }.disabled(someExprssion).buttonStyle(MyCustomStyle(disabled: someExpression))`


Clearly that's not desireable.


Also clearly the Apple supplied style has access to the information without requiring people to pass the disable state again to the style.


If you close the question, you will forever prevent anyone from providing a useful answer to this question on StackOverflow.


Please reconsider before you propose closing.


## EDIT2:


I have a guess that if you can get the `EnvironmentValues.isEnabled` of the `Button` view then it might be the right value.


So another way to ask the question is: is it possible in SwiftUI to get the environment of a view that you didn't define yourself?


","I found the answer thanks to this blog: <https://swiftui-lab.com/custom-styling/>


You can get the enabled state from the environment by creating a wrapper view and using it inside the style struct:



```
struct MyButtonStyle: ButtonStyle {
    func makeBody(configuration: ButtonStyle.Configuration) -> some View {
        MyButton(configuration: configuration)
    }

    struct MyButton: View {
        let configuration: ButtonStyle.Configuration
        @Environment(\.isEnabled) private var isEnabled: Bool
        var body: some View {
            configuration.label.foregroundColor(isEnabled ? Color.green : Color.red)
        }
    }
}

```

This example demonstrates how to get the state and use it to change the appearance of the button. It changes the button text color to red if the button is disabled or green if it's enabled.


"
"How to replace different blocks of text in all combinations using awk?
I'm trying to replace blocks of lines like this pattern:


- A block of lines is formed by the lines bellow which has an minor number.
- When a line has the ""="", then this block of lines could replace the block named after the ""=""


Let's see an example, this input:



```
01 hello
    02 stack
    02 overflow
        04 hi
    02 friends = overflow
        03 this
        03 is 
        03 my = is
        03 life
    02 lol
    02 im
    02 joking = im
        03 filler

```

Would generate the following ouput (each hello block is one element of an array):



```
01 hello
    02 stack
    02 overflow
        04 hi
    02 lol
    02 im

01 hello
    02 stack
    02 overflow
        04 hi
    02 lol
    02 joking = im
        03 filler

01 hello
    02 stack
    02 friends = overflow
        03 this
        03 is 
        03 life
    02 lol
    02 im

01 hello
    02 stack
    02 friends = overflow
        03 this
        03 is 
        03 life
    02 lol
    02 joking = im
        03 filler

01 hello
    02 stack
    02 friends = overflow
        03 this
        03 my = is
        03 life
    02 lol
    02 im

01 hello
    02 stack
    02 friends = overflow
        03 this
        03 my = is
        03 life
    02 lol
    02 joking = im
        03 filler

```

I tried it by this way:



```
#!/bin/bash

awk '{

    if ($0~/=/){
      level=$1
      oc=1
    }else if (oc && $1<=level){
        oc=0
    }

    if (!oc){
        print
    }

}' input.txt

```

But it only returns the first output that I need, and I don't know how to skip the `03 life` word which are within `friends`.


How could I generate these outputs?


I wouldn't mind a python or perl solution if is more confortable to you.


","Here is a python script to read the cobol input file and print out all the possible combinations of defined and redefined variables:



```
#!/usr/bin/python
""""""Read cobol file and print all possible redefines.""""""
import sys
from itertools import product

def readfile(fname):
    """"""Read cobol file & return a master list of lines and namecount of redefined lines.""""""
    master = []
    namecount = {}
    with open(fname) as f:
        for line in f:
            line = line.rstrip(' .\t\n')
            if not line:
                continue
            words = line.split()
            n = int(words[0])
            if '=' in words or 'REDEFINES' in words:
                name = words[3]
            else:
                name = words[1]
            master.append((n, name, line))
            namecount[name] = namecount.get(name, 0) + 1
    # py2.7: namecount = {key: val for key, val in namecount.items() if val > 1}
    namecount = dict((key, val) for key, val in namecount.items() if val > 1)

    return master, namecount

def compute(master, skip=None):
    """"""Return new cobol file given master and skip parameters.""""""
    if skip is None:
        skip = {}
    seen = {}
    skip_to = None
    output = ''
    for n, name, line in master:
        if skip_to and n > skip_to:
            continue
        seen[name] = seen.get(name, 0) + 1
        if seen[name] != skip.get(name, 1):
            skip_to = n
            continue
        skip_to = None
        output += line + '\n' 
    return output

def find_all(master, namecount):
    """"""Return list of all possible output files given master and namecount.""""""
    keys = namecount.keys()
    values = [namecount[k] for k in keys]
    out = []
    for combo in product(*[range(1, v + 1) for v in values]):
        skip = dict(zip(keys, combo))
        new = compute(master, skip=skip)
        if new not in out:
            out.append(new)
    return out

def main(argv):
    """"""Process command line arguments and print results.""""""
    fname = argv[-1]
    master, namecount = readfile(fname)
    out = find_all(master, namecount)
    print('\n'.join(out))

if __name__ == '__main__':
    main(sys.argv)

```

If the above script is save in a file called `cobol.py`, then if can be run as:



```
python cobol.py name_of_input_file

```

The various possible combinations of defines and redefines will be displayed on stdout.


This script runs under either python2 (2.6+) or python3.


### Explanation


The code uses three functions:


- `readfile` reads the input file and returns two variables that summarize the structure of what is in it.
- `compute` takes two parameters and, from them, computes an output block.
- `find_all` determines all the possible output blocks, uses `compute` to create them, and then returns them as a list.


Let's look at each function in more detail:


1. `readfile`


`readfile` takes the input file name as an argument and returns a list, `master`, and a dictionary, `namecount`. For every non-empty line in the input file, the list `master` has a tuple containing (1) the level number, (2) the name that is defined or redefined, and (2) the original line itself. For the sample input file, `readfile` returns this value for `master`:



```
[(1, 'hello', '01 hello'),
 (2, 'stack', '    02 stack'),
 (2, 'overflow', '    02 overflow'),
 (4, 'hi', '        04 hi'),
 (2, 'overflow', '    02 friends = overflow'),
 (3, 'this', '        03 this'),
 (3, 'is', '        03 is'),
 (3, 'is', '        03 my = is'),
 (3, 'life', '        03 life'),
 (2, 'lol', '    02 lol'),
 (2, 'im', '    02 im'),
 (2, 'im', '    02 joking = im'),
 (3, 'filler', '        03 filler')]

```

`readfile` also returns the dictionary `namecount` which has an entry for every name that gets redefined and has a count of how many definitions/redefinitions there are for that name. For the sample input file, `namecount` has the value:



```
{'im': 2, 'is': 2, 'overflow': 2}

```

This indicates that `im`, `is`, and `overflow` each have two possible values.


`readfile` was of course designed to work with the input file format in the current version of the question. To the extent possible, it was also designed to work with the formats from the previous versions of this question. For example, variable redefinitions are accepted whether they are signaled with an equal sign (current version) or with the word `REFDEFINES` as in previous versions. This is intended to make this script as flexible as possible.


2. `compute`


The function `compute` is what generates each output block. It uses two parameters. The first is `master` which comes directly from `readfile`. The second is `skip` which is derived from the `namecount` dictionary that was returned by `readfile`. For example, the `namecount` dictionary says that there are two possible definitions for `im`. This shows how `compute` can be used to generate the output block for each:



```
In [14]: print compute(master, skip={'im':1, 'is':1, 'overflow':1})
01 hello
    02 stack
    02 overflow
        04 hi
    02 lol
    02 im

In [15]: print compute(master, skip={'im':2, 'is':1, 'overflow':1})
01 hello
    02 stack
    02 overflow
        04 hi
    02 lol
    02 joking = im
        03 filler

```

Observe that the first call to `compute` above generated the block that uses the first definition of `im` and the second call generated the block that uses the second definition.


3. `find_all`


With the above two functions available, it is clear that the last step is just to generate all the different combinations of definitions and print them out. That is what the function `find_all` does. Using `master` and `namecount` as returned by `readfile`, it systematic runs through all the available combinations of definitions and calls `compute` to create a block for each one. It gathers up all the unique blocks that can be created this way and returns them.


The output returned by `find_all` is a list of strings. Each strings is the block which corresponds to one combination of defines/redefines. Using the sample input from the question, this shows what `find_all` returns:



```
In [16]: find_all(master, namecount)
Out[16]: 
['01 hello\n    02 stack\n    02 overflow\n        04 hi\n    02 lol\n    02 im\n',
 '01 hello\n    02 stack\n    02 friends = overflow\n        03 this\n        03 is\n        03 life\n    02 lol\n    02 im\n',
 '01 hello\n    02 stack\n    02 overflow\n        04 hi\n    02 lol\n    02 joking = im\n        03 filler\n',
 '01 hello\n    02 stack\n    02 friends = overflow\n        03 this\n        03 is\n        03 life\n    02 lol\n    02 joking = im\n        03 filler\n',
 '01 hello\n    02 stack\n    02 friends = overflow\n        03 this\n        03 my = is\n        03 life\n    02 lol\n    02 im\n',
 '01 hello\n    02 stack\n    02 friends = overflow\n        03 this\n        03 my = is\n        03 life\n    02 lol\n    02 joking = im\n        03 filler\n']

```

As an example, let's take the fourth string returned by `find_all` and, for better format, we will `print` it:



```
In [18]: print find_all(master, namecount)[3]
01 hello
    02 stack
    02 friends = overflow
        03 this
        03 is
        03 life
    02 lol
    02 joking = im
        03 filler

```

In the complete script, the output from `find_all` is combined together and printed to stdout as follows:



```
out = find_all(master, namecount)              
print('\n'.join(out))

```

In this way, the output displays all possible blocks.


## Answers for Earlier Versions of the Question


### Answer for Original Question



```
awk 'f==0 && !/REDEFINES/{s=s""\n""$0;next} /REDEFINES/{f=1;print s t>(""output"" ++c "".txt"");t=""""} {t=t""\n""$0} END{print s t>(""output"" ++c "".txt"")}' input

```

### Explanation:


This program has the following variables:


- `f` is a flag which is zero before the first REDEFINE and one thereafter.
- `s` contains all the text up to the first REDEFINE.
- `t` contains the text of the current REDEFINE.
- `c` is a counter which is used to determine the name of the output name.


The code works as follows:


1. `f==0 && !/REDEFINES/{s=s""\n""$0;next}`


Before the first redefine is encountered, the text is saved in the variable `s` and we skip the rest of the commands and jump to the `next` line.
2. `/REDEFINES/{f=1;print s t>(""output"" ++c "".txt"");t=""""}`


Every time that we encounter a REDEFINE line, we set the flag `f` to one and print the prolog section `s` along with the current REDEFINE section to a file named `outputn.txt` where n is replaced by the value of the counter `c`.


Because we are at the start of a new REDEFINE section, the variable `t` is set to empty.
3. `{t=t""\n""$0}`


Save the current line of this REDEFINE to the variable `t`.
4. `END{print s t>(""output"" ++c "".txt"")}`


The output file for the last REDEFINE section is printed.


### A Minor Improvement


Each of the output files produced by the code above has a leading blank line. The code below removes that via the `awk` `substr` function:



```
awk '/REDEFINES/{f=1;print substr(s,2) t>(""output"" ++c "".txt"");t=""""} f==0 {s=s""\n""$0;next} {t=t""\n""$0} END{print substr(s,2) t>(""output"" ++c "".txt"")}' input

```

For variety, this version has slightly different logic but, otherwise, achieves the same result.


### Answer for Revised Question



```
awk 'f==1 && pre==$1 && !/REDEFINES/{tail=tail ""\n"" $0} /REDEFINES/{pre=$1;f=1;t[++c]=""\n""$0} f==0 {head=head""\n""$0;next} pre!=$1{t[c]=t[c]""\n""$0} END{for (i=0;i<=c;i++) {print head t[i] tail>(""output"" (i+1) "".txt"")}}' file

```

"
"How to skip transaction in ActiveRecord for INSERT ONLY statement?
Look at this example:



```
2.1.3 :001 > Stat.create!
   (0.1ms)  BEGIN
  SQL (0.3ms)  INSERT INTO `stats` (`created_at`, `updated_at`) VALUES ('2015-03-16 11:20:08', '2015-03-16 11:20:08')
   (0.4ms)  COMMIT
 => #<Stat id: 1, uid: nil, country: nil, city: nil, created_at: ""2015-03-16 11:20:08"", updated_at: ""2015-03-16 11:20:08"">

```

As you can see the `create!` method execute insert statement inside useless transaction. How to disable transation in this case only (without disabling them in whole application)?


","How it works:


The persistence module define `create`: <https://github.com/rails/rails/blob/4-2-stable/activerecord/lib/active_record/persistence.rb#L46>



```
def create!(attributes = nil, &block)
  if attributes.is_a?(Array)
    attributes.collect { |attr| create!(attr, &block) }
  else
    object = new(attributes, &block)
    object.save!
    object
  end
end

```

It create an object and call `#save!`


It is not documented in the public api, but calls <https://github.com/rails/rails/blob/4-2-stable/activerecord/lib/active_record/transactions.rb#L290>



```
def save!(*) #:nodoc:
  with_transaction_returning_status { super }
end

```

At this point the transaction wrap the save (super), which is at Persistence module again: <https://github.com/rails/rails/blob/4-2-stable/activerecord/lib/active_record/persistence.rb#L141>



```
def save!(*)
  create_or_update || raise(RecordNotSaved.new(nil, self))
end

```

Let's hack this with some new methods:



```
module ActiveRecord
  module Persistence
    module ClassMethods

      def atomic_create!(attributes = nil, &block)
        if attributes.is_a?(Array)
          raise ""An array of records can't be atomic""
        else
          object = new(attributes, &block)
          object.atomic_save!
          object
        end
      end

    end

    alias_method :atomic_save!, :save!
  end
end

module ActiveRecord
  module Transactions

    def atomic_save!(*)
      super
    end

  end
end

```

Perhaps you want to use the standard `create!` method, then you need to redefine it. I define a first optional parameter `:atomic`, and when it's present means you want to use the `atomic_save!` method.



```
module ActiveRecord
  module Persistence
    module ClassMethods

      def create_with_atomic!(first = nil, second = nil, &block)
        attributes, atomic = second == nil ? [first, second] : [second, first]
        if attributes.is_a?(Array)
          create_without_atomic!(attributes, &block)
        else
          object = new(attributes, &block)
          atomic == :atomic ? object.atomic_save! : object.save!
          object
        end
      end
      alias_method_chain :create!, :atomic

    end
  end
end

```

With this in `config/initializers/<any_name>.rb` it can work.


**How it runs at console**:



```
~/rails/r41example (development) > Product.atomic_create!(name: 'atomic_create')
  SQL (99.4ms)  INSERT INTO ""products"" (""created_at"", ""name"", ""updated_at"") VALUES (?, ?, ?)  [[""created_at"", ""2015-03-22 03:50:07.558473""], [""name"", ""atomic_create""], [""updated_at"", ""2015-03-22 03:50:07.558473""]]
=> #<Product:0x000000083b1340> {
            :id => 1,
          :name => ""atomic_create"",
    :created_at => Sun, 22 Mar 2015 03:50:07 UTC +00:00,
    :updated_at => Sun, 22 Mar 2015 03:50:07 UTC +00:00
}
~/rails/r41example (development) > Product.create!(name: 'create with commit')
  (0.1ms)  begin transaction
  SQL (0.1ms)  INSERT INTO ""products"" (""created_at"", ""name"", ""updated_at"") VALUES (?, ?, ?)  [[""created_at"", ""2015-03-22 03:50:20.790566""], [""name"", ""create with commit""], [""updated_at"", ""2015-03-22 03:50:20.790566""]]
  (109.3ms)  commit transaction
=> #<Product:0x000000082f3138> {
            :id => 2,
          :name => ""create with commit"",
    :created_at => Sun, 22 Mar 2015 03:50:20 UTC +00:00,
    :updated_at => Sun, 22 Mar 2015 03:50:20 UTC +00:00
}
~/rails/r41example (development) > Product.create!(:atomic, name: 'create! atomic')
  SQL (137.3ms)  INSERT INTO ""products"" (""created_at"", ""name"", ""updated_at"") VALUES (?, ?, ?)  [[""created_at"", ""2015-03-22 03:51:03.001423""], [""name"", ""create! atomic""], [""updated_at"", ""2015-03-22 03:51:03.001423""]]
=> #<Product:0x000000082a0bb8> {
            :id => 3,
          :name => ""create! atomic"",
    :created_at => Sun, 22 Mar 2015 03:51:03 UTC +00:00,
    :updated_at => Sun, 22 Mar 2015 03:51:03 UTC +00:00
}

```

Caveat: You will lose after\_rollback and after\_commit callbacks!


Note: on 4.1 the methods create! and save! are in module Validations. On Rails 4.2 are in Persistence.


**Edit**: Perhaps you think you can earn the transaction elapsed time. In my examples the commit time goes to the inserts (I have a standard HD and I think you have an SSD).


"
"How to solve ""Pending update of ""firefox"" snap. Close the app to avoid disruptions"" error?
Recently, this warning appears when open Firefox:



> 
> Pending update of ""firefox"" snap
> 
> 
> Close the app to avoid disruptions
> 
> 
> 


What does it mean? And what do I have to do? Is this a bug?


","If you have received a notification of a pending Firefox update it will appear as a small bell icon next to the clock in the top panel. Clicking on the bell will show a Firefox pending update notification as shown in the below screenshot.


[![enter image description here](https://i.stack.imgur.com/sffhw.png)](https://i.stack.imgur.com/sffhw.png)


As soon as you see the `Pending update of ""firefox"" snap. Close the app to avoid disruptions (13 days left)` notification (which means you have 13 days to upgrade Firefox yourself or else the Firefox snap package will get an unattended upgrade) close Firefox and then reopen it after updating the Firefox snap package is finished. If the pending update to the Firefox snap package has not started yet, it can be initiated by running `sudo snap refresh`. If you don't close Firefox after Firefox has finished being updated, then you will be unable to open any new tabs until you close Firefox and then reopen it.


"
"rails 3, heroku, how to use sessions
I saw a great answer re how to setup sessions for rails 3 
( [Rails sessions current practices](https://stackoverflow.com/questions/2588241/rails-sessions-current-practices-especially-for-rails-3) )


we gather a series of questions from the user and I'd like to maintain in session a list of their answers until we get to the end then write it all out...


but am not sure how to write and read info from the session... any quick pointer would be appreciated on how to save, for example, the contents of a hash


Also, does having our app hosted on a grid at Heroku change how we could/should handle sessions?


Cheers,
JP


","You do not have to change anything for Heroku. By default, Rails sessions are stored in an encrypted cookie, so there is no server-side configuration necessary.


However, a cookie can only store 4,096 bytes of data. If you are storing **a lot** of data in the session (which generally is not recommended), you may overflow the cookie. In this case, you may wish to set up ActiveRecord or Memcached-based cookies. Both of these are easy to do, and are not really Heroku-specific issues. If you need help with this, you can always ask another StackOverflow question. For now, it's not worrying about until you hit the limit.


Some rough code to store and read your answers in the session, assuming Question and Answer are ActiveRecord models:



```
def store_answer(question, answer)
  session[:answers] ||= {}
  session[:answers][question.id] = answer.id
end

def read_answer(question)
  Answer.find(session[:answers][question.id])
end

```

"
"How to count continuous numbers in numpy
I have a Numpy one-dimensional array of 1 and 0. for e.g



```
a = np.array([0,1,1,1,0,0,0,0,0,0,0,1,0,1,1,0,0,0,1,1,0,0])

```

I want to count the continuous 0s and 1s in the array and output something like this



```
[1,3,7,1,1,2,3,2,2]

```

What I do atm is



```
np.diff(np.where(np.abs(np.diff(a)) == 1)[0])

```

and it outputs



```
array([3, 7, 1, 1, 2, 3, 2])

```

as you can see it is missing the first count 1.


I've tried `np.split` and then get the sizes of each segments but it does not seem to be optimistic.


Is there more elegant ""pythonic"" solution?


","Here's one vectorized approach -



```
np.diff(np.r_[0,np.flatnonzero(np.diff(a))+1,a.size])

```

Sample run -



```
In [208]: a = np.array([0,1,1,1,0,0,0,0,0,0,0,1,0,1,1,0,0,0,1,1,0,0])

In [209]: np.diff(np.r_[0,np.flatnonzero(np.diff(a))+1,a.size])
Out[209]: array([1, 3, 7, 1, 1, 2, 3, 2, 2])

```

Faster one with `boolean` concatenation -



```
np.diff(np.flatnonzero(np.concatenate(([True], a[1:]!= a[:-1], [True] ))))

```

**Runtime test**


For the setup, let's create a bigger dataset with islands of `0s` and `1s` and for a fair benchmarking as with the given sample, let's have the island lengths vary between `1` and `7` -



```
In [257]: n = 100000 # thus would create 100000 pair of islands

In [258]: a = np.repeat(np.arange(n)%2, np.random.randint(1,7,(n)))

# Approach #1 proposed in this post
In [259]: %timeit np.diff(np.r_[0,np.flatnonzero(np.diff(a))+1,a.size])
100 loops, best of 3: 2.13 ms per loop

# Approach #2 proposed in this post
In [260]: %timeit np.diff(np.flatnonzero(np.concatenate(([True], a[1:]!= a[:-1], [True] ))))
1000 loops, best of 3: 1.21 ms per loop

# @Vineet Jain's soln    
In [261]: %timeit [ sum(1 for i in g) for k,g in groupby(a)]
10 loops, best of 3: 61.3 ms per loop

```

"
"Inconsistent font size rendering issue on Chrome
I am having an css issue on my site. The font size is not consistent across all pages. The difference is one version of bigger font size and one version of a regular small font size. The small font size is the one I prefer. **If the page is showing the content in big font, it will change it back to a smaller one after reload the page for a couple of times, it also changes back to the smaller one right away if you right click and select inspect element.** Thus, I am not sure how can I ping point the issue.
Here are the screenshots, one with the big font and one with the small font, and again the small font is the one I prefer.
![the big font](https://i.stack.imgur.com/zTLYw.jpg)


![the small font](https://i.stack.imgur.com/T0KMz.jpg)


","`body { font-size: 1.6rem;}` might be your issue... you can axe that from your style.css and it should be more consistent. It looks like chrome might not like rem on the body as can be seen in the linked question 


removing/changing the bad line from the following class definition in style.css should fix you up



```
body, button, input, select, textarea {
   color: #404040;
   font-family: sans-serif;
   font-size: 16px;
   font-size: 1.6rem;
   line-height: 1.5;
}

```

[Similar SO question](https://stackoverflow.com/questions/20099844/chrome-not-respecting-rem-font-size-on-body-tag)


Text + good links from answer on other question:


""This seems to be a Chrome bug; see [Issue 319623: Rendering issue when using % + REMs in CSS](https://code.google.com/p/chromium/issues/detail?id=319623), and/or a partly-merged duplicate: [Issue 320754: font-size does not inherit if html has a font-size in percentage, and body in rem](https://code.google.com/p/chromium/issues/detail?id=320754)""


"
"Overlapping triangles
Here is the shape I'm trying to create in pure CSS: 


[![CSS overlapping triangles](https://i.stack.imgur.com/BkFrn.png)](https://i.stack.imgur.com/BkFrn.png)


I have a more complete jsfiddle <http://jsfiddle.net/8Lxr5s57/7/>. Is there a better, more efficient way to achieve this same result?





```
.angled_container {
  background-color: #fff;
  height: 200px;
  position: relative;
  overflow: hidden;
  clear: both;
}
.angled_container:before,
.angled_container:after {
  content: """";
  width: 110%;
  height: 100%;
  display: block;
  position: absolute;
  top: 0;
}
.angled_container:before {
  background-color: #606060;
  transform: rotate(12deg);
  -webkit-transform-origin: left top;
  left: 0;
}
.angled_container:after {
  background-color: #6bb2c6;
  transform: rotate(-12deg);
  -webkit-transform-origin: right top;
  left: -10%;
}
.angled_container--open-left:before {
  background-color: #6bb2c6;
  z-index: 2;
}
.angled_container--open-left:after {
  background-color: #606060;
  z-index: 1;
}
```


```
<div class=""angled_container angled_container--open-right""></div>
```






","## CSS


I would suggest using `skewY()` instead of `rotate()` for the two triangles. it will avoid some positioning issues and prevent using wider pseudo elements than the container :





```
.angled_container {
  height: 200px;
  position: relative;
  overflow: hidden;
}
.angled_container:before,
.angled_container:after {
  content: """";
  width: 100%; height: 100%;
  display: block;
  position: absolute;
  top: 0; left: 0;
}
.angled_container:before {
  background-color: #606060;
  transform: skewY(12deg);
  transform-origin: left top;
}
.angled_container:after {
  background-color: #6bb2c6;
  transform: skewY(-12deg);
  transform-origin: right top;
}
```


```
<div class=""angled_container angled_container--open-right""></div>
```








---


## SVG


Alternatively, you can use an inline SVG with 2 [polygon elements](https://developer.mozilla.org/en/docs/Web/SVG/Element/polygon). This is totaly responsive and probably easier to make/maintain as you can style the triangles in CSS with the `fill` property :





```
svg{display:block; width:100%;}
.first{
  fill:#606060;
}
.second{
  fill:#6bb2c6;
}
```


```
<svg viewbox=""0 0 100 30"">
  <polygon class=""first"" points=""0 0 100 28 0 25 0 28""/>
  <polygon class=""second"" points=""0 28 0 25 100 0 100 28 52 28 50 30 48 28 0""/>
</svg>
```






"
"ImageMagick - convert multipage PDF into one image
I am using ImageMagick to convert PDF files into images. However, some of the PDF's have multiple pages, which is proving to be a real problem.


My local convert is below.



```
exec(""\""C:\\Program Files (x86)\\ImageMagick-6.8.5-Q16\\convert.exe\"" -density 300 -quality 75 \""{$path}{$filename}{$ext}[$page]\"" \""{$targetFile}\"""");

```

If i remove `[$page]` from the exec it works but creates an image per page, which isn't what I want.


I have been searching for a while now and i've ran out of hope and ideas. Is there any way I can get all of the new images into one final image, or convert the PDF straight into one image? Any help would be greatly appreciated, cheers.


","Check out the -append and +append options.


`-append` appends the images vertically, and `+append` appends them horizontally.


Usage (<http://linuxers.org/quick-tips/convert-pdf-file-single-image-using-imagemagick>):


According to that link, the output from a multi-page pdf `convert` would be ${targetFile}-0.png, ${targetFile}-1.png, ${targetFile}-n.png, etc. Once you have converted the pdf into multiple images, use the -append or +append option:


`convert ${targetFile}-* -append single_image.png`


To put it all together, try something like this (you may have to play with it a bit; I haven't used Imagemagick from the Windows' shell):


`// convert pages of pdf`


`exec(""\""C:\\Program Files (x86)\\ImageMagick-6.8.5-Q16\\convert.exe\"" -density 300 -quality 75 \""{$path}{$filename}{$ext}\"" \""{$targetFile}\"""");`


`// then append them`


`exec(""\""C:\\Program Files (x86)\\ImageMagick-6.8.5-Q16\\convert.exe\"" \""{$targetFile}-*\"" -append ""${someName}\"""");`


More resources:


<http://www.imagemagick.org/script/command-line-options.php#append>


<http://www.imagemagick.org/Usage/layers/>


"
"ls -ad \*/ won't display hidden directories
`pwd` is my home directory. `ls */ -d` displays the directories.



```
~$ ls */ -d
Blog/  Desktop/  Documents/  Downloads/  Music/  Pictures/  Public/  Templates/  Videos/

```

And `ls -a` displays all hidden directories and files. Hidden directories - `.cache`, `.local` etc.



```
$ ls -a
.                        .cache      .gnupg       .xsession-errors.old
..                       .config     .gtkrc-2.0   Templates
.bash_history            Desktop     .gtkrc-xfce  Pictures      .themes
.bash_logout             .linuxmint   .pki          Videos
.bashrc                  Documents   .local       .profile      .viminfo
Blog                     Downloads   .mozilla     Public        .Xauthority
Music        .ssh          .xsession-errors

```

But when I run `ls -ad */`, it won't displays any hidden directories. Anyone care to explain?



```
$ ls -ad */ 
Blog/  Desktop/  Documents/  Downloads/  Music/  Pictures/  Public/  Templates/  Videos/

```

","`*` **deliberately** doesn't match hidden files or directories, because most of the time that's exactly what is wanted (as dot files and directories are generally used for configuration, not data), and you can override the default by explicitly specifying a glob starting with `.` if/when you need to.


If you want them to be matched by a glob, use `ls -d .*` (or `ls -d .*/` to match only hidden directories without regular files). To match both hidden and non-hidden directories, use `ls -d -- */ .*/`


NOTE: as mentioned by @rexkogitans in a comment, globs like `*` are expanded by the shell **before** they are passed to a program. `ls` never sees a `*` or `.*` (or `.*/`), it sees the list of file and/or directory names that are the result of the shell expanding the glob. This is why you need to quote or escape (by prefixing with a backslash) glob characters if you want to pass them as string literals to a program.


Also note: if there are no files/dirs matching the glob then the glob is passed to the program as is, unexpanded (i.e. `ls` will see `*` as its argument). `bash` (and some other bourne-like shells) allow this behaviour to be over-ridden - e.g. in bash, you can use the `nullglob` or `failglob` shell options. nullglob causes it to be expanded to nothing (i.e. it is removed from the argument list), while failglob triggers an error.


"
"Intersecting texts to find common words
I'm trying to find out which would be the most optimal way of intersection a set of texts and find the common words in them. Given this scenario:



```
var t1 = 'My name is Mary-Ann, and I come from Kansas!';
var t2 = 'John, meet Mary, she comes from far away';
var t3 = 'Hi Mary-Ann, come here, nice to meet you!';

```

intersection result should be:



```
var result =[""Mary""];

```

It should be able to ignore punctuation marks like `.,!?-`


Would a solution with regular expressions be optimal?


","Here's a tested solution :



```
function intersect() {
   var set = {};
   [].forEach.call(arguments, function(a,i){
     var tokens = a.match(/\w+/g);
     if (!i) {
       tokens.forEach(function(t){ set[t]=1 });
     } else {
       for (var k in set){
         if (tokens.indexOf(k)<0) delete set[k];
       }
     }
   });
   return Object.keys(set);
}

```

This function is variadic, you can call it with any number of texts :



```
console.log(intersect(t1, t2, t3)) // -> [""Mary""] 

console.log(intersect(t1, t2)) // -> [""Mary"", ""from""] 

console.log(intersect()) // -> [] 

```

If you need to support non English languages, then this regex won't be enough because of the poor support of Unicode in JavaScript regexes. Either you use a [regex library](https://github.com/slevithan/xregexp/blob/master/README.md#unicode) or you define your regex by explicitly excluding characters as in `a.match(/[^\s\-.,!?]+/g);` (this will probably be enough for you) .




---


**Detailed explanation :**


The idea is to fill a set with the tokens of the first text and then remove from the set the tokens missing in the other texts.


1. The set is a JavaScript object used as a map. Some purists would have used `Object.create(null)` to avoid a prototype, I like the simplicity of `{}`.
2. As I want my function to be [variadic](http://en.wikipedia.org/wiki/Variadic_function), I use [`arguments`](https://developer.mozilla.org/en-US/docs/Web/JavaScript/Reference/Functions_and_function_scope/arguments) instead of defining the passed texts as explicit arguments.
3. `arguments` isn't a real array, so to iterate over it you need either a `for` loop or a trick like `[].forEach.call`. It works because `arguments` is [""array-like""](https://www.inkling.com/read/javascript-definitive-guide-david-flanagan-6th/chapter-7/array-like-objects).
4. To tokenize, I simply use [`match`](https://developer.mozilla.org/en-US/docs/Web/JavaScript/Reference/Global_Objects/String/match) to match words, nothing special here (see note above regarding better support of other languages, though)
5. I use `!i` to check if it's the first text. In that case, I simply copy the tokens as properties in the set. A value must be used, I use `1`. In the future, [ES6 sets](http://www.nczonline.net/blog/2012/09/25/ecmascript-6-collections-part-1-sets/) will make the intent more obvious here.
6. For the following texts, I iterate over the elements of the sets (the keys) and I remove the ones which are not in the array of tokens (`tokens.indexOf(k)<0`)
7. Finally, I return the elements of the sets because we want an array. The simplest solution is to use [`Object.keys`](https://developer.mozilla.org/en-US/docs/Web/JavaScript/Reference/Global_Objects/Object/keys).


"
"Java make a separate annotation that combines a others
Working with Play Framework 2.2, making a RESTfull API.


In a model I'm using, I wanted to output(Json with Jackson) only the Id of a related object, not the entire object. I found how to do that, as follows:



```
@JsonIdentityInfo(generator = ObjectIdGenerators.PropertyGenerator.class, property = ""id"")
@JsonIdentityReference(alwaysAsId = true)
@JsonIgnore
public Object myObject;

```

The Json output will contain a JsonNode ""myObjectId: 1"". For instance.


The IdentityInfo and IdentityReference take care of this.


The problem is, wherever I want this, I have to paste down these 3 lines, along with any other annotations that need to be present for a certain field. This get grow too large and I'm trying to create just 1 custom annotation that does all these things together.


Is this possible, and how? A link where I can read about it or an example would be appreciated.


","To solve your problem you have to create an annotation annotated with [JacksonAnnotationInside](https://github.com/FasterXML/jackson-annotations/blob/master/src/main/java/com/fasterxml/jackson/annotation/JacksonAnnotationsInside.java) and annotations you want it to 'include'.


The Javadoc of that annotation says that it is:



> 
> Meta-annotation (annotations used on other annotations) used for
>  indicating that instead of using target annotation (annotation
>  annotated with this annotation), Jackson should use meta-annotations
>  it has. **This can be useful in creating ""combo-annotations"" by having a
>  container annotation, which needs to be annotated with this annotation
>  as well as all annotations it 'contains'.**
> 
> 
> 


For example for your case you'd have something like this:



```
@Target(ElementType.FIELD)
@Retention(RetentionPolicy.RUNTIME)
@com.fasterxml.jackson.annotation.JacksonAnnotationsInside // this is important
@JsonIdentityInfo(generator = ObjectIdGenerators.PropertyGenerator.class, property = ""id"")
@JsonIdentityReference(alwaysAsId = true)
public @interface MyCustomAnnotation {    
}

```

Then you can use it normally as you would Jackson annotations



```
@MyCustomAnnotation
public Object myObject;

```

"
"Compression performance related to chunk size in hdf5 files
I would like to ask a question about the performance of compression
which is related to chunk size of hdf5 files.


I have 2 hdf5 files on hand, which have the following properties.
They both only contain one dataset, called ""data"".


File A's ""data"":


1. Type: HDF5 Scalar Dataset
2. No. of Dimensions: 2
3. Dimension Size: **5094125 x 6**
4. Max. dimension size: Unlimited x Unlimited
5. Data type: 64-bit floating point
6. Chunking: **10000 x 6**
7. Compression: GZIP level = 7


File B's ""data"":


1. Type: HDF5 Scalar Dataset
2. No. of Dimensions: 2
3. Dimension Size: **6720 x 1000**
4. Max. dimension size: Unlimited x Unlimited
5. Data type: 64-bit floating point
6. Chunking: **6000 x 1**
7. Compression: GZIP level = 7


**File A's size: 
HDF5----19 MB
CSV-----165 MB**


**File B's size: 
HDF5----60 MB
CSV-----165 MB**


Both of them shows great compression on data stored when comparing to csv files.
However, the compression rate of file A is about 10% of original csv,
while that of file B is only about 30% of original csv.


I have tried different chunk size to make file B as small as possible, but it seems that 30% is the optimum compression rate. I would like to ask why file A can achieve a greater compression while file B cannot. 


If file B can also achieve, what should the chunk size be?


Is that any rule to determine the optimum chunk size of HDF5 for compression purpose?


Thanks!


","Chunking doesn't really affect the compression ratio *per se*, except in the manner @Ümit describes. What chunking does do is affect the I/O performance. When compression is applied to an HDF5 dataset, it is applied to whole chunks, individually. This means that when reading data from a single chunk in a dataset, the entire chunk must be decompressed - possibly involving a whole lot more I/O, depending on the size of the cache, shape of the chunk, etc.


What you should do is make sure that the chunk *shape* matches how you read/write your data. If you generally read a column at a time, make your chunks columns, for example. [This is a good tutorial on chunking.](http://www.hdfgroup.org/HDF5/doc/Advanced/Chunking/Chunking_Tutorial_EOS13_2009.pdf)


"
"Get definitive names for columns from table variable
I can declare a table variable as such:



```
DECLARE @tv_source TABLE(c1 int, 
providerName varchar(50),
providerSMS varchar(50))

```

If I then execute the following, I see the table name similar to: ""#468862B0""



```
select top 1 * from tempdb.sys.tables where type = 'U' order by create_date desc
select TOP 1 name,* from tempdb.sys.sysobjects ORDER BY CRDATE desc

```

If I then immediately execute:



```
select TOP 3 * 
from tempdb.sys.columns 
where object_id in (select TOP 1 object_id from tempdb.sys.tables ORDER BY Create_date desc)

```

I see the columns I declared above for the table variable.


My question is, **is there any way to definitively associate those columns with the name I declared in the table declaration above ""@tv\_source""?** 


In a normal table, you would see the actual name but, as noted above, table variables get morphed into a hex value (which, btw is the hex value of the object\_id).


","You can query your table variable `top(0)` with an `outer apply` from one row using `for xml path('')` and then query the XML for the element names.


This will work as long as your column names does not have names that is invalid XML element names. The column names can for instance not use ampersand or space.



```
declare @tv_source table
(
  c1 int, 
  providerName varchar(50),
  providerSMS varchar(50)
)

select TN.N.value('local-name(.)', 'sysname') as ColumnName
from 
  (
  select TV.*
  from (select 1) as D(N)
    outer apply (
                select top(0) *
                from @tv_source
                ) as TV
  for xml path(''), elements xsinil, type
  ) as TX(X)
cross apply TX.X.nodes('*') as TN(N)

```

Another option would be to use the `xmlschema` directive of `for xml auto`. This solution does handle invalid XML characters but they are escaped so if you have a column name with a space like `[provider Name]` the result will be `provider_x0020_Name`.  

You need to store the resulting XML to a variable and query that for the information you want.



```
declare @XML xml;

set @XML = 
  (
  select top(0) *
  from @tv_source
  for xml auto, xmlschema, type
  );

with xmlnamespaces('http://www.w3.org/2001/XMLSchema' as xsd)
select T.X.value('@name', 'sysname')
from @XML.nodes('//xsd:attribute') as T(X);

```

The XML created by `xmlschema` contains more information that might be of interest. You can retrieve the table variable name and the datatypes as well.



```
<xsd:schema xmlns:schema=""urn:schemas-microsoft-com:sql:SqlRowSet12"" xmlns:xsd=""http://www.w3.org/2001/XMLSchema"" xmlns:sqltypes=""http://schemas.microsoft.com/sqlserver/2004/sqltypes"" targetNamespace=""urn:schemas-microsoft-com:sql:SqlRowSet12"" elementFormDefault=""qualified"">
  <xsd:import namespace=""http://schemas.microsoft.com/sqlserver/2004/sqltypes"" schemaLocation=""http://schemas.microsoft.com/sqlserver/2004/sqltypes/sqltypes.xsd"" />
  <xsd:element name=""_x0040_tv_source"">
    <xsd:complexType>
      <xsd:attribute name=""c1"" type=""sqltypes:int"" />
      <xsd:attribute name=""providerName"">
        <xsd:simpleType>
          <xsd:restriction base=""sqltypes:varchar"" sqltypes:localeId=""1035"" sqltypes:sqlCompareOptions=""IgnoreCase IgnoreKanaType IgnoreWidth"">
            <xsd:maxLength value=""50"" />
          </xsd:restriction>
        </xsd:simpleType>
      </xsd:attribute>
      <xsd:attribute name=""providerSMS"">
        <xsd:simpleType>
          <xsd:restriction base=""sqltypes:varchar"" sqltypes:localeId=""1035"" sqltypes:sqlCompareOptions=""IgnoreCase IgnoreKanaType IgnoreWidth"">
            <xsd:maxLength value=""50"" />
          </xsd:restriction>
        </xsd:simpleType>
      </xsd:attribute>
    </xsd:complexType>
  </xsd:element>
</xsd:schema>

```

"
"Can't change UINavigationBar prompt color
I am unable to change the prompt color on my navigation bar. I've tried the code below in `viewDidLoad`, but nothing happens.



```
self.navigationController?.navigationBar.titleTextAttributes = [NSAttributedStringKey.foregroundColor: UIColor.white]

```

[![navigation bar image](https://i.stack.imgur.com/XPI6d.png)](https://i.stack.imgur.com/XPI6d.png)


Am I missing something? Is the code above wrong?


","It seems like you're right about this one. You need to use `UIAppearance` to style the prompt text on iOS 11.


I've filed radar #34758558 that the `titleTextAttributes` property just stopped working for prompt in iOS 11.


The good news is that there are a couple of workarounds, which we can uncover by using Xcode's view hierarchy debugger:


[![Screenshot of UINavigation Bar Prompt on iOS 11](https://i.stack.imgur.com/jDS2i.png)](https://i.stack.imgur.com/jDS2i.png)



```
// 1. This works, but potentially changes *all* labels in the navigation bar. 
// If you want this, it works.
UILabel.appearance(whenContainedInInstancesOf: [UINavigationBar.self]).textColor = UIColor.white

```

The prompt is just a UILabel. If we use UIAppearance's `whenContainedInInstancesOf:`, we can pretty easily update the color the way we want.


If you look closely, you'll notice that there's also a wrapper view on the UILabel. It has its own class that might respond to UIAppearance...


[![Screenshot of private UINavigationBar subview](https://i.stack.imgur.com/5507U.png)](https://i.stack.imgur.com/5507U.png)



```
// 2. This is a more precise workaround but it requires using a private class.

if let promptClass = NSClassFromString(""_UINavigationBarModernPromptView"") as? UIAppearanceContainer.Type
{
  UILabel.appearance(whenContainedInInstancesOf: [promptClass]).textColor = UIColor.white
}

```

I'd advise sticking to the more general solution, since it doesn't use private API. (App review, etc.) Check out what you get with either of these two solutions:


[![Properly colored prompt text](https://i.stack.imgur.com/d224m.png)](https://i.stack.imgur.com/d224m.png)


"
"How to calculate close crop rectangle of a bitmap?
I need to calculate a close crop rectangle for a bitmap by a given background color key. In the following pictures you can see what is meant to be the close crop. On the left side is the source, on the right side output of the close crop:


![enter image description here](https://i.stack.imgur.com/5ZpIH.png)


![enter image description here](https://i.stack.imgur.com/QAnkg.png)


As you can see, I need to find the topmost, leftmost, lowermost and rightmost pixels which differ in color from the background to build the close crop rectangle. So, how to find those differing outer pixels to get the close crop rectangle ? Or, in other words, how to calculate close crop rectangle of a bitmap ?


","You can use this code (you may follow the [`commented version`](https://stackoverflow.com/revisions/10533819/4) of this post as well):



```
procedure CalcCloseCrop(ABitmap: TBitmap; const ABackColor: TColor;
  out ACropRect: TRect);
var
  X: Integer;
  Y: Integer;
  Color: TColor;
  Pixel: PRGBTriple;
  RowClean: Boolean;
  LastClean: Boolean;
begin
  if ABitmap.PixelFormat <> pf24bit then
    raise Exception.Create('Incorrect bit depth, bitmap must be 24-bit!');

  LastClean := False;
  ACropRect := Rect(ABitmap.Width, ABitmap.Height, 0, 0);

  for Y := 0 to ABitmap.Height-1 do
  begin
    RowClean := True;
    Pixel := ABitmap.ScanLine[Y];
    for X := 0 to ABitmap.Width - 1 do
    begin
      Color := RGB(Pixel.rgbtRed, Pixel.rgbtGreen, Pixel.rgbtBlue);
      if Color <> ABackColor then
      begin
        RowClean := False;
        if X < ACropRect.Left then
          ACropRect.Left := X;
        if X + 1 > ACropRect.Right then
          ACropRect.Right := X + 1;
      end;
      Inc(Pixel);
    end;

    if not RowClean then
    begin
      if not LastClean then
      begin
        LastClean := True;
        ACropRect.Top := Y;
      end;
      if Y + 1 > ACropRect.Bottom then
        ACropRect.Bottom := Y + 1;
    end;
  end;

  if ACropRect.IsEmpty then
  begin
    if ACropRect.Left = ABitmap.Width then
      ACropRect.Left := 0;
    if ACropRect.Top = ABitmap.Height then
      ACropRect.Top := 0;
    if ACropRect.Right = 0 then
      ACropRect.Right := ABitmap.Width;
    if ACropRect.Bottom = 0 then
      ACropRect.Bottom := ABitmap.Height;
  end;
end;

procedure TForm1.Button1Click(Sender: TObject);
var
  R: TRect;
  Bitmap: TBitmap;
begin
  CalcCloseCrop(Image1.Picture.Bitmap, $00FFA749, R);
  Bitmap := TBitmap.Create;
  try
    Bitmap.SetSize(R.Width, R.Height);
    Bitmap.Canvas.CopyRect(Rect(0, 0, R.Width, R.Height), Image1.Canvas, R);
    Image1.Picture.Bitmap.Assign(Bitmap);
  finally
    Bitmap.Free;
  end;
end;

```

"
"Determine which DOM elements the current text selection contains
I want to be able to find out which DOM elements are present in the text currently selected by the user in the browser. 


document.getSelection() would get us the currently selected text. But how can we determine which DOM elements are contained in this text selection?


","`window.getSelection()` gives you a [Selection](http://www.w3.org/TR/html5/editing.html#selection-0) object. Use `selection.rangeCount` and `selection.getRangeAt()` to get a [Range](http://www.w3.org/TR/DOM-Level-2-Traversal-Range/ranges.html) object representing which selection you want.


Now you can get the first and last nodes in the selection from `range.startContainer`/`startOffset` and `range.endContainer`/`endOffset`. You could then walk up and down the tree to pick up all the elements in between those two positions, eg. using the `getElementsBetweenTree()` function from [this answer](https://stackoverflow.com/questions/2514125/jquery-select-elements-between-two-elements-that-are-not-siblings/2515169#2515169).


Unfortunately, IE's [TextRange](http://msdn.microsoft.com/en-us/library/ms535872%28v=VS.85%29.aspx) model is totally different to the W3 and HTML5 standard, and in general much worse. It does't give up the start and end positions nearly as easily, and not in any reliable way. All you get is `parentElement` to tell you the common ancestor, and from there you're limited to guessing where the range is based on creating a new Range and calling `moveToElementText` on it then `compareEndPoints` with the selection range. And if you need to know the exact text selection you're then guessing based on `moveStart`/`moveEnd`​ing the range and comparing, which isn't fun at all.


"
"'array' in namespace 'std' does not name a template type
I receive the following error:



> 
> 'array' in namespace 'std' does not name a template type.
> 
> 
> 


I changed my compiler to g++ 4.9. Still having issues. I think that I may have an old version of the std library but am not sure on how to proceed to fix that.



```
#ifndef DBINTERFACE_H_
#define DBINTERFACE_H_

#include ""mysql_connection.h""
#include <cppconn/driver.h>
#include <cppconn/exception.h>
#include <cppconn/resultset.h>
#include <cppconn/statement.h>
#include ""mysql_driver.h""
#include <array>

class DBInterface {
private:
    double Found_Buffer[100][59];
    int Found_Count;
    double New_Buffer[100][59];
    std::array<double, 5> BLAH;

    int New_Count;
    double M[59];
    double B[59];
    sql::mysql::MySQL_Driver *driver;
    sql::Connection *con;

public:
    DBInterface();
    virtual ~DBInterface();
    void Update();
    void Search();
    void Add2DB();
    void Add2Buffer(double Found_Objects[][59], double New_Objects[][59]);
    void Build();

    /*
     * To be added:
     * void CollapseBuffer();
     * void DetDynamic();
     *
     */
};

#endif /* DBINTERFACE_H_ */

```

Error messages:



```
17:20:06 **** Incremental Build of configuration Debug for project CANS ****
make all 
Building file: ../src/DBInterface.cpp
Invoking: Cross G++ Compiler
g++ -I/home/derek/soci/include -O0 -g3 -Wall -c -fmessage-length=0 -MMD -MP -MF""src/DBInterface.d"" -MT""src/DBInterface.d"" -o ""src/DBInterface.o"" ""../src/DBInterface.cpp""
In file included from /usr/include/c++/4.9/array:35:0,
                 from ../src/DBInterface.h:17,
                 from ../src/DBInterface.cpp:8:
/usr/include/c++/4.9/bits/c++0x_warning.h:32:2: error: #error This file requires compiler and library support for the ISO C++ 2011 standard. This support is currently experimental, and must be enabled with the -std=c++11 or -std=gnu++11 compiler options.
 #error This file requires compiler and library support for the \
  ^
In file included from ../src/DBInterface.cpp:8:0:
../src/DBInterface.h:24:7: error: ‘array’ in namespace ‘std’ does not name a template type
  std::array<double, 5> BLAH;
       ^
../src/DBInterface.cpp: In function ‘void Add2Buffer(double (*)[59], double (*)[59])’:
../src/DBInterface.cpp:44:6: warning: unused variable ‘NoOfFO’ [-Wunused-variable]
  int NoOfFO;
      ^
../src/DBInterface.cpp:45:6: warning: unused variable ‘NoOfNO’ [-Wunused-variable]
  int NoOfNO;
      ^
make: *** [src/DBInterface.o] Error 1

17:20:08 Build Finished (took 2s.464ms)

```

","Just read what the compiler is telling you!


`/usr/include/c++/4.9/bits/c++0x_warning.h:32:2: error: #error This file requires compiler and library support for the ISO C++ 2011 standard. This support is currently experimental, and must be enabled with the -std=c++11 or -std=gnu++11 compiler options.`


That tells you that you have included a C++11 header (in this case `<array>`) but you are not compiling in C++11 mode.


It even tells you how to fix the problem (use `-std=c++11` or `-std=gnu++11`).


The later error saying `array` doesn't name a type is because you've included a C++11 header without enabling C++11, so if you fix the **first** error then the later ones will go away.


When reading compiler errors start at the top and fix the first one, don't just pick a random part of the output to fix. Later errors are often caused by the earlier ones, so fix the earlier ones first.


"
"Auth0 JWT Access Tokens
I am having difficulty getting Auth0 to return access tokens in JWT format. I need them in JWT format in order that I can verify them using a Java JWT library.


I am using Auth0 `lock` to login, and use `/oauth/token` to get the access token - I have tried setting the audience to our API identifier (in multiple places including the lock auth params, and the `/oauth/token` payload), without success - the access token is returned however not a JWT.


Alternatively, is there a Java library for validating ""native"" Auth0 access tokens?



```
var options = {
    auth: {
        redirectUrl: '<redirect_link>',
        responseType: 'code',
        params: {
            audience: '<API_identifier>',
            scope: 'openid offline_access'
        }
    },
    allowedConnections: ['Username-Password-Authentication']
};

var lock = new Auth0Lock('<clientId>', '<auth0_Host>', options); 
lock.show();

```

The returned code is used to POST to `https://<host>/oauth/token` with data:



```
client_id=<client_id>&redirect_uri=<redirect_url>&client_secret=<client_secret>&code=<returned_code>&grant_type=authorization_code&audience=<API_identifier>

```

Which is successful but the access token is not JWT, for example: `""access_token"":""sG99DAJI789SYgTj""`


Using the scope openid returns an id\_token in JWT format, but from reading the documentation this token should not be used for API authorisation.


","It seems Auth0 is using [OpenID connect](http://openid.net/specs/openid-connect-core-1_0.html), that is an extension of OAuth2. After a successful end-user authentication, the server returns an ID Token in JWT format and an access token



> 
> **ID Token**
> 
> 
> The ID Token is a security token that contains Claims about the Authentication of an End-User by an Authorization Server when using a Client, and potentially other requested Claims. The ID Token is represented as a JSON Web Token (JWT) [JWT].
> 
> 
> 


You can validate the ID Token at client side using any JWT library, but the validation rules for access tokens are different



> 
> **3.2.2.9. Access Token Validation**
> 
> 
> To validate an Access Token issued from the Authorization Endpoint with an ID Token, the Client SHOULD do the following:
> 
> 
> 1. Hash the octets of the ASCII representation of the access\_token with the hash algorithm specified in JWA for the alg Header Parameter of the ID Token's JOSE Header. For instance, if the alg is RS256, the hash algorithm used is SHA-256.
> 2. Take the left-most half of the hash and base64url encode it.
> 3. The value of `at_hash` in the ID Token MUST match the value produced in the previous step.
> 
> 
> 


So, basically to validate it you need to compute the digest of the `access_token` using the hash algorithm of ID token, and check that it matches with the `at_hash` claim of ID Token


"
"Debugging weirdness with BeginInvoke
I have the following method:



```
protected void OnBarcodeScan(BarcodeScannerEventArgs e)
{
    if (BarcodeScan != null)
    {
        //BarcodeScan.BeginInvoke(e, null, null);
        BarcodeScan(e);
    }
}

```

When I try to step into the above method it works fine. I am able to step in and over all the parts of the method. 


However, if I switch the comment (so `BarcodeScan(e)` is commented out and remove the comment on `BarcodeScan.BeginInvoke(e, null, null)` then I cannot step into any part of the OnBarcodeScan method (ie a break point on `if (BarcodeScan != null)` does not get hit.


I tried putting some debug statements in there too. But as long as that begin invoke call is in there it does not let me step into the method.


I checked the output and when I try to step in it says this:



> 
> A first chance exception of type 'System.NotSupportedException' occurred in ScannerTest.exe
>  Step into: Stepping over method without symbols 'Symbol.Marshaller.SymbolMessageWindow.WndProc'
>  Step into: Stepping over method without symbols 'Microsoft.WindowsCE.Forms.MessageWindow.\_WndProc'
> 
> 
> 


Why would the whole method be un step able when there is a BeginInvoke in it?


Any help would be great!


","Asynchronous delegate calls (i.e., `BeginInvoke`) are not supported by the Compact Framework.


As to the reason why the debugger will not even break into the method, I believe it is because of the following:


- `BeginInvoke`/`EndInvoke` methods are generated by the C# compiler (it is [required to do this](http://msdn.microsoft.com/en-us/library/22t547yb%28v=VS.90%29.aspx)), marked as ""native"". This means the CLR will provide the implementation.
- The Compact Framework CLR [does not provide the implementation](http://msdn.microsoft.com/en-us/library/2e08f6yc%28v=VS.90%29.aspx).
- When the JIT compiler executes a method for the first time, it looks up all of the methods it may call (loading other assemblies, etc).
- Since the CLR doesn't support `Delegate.BeginInvoke`, any method that calls it cannot be JIT-compiled, and therefore cannot be executed.
- The `NotSupportedException` is thrown when `OnBarcodeScan` is first called (and the JIT-compiler attempts to compile it and fails). This is why it cannot be stepped into by the debugger.


"
"What are correlations among between test statistics in multiple comparison procedures
In the book [Multiple Comparisons Using R](http://www.crcpress.com/product/isbn/9781584885740) Bretz et al write that their R package [multcomp](http://cran.r-project.org/web/packages/multcomp/index.html) improves power above other multiple comparison procedures by ""accounting for the structural correlations between the test statistics"" (pg 41). The default option for this multiple comparison procedure implements this and is called ""`single-step`.""  

My questions are 
1)What do they mean by ""structural correlations"" (their explanation eludes me)
2)Is their a standard way to refer to or cite this approach. ""Single-step"" is what it is called in the call to their R function, but this is a general description of [multiple comparisons](https://stats.stackexchange.com/questions/59999/what-do-single-step-and-multi-step-mean-in-post-hoc-testing-of-anovas). 
In my field people are mostly familiar with the usual Bonferroni corrections and I'd like to avoid the critical eye of reviewers and also point readers toward work that can help them understand what is going on.


","The Bonferroni procedure works for the worst case dependency between the involved single test statistics which is somehow negatively correlated. In contrast, the Sidák procedure works for independency between them. Clearly: You don't need to adjust if the multiple tests are copies from each other. Conversely, you have to adjust more, if one test tends to decide the opposite of the other one in order to meet the FWER.


In the special case of linear test statistics in normally distributed random variables (t-test), the dependency can pretty well be decribed by the correlation of the effect estimators. E.g. if you test simultaneously $ \mu\_1 - \mu\_2$ and $\mu\_1 - \mu\_3$, your test statistics will base on the respective means: $\bar{x}\_1 - \bar{x}\_2$ and $\bar{x}\_1 - \bar{x}\_3$. Both differences are correlated due to the *structure* of the hypotheses. Their covariance is the variance of $ \bar{x}\_1$, and so the test statistics will likewise be (asyptotically, i.e. up to studentization) correlated. That's why Sidák cannot be used. But if you incorporate the dependency as Bretz et al. suggest, you will meet your FWER(strong) less conservative than if you would use Bonferroni.


Single step means that the local alpha will be calculated for all tests instateously. You can use these local alphas for simultaneous confidence intervals that convey the same decision as the tests would do. Stepwise (like Holm) means, that the local alphas depend on preceding test decisions. This means that the simultaneous confidence intervals would alsodepend on single test decisions! But confidence intervals are supposed to give information about their respective parameters, not about So if you have to supply confidence intervals, stepwise procedures are a bad choice.


"
"Partial<> on specific key
I have a typescript interface describing mongodb database data like :



```
interface Foo {
  _id: string,
  name: string,

  fish: {
    _id: string,
    name: string,   
  },
}

```

I have a method that is building one data to insert in the database, like :



```
function pushNewFish() {
   const obj: Foo = {
      name: 'obiwan',

      fish: {
         name: 'kenobi',
      },
  };

  // ...
}

```

**The problem is** : I do not implements every `_id` elements that are specific to database.


I have the possibility to use `Partial<Foo>` but it's not what I'm looking for.


I also have the possibility to turn every `_id: string` into `_id ?: string`, but it feels wrong.


Is there a way to do : 



```
interface FooDatabase { ... }

type Foo = Partial<'_id', FooDatabase>;

```

? Thanks


","Maybe you want types like these:



```
type PartialK<T, K extends PropertyKey = PropertyKey> =
  Partial<Pick<T, Extract<keyof T, K>>> & Omit<T, K> extends infer O ?
  { [P in keyof O]: O[P] } : never;

```

Here, `PartialK<T, K>` acts like `Partial<T>` but only for the keys in `K`, leaving the rest alone. So `PartialK<T, keyof T>` or `PartialK<T, PropertyKey>` should act like `Partial<T>`. This should work reasonably well as long as `T` is an object type without index signatures, call signatures, or construct signatures. I've done a little trick with conditional types so that the resulting object type does not have an intersection in it, like this:



```
type TestPartialK = PartialK<{ a: string, b: number, c: boolean }, ""b"" | ""c"">
/* type TestPartialK = {
    b?: number | undefined;
    c?: boolean | undefined;
    a: string;
} */

```

Now that only works one level deep, and it looks like your question wants to do this recursively down through nested objects. So here's `NestedPartialK<T, K>`:



```
type NestedPartialK<T, K extends PropertyKey = PropertyKey> =
  T extends Function ? T :
  T extends Array<any> ? Array<NestedPartialK<T[number], K>> :
  T extends object ? PartialK<{ [P in keyof T]: NestedPartialK<T[P], K> }, K> :
  T;

```

It uses `PartialK` in its definition; it leaves functions and primitives alone, and uses recursive types to map arrays to arrays, and objects to objects. Note that it doesn't iterate tuples, but partial tuples are weird anyway and your use case doesn't even mention arrays, so I'm not worried about that much. Let's see if it works on your `FooDatabase`:



```
type Foo = NestedPartialK<FooDatabase, ""_id"">
/* type Foo = {
    _id?: string | undefined;
    foo: string;
    bar: string;
    tree: {
        _id?: string | undefined;
        again: {
            _id?: string | undefined;
            value: string;
        };
    };
} */

```

Looks good to me. All the `_id` props are now optional, and the other props are left alone. Okay, hope that helps; good luck!


[Playground link to code](https://www.typescriptlang.org/play/#code/C4TwDgpgBACghgJ2ASzgGwNIB4AqAaKDKCAD2AgDsATAZ1gQHtIkQMIQoBeepiFtkAD4uAKCixEKdFhjIAxgGtcBAKJkEcOcCwL2DAGZR8hQaagAyKAHkAtsm3GMw0uWp1kFfX2tQA-GKgAbygAbRgoDyhdEANrAF0ALmswuKgAXygkiggANz4AbhERUEgoADkIGnIqeCRUTGVCYjJKWh5mUAEudr5O9mFOAJxm1zaAMQBXCi1kBgo-I0yhkda6AEEEDRAsOAohBY2trAqqiBrJeuwcEIoJmwAjPjiCJ2EE5ZdVqAZ7gCsILQLWpSBrBMIRebRWI4RLlSrVYGXXApF7CNKopbiHCFYrgaA4eGI9BEbhE0FQOBJKoIDwAcwI9yyd0eCAIciS9wYDDQEF26QIACJ7gKoAAfKACuQCwQiAD0ACooCV8YSLsTuoEAuJ7r4mQ9vOKplQIPoPGdCuJxHJdVBOdzefNDdQTWaqBbLZSoNS6YUMvLZUUPOQEPpNNAxlyACJwYBwe5wGjQTXiAD6yCoVOANIotPd+i5mezuagAXjCELPqK4izEAgSWTlrTGa9WbpeCrHtpcA89a1jfTFZz7ctlpy6AmdZbReHlrSATnc9xpQjDG6JwRaoaK+jsfjicFTelcsVyqgK41fabNu9ObFUCNLuybr7+YYg9zfbL7-d1YQtd7I4jle753g+ppPj+gFwF2PZBH2gH9lQ16treTrGuB5rwQhUBjmgE7flh4hpJBxEiH6AZFHIcxVFA+hJOe3ANq+SQAOQsTOX5QGx7YBDWk4NuI0HdhQAGAbh+FcSxfZzkRZH5EAA)


"
"How to enable events so Workbook\_BeforeSave gets called
My `Workbook_BeforeSave` event is not called before saving


This is my code:



```
Option Explicit

Private Sub Workbook_BeforeSave(ByVal SaveAsUI As Boolean, Cancel As Boolean)
   a = MsgBox(""Do you really want to save the workbook?"", vbYesNo)
   If a = vbNo Then Cancel = True
End Sub

```

This is probably normal, because events are probably not enabled.
Now I tried to put `Application.Events = True` like this:



```
Option Explicit
Application.Events = True

Private Sub Workbook_BeforeSave(ByVal SaveAsUI As Boolean, Cancel As Boolean)
   a = MsgBox(""Do you really want to save the workbook?"", vbYesNo)
   If a = vbNo Then Cancel = True
End Sub

```

This doesn't change anything, Workbook\_BeforeSave is still not called up on saving. But when I close the excel file, following error message is displayed :


![enter image description here](https://i.stack.imgur.com/oopM8.png)


The english translation is ""Compilation error: Incorrect instruction outside of a procedure.""


Apparently the `Application.Events = True` is not at the right place, but where should I put it ?


","Hope these will help:


1. `Private Sub Workbook_BeforeSave(ByVal SaveAsUI As Boolean, Cancel As Boolean)` must be inside `ThisWorkbook` in a VBA project.
2. `Application.EnableEvents = True` can not be inserted outside procedure or function.
3. Events are enabled by default. So, there must be somewhere inside vba project Events are getting disabled. This can be searched by :


Once you are inside VBA project, Press Ctrl+F to open the Find dialog box. Then search for application.enableevents in the current project. Press Find Next. See the image below.
4. You can use a little sub to change and view the Application.EnableEvents status (ON/OFF). Place the sub under any standard module. See the image below.


![enter image description here](https://i.stack.imgur.com/ZqlUT.png)


"
"When to use promise.all()?
This is more of a conceptual question. I understand the Promise design pattern, but couldn't find a reliable source to answer my question about `promise.all()`:


### What is(are) the correct scenario(s) to use `promise.all()`


OR


### Are there any best practices to use `promise.all()`? Should it be ideally used only if all of the promise objects are of the same or similar types?


The only one I could think of is:


- Use `promise.all()` if you want to resolve the promise **only** if **all** of the promise objects resolve and reject if even one rejects.


","I'm not sure anyone has really given the most general purpose explanation for when to use `Promise.all()` (and when not to use it):



> 
> What is(are) the correct scenario(s) to use promise.all()
> 
> 
> 


`Promise.all()` is useful anytime you have more than one promise and your code wants to know when all the operations that those promises represent have finished successfully. It does not matter what the individual async operations are. If they are async, are represented by promises and your code wants to know when they have all completed successfully, then `Promise.all()` is built to do exactly that.


For example, suppose you need to gather information from three separate remote API calls and when you have the results from all three API calls, you then need to run some further code using all three results. That situation would be perfect for `Promise.all()`. You could so something like this:



```
Promise.all([apiRequest(...), apiRequest(...), apiRequest(...)]).then(function(results) {
    // API results in the results array here
    // processing can continue using the results of all three API requests
}, function(err) {
    // an error occurred, process the error here
});

```

`Promise.all()` is probably most commonly used with similar types of requests (as in the above example), but there is no reason that it needs to be. If you had a different case where you needed to make a remote API request, read a local file and read a local temperature probe and then when you had data from all three async operations, you wanted to then do some processing with the data from all three, you would again use `Promise.all()`:



```
Promise.all([apiRequest(...), fs.promises.readFile(...), readTemperature(...)]).then(function(results) {
    // all results in the results array here
    // processing can continue using the results of all three async operations
}, function(err) {
    // an error occurred, process the error here
});

```

On the flip side, if you don't need to coordinate among them and can just handle each async operation individually, then you don't need `Promise.all()`. You can just fire each of your separate async operations with their own `.then()` handlers and no coordination between them is needed.


In addition `Promise.all()` has what is called a ""fast fail"" implementation. It returns a master promise that will reject as soon as the first promise you passed it rejects or it will resolve when all the promises have resolved. So, to use `Promise.all()` that type of implementation needs to work for your situation. There are other situations where you want to run multiple async operations and you need all the results, even if some of them failed. `Promise.all()` will not do that for you directly. Instead, you would likely use something like `Promise.settle()` for that situation. You can see [an implementation of `.settle()` here](https://stackoverflow.com/questions/36605253/es6-promise-all-error-handle-is-settle-needed/36605453#36605453) which gives you access to all the results, even if some failed. This is particularly useful when you expect that some operations might fail and you have a useful task to pursue with the results from whatever operations succeeded or you want to examine the failure reasons for all the operations that failed to make decisions based on that.



> 
> Are there any best practices to use promise.all()? Should it be
> ideally used only if all of the promise objects are of the same or
> similar types?
> 
> 
> 


As explained above, it does not matter what the individual async operations are or if they are the same type. It only matters whether your code needs to coordinate them and know when they all succeed.




---


It's also useful to list some situations when you would **not** use `Promise.all()`:


1. When you only have one async operation. With only one operation, you can just use a `.then()` handler on the one promise and there is no reason for `Promise.all()`.
2. When you don't need to coordinate among multiple async operations.
3. When a fast fail implementation is not appropriate. If you need all results, even if some fail, then `Promise.all()` will not do that by itself. You will probably want something like `Promise.allSettled()` instead.
4. If your async operations do not all return promises, `Promise.all()` cannot track an async operation that is not managed through a promise.


"
"pyglet - loading/blitting image with alpha
I'm trying to make a simple application with pyglet. My main problem so far is that I can't seem to blit an image with alpha - all of the transparent pixels are converted into black pixels. I'm not sure whether the problem is with the loading of the image or the blitting. Here is a very basic overview of how I'm trying to render the image:



```
import pyglet
import pyglet.clock

window = pyglet.window.Window()

window.config.alpha_size = 8

#fancy text
text = pyglet.resource.image(""text.png"")

#background image
bg = pyglet.resource.image(""bg.png"")

bg.blit(0, 0)
text.blit(100, 100)

pyglet.app.run()

```

Any help is appreciated. Thanks in advance.


","You most likely just need to enable GL ALPHA blends.



```
from pyglet.gl import *
glEnable(GL_BLEND)
glBlendFunc(GL_SRC_ALPHA, GL_ONE_MINUS_SRC_ALPHA)

```

But first of all, your code is not able to run.
Mostly because you don't declare a `window.event` function to handle the `on_draw` where you normally render things.


Secondly, you never clear your window (which will cause a mess).


Here's a minimal working example of your code:



```
import pyglet
import pyglet.clock

window = pyglet.window.Window()

window.config.alpha_size = 8

#fancy text
text = pyglet.resource.image(""text.png"")

#background image
bg = pyglet.resource.image(""bg.png"")

@window.event
def on_draw():
    window.clear()

    bg.blit(0, 0)
    text.blit(100, 100)

pyglet.app.run()

```

Now this generates this:


[![enter image description here](https://i.stack.imgur.com/zMxyl.png)](https://i.stack.imgur.com/zMxyl.png)


And here's a working example of how you use the GL\_BLEND feature:



```
import pyglet
import pyglet.clock
from pyglet.gl import *

window = pyglet.window.Window()

window.config.alpha_size = 8

#fancy text
text = pyglet.resource.image(""text.png"")

#background image
bg = pyglet.resource.image(""bg.png"")

@window.event
def on_draw():
    window.clear()
    glEnable(GL_BLEND)

    glBlendFunc(GL_SRC_ALPHA, GL_ONE_MINUS_SRC_ALPHA)
    bg.blit(0, 0)
    glBlendFunc(GL_SRC_ALPHA, GL_ONE_MINUS_SRC_ALPHA)
    text.blit(100, 100)

pyglet.app.run()

```

This yields a result like so:


[![enter image description here](https://i.stack.imgur.com/PKPOo.png)](https://i.stack.imgur.com/PKPOo.png)


However, this code will quickly become messy.  

So there's two things you can do. You can first, put your images into sprite objects. Secondly, make this a bit more object oriented.


First, we'll use sprites.



```
self.fancy_background = pyglet.sprite.Sprite(pyglet.image.load('bg.png'))
self.fancy_background.draw()  # not blit!

```

Sprites automatically uses transparency, which makes your life (and code) a lot easier.


Secondly, we'll put these into a batch.  

Batches are made to bunch A LOT of sprites so you can call `.draw()` on the batch, and all sprites in that batch gets insta-rendered.



```
self.background = pyglet.graphics.Batch()
self.fancy_background = pyglet.sprite.Sprite(pyglet.image.load('bg.png'), batch=self.background)
self.background.draw() # background, not fancy_background! And also not blit!!

```

Last and most certainly not least.  

We'll put this into a class so we can do cool stuff later on.



```
import pyglet
import pyglet.clock
from pyglet.gl import *

key = pyglet.window.key

class main(pyglet.window.Window):
    def __init__ (self, width=800, height=600, fps=False, *args, **kwargs):
        super(main, self).__init__(width, height, *args, **kwargs)
        self.x, self.y = 0, 0

        self.background = pyglet.graphics.Batch()
        self.texts = pyglet.graphics.Batch()

        self.fancy_background = pyglet.sprite.Sprite(pyglet.image.load('bg.png'), batch=self.background)
        self.fancy_text = pyglet.sprite.Sprite(pyglet.image.load('text.png'), batch=self.texts)

        self.mouse_x = 0
        self.mouse_y = 0
        self.alive = 1

    def on_draw(self):
        self.render()

    def on_close(self):
        self.alive = 0

    def on_mouse_motion(self, x, y, dx, dy):
        self.mouse_x = x
        self.mouse_y = y

    def on_mouse_press(self, x, y, button, modifiers):
        if button == 1: # Left click
            pass

    def on_key_press(self, symbol, modifiers):
        if symbol == key.ESCAPE: # [ESC]
            self.alive = 0

    def render(self):
        self.clear()

        self.background.draw()
        self.texts.draw()

        self.flip()

    def run(self):
        while self.alive == 1:
            self.render()

            # -----------> This is key <----------
            # This is what replaces pyglet.app.run()
            # but is required for the GUI to not freeze
            #
            event = self.dispatch_events()

if __name__ == '__main__':
    x = main()
    x.run()

```

[![enter image description here](https://i.stack.imgur.com/2E78S.png)](https://i.stack.imgur.com/2E78S.png)


BAM.  



This code will enable you to create custom functions and custom ""player objects"" later on for instance. Also you can do collision detection easier and the code just looks a lot more structured (I threw in a little bonus features such as keyboard and mouse events).


Note tho, that the position of the sprites will default to `x=0, y=0` as shown in the last picture. You can set the position with `x=100` either on the variable/handle or when creating the sprite.


"
"Good approach to raise an exception
I have class `House` and module `Lockable`. Locking and unlocking `House` should reflect the real world, so you can't lock twice.


Do you think this is a good approach to using a module and raising an exception?



```
module Lockable
  attr_reader :locked

  def lock!
    raise StandardError, ""Already locked"" if @locked == true
    @locked = true
  end

  def unlock!
    raise StandardError, ""Already unlocked"" if @locked == false
    @locked = false
  end
end

class House
  include Lockable
end

```

Usage:



```
house = House.new
house.lock!
house.unlock!
house.unlock! # raise an exception

```

","I would define domain specific error classes like so:



```
module Lockable
  Error           = Class.new StandardError
  AlreadyLocked   = Class.new Error
  AlreadyUnlocked = Class.new Error

  def lock!
    raise AlreadyLocked if locked?
    @locked = true
  end

  def unlock!
    raise AlreadyUnlocked unless locked?
    @locked = false
  end

  def locked?
    # unless already defined, this assumes an
    # initial unlocked state (!!nil == false)
    !!@locked
  end
end

```

By having a common ancestor (`LockedError`), users can simply catch that one if they don't care about the inner state of the lockable object:



```
begin
  house.lock!
  house.lock!
rescue Lockable::Error
  # meh...
rescue StandardError
  # $! needs handling
end

```

*On the other hand*, if I try to lock an already locked lock, the key just won't move (and the lock doesn't blow up in my face). So, a more realistic reflection of the world would be this implementation:



```
module Lockable
  def lock!
    @locked = true
  end

  def unlock!
    @locked = false
  end
end

```

"
"TestNG: Is running classes in parallel thread safe?
When i run my test classes in parallel, the number of failed tests are higher than if i run the test classes serialized. I log a lot to the console for debugging purposes, but it is almost useless when running in parallel as everything logged is jumbled/out of order.


Before i try and fix the logging issue so i can debug better, i was wondering, if I set `parallel=""classes""`, does it guarantee thread safety? I assume if I set `parallel=""tests""` two methods could try to access the same object in scope, leading to thread safety issues.


ie:



```
<suite name=""test"" parallel=""classes"" thread-count=""3"">

```

","From TestNG Documentation [5.10.2 - Parallel tests, classes and methods](http://testng.org/doc/documentation-main.html#parallel-tests):



> 
> The *parallel* attribute on the tag can take one of following values:
> 
> 
> 
> ```
> <suite name=""My suite"" parallel=""methods"" thread-count=""5"">
> 
> ```
> 
> 
> ```
> <suite name=""My suite"" parallel=""tests"" thread-count=""5"">
> 
> ```
> 
> 
> ```
> <suite name=""My suite"" parallel=""classes"" thread-count=""5"">
> 
> ```
> 
> 
> ```
> <suite name=""My suite"" parallel=""instances"" thread-count=""5"">
> 
> ```
> 
> - **`parallel=""methods""`**: TestNG will run all your test methods in separate threads. Dependent methods will also run in separate threads but they will respect the order that you specified.
> - **`parallel=""tests""`**: TestNG will run all the methods in the same tag in the same thread, but each tag will be in a separate thread. This allows you to group all your classes that are not thread safe in the same and guarantee they will all run in the same thread while taking advantage of TestNG using as many threads as possible to run your tests.
> - **`parallel=""classes""`**: TestNG will run all the methods in the same class in the same thread, but each class will be run in a separate thread.
> - **`parallel=""instances""`**: TestNG will run all the methods in the same instance in the same thread, but two methods on two different instances will be running in different threads.
> 
> 
> 


As such, if by ""thread safe"" you mean there won't be more than one TestNG thread interacting with any given test class instance then yes, using ""classes"" will get you what you want. However, ""tests"" does not run each test method in its own thread but each `<test>` tag from you testng.xml. The descriptions above should help.


"
"IntelliJ IDEA 2016.1(64) : Unsupported java version Cannot start under Java 1.7.0\_79-b15: Java 1.8 or later is required
Currently I set `JAVA_HOME` environment variable to 1.7


When I try to open the file **`idea64.exe`** (path : `..\JetBrains\IntelliJ IDEA 2016.1\bin\idea64.exe`) below error is populated


[![enter image description here](https://i.stack.imgur.com/mlgbc.png)](https://i.stack.imgur.com/mlgbc.png)



> 
> *Unsupported java version  
> 
>  Cannot start under Java 1.7.0\_79-b15: Java 1.8 or later is required.*
> 
> 
> 


From the above error I understood that I have to set my `JAVA_HOME` to 1.8, but I don't want to set `JAVA_HOME` to 1.8.


I believe there should be a configuration file where in I can point to Java-8 and fix this, can anyone please let me know where can I do the configuration changes?


","You should create `IDEA_JDK_64` environment variable and set its value to `JDK8` instance location on your computer.


This `JDK8` instance can be used only for `idea` and `JAVA_HOME` can point to the Java 7 at the same time without any problems.


Via [documentation](https://intellij-support.jetbrains.com/hc/en-us/articles/206544879-Selecting-the-JDK-version-the-IDE-will-run-under):



> 
> idea64.exe uses this JDK search sequence:
> 
> 
> - IDEA\_JDK\_64 environment variable
> - ..\jre64 directory
> - system Registry
> - JDK\_HOME environment variable
> - JAVA\_HOME environment variable
> 
> 
> 


"
"Using Vault with multiple dynamic namespaces in Kubernetes
Is it possible to share a ServiceAccount between namespaces or somehow start a pod with a ServiceAccount from a different namespace? 


We are looking to use vault to store common secret data between dynamic development environments. Following the very good walk though [HERE](https://medium.com/@gmaliar/dynamic-secrets-on-kubernetes-pods-using-vault-35d9094d169) we were able to authenticate and pull secrets for a single namespace. However in our use case we will be creating a new namespace for each development environment during it's lifetime. 


If possible we would like to avoid having to also configure vault with a new auth backend for each namespace.


","When you create the Vault role, you can [configure `bound_service_account_namespaces` to be the special value `*`](https://www.vaultproject.io/api/auth/kubernetes/index.html#bound_service_account_namespaces), and allow a fixed service account name from any namespace. To adapt the ""create role"" [example from the documentation](https://www.vaultproject.io/docs/auth/kubernetes.html#configuration):



```
vault write auth/kubernetes/role/demo \
    bound_service_account_names=vault-auth \
    bound_service_account_namespaces='*' \
    policies=default \
    ttl=1h

```

You have to recreate the Kubernetes service account in every namespace, and it must have the exact name specified in the role. However, the Kubernetes service account is a single k8s object and it's not any harder than the Deployments, Services, ConfigMaps, and Secrets you already have; this pattern doesn't require any Vault reconfiguration.


(If you're using a templating tool like Helm, the service account can't follow a naming convention like `{{ .Release.Name }}-{{ .Chart.Name }}`: Vault doesn't appear to have any sort of pattern matching on this name.)


"
"Unix join command complexity
I was wondering if anyone would know the complexity of the Unix `join` command? I had assumed that it might be linear since both files need to be sorted.


Someone insisted to me that it was logarithmic which I doubt. Or perhaps it depends on the files and can be logarithmic (or `N*log(N)`) when one of the files is small and approach linear when both are large?


","The BSD `join` implementation is quite simple to follow and seems to be linear with regards to the number of lines in the files. This has gone mostly unchanged in all BSD systems since at least BSD 4.4 Lite2. The snippet below comes from [a current OpenBSD system](https://cvsweb.openbsd.org/cgi-bin/cvsweb/src/usr.bin/join/join.c?rev=1.27), for comparison, [this is a link to the BSD 4.4 Lite2 code](https://github.com/dspinellis/unix-history-repo/blob/BSD-Release/usr/src/usr.bin/join/join.c#L227) originally committed by Keith Bostic in 1991 (replacing an earlier version of the utility):



```
/*
 * We try to let the files have the same field value, advancing
 * whoever falls behind and always advancing the file(s) we output
 * from.
*/
while (F1->setcnt && F2->setcnt) {
        cval = cmp(F1->set, F1->joinf, F2->set, F2->joinf);
        if (cval == 0) {
                /* Oh joy, oh rapture, oh beauty divine! */
                if (joinout)
                        joinlines(F1, F2);
                slurp(F1);
                slurp(F2);
        }
        else {
                if (F1->unpair
                && (cval < 0 || F2->set->cfieldc == F2->setusedc -1)) {
                        joinlines(F1, NULL);
                        slurp(F1);
                }
                else if (cval < 0)
                        /* File 1 takes the lead... */
                        slurp(F1);
                if (F2->unpair
                && (cval > 0 || F1->set->cfieldc == F1->setusedc -1)) {
                        joinlines(F2, NULL);
                        slurp(F2);
                }
                else if (cval > 0)
                        /* File 2 takes the lead... */
                        slurp(F2);
        }
}

```

I looked at [the code for `join` in GNU coreutils](http://git.savannah.gnu.org/gitweb/?p=coreutils.git;a=blob_plain;f=src/join.c;hb=HEAD), but GNU code has so much going on in it that I really only can *guess*, based on the comments in the code, that it more or less also implements the same sort of algorithm:



```
/* Keep reading lines from file1 as long as they continue to
   match the current line from file2.  */

[...]

/* Keep reading lines from file2 as long as they continue to
   match the current line from file1.  */

[...]

```

If you take the sorting into account, and assume an `N*log(N)` sorting algorithm, then the complete time complexity would be `N*(1 + log(N))`, or `N*log(N)` for large `N` values. That is, the JOIN operation is faster than the sorting.


You can't do better than linear for a JOIN operation, because you can't skip lines (unless you have a precalculated index of some description and don't include the indexing in the time complexity). The best case scenario is that none of the lines join, in which case you need to read all lines from one of the two files and compare these to first line of the other file. The worst case scenario is that all lines join, in which case you need to read both files and do pairwise comparisons between both sets of lines (a linear operation on sorted files). If the user requests to see unpaired lines, then you are forced to read both files completely.


If you manage to do worse than linear for the JOIN alone, then you're doing something wrong.


"
"local (file://) website favicon works in Firefox, not in Chrome or Safari- why?
## The Question


When my website is local (using the `file://` protocol), the favicon doesn't show up in Chrome or Safari, but it works in Firefox (all on a mac). However, when the exact same site is actually hosted, the favicon works fine in all of the browsers I tried. Why do the webkit browsers not display local favicons?


## Details


The favicon.ico file is in the same directory as the index.html page. I am using the following code, although I've tried several variations of it:



```
   <link rel=""shortcut icon"" href=""favicon.ico"" type=""image/x-icon"" />

```

Since the favicon works when accessed over `http://`, I doubt that there's something wrong with the way I'm including it, but you never know...


## Update


I've found a few forum posts (nothing too official) that suggest that this is by design, but none of them give the rationale. Also, it sounds like some versions of IE have the same problem. Is there some security risk with local favicons? (I know that some browsers have issues with local cookies, for example.)


**Disclaimer:** There seem to be a ton of questions in a similar vein, but I haven't found exactly this one. (In fact I have basically the same question as [this one](https://stackoverflow.com/questions/3054747/favicon-issue-with-webkit-browsers), but none of the answers there worked for me.)


","You can use Base64 data of an image. Put something like this:



```
<link href=""data:image/x-icon;base64,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"" rel=""icon"" type=""image/x-icon"" />

```

into the `<head>` section of an html file.  

Base64 data can be received, for example, by [favicon.cc](http://www.favicon.cc)


"
"Difference between JE/JNE and JZ/JNZ
In x86 assembly code, are `JE` and `JNE` exactly the same as `JZ` and `JNZ`?


","`JE` and `JZ` are just different names for exactly the same thing: a
conditional jump when `ZF` (the ""zero"" flag) is equal to 1.


(Similarly, `JNE` and `JNZ` are just different names for a conditional jump
when `ZF` is equal to 0.)


You *could* use them interchangeably, but you *should* use them depending on
what you are doing:


- `JZ`/`JNZ` are more appropriate when you are explicitly testing
for something being equal to zero:



```
dec  ecx
jz   counter_is_now_zero

```
- `JE` and `JNE` are more appropriate after a `CMP` instruction:



```
cmp  edx, 42
je   the_answer_is_42

```

(A `CMP` instruction performs a subtraction, and throws the value of the result away, while keeping the flags; which is why you get `ZF=1` when the operands are equal
and `ZF=0` when they're not.)


"
"Hex combination of binary flags
Which of the following give back 63 as long (in Java) and how?



```
0x0
0x1
0x2
0x4
0x8
0x10
0x20

```

I'm working with NetworkManager API flags if that helps. I'm getting 63 from one of the operations but don't know how should I match the return value to the description.


Thanks


","63 is `32 | 16 | 8 | 4 | 2 | 1`, where `|` is the *binary or* operator.


Or in other words (in hex): 63 (which is 0x3F) is `0x20 | 0x10 | 0x8 | 0x4 | 0x2 | 0x1`. If you look at them all in binary, it is obvious:



```
0x20 : 00100000
0x10 : 00010000
0x08 : 00001000
0x04 : 00000100
0x02 : 00000010
0x01 : 00000001

```

And 63 is:



```
0x3F : 00111111

```



---


If you're getting some return status and want to know what it means, you'll have to use *binary and*. For example:



```
if (status & 0x02) 
{
}

```

Will execute if the flag 0x02 (that is, the 2nd bit from the right) is turned on in the returned status. Most often, these flags have names (descriptions), so the code above will read something like:



```
if (status & CONNECT_ERROR_FLAG)
{
}

```

Again, the status can be a combination of stuff:



```
// Check if both flags are set in the status
if (status & (CONNECT_ERROR_FLAG | WRONG_IP_FLAG))
{
}

```

P.S.: To learn why this works, [this](http://www.nitrogen.za.org/viewtutorial.asp?id=17) is a nice article about binary flags and their combinations.


"
"Share Node Addons
I have created a node addon and I am able to use it without any issues. I want to share the addon to other developers. What is the best way to share the addon so that other developers can use the same. Is it enough to share .node file only ?


","In general, a native add-on can be distributed like any other package on [npm](https://www.npmjs.com/)


However, there are some things to be aware of:


1. `*.node` files are shared libraries, so they are bound to a specific target platform. Your `*.node` file would only be suitable for the platform you built it for. For broader use, you'd have to build a dedicated `*.node` file for ever platform you want to support.
2. Used 3rd-party libraries: If your native add-on uses any 3rd-party libraries, you have to ensure that these libraries are present on the target system *and* can be found. You'd have to ship all required libs (and take care of during build) if you want to make sure they are available. Alternatively: Hard requirements for your package
3. Native add-on type: Depending on how you built your native add-on (e.g. NAN or N-API), your add-on might be suitable for certain node versions.


3.1. N-API: Your native add-on is usable with various node versions (see [this overview](https://nodejs.org/api/n-api.html#n_api_n_api_version_matrix))


3.2. NAN: NAN add-ons are built for specific node ABI versions, so an add-on built against a certain node ABI (e.g. 64) won't be usable with another node ABI (e.g. 56). You'd have to provide a dedicated `*.node` file for every node ABI you want to support


There's some useful tooling to help distributing native add-ons: [N-API docs](https://nodejs.org/api/n-api.html#n_api_build_tools)


- [prebuild](https://www.npmjs.com/package/prebuild) and [prebuild-install](https://www.npmjs.com/package/prebuild-install) will help you distribute your native add-ons via GitHub releases
- [node-pre-gyp](https://www.npmjs.com/package/node-pre-gyp) is able to retrieve add-ons from any server you provide


"
"Extending the range of bins in seaborn histogram
I'm trying to create a histogram with seaborn, where the bins start at 0 and go to 1. However, there is only date in the range from 0.22 to 0.34. I want the empty space more for a visual effect to better present the data.


I create my sheet with 



```
import pandas as pd
import matplotlib as mpl
import matplotlib.pyplot as plt
import numpy as np
import seaborn as sns

%matplotlib inline
from IPython.display import set_matplotlib_formats
set_matplotlib_formats('svg', 'pdf')

df = pd.read_excel('test.xlsx', sheetname='IvT')

```

Here I create a variable for my list and one that I think should define the range of the bins of the histogram.



```
st = pd.Series(df['Short total'])
a = np.arange(0, 1, 15, dtype=None)

```

And the histogram itself looks like this



```
sns.set_style(""white"")
plt.figure(figsize=(12,10))
plt.xlabel('Ration short/total', fontsize=18)
plt.title ('CO3 In vitro transcription, Na+', fontsize=22)

ax = sns.distplot(st, bins=a, kde=False)

plt.savefig(""hist.svg"", format=""svg"")
plt.show()

```

[Histogram](https://i.stack.imgur.com/0RfvH.jpg)


It creates a graph bit the range in x goes from 0 to 0.2050 and in y from -0.04 to 0.04. So completely different from what I expect. I google searched for quite some time but can't seem to find an answer to my specific problem.


Already, thanks for your help guys.


","There are a few approaches to achieve the desired results here. For example, you can change the xaxis limits after you have plotted the histogram, or adjust the range over which the bins are created.



```
import seaborn as sns

# Load sample data and create a column with values in the suitable range
iris = sns.load_dataset('iris')
iris['norm_sep_len'] = iris['sepal_length'] / (iris['sepal_length'].max()*2)
sns.distplot(iris['norm_sep_len'], bins=10, kde=False)

```

[![enter image description here](https://i.stack.imgur.com/Yl9vA.png)](https://i.stack.imgur.com/Yl9vA.png)


Change the xaxis limits (the bins are still created over the range of your data):



```
ax = sns.distplot(iris['norm_sep_len'], bins=10, kde=False)
ax.set_xlim(0,1)

```

[![enter image description here](https://i.stack.imgur.com/9lPxe.png)](https://i.stack.imgur.com/9lPxe.png)


Create the bins over the range 0 to 1:



```
sns.distplot(iris['norm_sep_len'], bins=10, kde=False, hist_kws={'range':(0,1)})

```

[![enter image description here](https://i.stack.imgur.com/v2Kz0.png)](https://i.stack.imgur.com/v2Kz0.png)


Since the range for the bins is larger, you now need to use more bins if you want to have the same bin width as when adjusting the xlim:



```
sns.distplot(iris['norm_sep_len'], bins=45, kde=False, hist_kws={'range':(0,1)})

```

[![enter image description here](https://i.stack.imgur.com/IJVOH.png)](https://i.stack.imgur.com/IJVOH.png)


"
"Geometry.getarea() and Polygon.getarea() what are the units?
So I am working with a set of images where the the corners are the lat long on the map. I am creating Polygons to check if there is overlap of the given images. I need to know the units associated with the .getarea() methods for both Polygon and geometry. I'm using the following objects to create my polygons and geometry


<http://tsusiatsoftware.net/jts/javadoc/com/vividsolutions/jts/geom/Geometry.html#getArea()>
<http://www.vividsolutions.com/jts/javadoc/com/vividsolutions/jts/geom/Polygon.html>


When using both of the above objects I get a number back however I have found no indication on what the units associated with the number are. So are we talking meters, kilometers, miles?


a sample of cordinates i'm using are 
+30.658739 -086.345670, +30.659997 -086.34002, +30.664041 -086.345082, +30.662783 -086.342750 I'm looking for the area between these 4 points.  

the value i get back from the .getArea() is 1.31039680000139E-5 The points are realitively close so i'm thinking its in meters which would be 1310.4 meters


","This is difficult to tell without knowing what the underlying coordinate system is. After taking a quick glance at the Javadoc API you linked to, I suspect that your geometry package is probably dealing in raw cartesian coordinates and is agnostic about the units of measurement being used. If this is the case, you are on a very slippery slope. Here's the problem:


**Not All Degrees Are Created Equal**


Degrees in latitude all have the same spatial resolution. Each degree of latitude corresponds to roughly 111km of distance. This is not the case for degrees of longitude. At the equator, degrees of longitude correspond to 111km of distance, but at the poles, 1 degree of longitude has ZERO km of distance. In other words, if you have a Lat/Lon box with the upper left at 10:0 and lower-right at 0:10, it will have more surface area than a Lat/Lon box with upper left at 20:0 and lower-right at 10:10, even though the sides of those boxes are all 10 degrees long.


A second issue is the curvature of the earth. Because of the earth's curvature, a 100km by 100km square on the earth's surface will have more surface area than 10000 km^2, because the shortest distance from one point on the earth's surface to another is not actually straight line, but an arc.


A third and often-overlooked but less-important issue is that the Earth is not actually a sphere, but an ellipsoid. It tends to bulge near the equator, which breaks our assumption that a degree of latitude anywhere on earth is the same distance as another degree of latitude anywhere else. However, this issue does not introduce as much error into our surface area estimates as the first two.


In other words, spherical (or in reality, *ellipsoidal*) surface area isn't an easy problem to solve, at least not as easy as mapping to cartesian coordinates and using them to find euclidean measurements of surface area. You can get away with it if the surface area you are dealing with spans a very small angular distance, but the larger your lat/lon box is, the more distortion you get.


**Possible Solution**


This only works if your images are rectangular and top/bottom of the image have constant latitude from left-to right. This still introduces more error with wider ranges of latitude because it still ignores the curvature of the earth, but does a better job than assuming all degrees are created equal in a cartesian coordinate system. If this is the case, then the intersection of your image will be bounded by the following coordinates:


*topLat*:\*leftLon\*, *bottomLat*:\*rightLon\*


Calculate the average latitude, then use it to find distance per degree of longitude at that average latitude, *kmPerLon*. We now have the following equation:


*area* = ((*topLat* - *bottomLat*) \* 111km) \* (*rightLon* - *leftLon*) \* *kmPerLon*)


The area you get from this will be measured in square kilometers, but again, I'd like to reiterate that this only works reasonably well if your images are rectangularly aligned to parallel latitudes and do not span much angular distance.


"
"The number of functions in C library
I want to know there are how many functions I can call in C library, e.g. Gnu C library? or Approximately?


","If you are in unix like OS, you can use `nm` utility, part of [gnu binutils](https://www.gnu.org/software/binutils/). In my cygwin environment, `nm` lists 1570 symbols defined in the text section.



```
% nm -C /usr/lib/libc.a  | grep -w T | wc -l
1570

```

Let's exclude the `_` variants



```
% nm -C /usr/lib/libc.a  | grep -w T | grep -v _ | wc -l
751

```

If you are on windows, try [dumpbin](https://support.microsoft.com/en-us/kb/177429) utility.



```
% dumpbin /exports msvcr110.dll

```

`-C` in the `nm` command demangles the symbol names. I do not know how to get `dumpbin` to print original symbol names. If anyone knows how to, please suggest.


Some functions might be defined in another object file - like libm for math, libnsl for network services etc. To be sure, also look at the library documentation/source.


"
"cygwin - file copy to UNC share
First, I know absolutely nothing about CYGWIN so bear with me...


Our developers need to procure some text dumps from our DBAs. They work with Oracle on Windows via a CYGWIN environment, and kornshell.


**Question:** How can we use kornshell/cygwin to copy an old fashioned text file to a standard Windows network share (UNC address)?




---


**Note:** There is a workaround we have discussed, where the DBAs can drop the text files in a local temp folder, and then setup a Windows Task to copy/move the text file to a UNC address. We would like to avoid this kluge.


","UNC paths are mapped to Cygwin paths by replacing `\` characters with `/` characters.


For example, a file that appears in Windows as



```
\\server\dir1\dir2\foo.txt

```

should appear in Cygwin as:



```
//server/dir1/dir2/foo.txt

```

If that doesn't work (it should), you can also map a drive letter to a UNC path. For example, if you map `\\server\dir1` to `S:\`, then the same file (`S:\dir2\foo.txt`) should appear in Cygwin as `/cygdrive/s/dir2/foo.txt`.


In short, a Cygwin command can treat a directory with a UNC path as an ordinary directory.


See <https://cygwin.com/cygwin-ug-net/using.html#unc-paths>


"
"Using vim's make command with a makefile that runs make on subdirectories
I have a project setup that has multiple sub-projects all with their own make files. There is a master make file in the main directory that cd's into each project and runs make in the proper order.


The makefile command looks like this:



```
sub-project/sub-project.a:
    cd sub-project ; make

```

The problem I have is that when I run make in vim and an error is reported, vim tries to open the file based on the relative path from inside whatever project had the error. Is there any way to have vim get the relative path from the root directory instead of the sub-project directory?


","Make sure your makefile is reporting when it enters and leaves the subdirectories, as follows:



```
make[1]: Entering directory `/home/username/projects/projectname/subdir1'
[... compile commands ...]
make[1]: Leaving directory `/home/username/projects/projectname/subdir1'

```

In my experience, vim will parse this portion of the output to find the correct paths for the files with errors. Also in my (frustrated) experience as of a few years ago, without the messages vim has no idea where to look for the files referenced by the error messages.


Note that this works even with recursive subdirectory building, as in this example:



```
make[1]: Entering directory `/home/username/projects/projectname/subdir1'
make[2]: Entering directory `/home/username/projects/projectname/subdir1/blah'
[... compile commands ...]
make[2]: Leaving directory `/home/username/projects/projectname/subdir1/blah'
make[1]: Leaving directory `/home/username/projects/projectname/subdir1'

```

"
"If an ASP.NET application uses many DLLs, what's the best approach for keeping configurations of each dll separate?
After creating settings for each `dll`, a `.dll.config` file is generated. If that `dll` be part of an asp.net application, how to keep this configurations separate for each `dll` and don't merge them into `web.config` ?


Example:
I have a GMailSender class library (DLL) which sends email via GMail servers. All you need is an instance of GMailSender, like this:



```
GMailSender gms = new GMailSender();
gms.To = ""myfriend@yahoo.com"";
gms.Subject = ""System.Configuration dilemma"";
gms.Body = ""Can anyone help me with this question ?"";
gms.Send();

```

Consider `GMailSender` is inside `GMailSender.dll` and it's configuration file is `GMailSender.dll.config` and the username and password of GMail account are inside it.


I want this `DLL` use the config file of itself (both dll and config in the same directory, i.e in `Bin` folder of an ASP.NET application) or beside a desktop application. This way the GMailSender is independent of who is using him for retrieving it's configurations (current `AppDomain` who has loaded this `DLL`).


I want this without reconstructing the wheel (no custom configuration classes). I guess its possible with `System.Configuration` but this namespace is probably the worst designed of .NET !


Please don't tell why you designed so that ...


This is a `plug-in` based design and finally `MEF` do some kind of it now in .NET 4.0, but there's the same problem for `Parts` configurations. With `MEF` at least I don't need anymore to argue for `plugin-based` design advantages.


","What you need can be achieved using **Custom Configuration Sections**. This [MSDN article](http://msdn.microsoft.com/en-us/library/2tw134k3.aspx) provides some details and samples on how to implement such config sections.


If you need simple **key/value** configs, like in the appSettings section, then creating a custom configuration section is quite simple.


The first thing you need to do is define your **config section** inside web.config:



```
<configSections>
  <section name=""GmailSettings"" restartOnExternalChanges=""true"" type=""System.Configuration.NameValueFileSectionHandler"" />
</configSections>

```

Now inside the web.config file you can **declare** the GmailSettings section and set the location of the external config file that you are going to use for that setting:



```
<GmailSettings configSource=""GmailSettings.config""></GmailSettings>

```

**configSource** specifies the name and location of the external config file used for defining the GmailSettings section. Please note the `restartOnExternalChanges` attribute used when defining the section. Set it to true if you want the application to be automatically restarted when you modify the GmailSettings.config file (as it does when you modify the web.config file).


Below is an example of how to implement the **GmailSettings.config** file:



```
<?xml version=""1.0""?>
<GmailSettings>
  <add key=""userName"" value=""blabla""/>
  <add key=""password"" value=""moreBla""/>
</GmailSettings>

```

You can access the settings from GmailSettings using the `ConfigurationManager.GetSection()` method or by implementing a **helper class** like the one below:



```
public class GmailSettings
{
    private static readonly GmailSettings _instance = new GmailSettings();
    private NameValueCollection _settings = ConfigurationManager.GetSection(""GmailSettings"") as NameValueCollection;

    public static GmailSettings Instance
    {
        get { return _instance; }
    }

    public string this[string key]
    {
        get { return _settings[key]; }
    }
}

```

The settings can now be accessed like `GmailSettings.Instance[""userName""]`.


Hope this helps.


"
"Function template overload resolution, dependent and non-dependent parameters
Given the following program



```
#include <iostream>

template<class T> struct id { using type = T; };

template<class T1, class T2>
int func(T1, T2) { return 0; }

template<class T1, class T2>
int func(typename id<T1>::type, typename id<T2>::type) { return 1; }

int main()
{
    std::cout << func<int, int>(0, 0) << std::endl;
}

```

GCC and Clang both prints `1` for this program. Is this program guaranteed to print `1` by the standard?


I tried finding the answer [here](https://en.cppreference.com/w/cpp/language/function_template#Function_template_overloading) but couldn't decipher it. It looks like the function templates might be equivalent and therefore breaks ODR but I'm not sure.


Does changing the second function template to 



```
template<class T>
using id_type = typename id<T>::type;

template<class T1, class T2>
int func(id_type<T1>, id_type<T2>) { return 1; }

```

make a difference?


","This is bog-standard partial ordering. We substitute unique types into one of the function templates and try to deduce the other against it. Do it both ways and if deduction only succeeds in one direction, we have an order. If you want to read the arcane rules, see [temp.func.order] and [temp.deduct.partial].


So here,


- Substituting `T1=U1, T2=U2` into the first overload's function type produces `int f(U1, U2);` Can we deduce `T1` and `T2` in the second overload from this? No; both are in non-deduced contexts. Ergo, deduction fails.
- Substituting `T1=U1, T2=U2` into the second overload produces `int f(id<U1>::type, id<U2>::type)` (this is conducted in the definition context so we can't substitute further into `id` - there may be a specialization somewhere). Can we deduce the `T1` and `T2` in the first overload from this? Yes, by deducing `T1 = id<U1>::type` and `T2 = id<U2>::type`. Deduction succeeds.


Since deduction succeeds only in one direction - deducing the first from transformed second - the second is more specialized than the first and is preferentially picked by overload resolution.


The alias template case changes nothing.




---


These templates are neither equivalent nor functionally equivalent.


"
"How to create a database with UTF-8 collation in PostgreSQL on Windows?
I'm configuring PostgreSQL db for the Bitbucket Server on Windows. In the official [guide](https://confluence.atlassian.com/bitbucketserver/connecting-bitbucket-server-to-postgresql-776640389.html) it says that:



> 
> The database must be configured to use the UTF-8 character set.
> 
> 
> 


It doesn't strictly say that you have to set collation to UTF-8, but for other atlassian procucts it's recommended so I assume that's the same case for Bitbucket Server. Exmaple from Confluence documentation:



> 
> - Character encoding must be set to utf8 encoding.
> - Collation must also be set to utf8. Other collations, such as ""C"", are known to cause issues with Confluence.
> 
> 
> 


This is what I have now, the problem is that it sets the collation to `English_United States.1252`:



```
CREATE DATABASE test
WITH OWNER ""postgres""
ENCODING 'UTF8'
LC_COLLATE = 'american_usa'
LC_CTYPE = 'american_usa'
TEMPLATE template0;

```

Is setting collation to UTF-8 actually necessary and if yes, how can I do it?


","Assuming that you are trying to create a PostgreSQL database with US locale sort order and character classification with UTF-8 encoding on Windows, following is a modification to the code example posted in the original question that may be used to achieve that result.



```
CREATE DATABASE ""example_db""
WITH OWNER ""postgres""
ENCODING 'UTF8'
LC_COLLATE = 'en-US'
LC_CTYPE = 'en-US'
TEMPLATE template0;

```

One liner format for terminal copy / paste:



```
CREATE DATABASE ""example_db"" WITH OWNER ""postgres"" ENCODING 'UTF8' LC_COLLATE = 'en-US' LC_CTYPE = 'en-US' TEMPLATE template0;

```

For anyone trying to create a similar database in a Linux environment such as Ubuntu on Windows Subsystem for Linux, you can do the following (depending on the specific environment, you may need to use `'en_US.UTF8'` as the locale instead):



```
CREATE DATABASE ""example_db""
WITH OWNER ""postgres""
ENCODING 'UTF8'
LC_COLLATE = 'en_US.UTF-8'
LC_CTYPE = 'en_US.UTF-8'
TEMPLATE template0;

```

One liner format for terminal copy / paste:



```
CREATE DATABASE ""example_db"" WITH OWNER ""postgres"" ENCODING 'UTF8' LC_COLLATE = 'en_US.UTF-8' LC_CTYPE = 'en_US.UTF-8' TEMPLATE template0;

```

"
"Autolayout and Scrollview width fixed to screen (not content)
I have followed and number of exmaples (most of which are the same) such as <https://www.youtube.com/watch?v=UnQsFlMGDsI>


I opena new project, drop a scrollview in and then a ""content"" view in that, constrain the scrollview to it's superview (top, bottom, left and right) then the Content view is constrained to the inside of the scrollview and then it's width and hight to the scrollview superview. Scrollview in red and content view black.


I drop a Textfield in and then constraint it to the top, leading and trailing of the content view. It all looks good in storyboard but when I run it the content view is not there and the textfield is the width of the objects in it (which are based on the textfield) in it.


[![None fixed height of content view](https://i.stack.imgur.com/kIITp.png)](https://i.stack.imgur.com/kIITp.png)


I then change the content view hight to be a fixed 800 and I can now see the content view but width is fixed by the size of the textfield (not the other way around), i.e. if I add more ""Test info"" text it all gets wider.


[![enter image description here](https://i.stack.imgur.com/VvtVS.png)](https://i.stack.imgur.com/VvtVS.png)


This is not my real project but I just wanted to eliminate reasons so did this from scratch and it does the same. I can not see what I am missing (hoping a tick box).


If I force the textfield to be 700 wide it scrolls left horizontally, I can not see how to pin the content view width to the screen size.


","**UPDATED ANSWER**


Sorry about the answer that I previously posted. I missed some points in that. Here I will explain to you step by step what need to be done(refer screenshots).


Here is an overview of the entire view with constraints added:


[![enter image description here](https://i.stack.imgur.com/uZbfW.png)](https://i.stack.imgur.com/uZbfW.png)


First thing to do, add the scrollview and add the constraint for leading, trailing, top and bottom with respect to superview.


[![enter image description here](https://i.stack.imgur.com/AIDbi.png)](https://i.stack.imgur.com/AIDbi.png)


Next thing to do, add a `UIView` inside the scroll view. Now this part is tricky, the view you just added need to act as the content view of the scroll view. So whatever views you need to add to the scroll view should actually go inside this view. For this view, set the constraints like, top, bottom, leading and trailing. But the tricky part here is, set the bottom constraint to superview(ie, the scrollview) as zero with low priority(250). Also set the width of this view as equal to the scroll view's superview, that is the ViewController's view as well as the scroll view.


[![enter image description here](https://i.stack.imgur.com/uAb3S.png)](https://i.stack.imgur.com/uAb3S.png)


Constraints for content view


[![enter image description here](https://i.stack.imgur.com/PuRCP.png)](https://i.stack.imgur.com/PuRCP.png)


Bottom constraint set to zero low priority.


[![enter image description here](https://i.stack.imgur.com/5SBk8.png)](https://i.stack.imgur.com/5SBk8.png)


Equal width with reference to scrollview's superview as well as the scroll view.


The bottom constraint is set to low priority since the constraints of the inner views need to be given higher priority when it comes to bottom constraint of the view for the sake of scrolling to work.


Now add the required view inside the content view, in your case the text field. Add all the required constraints with respect to the content view(top, leading, trailing, bottom). Add any number of subviews you wish to add, but ensure you specify the bottom constraint for the bottom most view with respect to the content view, that too with required priority.
Here are the constraints for the textfield that I added inside the content view:


[![enter image description here](https://i.stack.imgur.com/4ZSSF.png)](https://i.stack.imgur.com/4ZSSF.png)


Note that the bottom constraint in this case is required priority.


Final output:


[![enter image description here](https://i.stack.imgur.com/Sumuu.png)](https://i.stack.imgur.com/Sumuu.png)


Yellow = ViewController's view, Green = scroll view, Red = container view, pink = text field.


A note for anyone who need to make a horizontal scrollview, just switch the width of the content view with height and its bottom constraint with trailing constraint. That is, set equal heights with superview and scrollview and pin trailing space to scroll view with low priority.


"
"Use Linq and lambda to flatten a list
I have a class.



```
public class MedicalRequest
{
    private int id
    private IList<MedicalDays> Days 
    private string MedicalUser
    ...
}

```

and another



```
public class MedicalDays
{
    private int id;
    private DateTime? day
    private MedicalRequest request
    ...
}

```

I have the MedicalUser so I'm able to select an



```
IList<MedicalRequest> reqList = dao.FindAll(example);

```

What I would love to be able to do at this point is flatten out the Lists of MedicalDays and return the DateTime day. 


Something like



```
IList<DateTime> dateList = reqList.SelectMany(i => i.MedicalDays.day);

```

Can someone give me a push in the right direction? 


Thanks for your time. 


","You're nearly there:



```
IEnumerable<DateTime?> dateList = reqList.SelectMany(i => i.MedicalDays)
                                         .Select(m => m.day);

```

Or:



```
IEnumerable<DateTime?> dateList = reqList.SelectMany(i => i.MedicalDays,
                                                     (i, m) => m.day);

```

- If you need an `IList<T>` instead of `IEnumerable<T>` you can call `ToList()` on the result
- If you need to work with `DateTime` instead of `DateTime?` you can filter out null days like this:



```
IEnumerable<DateTime?> dateList = reqList.SelectMany(i => i.MedicalDays)
                                         .Select(m => m.day)
                                         .Where(x => x.HasValue)
                                         .Select(x => x.Value);

```


"
"TypeScript + React: defining defaultProps correctly
Say you define your component like so:



```
interface IProps {
  req: string;
  defaulted: string;
}

class Comp extends React.Component<IProps, void> {
  static defaultProps = {
    defaulted: 'test',
  };

  render() {
    const { defaulted } = this.props;

    return (
      <span>{defaulted.toUpperCase()}</span>
    );
  }
}

```

when you want to use it, TypeScript wants the `defaulted` prop from you, even though it's defined in `defaultProps`:



```
<Comp req=""something"" />  // ERROR: TypeScript: prop 'defaulted' is required

```

However, if you define the props interface like so:



```
interface IProps {
  req: string;
  defaulted?: string;  // note the ? here
}

```

then you cannot use it in:



```
render() {
  const { defaulted } = this.props;  // ERROR: prop 'defaulted' possibly undefined

  return (
    <span>{defaulted.toUpperCase()}</span>
  );
}

```

How to define the IProps, defaultProps and component correctly so the types make sense?


**EDIT:**


I'm using the `strictNullChecks` flag.


","I have an example with the following code (ComponentBase is just my wrapper around React.Component).


**Edit:** updated code to work with 'strictNullChecks' setting



```
interface IExampleProps {
    name: string;
    otherPerson?: string;
}

/**
 * Class with props with default values
 *
 * @class Example
 * @extends {ComponentBase<IComponentBaseSubProps, {}>}
 */
export class Example extends ComponentBase<IExampleProps, {}> {
    public static defaultProps: IExampleProps = {
        otherPerson: ""Simon"",
        name: ""Johnny""
    };

    constructor(props: IExampleProps) {
        super(props);
    }

    public render(): JSX.Element {
        const person: string = this.props.otherPerson === undefined ? """" : this.props.otherPerson;
        return(
            <div>
                <h1><small>Message by ComponentBaseSub: Hello {this.props.name} and {person} </small></h1>
            </div>
        );
    }
}

```

I have no issues using Visual Studio Code, TypeScript 2.0.3, TSLint 0.5.39.


"
"Jquery ajax button click event firing twice?
I have a Employee page which shows list of employees with an edit option. On clicking the edit button `jquery-ajax` is used to fetch the data from the server.
The problem is when I click the edit button the event is firing twice.


I am using a seperate js file and is referring the file to the main page.The script was working fine until i moved it to the seperate js file.


[![enter image description here](https://i.stack.imgur.com/dkZNL.png)](https://i.stack.imgur.com/dkZNL.png)


**The Jquery script is** 



```
  //ajaxGet on edit button click
$(document).on('click', '.editRole', ajaxGet);

 var ajaxGet = function (e) {       


    var spinner = $(this).parent('div').find('.spinner');
    var href = $(""#editMenuSettings"").data(""url"");
    var menuRoleId = $(this).data('id');

    spinner.toggle(true);

    var options = {
        type: ""GET"",
        url: href,
        data: { menuRoleId: menuRoleId }
    };

    $.ajax(options).success(function (data) {
        spinner.toggle(false);
        $("".modal-body"").html(data);
        $("".modal"").modal({
            backdrop: 'static'
        });
    });

    $.ajax(options).error(function (data) {
        spinner.toggle(false);
        toastr.error(""Oops..Some thing gone wrong"");
    });

    return false;

};

```

","You call `$.ajax` twice. 


At lines 



```
$.ajax(options).success(function(data)...

$.ajax(options).error(function(data)...

```

you actually make two **different** AJAX calls - one with `success` callback only, another one with `error` callback.


In your case, your call should look like this:



```
var options = {
    type: ""GET"",
    url: href,
    data: { menuRoleId: menuRoleId }
};

$.ajax(options)
    .success(function (data) {
        spinner.toggle(false);
        $("".modal-body"").html(data);
        $("".modal"").modal({
            backdrop: 'static'
        });
    })
    .error(function (data) {
        spinner.toggle(false);
        toastr.error(""Oops..Some thing gone wrong"");
    });

return false;

```

It will set both callbacks to the **single** AJAX call and execute this one.


"
"'async'/'await' in Node.js is not working in Node.js v8.1.0
I am trying to flush the Redis cache database and return the status in the response. But before the cache is cleared it returns the response.


In the below code the *console.log()* call will always print undefined since flushRedisDB is not awaited.


My Node.js version is v8.1.0


### File *myfile.js*



```
async function flushRedisapi(request, response)
{
    try {
        var value = await redisModules.flushRedisDB();
        console.log(""the value is: "" + value);
        if(value)
            response.status(200).send({status : 'Redis Cache Cleared'});
        else
            response.status(400).send({status : ""Redis Cache could not be flushed""});

    } catch (error) {
        response.status(400).send({status : ""Redis Cache could not be flushed""});
    }
}

```

### File *redismodule.js*



```
var redisClient;        // Global (avoids duplicate connections)

module.exports =
{
    openRedisConnection : function()
    {
        if (redisClient == null)
        {
            redisClient = require(""redis"").createClient(6379, 'localhost');
            redisClient.selected_db = 1;
        }
    },
    isRedisConnectionOpened : function()
    {
        if (redisClient && redisClient.connected == true)
        {
            return true;
        }
        else
        {
            if(redisClient)
                redisClient.end();  // End and open once more

            module.exports.openRedisConnection();

            return true;
        }
    },
    flushRedisDB: async function()
    {
        if(!module.exports.isRedisConnectionOpened())
            return false;

        await redisClient.flushall(function (err, result)
        {
            return (result == 'OK') ? true : false;
        });
    }
};

```

How can I solve this issue?


","Async/await only works with promises (as said in the comment). So just wrap your callback into a Promise.



```
function cbToPromise(asyncFunc, ...params) {
    return new Promise((resolve, reject) => {
        asyncFunc(...params, (err, result) => {
            if (err) reject(err);
            else resolve(result);
        });
    });
};

try {
    const result = await cbToPromise(redisClient.flushall);
    return result == 'OK';
}
catch(err) {
    console.error(err);
} 

```

**Addendum:**


It will work only if the signature of the callback is `function(err, result)`. Which, according to the other answer, would not be the case (no error passed as first parameter), as it never fail. So in this case just strip the err parameter, reject, and try/catch handler.


I let my answer in place here for convenience because it will most probably help you solve it for other Redis related methods.


"
"Can I Add GIF format Image as a Splash Screen
How can i add GIF format image as a splash screen. I tried to add but when i add it. It acts as a normal image. I don't know whether we can use Gif image as splash Screen. if there is any option to work that GIF format image. 


","Yes, you can Add Gif Image for the Splash Screen. To use Gif Image in your View.


add this dependency to your `build.gradle` file in the app module.


For the latest version.



```
dependencies {
    compile 'pl.droidsonroids.gif:android-gif-drawable:1.2.10'
}

```

and In the XML file add this way



```
<pl.droidsonroids.gif.GifTextView
    android:layout_width=""match_parent""
    android:layout_height=""match_parent""
    android:background=""@drawable/yourgif_image""
    />

```

and add your Gif Image to Drawable.


For more detail visit this: <https://github.com/koral--/android-gif-drawable>


**Or**


you can Create different Frame for Image and add it one after another in short interval of time.


"
"Calculating gravity in XNA
Alright, so I'm trying to create my own Physics engine for a 3D XNA game, and I'm having troubles calculating how much I should move my object by for gravity. 


XNA's game timer occurs every 16 milliseconds, so after some calculations, and using 9.81m/s as my gravitational velocity, you can see that you should increase the velocity of the object that has gravity by:


0.15696 meters/16milliseconds  

 - basically every update call should increase the object by 0.15696 meters


The questions is, how do I convert 0.15696 meters into pixels. Obviously if I just use a 1:1 relationship the object will only move 9.81 pixels/second. Which does not really simulate gravity :P Does anyone have a good idea on how I can determine how many pixels I should move the object by?


Thanks for all the help!


","Although 2d game dev is more pixel-centric, 3d game dev doesn't concern itself with pixels for this kind of stuff, but rather units. Unit size is completely arbitrary. 


Typically, you have a velocity vector whose magnitude is equivalent to the meters(or feet or ??) per second that your object is going. That way the position is updated each frame by adding this velocity \* the elapsed time since last frame (the 16.6666 ms). The acceleration from gravity is added to the velocity vector in the same way:



```
Vector3 gravity = Vector3.Down * 9.81f;


//in the update loop
float elapsed = (float)gameTime.ElapsedGameTime.TotalSeconds;//0.01666667 (or 16.67 ms) @ 60FPS
velocity += gravity * elapsed; 
position += velocity * elapsed;

```

In this example, I've arbitrarily established that 1 xna unit == 1 meter.


"
"confused about the need for $scope.$apply
I have an angular controller:



```
.controller('DashCtrl', function($scope, Auth) {
    $scope.login = function() {
        Auth.login().then(function(result) {
            $scope.userInfo = result;
        });
    };
});

```

Which is using a service I created:



```
.service('Auth', function($window) {
    var authContext = $window.Microsoft.ADAL.AuthenticationContext(...);

    this.login = function() {
        return authContext.acquireTokenAsync(...)
            .then(function(authResult) {
                return authResult.userInfo;
            });

    };
});

```

The Auth service is using a Cordova plugin which would be outside of the angular world. I guess I am not clear when you need to use a `$scope.$apply` to update your $scope and when you don't. My incorrect assumption was since I had wrapped the logic into an angular service then I wouldn't need it in this instance, but nothing gets updated unless I wrap the `$scope.userInfo =` statement in a `$timeout` or `$scope.$apply`.


Why is it necessary in this case?


","From [angular's wiki](https://github.com/angular/angular.js/wiki/When-to-use-$scope.$apply()):



> 
> AngularJS provides wrappers for common native JS async behaviors:
> 
> 
> ...
> 
> 
> jQuery.ajax() => $http
> 
> 
> This is just a traditional async function with a $scope.$apply()
>  called at the end, to tell AngularJS that an asynchronous event just
>  occurred.
> 
> 
> 


So i guess since your `Auth` service does not use angular's `$http`, `$scope.$apply()` isn't called by angular after executing the Async `Auth` function.



> 
> Whenever possible, use AngularJS services instead of native. If you're
>  creating an AngularJS service (such as for sockets) it should have a
>  $scope.$apply() anywhere it fires a callback.
> 
> 
> 


EDIT:


In your case, you should trigger the [digest cycle](https://www.ng-book.com/p/The-Digest-Loop-and-apply/) once the model is updated by wrapping (as you did):



```
Auth.login().then(function(result) {
   $scope.$apply(function(){
      $scope.userInfo = result;
   });
});

```

Or



```
Auth.login().then(function(result) {
    $scope.userInfo = result;
    $scope.$apply();
});

```

"
"How to merge sets which have intersections (connected components algorithm)?
Is there any effective way to merge sets which have intersections. For example:



```
l = [{1, 3}, {2, 3}, {4, 5}, {6, 5}, {7, 5}, {8, 9}]

```

The expected result is:



```
r = [{1, 2, 3}, {4, 5, 6, 7}, {8, 9}]

```

All sets which have intersection (common components) should be merged. For example:



```
{1, 3} & {2, 3}
# {3}

```

So these two sets should be merged:



```
{1, 3} | {2, 3}
# {1, 2, 3}

```

Unfortunately I don't have any working solution.


UPDATE: The order of sets in the result is not important.


","An efficient way to implement the [connected components algorithm](https://en.wikipedia.org/wiki/Connected_component_(graph_theory)) as mentioned by @mkrieger1 in the comments is to convert the list of sets to a set of hashable frozensets so that as you iterate through it and find a frozenset that intersects with the current one you can easily remove it from the pool:



```
pool = set(map(frozenset, l))
groups = []
while pool:
    groups.append(set(pool.pop()))
    while True:
        for candidate in pool:
            if groups[-1] & candidate:
                groups[-1] |= candidate
                pool.remove(candidate)
                break
        else:
            break

```

Given `l = [{1, 3}, {2, 3}, {4, 5}, {6, 5}, {7, 5}, {8, 9}]`, `groups` will become:



```
[{1, 2, 3}, {4, 5, 6, 7}, {8, 9}]

```

And given `l = [{1, 2}, {3, 4}, {2, 3}]`, `groups` will become:



```
[{1, 2, 3, 4}]

```

And given `l = [{1}, {2}, {1, 2}]`, `groups` will become:



```
[{1, 2}]

```

"
"Understanding java's native threads and the jvm
I understand that the jvm is itself an application that turns the bytecode of the java executable into native machine code, but when using native threads I have some questions that I just cannot seem to answer.


- Does every thread create their own
instance of the jvm to handle their
particular execution?
- If not then does the jvm have to have some way to schedule which thread it will handle next, if so wouldn't this render the multi-threaded nature of java useless since only one thread can be ran at a time?


","
> 
> Does every thread create their own instance of the JVM to handle their particular execution?
> 
> 
> 


No. They execute in the same JVM so that (for example) they can share objects and values of `static` fields.




---



> 
> If not then does the JVM have to have some way to schedule which thread it will handle next
> 
> 
> 


There are two kinds of thread implementation in Java. Native threads are mapped onto a thread abstraction which is implemented by the host OS. The OS takes care of native thread scheduling, and time slicing.


The second kind of thread is ""green threads"". These are implemented and managed by the JVM itself, with the JVM implementing thread scheduling. Java green thread implementations have not been supported by Sun / Oracle JVMs since Java 1.2. (See [Green Threads vs Non Green Threads](https://stackoverflow.com/questions/5713142/green-threads-vs-non-green-threads))




---



> 
> If so wouldn't this render the multi-threaded nature of Java useless since only one thread can be ran at a time?
> 
> 
> 


We are talking about green threads now, and this is of historic interest (only) from the Java perspective.


- Green threads have the advantage that scheduling and context switching are faster in the non-I/O case. (Based on measurements made with Java on Linux 2.2; <http://citeseerx.ist.psu.edu/viewdoc/summary?doi=10.1.1.8.9238>)
- With pure green threads, N programming language threads are mapped to a single native thread. In this model you don't get true parallel execution, as you noted.
- In a hybrid thread implementation, N programming language threads are mapped onto M native threads (where N > M). In this model, the in-process thread scheduler is responsible for the green thread to native thread scheduling AND you get true parallel execution (if M > 1); see <https://stackoverflow.com/a/16965741/139985>.


But even with the pure green threads, you still get concurrency. Control is switched to another threads a thread blocks on an I/O operation, whick acquiring a lock, and so on. Furthermore, the JVM's runtime could implement periodic thread preemption so that a CPU intensive thread doesn't monopolize the (single) core to the exclusion of other threads


"
"Pandas - how to avoid map converting int to floats

```
I have a dictionary:

matches = {282: 285,
 266: 277,
 276: 293,
 263: 264,
 286: 280,
 356: 1371,
 373: 262,
 314: 327,
 294: 290,
 285: 282,
 277: 266,
 293: 276,
 264: 263,
 280: 286,
 1371: 356,
 262: 373,
 327: 314,
 290: 294}

```

And a df, like so:



```
    team_id 
0   327 
1   293 
2   373 
3   282 
4   314 
5   263 
6   280
7   354 
8   264 
9   294 
10  1371    
11  262 
12  266 
13  356 
14  290 
15  285 
16  286 
17  275 
18  277 
19  276 

```



---


Now I'm trying to create an 'adversary\_id' column, mapped from the dict, like so:



```
df['adversary_id'] = df['team_id'].map(matches)

```

But this new column adversary\_id is being converted to `type` `float`, and two rows are ending up with `NaN`:


Why, if all data is type int?


How do I fix this?


","This is because the np.nan or NaN (they are not exact same) values you see in the dataframe are of type float. It is a limitation that pitifully can't be avoided as long as you have NaN values in your code.


Kindly read more in pandas' documentation [here](https://pandas.pydata.org/pandas-docs/stable/user_guide/missing_data.html).



> 
> Because NaN is a float, a column of integers with even one missing values is cast to floating-point dtype (see Support for integer NA for more). pandas provides a nullable integer array, which can be used by explicitly requesting the dtype:
> 
> 
> 


The proposed solution is to force the `type` with:



```
df['team_id'] = pd.Series(df['team_id'],dtype=pd.Int64Dtype())

```

Returning:



```
<class 'pandas.core.frame.DataFrame'>
RangeIndex: 5 entries, 0 to 4
Data columns (total 1 columns):
 #   Column   Non-Null Count  Dtype
---  ------   --------------  -----
 0   Example  4 non-null      Int64
dtypes: Int64(1)
memory usage: 173.0 bytes

```

"
"Unexpected results when comparing list comprehension with generator expression
I think I'm overlooking something simple, but I can't seem to figure out what exactly. Please consider the following code:



```
a = [2, 3, 4, 5]

lc = [ x for x in a if x >= 4 ] # List comprehension
lg = ( x for x in a if x >= 4 ) # Generator expression

a.extend([6,7,8,9])

for i in lc:
    print(""{} "".format(i), end="""")

for i in lg:
    print(""{} "".format(i), end="""")

```

I expected that both for-loops would produce the same result, so `4 5`. However, the for-loop that prints the generator exp prints `4 5 6 7 8 9`. I think it has something to do with the declaration of the list comprehension (Which is declared before the extend). But why is the result of the generator different, as it is also declared before extending the list? E.g. what is going on internally?


","Generators aren't evaluated until you call `next()` on them which is what makes them useful, while list comprehensions are evaluated immediately. 


So `lc = [4,5]` before extend and is therefore done. 


`lg` is still the same value at the start so the `extend` still applies to the `a` which hasn't finished being evaluated within the generator, meaning that `a` gets extended before you start printing it which is why it will print out longer with the rest of the numbers as well.


Check it out like this:



```
>>> a = [2, 3, 4, 5]
>>> lg = ( x for x in a if x >= 4 )
>>> next(lg)
4
>>> next(lg)
5
>>> a.extend([6,7,8,9])
>>> next(lg)
6

```

However, if you were to try calling an extra `next()` before extend you'll get `StopIteration` because the generator is exhausted at that point and then you won't be able to call it any longer.



```
>>> a = [2, 3, 4, 5]
>>> lg = ( x for x in a if x >= 4 )
>>> next(lg)
4
>>> next(lg)
5
>>> next(lg)
Traceback (most recent call last):
  File ""<stdin>"", line 1, in <module>
StopIteration
>>> a.extend([6,7,8,9])
>>> next(lg)
Traceback (most recent call last):
  File ""<stdin>"", line 1, in <module>
StopIteration

```

"
"Solve systems of ODEs in Julia with initial condition as 2D array
<http://diffeq.sciml.ai/latest/tutorials/ode_example.html>


I am trying to use the ODE solver in Julia (`DifferentialEquations.jl`) to solve a system of n interacting particles. Let's say that the system is in 2D, and the equation of motion of each particle is described by a second-order ODE of its position with respect to time. Then four variables are needed for each particle, two for positions and two for velocities. Then 4n variables are needed to be declared. Is there a way to generalize, such that one does not need to list all 4n equations one by one?


For example:


<http://diffeq.sciml.ai/latest/tutorials/ode_example.html#Example-2:-Solving-Systems-of-Equations-1>


I try to modify the Lorenz equation in the link above a little bit to n particles (which is a very very rough attempt since I actually have no idea how to do it) by trying to extend `u` and `du` to 2D arrays.



```
using DifferentialEquations
using Plots
n = 4
function lorenz(du,u,p,t,i)
    du[i,1] = 10.0*(u[i,2]-u[i,1])*sum(u[1:n,1])
    du[i,2] = (u[i,1]*(28.0-u[i,3]) - u[i,2])*sum(u[1:n,1])
    du[i,3] = (u[i,1]*u[i,2] - (8/3)*u[i,3])*sum(u[1:n,1])
end

u0 = hcat([1.0;0.0;0.0], [0.0;1.0;0.0], [0.0;0.0;1.0])
tspan = (0.0,100.0)
prob = ODEProblem(lorenz,u0,tspan)
sol = solve(prob)

```

This, without surprise, does not work, but I hope that you get the idea what I am trying to do. Is there anyway for the ODE solver to solve `u` as a 2D array (or other ways that can achieve similar purposes?)


","The problem is not the 2D Array. For example replacing your lorenz definition with 



```
function lorenz(du,u,p,t)
    du[1,1] = 10.0*(u[1,2]-u[1,1])
    du[1,2] = (u[1,1]*(28.0-u[1,3]) - u[1,2])
    du[1,3] = (u[1,1]*u[1,2] - (8/3)*u[1,3])
end

```

will work.


The problem is the function signature, the additional `i` is not supported. If you want to solve a network of Lorenz oscillators you have to code it with a function with the same signature, e.g. `lorenz_network!(du, u, p, t)` for the inplace version. Put a loop over the individual oscillators in your function and you are almost there. 


"
"CSS3 - style part of div after   
 tag
I have following div:



```
<div class=""mydiv"">
     This is some introduction text
     <br />
     click me
</div>

```

Is it possible to style the `click me` differently than the rest of the div using CSS only?


","You can style the `:first-line` differently. If there's only 2 lines, it's **kind of** feasible to style the 2nd and last line than the first one in pure CSS.


[Codepen](http://codepen.io/PhilippeVay/pen/qRqPeq)


BUT 


- you can't style [every property (MDN)](https://developer.mozilla.org/en-US/docs/Web/CSS/::first-line)
- being certain that a text will occupy exactly 2 lines would be ignoring narrow devices like smartphones (hello RWD) or zooming at the will of each user (graphical or text zooming). The web is not a PDF :)


+1 to Pevara suggestion: it should be a link or a button and then it can easily be styled





```
div {
  text-transform: uppercase;
  font-size: 2rem;
  color: red;
}
div::first-line {
  text-transform: initial;
  font-size: 1rem;
  color: initial;
}
```


```
<p>OK</p>
<div class=""mydiv"">
     This is some introduction text
     <br />
     click me
</div>
<hr>
<p>#fail</p>
<div class=""mydiv"" style=""width: 100px; border: 1px dotted blue"">
     This is some introduction text
     <br />
     click me
</div>
```






"
"CheerioJS, looping through  with same class name
I'm trying to loop through each `<ul>` and get the value of each `<li>`. The thing is, it only takes the first `<ul>` and skips the rest.


HTML



```
<div id=""browse-results"">
    <ul class=""tips cf"">
        <li>tip11</li>
        <li>tip12</li>
        <li>tip13</li>
    </ul>
    <ul class=""tips cf"">
        <li>tip21</li>        
        <li>tip22</li>        
        <li>tip23</li>        
    </ul>
    <ul class=""tips cf"">
        <li>tip31</li>        
        <li>tip32</li>        
        <li>tip33</li>        
    </ul>
    <ul class=""tips cf"">
        <li>tip41</li>        
        <li>tip42</li>        
        <li>tip43</li>        
    </ul>
</div>

```

Cheerio Parsing



```
$('#browse-results').find('.tips.cf').each(function(i, elm) {
    console.log($(this).text()) // for testing do text() 
});

$('#browse-results').children().find('.tips').each(function(i, elm) {
    console.log($(this).text())
});
I've tried many more

```

The output is just the values of the first `<ul>`.



```
tip11
tip12
tip13

```

Please note guys that this is just a snippet example with the same structure of what I'm trying to parse. 


I spent nearly 2 hours on this, and I can't find a way to do it.


","Try this:



```
var html = '<div id=""browse-results""> \
    <ul class=""tips cf""> \
        <li>tip11</li> \
        <li>tip12</li> \
        <li>tip13</li> \
    </ul> \
    <ul class=""tips cf""> \
        <li>tip21</li> \
        <li>tip22</li> \
        <li>tip23</li> \
    </ul> \
    <ul class=""tips cf""> \
        <li>tip31</li> \
        <li>tip32</li> \
        <li>tip33</li> \
    </ul> \
    <ul class=""tips cf""> \
        <li>tip41</li> \
        <li>tip42</li> \
        <li>tip43</li> \
    </ul> \
</div>';

var $ = cheerio.load(html);

$('#browse-results li').each(function(i, elm) {
    console.log($(this).text()) // for testing do text() 
});

```

This will select all `li` elements which are descendants of `#browse-results`.


"
"Read from word document line by line
  

I'm trying to read a word document using C#. I am able to get all text but I want to be able to read **line by line** and **store in a list and bind to a gridview**. Currently my code returns a list of one item only with all text (not line by line as desired). I'm using the **Microsoft.Office.Interop.Word** library to read the file. Below is my code till now:



```
    Application word = new Application();
    Document doc = new Document();

    object fileName = path;
    // Define an object to pass to the API for missing parameters
    object missing = System.Type.Missing;
    doc = word.Documents.Open(ref fileName,
            ref missing, ref missing, ref missing, ref missing,
            ref missing, ref missing, ref missing, ref missing,
            ref missing, ref missing, ref missing, ref missing,
            ref missing, ref missing, ref missing);

    String read = string.Empty;
    List<string> data = new List<string>();
    foreach (Range tmpRange in doc.StoryRanges)
    {
        //read += tmpRange.Text + ""<br>"";
        data.Add(tmpRange.Text);
    }
    ((_Document)doc).Close();
    ((_Application)word).Quit();

    GridView1.DataSource = data;
    GridView1.DataBind();

```

","Ok. I found the solution [here](http://mantascode.com/c-how-to-parse-the-text-content-from-microsoft-word-document/).


  

The final code is as follows:



```
Application word = new Application();
Document doc = new Document();

object fileName = path;
// Define an object to pass to the API for missing parameters
object missing = System.Type.Missing;
doc = word.Documents.Open(ref fileName,
        ref missing, ref missing, ref missing, ref missing,
        ref missing, ref missing, ref missing, ref missing,
        ref missing, ref missing, ref missing, ref missing,
        ref missing, ref missing, ref missing);

String read = string.Empty;
List<string> data = new List<string>();
for (int i = 0; i < doc.Paragraphs.Count; i++)
{
    string temp = doc.Paragraphs[i + 1].Range.Text.Trim();
    if (temp != string.Empty)
        data.Add(temp);
}
((_Document)doc).Close();
((_Application)word).Quit();

GridView1.DataSource = data;
GridView1.DataBind();

```

"
"Uninitialized variable behaviour in C++
I've checked myself, I wrote a program like this



```
int main() {
 int i;
 cout << i;
 return 0;
}

```

I ran the program a few times and the result was same all the time, zero.
I've tried it in C and the result was the same.


But my textbook says 



> 
> If you don’t initialize an variable that’s defined
>  inside a function, the variable value remain undefined.That means the element takes on
>  whatever value previously resided at that location in memory.
> 
> 
> 


How's this possible when the program always assign a free memory location to a variable? How could it be something other than zero(I assume that default free memory value is zero)?


","
> 
> How's this possible when the program always assign a free memory
>  location to a variable? How could it be something rather than zero?
> 
> 
> 


Let's take a look at an example practical implementation.


Let's say it utilizes stack to keep local variables.



```
void
foo(void)
{
        int foo_var = 42;
}

void
bar(void)
{
        int bar_var;
        printf(""%d\n"", bar_var);
}

int
main(void)
{
        bar();
        foo();
        bar();
}

```

Totally broken code above illustrates the point. After we call foo, certain location on the stack where foo\_var was placed is set to 42. When we call bar, bar\_var occupies that exact location. And indeed, executing the code results in printing 0 and 42, showing that bar\_var value cannot be relied upon unless initialized.


Now it should be clear that local variable initialisation is required. But could **main** be an exception? Is there anything which could play with the stack and in result give us a non-zero value?


Yes. **main is not the first function executed in your program**. In fact there is **tons** of work required to set everything up. Any of this work could have used the stack and leave some non-zeros on it. Not only you can't expect the same value on different operating systems, it may very well suddenly change on the very system you are using right now. Interested parties can google for ""dynamic linker"".


Finally, the C language standard does not even have the term stack. Having a ""place"" for local variables is left to the compiler. It could even get random crap from whatever happened to be in a given register. It really can be totally **anything**. In fact, if an undefined behaviour is triggered, the compiler has the freedom to do whatever it feels like.


"
"How to program in a type system that mismatches Haskell's System-Fw?
I'm studying optimal implementations of the λ-calculus. There is a particular subset of lambda terms that is very efficient. It corresponds to the type system of **Elementary Affine Logic** with fixed points. In order to test my implementations of that algorithm, I have to write moderately complex terms on that system. This is difficult without a infrastructure. I have to use the untyped lambda calculus and then manually add the types; no checking, unification, no type errors.


One idea would be to write programs in Haskell - taking benefit from its mature type-checker - and then translate to EAL. Unfortunately, there is a mismatch between System-Fw and EAL. For example, you can't express Scott-encoded ADTs in Haskell without `newtype`, due to the lack of a type-level `fix`. Moreover, Haskell is a complex language, and writing a `Haskell->EAl` compiler would not be trivial. 


Is there any quick/dirty way to get a working type checker/inferencer/unifier for that system - or at least something close enough - without having to program it all myself?


","Probably the quickest and easiest way is to embed your system into Haskell as an EDSL. The [finally, tagless](http://okmij.org/ftp/tagless-final/course/) approach may be ideal and there's an [example of encoding a linear type system](http://okmij.org/ftp/tagless-final/course/#linear) with it. In particular, I'd recommend using the [HOAS variation by Jeff Polakow](https://mail.haskell.org/pipermail/haskell-cafe/2013-March/107215.html). This will give you syntax like:



```
*Main> :t eval $ llam $ \x -> add x (int 1)
eval $ llam $ \x -> add x (int 1) :: Int -<> Int

```

which isn't too terrible. One aspect of the finally, tagless approach is that you can can have multiple interpretations for the same term, so you can have an interpretation that translates the term to some AST representing EAL, or have an interpretation that performs some additional typechecking if you aren't able to capture all of it in Haskell's type system.


"
"How to use ccache to speed up compiling of aosp?
I found the ccache aosp repo here <https://android.googlesource.com/toolchain/ccache/>, but I have no idea how to use ccache on aosp v10, the official aosp site doesn't have any info abouut ccache anymore, only multiple builds. Does ccache still work with recent versions of aosp ?


","
> 
> We no longer provide a ccache prebuilt.
> 
> 
> Ours was old, and had a number of issues that triggered non-reproducible
>  results and other failures. Newer ccache versions may fix some of those
>  issues, but at the large scale of our build servers, we weren't seeing
>  significant performance gains from using ccache -- you end up needing very
>  good locality and/or very large caches if you're building many different
>  configurations.
> 
> 
> Local no-change full rebuilds were showing better results, but why not just
>  use incremental builds at that point?
> 
> 
> So if you still want to use ccache, continue setting USE\_CCACHE, but also set
>  the CCACHE\_EXEC environment variable to the path to your ccache executable.
> 
> 
> 


<https://cs.android.com/android/platform/superproject/+/android10-release:build/make/core/ccache.mk;l=17>


"
"Python update Matplotlib from threads
I'm pretty new to the python world and unfortunately I could not find any solution to this yet.


I'm running python 3.6 with matplotlib==1.3.1 on Mac OS X 10.13.2


Currently I'm trying to build a small software which fetches data at 4Hz and displays the fetched data in a plot at 1Hz. Therefore I created a class which runs in a thread to fetch the data and another class to update the actual plots. In-between there is a data class which will hold the data and is used as interface inbetween the two classes.



```
import matplotlib.pyplot as plt
import numpy as np
import threading
import random
import time

class MyDataClass():

    def __init__(self):

        self.XData = [0]
        self.YData = [0]


class MyPlotClass(threading.Thread):

    def __init__(self, dataClass):

        threading.Thread.__init__(self)

        self._dataClass = dataClass
        self._period = 1
        self._nextCall = time.time()

        self.hLine, = plt.plot(0, 0)

        plt.ion()

    def run(self):

        while True:
            self.hLine.set_data(self._dataClass.XData, self._dataClass.YData)
            plt.draw()
            print(""updated %i datapoints"" % len(self._dataClass.XData))
            # sleep until next execution
            self._nextCall = self._nextCall + self._period
            time.sleep(self._nextCall - time.time())


class MyDataFetchClass(threading.Thread):

    def __init__(self, dataClass):

        threading.Thread.__init__(self)

        self._dataClass = dataClass
        self._period = 0.25
        self._nextCall = time.time()

    def run(self):

        while True:
            # add data to data class
            self._dataClass.XData.append(self._dataClass.XData[-1] + 1)
            self._dataClass.YData.append(random.randint(0, 256))
            print(""Added (%i, %i)"" % (self._dataClass.XData[-1], self._dataClass.YData[-1]))
            # sleep until next execution
            self._nextCall = self._nextCall + self._period
            time.sleep(self._nextCall - time.time())


data = MyDataClass()
fetcher = MyDataFetchClass(data)
plotter = MyPlotClass(data)

fetcher.start()
plotter.start()

fetcher.join()
plotter.join()

```

I can see that the threads are running due to the command line output. But for some reason the figure holding the plots won't show up.


The rocket symbol will just bounce up and down instead of showing up. See attached Screenshot.


Simple examples where the plot is only created ones and the plt.show() command is used works fine.


I can't figure out what I'm doing wrong. I hope anyone of you might have an idea.


Thanks!


Edit: Solutions presented here [How to update a plot in matplotlib?](https://stackoverflow.com/questions/4098131/how-to-update-a-plot-in-matplotlib) will not work for me because I don't want to be limited to a certain number of frames (using the animation framework of Matplotlib). I need to update the plot with 1Hz continiously.


[Figure not showing up](https://i.stack.imgur.com/xFzUe.png)


","I don't think you can run matplotlib GUI outside the main thread. So keeping the plotting in the main thread and using a `FuncAnimation` to steer the plotting, the following seems to work fine. 


Due to the `while True` loop it will run forever, even after closing the window, so for any real world application this should still be adjusted.



```
import matplotlib.pyplot as plt
from matplotlib.animation import FuncAnimation
import numpy as np
import threading
import random
import time

class MyDataClass():

    def __init__(self):

        self.XData = [0]
        self.YData = [0]


class MyPlotClass():

    def __init__(self, dataClass):

        self._dataClass = dataClass

        self.hLine, = plt.plot(0, 0)

        self.ani = FuncAnimation(plt.gcf(), self.run, interval = 1000, repeat=True)


    def run(self, i):  
        print(""plotting data"")
        self.hLine.set_data(self._dataClass.XData, self._dataClass.YData)
        self.hLine.axes.relim()
        self.hLine.axes.autoscale_view()


class MyDataFetchClass(threading.Thread):

    def __init__(self, dataClass):

        threading.Thread.__init__(self)

        self._dataClass = dataClass
        self._period = 0.25
        self._nextCall = time.time()


    def run(self):

        while True:
            print(""updating data"")
            # add data to data class
            self._dataClass.XData.append(self._dataClass.XData[-1] + 1)
            self._dataClass.YData.append(random.randint(0, 256))
            # sleep until next execution
            self._nextCall = self._nextCall + self._period;
            time.sleep(self._nextCall - time.time())


data = MyDataClass()
plotter = MyPlotClass(data)
fetcher = MyDataFetchClass(data)

fetcher.start()
plt.show()
#fetcher.join()

```

"
"Ckeditor character limitation with charcount plugin
How can i **prevent** users to enter new characters **after max character** limit is reached ?


Ckeditor charcount plugin just shows me the remaining characters, i want it to stop at 0. But it goes minus integers.


Here's my html code.



```
<textarea id=""myTextEditor1"" name=""myTextEditor""></textarea>
<script type=""text/javascript"">
        CKEDITOR.replace('myTextEditor1', { 
            height: 200, 
            extraPlugins: 'charcount', 
            maxLength: 10, 
            toolbar: 'TinyBare', 
            toolbar_TinyBare: [
                 ['Bold','Italic','Underline'],
                 ['Undo','Redo'],['Cut','Copy','Paste'],
                 ['NumberedList','BulletedList','Table'],['CharCount']
            ] 
        });
</script>

```

Do i have to use **onChange plugin** ? If i have to how can i limit users entering new characters ?


","I used the ckeditor jQuery adapter for this.



```
<textarea id=""myTextEditor1"" name=""myTextEditor""></textarea>
<script type=""text/javascript"">
    $(function () {
        var myEditor = $('#myTextEditor1');
        myEditor.ckeditor({ 
        height: 200, 
        extraPlugins: 'charcount', 
        maxLength: 10, 
        toolbar: 'TinyBare', 
        toolbar_TinyBare: [
             ['Bold','Italic','Underline'],
             ['Undo','Redo'],['Cut','Copy','Paste'],
             ['NumberedList','BulletedList','Table'],['CharCount']
        ] 
        }).ckeditor().editor.on('key', function(obj) {
            if (obj.data.keyCode === 8 || obj.data.keyCode === 46) {
                return true;
            }
            if (myEditor.ckeditor().editor.document.getBody().getText().length >= 10) {
                alert('No more characters possible');
                return false;
            }else { return true; }

        });
    });
</script>

```

The keyCode check is to allow backspace and delete key presses. To use the jQuery adapter, don't forget to insert it:



```
<script src=""/path-to/ckeditor/adapters/jquery.js""></script>

```

"
"Bumping package.json version without invalidating docker cache
I'm using a pretty standard Dockerfile to containerize a Node.js application:



```
# Simplified version
FROM node:alpine

# Copy package.json first for docker build's layer caching
COPY package.json package-lock.json foo/
RUN npm install

COPY src/ foo/
RUN npm run build

```

Breaking up my `COPY` into two parts was advantageous because it allowed Docker to cache the (long) `npm install` step.


Recently, however, I started bumping my `package.json` version using semver. This had the side effect of invalidating the Docker cache for the `npm install` step, lengthening my build times significantly.


Is there an alternative caching strategy I can use so that `npm install` only runs when my dependencies change?


","Here's my take on this, based off other answers, but shorter and with usage of `jq`:


**Dockerfile:**



```
FROM endeveit/docker-jq AS deps

# https://stackoverflow.com/a/58487433
# To prevent cache invalidation from changes in fields other than dependencies

COPY package.json /tmp

RUN jq '{ dependencies, devDependencies }' < /tmp/package.json > /tmp/deps.json

FROM node:12-alpine

WORKDIR /app

COPY --from=deps /tmp/deps.json ./package.json
COPY package-lock.json .

RUN npm ci
# https://docs.npmjs.com/cli/ci.html#description

COPY . .

RUN npm run build

LABEL maintainer=""Alexey Vishnyakov <n3tn0de@gmail.com>""

```

I extract `dependencies` and `devDependencies` fields to a separate file, then on next build step I copy it from the previous step as `package.json` (`COPY --from=deps /tmp/deps.json ./package.json`).


After `RUN npm ci`, `COPY . .` will overwrite gutted `package.json` with the original one (you can test it by adding `RUN cat package.json` after `COPY . .` command.


Note that [npm-scripts commands](https://docs.npmjs.com/misc/scripts) like `postinstall` won't run since they're not present in the file during `npm ci` and also if `npm ci` [is running from `root`](https://stackoverflow.com/questions/47748075/npm-postinstall-not-running-in-docker) and without `--unsafe-perm`


Either run commands after `COPY . .` or/and (if needed) include them via `jq` (changing command will invalidate cache layer) or add `--unsafe-perm`


**Dockerfile:**



```
FROM endeveit/docker-jq AS deps

COPY package.json /tmp

RUN jq '{ dependencies, devDependencies, peerDependencies, scripts: (.scripts | { postinstall }) }' < /tmp/package.json > /tmp/deps.json
# keep postinstall script 

FROM node:12-alpine

WORKDIR /app

COPY --from=deps /tmp/deps.json ./package.json
COPY package-lock.json .

# RUN npm ci --unsafe-perm 
# allow postinstall to run from root (security risk)

RUN npm ci
# https://docs.npmjs.com/cli/ci.html#description

RUN npm run postinstall

...

```

"
"bufferedreader won't move to the next line
Good evening,


I need to read some lines from a file, the file is csv and the file is structured like so:



```
4,
Mo Farah,30,
Jessica Ennis,27,

```

i need to read those values and put them into variables, that's how i tried to do this:



```
  while(nextline != null){
           StringTokenizer st = new StringTokenizer(nextline,"","");
           int size = Integer.parseInt(st.nextToken());
           System.out.println(size);

           nextline = reader.readLine();

           StringTokenizer st2 = new StringTokenizer(nextline, "",""); //why???


           String name = st2.nextToken();
           System.out.println(name);
           int age = Integer.parseInt(st2.nextToken());

           System.out.println(age);

```

the first int (4) is read just fine, however if i want to move to the next line it throws a noSuchElementException, so i had to write the line next to ""//why???"" in order to move to the next line, do i really have to instantiate a new tokenizer each time i want to move to the next line? or is there a better way of doing this?


Thank you very much


","## Lets have a look at what goes on



```
nextline = reader.readLine(); // I'm assuming you have this here
while(nextline != null){

```

The following will parse the line `""4,""` and split it into tokens separated by `,` it will contain 1 token: `""4""`:



```
       StringTokenizer st = new StringTokenizer(nextline,"","");

```

The following will succeed but you really should check that `st.hasMoreTokens()` returns true before doing this. It will consume the token `""4""` from the `st` which now contains no more tokens:



```
       int size = Integer.parseInt(st.nextToken());
       System.out.println(size);

       // We read the next line into ""nextline"", this is okay but not pretty.
       // Also you need to check that nextline is not `null` here.
       nextline = reader.readLine();

       StringTokenizer st2 = new StringTokenizer(nextline, "",""); //why???

```

## Why?


Because `st` was made to parse the line `""4,""`. Assigning `nextline` after the line has been parsed does not update the internal state of `st` (it's impossible to achieve). It simply doesn't work that way. So if you call `st.nextToken()` here, you will get `NoSuchElementException` as you said. Because you already read the `""4""` from `st` and there are no tokens left.


The following is the right thing to do if you want to use `StringTokenizer`, it will cause the line `""Mo Farah,30,""` to be parsed to tokens: `""Mo Farah""` and `""30""`:



```
       st = new StringTokenizer(nextline, "",""); // Just parse the new line.
       // You can reuse the old st variable as it is now depleted.

```

The following will now return the first token from the next line which is `""Mo Farah""`. Although again, you really should check `st.hasMoreTokens()` first and give a reasonable error message:



```
       String name = st.nextToken(); // Change from st2 to st
       System.out.println(name);
       int age = Integer.parseInt(st.nextToken()); // Ditto

       System.out.println(age);

```

So the short answer is, yes, if you want to use `StringTokenizer` you have to make a new `StringTokenizer` for every line. If you don't care above line endings you can just read all lines and concatenate them before passing all of it to the `StringTokenizer`.


But the easiest thing to do is just to use String tokens[] = newline.split("",""); as Mark did in his answer.


"
"How to execute a remote command to an existing ssh session
I'm trying to get basic informations on my local machine from ssh connections. The purpose here is to get the current user, hostname, current working directory in my tmux status bar.


I've enabled `ControlMaster` in my `~/.ssh/config`:



```
Host *
  ControlMaster auto
  ControlPath ~/.ssh/socket-%r@%h:%p

```

Now, I've managed to get the `ControlPath` from a homemade tmux plugin and I can send commands through the socket:



```
echo $(ssh -S ~/.ssh/socket-$user@$host:$port $host ""$command"")

```

This is working fine to get the remote hostname and the user. The issue is that when I change my remote user, `whoami` is still returning the user that was used to log in.


For example with `whoami` (`~` is local, `>` is executed on the ssh session):



```
~ ssh soyuka@host
> whoami
  soyuka
~ ssh -S ~/.ssh/socket-soyuka@host:22 host whoami
  soyuka

> sudo -s
> whoami
  root

~ ssh -S ~/.ssh/socket-soyuka@host:22 host whoami
  soyuka #Here I want to get root because the remote user is now root

```

For example with `pwd` (`~` is local, `>` is executed on the ssh session):



```
~ ssh soyuka@host
~ ssh -S ~/.ssh/socket-soyuka@host:22 host pwd
  /home/soyuka

> cd downloads
> pwd
  /home/soyuka/downloads

~ ssh -S ~/.ssh/socket-soyuka@host:22 host pwd
  /home/soyuka #Here I want to get /home/soyuka/downloads

```

Is this possible somehow to execute the command on the current ssh session by it's socket?


","What you're asking for does not seem to make sense.



> 
> Here I want to get root because the remote user is now root
> 
> 
> 


No, the remote user is now root. The fact that you've done `sudo -s` on the local system doesn't change that you're ssh'ing into the `soyuka` account remotely.


The answer for ""Here I want to get /home/soyuka/downloads"" is similar.



> 
> Is this possible somehow to execute the command on the current ssh session by it's socket?
> 
> 
> 


I believe that you may have a misconception that there is any persistent long-lived SSH session. In fact, every time you do:



```
ssh -l remoteuser host some-command

```

...a *new* SSH session is opened, `some-command` is executed in that session (without a pty, unlike SSH sessions opened without a remote command specified), and then that session is closed.


Even if there were such a persistent remote session, there is no reason to expect that doing things locally such as `sudo` or `cd` would affect that remote session in any way.


The fact that you are using SSH control sockets does not change this at all. The only thing that is different with a slave SSH session compared to a normal one is that the SSH **session** is piggybacked on an existing SSH **connection** instead of a brand new connection.


"
"How to use NSTimeInterval properly?
I'm new to Xcode and I have the following problem:


I want a method that gives me the time that as passed since a given date till now.
If it was 2 years ago I want to return a string ""2 years ago"", but if it was 5 minutes ago I want it to return a string with ""5 minutes ago"".


I've already checked NSTimeInterval and NSDateFormatter but I couldn't find a way to make it work. 


","Just to get you started...


You would create a reference date like this:



```
NSCalendar *gregorian = [[NSCalendar alloc]
                             initWithCalendarIdentifier:NSGregorianCalendar];

NSDateComponents *dateComponents = [[NSDateComponents alloc] init];
dateComponents.year = 2012;
dateComponents.month = 1;
dateComponents.day = 1;
dateComponents.hour = 0;
dateComponents.minute = 0;
dateComponents.second = 0;

NSDate *referenceDate = [gregorian dateFromComponents: dateComponents];

```

Current date can be obtained like this:



```
NSDate *now = [NSDate date];

```

Time interval is in seconds, so...



```
NSTimeInterval interval = [now timeIntervalSinceDate:referenceDate];
NSLog (@""reference date was %.0f seconds ago"", interval);

```

I'm sure you'll be able to figure it out on yourself from here on...


[This answer](https://stackoverflow.com/a/1386746/653513) might help you out.


"
"Typecasting an array to pointer?
Is is possible to typedef an array?


I have a set of vector function which all accept a pointer to a float which is an array of three floats. I can typedef float\* vec3\_t, however it will not let me create an object by simply setting it equal to an array in brackets.



```
typedef float* vec3_t;

vec3_t a = {1,1,1}; // Does not work
vec3_t b = (float[]){1,1,1}; // Works
float c[] = {1,1,1}; // Works

void f(vec3_t x);

f({1,1,1}); // Error
f((float[]){1,1,1}; // OK

```

Could someone please explain why this works this way?


","A pointer and an array are not the same thing. While they often behave in the same way, there are major differences, one of which you have just discovered.


## What your code actually does


I have substituted the typedefed type with real type, to make the explaination clearer.


`float c[] = {1,1,1};`


You have simply created and initialized an array


`f({1,1,1});`


The code above is neither a lvalue nor rvalue. The `{val1,...,valn}` syntax is *nothing more than an initializer* and can not be used elsewehere.


`float* b = (float[]){1,1,1};`


In here you have created and initialized an array and **then** stored it's location in a pointer.


`f((float[]){1,1,1};`


This case is the same as the one above, but instead of storing the pointer you pass it as an argument to a function.


`float* a = {1,1,1};`


You are attempting to write three variables to a memory location that is not allocated yet.
In general, `{valn,...,valn}`, is an initializer. At this moment you have nothing to initialize. Hence this syntax is invalid. You are trying to pour gas into a canister that has not yet been manufactured.


While I understand what you wanted to achieve, you seem to be missunderstanding the whole concept of memory and pointers. Imagine this code, which (in some dirty logic) is an equivalent of what you're trying to do:



```
float* a = NULL;
a[0] = 1;
a[1] = 1;
a[2] = 1;

```

What would happen if you executed this code?


So now you know why the compiler forbids it.


"
"DatePicker doesn't update view model unless focus changes to another control
How can I make the DatePicker control update the view model without losing focus?


I'd like to be able to run this sample, type 12/9/11 for the date, click the close button for the window and have the Debug.Assert pass. If I tab to the text box before closing the window it works fine.



```
<Window x:Class=""DatePickerTest.MainWindow""
        xmlns=""http://schemas.microsoft.com/winfx/2006/xaml/presentation""
        xmlns:x=""http://schemas.microsoft.com/winfx/2006/xaml""
        Title=""MainWindow"" Height=""350"" Width=""525"" Closing=""MainWindow_OnClosing"">
    <StackPanel>
        <DatePicker Name=""TheDate"" SelectedDate=""{Binding Path=SelectedDate}"" />
        <TextBox/>
    </StackPanel>
</Window>

```


```
using System;
using System.ComponentModel;
using System.Diagnostics;

namespace DatePickerTest
{
   public partial class MainWindow
   {
      public MainWindow()
      {
         InitializeComponent();
         DataContext = new ViewModel();
      }

      private void MainWindow_OnClosing(object sender, CancelEventArgs e)
      {
         var dataContext = DataContext as ViewModel;
         Debug.Assert(dataContext.SelectedDate == new DateTime(2011, 12, 9));
      }
   }

   public class ViewModel
   {
      public DateTime SelectedDate { get; set; }
   }

}

```

","Try changing the `UpdateSourceTrigger` to `PropertyChanged`, e.g.



```
<DatePicker Name=""TheDate"" 
            SelectedDate=""{Binding Path=SelectedDate
                 , UpdateSourceTrigger=PropertyChanged
                 , Mode=TwoWay}"" />

```

**Update**


Ok, it seems as though this is a [known issue](http://tsells.wordpress.com/2011/03/22/wpf-datepicker-not-updating-selecteddate-property/) with the `DatePicker`. You can bind to the `Text` property instead, and set a `TargetNullValue`:



```
<DatePicker
   Name=""TheDate""
   Text=""{Binding Path=SelectedDate
          , Mode=TwoWay
          , UpdateSourceTrigger=PropertyChanged
          , ValidatesOnDataErrors=True
          , TargetNullValue=''}""
/>

```

"
"Mass Downloading of Webpages C#
My application requires that I download a large amount of webpages into memory for further parsing and processing. What is the fastest way to do it? My current method (shown below) seems to be too slow and occasionally results in timeouts.



```
for (int i = 1; i<=pages; i++)
{
    string page_specific_link = baseurl + ""&page="" + i.ToString();

    try
    {    
        WebClient client = new WebClient();
        var pagesource = client.DownloadString(page_specific_link);
        client.Dispose();
        sourcelist.Add(pagesource);
    }
    catch (Exception)
    {
    }
}

```

","The way you approach this problem is going to depend very much on how many pages you want to download, and how many sites you're referencing.


I'll use a good round number like 1,000. If you want to download that many pages from a single site, it's going to take a lot longer than if you want to download 1,000 pages that are spread out across dozens or hundreds of sites. The reason is that if you hit a single site with a whole bunch of concurrent requests, you'll probably end up getting blocked.


So you have to implement a type of ""politeness policy,"" that issues a delay between multiple requests on a single site. The length of that delay depends on a number of things. If the site's robots.txt file has a `crawl-delay` entry, you should respect that. If they don't want you accessing more than one page per minute, then that's as fast as you should crawl. If there's no `crawl-delay`, you should base your delay on how long it takes a site to respond. For example, if you can download a page from the site in 500 milliseconds, you set your delay to X. If it takes a full second, set your delay to 2X. You can probably cap your delay to 60 seconds (unless `crawl-delay` is longer), and I would recommend that you set a minimum delay of 5 to 10 seconds.


I wouldn't recommend using `Parallel.ForEach` for this. My testing has shown that it doesn't do a good job. Sometimes it over-taxes the connection and often it doesn't allow enough concurrent connections. I would instead create a queue of `WebClient` instances and then write something like:



```
// Create queue of WebClient instances
BlockingCollection<WebClient> ClientQueue = new BlockingCollection<WebClient>();
// Initialize queue with some number of WebClient instances

// now process urls
foreach (var url in urls_to_download)
{
    var worker = ClientQueue.Take();
    worker.DownloadStringAsync(url, ...);
}

```

When you initialize the `WebClient` instances that go into the queue, set their `OnDownloadStringCompleted` event handlers to point to a completed event handler. That handler should save the string to a file (or perhaps you should just use `DownloadFileAsync`), and then the client, *adds itself back to the `ClientQueue`*.


In my testing, I've been able to support 10 to 15 concurrent connections with this method. Any more than that and I run into problems with DNS resolution (`DownloadStringAsync' doesn't do the DNS resolution asynchronously). You can get more connections, but doing so is a lot of work.


That's the approach I've taken in the past, and it's worked very well for downloading thousands of pages quickly. It's definitely not the approach I took with my high performance Web crawler, though.


I should also note that there is a *huge* difference in resource usage between these two blocks of code:



```
WebClient MyWebClient = new WebClient();
foreach (var url in urls_to_download)
{
    MyWebClient.DownloadString(url);
}

---------------

foreach (var url in urls_to_download)
{
    WebClient MyWebClient = new WebClient();
    MyWebClient.DownloadString(url);
}

```

The first allocates a single `WebClient` instance that is used for all requests. The second allocates one `WebClient` for each request. The difference is huge. `WebClient` uses a lot of system resources, and allocating thousands of them in a relatively short time is going to impact performance. Believe me ... I've run into this. You're better off allocating just 10 or 20 `WebClient`s (as many as you need for concurrent processing), rather than allocating one per request.


"
"jQuery to submit form to new tab?
I have an action link in a form that needs to be updated with a token every minute. I get the new url/token from an api call when the user clicks on the submit button. I'm using something like this



```
 <form id=""somelink"" action=""http://some.external.url/"" target=""_blank"" />

 $('#somebutton').click(function(e) {
      e.preventDefault();
      var url = '/apicall?id=' + $('#someid').val();                  
      $.post(url, function(data) {
           $('#somelink').attr('action', data);
           $('#somelink').submit();

```

problem is, using js/jQuery when it submits it opens in a new window without the typical menu and buttons in the browser and it's the size of a popup window.


Without the jQuery if I just clicked on the button it'll open in a new tab.


How can I accomplish this with jQuery?


","Instead of catching the `click` event, try catching the `submit` event of the form and in case you find any error you can use event.preventDefault();  

**Update**
Here i tested with this code and it's working fine,



```
<!doctype html />
<html>
  <head>        
    <title></title>
    <script type=""text/javascript"" src=""jquery.js""/></script>
</head>
<body>
  <form action='http://someSite.com' id='somelink' target='_blank'>
    <input type='text' name='lol' id='lol'/>
    <input type='submit'/>
  </form>
  <script type=""text/javascript""/>
      $('#somelink').on('submit',function(e) {
          if($('#lol').val() == ""poop""){ //Sample error check
              e.preventDefault();
              alert($('#lol').val());
          }
      });
  </script>
</body>
</html>

```

What this does is, if you enter ""poop"" it does not submit the form, for anything else it submits the form to new tab. Hope this helps :)   

**Update:**   

Regarding your comment, if in case of error you want to open some other tab, then replace `e.preventDefault()` with `$(this).attr({""action"":""http://www.someothersite.com""});` . Hope this helps :)  

**Update:**  

If you want to open a tab asynchronously, then you would have a hard time, this is the closest it can get.   

HTML Part  




```
<form id=""someForm"" action=""about:blank"" target=""newStuff"" />
<input type=""submit"" value=""12345"" id=""someid"" />

```

Javascript Part



```
$(document).ready(function() {
 $('#someForm').submit(function(e) {
    $.ajax({
        url: 'http://search.twitter.com/search.json?',
        dataType: 'jsonp',
        success: function(data) {
            if (data != """") {     
                var link = ""http://www.google.com/"";
                window.open(link,'newStuff'); //open's link in newly opened tab!
            }
        }
    });
 });
 });

```

"
"bootstrap not working in laravel 5.6
I tried adding a navigation bar to my laravel version 5.6 project but it doesn't add styles and nor JS 


[![image](https://i.stack.imgur.com/dauLD.png)](https://i.stack.imgur.com/dauLD.png)


**my code :**



```
<nav class=""navbar navbar-inverse "">
    <div class=""container"">
      <div class=""navbar-header"">
        <button type=""button"" class=""navbar-toggle collapsed"" data-toggle=""collapse"" data-target=""#navbar"" aria-expanded=""false"" aria-controls=""navbar"">
          <span class=""sr-only"">Toggle navigation</span>
          <span class=""icon-bar""></span>
          <span class=""icon-bar""></span>
          <span class=""icon-bar""></span>
        </button>
        <a class=""navbar-brand"" href=""#"">LSAPP</a>
      </div>
      <div id=""navbar"" class=""collapse navbar-collapse"">
        <ul class=""nav navbar-nav"">
          <li><a href=""/"">Home</a></li>
          <li><a href=""/about"">About</a></li>
          <li><a href=""/services"">services</a></li>
        </ul>
      </div><!--/.nav-collapse -->
    </div>
  </nav>

```

","Considering you have a fresh laravel app you can add on your head



```
<!-- Styles -->
<link href=""{{ asset('css/app.css') }}"" rel=""stylesheet"">

```

and at the end of your body



```
<!-- Scripts -->
<script src=""{{ asset('js/app.js') }}""></script>

```

to use the Laravel css and javascript boilerplate. 


**Remember that Laravel 5.6 uses Bootstrap 4.0, so you need to use its syntax:**



```
<nav class=""navbar navbar-expand-lg navbar-light bg-light"">
  <a class=""navbar-brand"" href=""#"">LSAPP</a>
  <button class=""navbar-toggler"" type=""button"" data-toggle=""collapse"" data-target=""#navbarNav"" aria-controls=""navbarNav"" aria-expanded=""false"" aria-label=""Toggle navigation"">
    <span class=""navbar-toggler-icon""></span>
  </button>
  <div class=""collapse navbar-collapse"" id=""navbarNav"">
    <ul class=""navbar-nav"">
      <li class=""nav-item active"">
        <a class=""nav-link"" href=""#"">Home <span class=""sr-only"">(current)</span></a>
      </li>
      <li class=""nav-item"">
        <a class=""nav-link"" href=""/about"">About</a>
      </li>
      <li class=""nav-item"">
        <a class=""nav-link"" href=""/services"">Services</a>
      </li>
    </ul>
  </div>
</nav>

```

Check the component documentation [here](https://getbootstrap.com/docs/4.0/components/navbar/).


Also bair in mind that the default Laravel assets have more stuff than only JQuery and Bootstrap.


"
"Wifi losing connection, weak signal, Intel 7260 adapter
I yesterday upgraded from Windows to Ubuntu 15.04 but I have problem with my wireless adapter Intel Corporation Wireless 7260 (rev 73). It randomly loses internet connection. Restarting adapter always helps. (same it did with Windows, except that sometimes restarting adapter didn't help). 


Additionally, I have weaker connection then I had with Windows. 
Thank you!


","You can try this option. Run in terminal



```
sudo tee /etc/modprobe.d/iwlwifi-opt.conf <<< ""options iwlwifi 11n_disable=1""

```

and reboot.


If this does not help, some other options can be used.


It looks like this adapter does not work well in 802.11n mode. So disabling it improves connection.


There are some other options, that may be tuned.



```
power_save=0
bt_coex_active=0
swcrypto=1

```

You can also try to disable not all 802.11n, but partially. It can be selected by `11n_disable=2`, `11n_disable=4` or `11n_disable=8` instead of `11n_disable=1`.


You can apply them the same way as with `11n_disable=1`. You can use any combination of these options, e.g.



```
sudo tee /etc/modprobe.d/iwlwifi-opt.conf <<< ""options iwlwifi power_save=0 swcrypto=1""

```

You can enable all of them and remove one by one to test which one does the trick.


Every new setting overwrites the previous one.


To completely remove the options, remove `/etc/modprobe.d/iwlwifi-opt.conf` by running



```
sudo rm /etc/modprobe.d/iwlwifi-opt.conf

```

***Please do not forget to comment on which options worked for you.***


Explanation of the options can be found by



```
 modinfo iwlwifi | grep parm

```

"
"Java FizzString
I am testing what knowledge I have with the [FizzBuzz](http://codingbat.com/doc/practice/fizzbuzz-code.html) tests in Java. This is the one I just completed:



> 
> Given a string str, if the string starts with ""f"" return ""Fizz"". If the string ends with ""b"" return ""Buzz"". If both the ""f"" and ""b"" conditions are true, return ""FizzBuzz"". In all other cases, return the string unchanged. 
> 
> 
> 


My code (working):



```
    public String fizzString(String str){

    if (str.startsWith(""f"") && str.endsWith(""b"")) {
        return ""FizzBuzz"";
    }

    if (!str.startsWith(""f"")) {
        if (!str.endsWith(""b"")) {
            return str;
        } else return ""Buzz"";
    } else return ""Fizz"";
}

```

Questions: How can I make this code more efficient? Should I be assigning ""Fizz"" and ""Buzz"" to separate `string` variables, so that there is just one `return` statement at the end of the method? Should there be fewer `if` statements - if so, I'm not sure how it would be done. How would you go about the problem?


","Here is a very direct translation of the problem into code.



```
public String fizzString(String str) {
    if (str.startsWith(""f"") && str.endsWith(""b"")) {
        return ""FizzBuzz"";
    } else if (str.startsWith(""f"")) {
        return ""Fizz"";
    } else if (str.startsWith(""b"")) {
        return ""Buzz"";
    } else {
        return str;
    }
}

```

Compared to your original code, it is easier to read because:


- There is no nesting
- All conditions are affirmative, not a mixture of affirmative and negative


There is no advantage to assigning the result to a variable and returning that variable.


If you want a single `return` statement, you can write:



```
public String fizzString(String str) {
    return str.startsWith(""f"") && str.endsWith(""b"") ? ""FizzBuzz""
         : str.startsWith(""f"")                      ? ""Fizz""
         :                        str.endsWith(""b"") ? ""Buzz""
                                                    : str;
}

```

… which likely compiles to exactly the same code. Choose whichever version you think is more readable.


"
"Get CSS class definition from chrome developer tools console
I would like to programatically retrieve a set of CSS class definitions from chrome developer tools. In effect similar to what is displayed in styles tab in the right hand side. The input needs to be a class name and the output should be all the styles defined in it. 


I'm aware of getComputedStyle DOM method, but this doesn't separate into separate classes which I need.


","This approach worked for me ([stackoverflow.com/a/27527462/1023562](https://stackoverflow.com/a/27527462/1023562)): 



```
 /**
     * Gets styles by a classname
     * 
     * @notice The className must be 1:1 the same as in the CSS
     * @param string className_
     */
    function getStyle(className_) {
    var styleSheets = window.document.styleSheets;
    var styleSheetsLength = styleSheets.length;
    for(var i = 0; i < styleSheetsLength; i++){
        var classes = styleSheets[i].rules || styleSheets[i].cssRules;
        var classesLength = classes.length;
        for (var x = 0; x < classesLength; x++) {
            if (classes[x].selectorText == className_) {
                var ret;
                if(classes[x].cssText){
                    ret = classes[x].cssText;
                } else {
                    ret = classes[x].style.cssText;
                }
                if(ret.indexOf(classes[x].selectorText) == -1){
                    ret = classes[x].selectorText + ""{"" + ret + ""}"";
                }
                return ret;
            }
        }
    }

}

```

It lets you invoke the `javascript` code in Chrome console like this: 



```
console.log(getStyle('#heder_logo a')); 

```

and get results like this: 



```
> #heder_logo a { width: 200px; height: 114px; display: block; }.

```

I did have issues with some CSS files which were not on the same domain (they were pulled from CDN), but there are variety of proposals in that thread, so some should work for you.


"
"How do I reset the useMutation hook of Apollo Client
I am developing a React form that's tied to a GraphQL mutation using the [`useMutation`](https://www.apollographql.com/docs/react/data/mutations/) of [Apollo Client](https://www.apollographql.com/docs/react/). On the server, I perform some validation and in case of errors, I reject the mutation. On the client-side, I use the `error` object to receive the validation errors. My hook looks like this:



```
const [addDrone, { error }] = useMutation(ADD_DRONE)

```

So I unpack the `error` object and present it to the user in a dialog to let him/her know what went wrong. After the user dismisses the dialog, I want to give the user a chance to fix the error so he/she can resubmit the form. This is where things get hairy. I want to clear the `error` object when the user dismisses the dialog, but since this variable comes from the `useMutation` hook there is no way for me to mutate or reset it. It looks like the `useMutation` was designed to be fired once, and not used again.


So my question is, is there a way to ""reset"" a `useMutation` hook back to it's original state?


","**UPDATE 12/22/2021**: As of version [3.5.0](https://github.com/apollographql/apollo-client/blob/main/CHANGELOG.md#apollo-client-350-2021-11-08) of `@apollo/client` the `useMutation` hook now provides a `reset` method. You can use this method to reset the hook back to its initial state. E.g.:



```
const [addDrone, { error, reset }] = useMutation(ADD_DRONE)

```

Below are some notes from [the official documentation on this method](https://www.apollographql.com/docs/react/data/mutations/#resetting-mutation-status).



> 
> Call reset to reset the mutation's result to its initial state (i.e.,
> before the mutate function was called). You can use this to enable
> users to dismiss mutation result data or errors in the UI.
> 
> 
> Calling reset does not remove any cached data returned by the
> mutation's execution. It only affects the state associated with the
> useMutation hook, causing the corresponding component to rerender.
> 
> 
> 




---


Old answer, kept for those who use an older version of `@apollo/client`:


As of 12/26/2020 there's no way to reset a `useMutation`, however there are two feature requests [#157](https://github.com/apollographql/apollo-feature-requests/issues/157) [#170](https://github.com/apollographql/apollo-feature-requests/issues/170) on the Apollo Feature Request repo asking for one, so this may change in the future.


The easiest way to implement this behavior right now is to drop the `error` variable and use the [`onError`](https://www.apollographql.com/docs/react/data/mutations/#usemutation-api) handler instead. This is a callback provided by `useMutation` which uses the same `error` object. You can then wire that to your own stateful variable, which gives you the exact same `error` variable *and* the ability to use `setError` to clear it. See this example below:



```
  const [error, setError] = React.useState(null)
  const [addDrone] = useMutation(ADD_DRONE, {
    onError: setError,
  })

```

If you're looking to reset the `data` object as well you can introduce another stateful tuple `[data, setData]` and wire the `setData` to the `onCompleted` callback.


"
"Modify OHLC resample code as per deprecated warning
**Issue:**


When working with market data and resampling intra-day data to the daily timeframe as follows:



```
ohlc_dict = {
'Open':'first',
'High':'max',
'Low':'min',
'Last': 'last',
'Volume': 'sum'}

data.resample('1D',how=ohlc_dict).tail().dropna()

                Open    High    Last    Low     Volume
    Timestamp                   
    2016-12-27  163.55  164.18  164.11  163.55  144793.00
    2016-12-28  164.18  164.33  164.22  163.89  215288.00
    2016-12-29  164.44  164.65  164.49  164.27  245538.00
    2016-12-30  164.55  164.56  164.18  164.09  286847.00

```

Which seems to gives me the output I need (still need to verify)... 


I get the following warning:



```
FutureWarning: how in .resample() is deprecated
the new syntax is .resample(...)..apply(<func>)

```

**Question:**


How would this `resample` code be replicated using the new syntax to align with the current best practice using `apply`?


**What I have tried:**


Just using data['Low'] as an example:



```
def ohlc (df):
    return df['Low'].min()

data.resample('1D').dropna().apply(ohlc,axis=1).tail(2)

Timestamp
2016-12-29   164.45
2016-12-30   164.26
dtype: float64

```

Does not give me the same results and Im not sure where to insert the `apply`.


Here is a slice of the [data](https://www.dropbox.com/s/obiwqesby5dzt22/stacktest?dl=0) to test this with if required:


thanks


","`.resample()` works like `groupby` so you can pass that dictionary to `resample().agg()`:



```
df.resample('1D').agg(ohlc_dict).tail().dropna()
Out: 
              Volume    Last    High    Open     Low
Timestamp                                           
2016-12-27  144793.0  164.11  164.18  163.55  163.55
2016-12-28  215288.0  164.22  164.33  164.18  163.89
2016-12-29  245538.0  164.49  164.65  164.44  164.27
2016-12-30  286847.0  164.18  164.56  164.55  164.09

```

"
"Upgrade Apache Solr from 4.10.4 to 6.6.0 on Cent os 7
How to upgrade Apache solr from existing version 4.10.4 to 6.6.0 (A latest one)
On linux centos 7 operating system.


","**stop existing running solr**



```
sudo service solr stop

```

**confirm its stopped or not**



```
sudo service solr status

```


> 
> You can take backup of existing solr data from
>  /var/solr/data/< core\_name >, But reindexing solr data after upgrading
>  solr version will give best performance without any unknown issues.
> 
> 
> 


**now remove existing solr from centos**



```
sudo rm -r /var/solr

```

**remove solr from following location if its there**



```
sudo rm -r /opt/solr
sudo rm /etc/init.d/solr

```

**In my case following path was not exist to delete, Delete if exists in your case**



```
sudo rm -r /opt/solr/solr-4.10.4

```

**remove existing solr user with home directory**



```
sudo userdel -r solr

```

**for ubuntu**



```
sudo deluser --remove-home solr

```

**remove existing group**



```
sudo groupdel solr

```

**for ubuntu**



```
sudo deluser --group solr

```

**restart your system**



```
sudo reboot

```

# Check supporting application and download solr 6.6


**Check which version Java version is installed on your machine**



```
java -version

```

**Check available java versions, and update to latest version**



```
yum list available java*

```

**Install latest java version**



```
sudo yum install <jdk version>

```

**download and extract installation folder**



```
mkdir /home/user/temp
cd temp
wget http://apache.org/dist/lucene/solr/6.6.0/solr-6.6.0.tgz
tar -zxvf solr-6.6.0.tgz

```

**install solr**



```
cd solr-6.6.0/
sudo bin/install_solr_service.sh /home/user/temp/solr-6.6.0.tgz

```

# after successful installation you can see following output



> 
> id: solr: no such user Creating new user: solr
> 
> 
> Extracting /home/user/temp/solr-6.6.0.tgz to /opt
> 
> 
> Installing symlink /opt/solr -> /opt/solr-6.6.0 ...
> 
> 
> Installing /etc/init.d/solr script ...
> 
> 
> Installing /etc/default/solr.in.sh ...
> 
> 
> Service solr installed. Customize Solr startup configuration in
>  /etc/default/solr.in.sh Waiting up to 180 seconds to see Solr running
>  on port 8983 [] Started Solr server on port 8983 (pid=4802). Happy
>  searching!
> 
> 
> 
> ```
>  Found 1 Solr nodes: 
> 
> ```
> 
> Solr process 4802 running on port 8983 {  
> 
>  ""solr\_home"":""/var/solr/data"", ""version"":""6.6.0
>  5c7a7b65d2aa7ce5ec96458315c661a18b320241 - ishan - 2017-05-30
>  07:32:53"", ""startTime"":""2017-06-26T07:02:51.523Z"", ""uptime"":""0
>  days, 0 hours, 0 minutes, 13 seconds"", ""memory"":""34.4 MB (%7) of
>  490.7 MB""}
> 
> 
> 


**Now check if you can access solr admin panel**



```
http://localhost:8983/solr/

```

**Check if solr is running, start solr if not running and then try**



```
sudo service solr status

```

**If you are still not able to access solr admin panel, then may you OS firewall is blocking to port 8983**


use following commands to allow port 8983 in firewall



```
sudo firewall-cmd --zone = public --add-port = 8983 / tcp --permanent
sudo firewall-cmd --reload

```


> 
> # For further data migration please see [Migration of solr cores after solr version upgradation from 4.10.4 to 6.6.0 on Cent os 7](https://stackoverflow.com/questions/44760114/migration-of-solr-cores-after-solr-version-upgradation-from-4-10-4-to-6-6-0-on-c/44760115#44760115)
> 
> 
> 


"
"E-mail crawler for yellowpages
I've written a script in python to scrape e-mail addresses from different pizza shops located in los-angeles available in yellowpage traversing multiple pages. It is able to go one-layer deep and dig out email addresses. I believe, this crawler has got the ability to parse all the emails from any link no matter how many pages it has spread across. Just needed to adjust the last page number in the crawler.


Here is what I've written:



```
import requests
from lxml import html

link = ""https://www.yellowpages.com""
for page_num in range(1,10):
  for item in html.fromstring(requests.get(""https://www.yellowpages.com/search?search_terms=pizza&geo_location_terms=Los%20Angeles%2C%20CA&page={0}"".format(page_num)).text).cssselect('div.info h2.n a:not([itemprop=""name""]).business-name'):
    for data in html.fromstring(requests.get(link + item.attrib['href']).text).cssselect('a.email-business'):
      print(data.attrib['href'].replace(""mailto:"",""""))

```

","Knowing that you had a ton of scraper questions, this doesn't look nice at all. It looks like you were in a hurry and didn't much care about how the code looks.




---


### Styling


- Use 4 spaces per indentation level
- Constants should be UPPERCASED
- You should have a space after `,`
- Your lines are way too long. Stick to 72 characters or 120 (at most)




---


### Code


First, you don't have to hardcode your search items. Instead, you might as well define a function which returns the data you need (the entire URL):



```
def build_url(item, location, page):
    params = urllib.parse.urlencode({
        'search_terms': item,
        'geo_location_terms': location,
        'page': page
    })

    return '{}/search?{}'.format(URL, params)

```

With that in mind, your code might be restructured nicer, like this:



```
import requests
import urllib

from lxml import html

URL = ""https://www.yellowpages.com""
LOWER_LIMIT = 1
UPPER_LIMIT = 10


def build_url(item, location, page):
    params = urllib.parse.urlencode({
        'search_terms': item,
        'geo_location_terms': location,
        'page': page
    })

    return '{}/search?{}'.format(URL, params)


def scrape():
    for page in range(LOWER_LIMIT, UPPER_LIMIT):
        url = build_url('pizza', 'Los Angeles CA', page)
        html_ = requests.get(url).text
        for item in html.fromstring(html_).cssselect('div.info h2.n a:not([itemprop=""name""]).business-name'):
            for data in html.fromstring(requests.get(URL + item.attrib['href']).text).cssselect('a.email-business'):
                print(data.attrib['href'].replace(""mailto:"", """"))


if __name__ == '__main__':
    scrape()

```

Other changes that I did above:


- Move the logic into separate functions
- Make the range arguments constants so that they can be easily modified
- Add the guard check [`if __name__ == '__main__'`](https://stackoverflow.com/questions/419163/what-does-if-name-main-do)




---


As is, your scraper is pretty slow because it has to search in each page the entire html for that specific `a` info. You might consider using [`Scrapy`](https://scrapy.org/).


"
"How to modify a PKGBUILD which uses git sources to pull only a shallow clone?
The other day I tried installing [`opencv-git`](https://aur.archlinux.org/packages/opencv-git/) from the *AUR* with `makepkg` on Arch Linux. Of course it pulls from the *git* repository as the name indicates. This pulls 1Gb. I am reading about making a [*shallow clone*](https://git.wiki.kernel.org/index.php/GitFaq#How_do_I_do_a_quick_clone_without_history_revisions.3F) with [`git`](http://git-scm.com/book). When I look at the `PKGBUILD` file, using `grep git PKGBUILD`, I see:



```
pkgname=""opencv-git""
makedepends=('git' 'cmake' 'python2-numpy' 'mesa' 'eigen2')
provides=(""${pkgname%-git}"")
conflicts=(""${pkgname%-git}"")
source=(""${pkgname%-git}::git+http://github.com/Itseez/opencv.git""
    cd ""${srcdir}/${pkgname%-git}""
    git describe --long | sed -r 's/([^-]*-g)/r\1/;s/-/./g'
    cd ""${srcdir}/${pkgname%-git}""
    cd ""${srcdir}/${pkgname%-git}""
    cd ""${srcdir}/${pkgname%-git}""
    install -Dm644 ""LICENSE"" ""${pkgdir}/usr/share/licenses/${pkgname%-git}/LICENSE""

```

Is there a way to modify the recipe or the `makepkg` command to pull only a shallow clone (the latest version of the source is what I want) and not the full repository to save space and bandwidth? Reading [`man 5 PKGBUILD`](https://www.archlinux.org/pacman/PKGBUILD.5.html) doesn't provide the insight I'm looking for. Also looked quickly through the `makepkg` and `pacman` [manpages](https://www.archlinux.org/pacman/) - can't seem to find how to do that.


","This can be done by using a custom [dlagent](https://www.archlinux.org/pacman/makepkg.conf.5.html). I do not really understand Arch packaging or how the dlagents work, so I only have a hack answer, but it gets the job done.


The idea is to modify the PKGBUILD to use a custom download agent. I modified the source



```
""${pkgname%-git}::git+http://github.com/Itseez/opencv.git""

```

into



```
""${pkgname%-git}::mygit://opencv.git""

```

and then defined a new dlagent called `mygit` which does a shallow clone by



```
makepkg DLAGENTS='mygit::/usr/bin/git clone --depth 1 http://github.com/Itseez/opencv.git'

```

Also notice that the repository that is being cloned is hard-coded into the command. Again, this can probably be avoided. Finally, the download location is not what the PKGBUILD expects. To work around this, I simply move the repository after downloading it. I do this by adding



```
mv ""${srcdir}/../mygit:/opencv.git"" ""${srcdir}/../${pkgname%-git}""

```

at the beginning of the `pkgver` function.


I think the cleaner solution would be to figure out what the `git+http` dlagent is doing and redfine that temporarily. This should avoid all the hack aspects of the solution.


"
"Erlang Recursive end loop
I just started learning Erlang and since I found out there is no for loop I tried recreating one with recursion:



```
display(Rooms, In) ->
    Room = array:get(In, Rooms)
    io:format(""~w"", [Room]),
    if
        In < 59 -> display(Rooms, In + 1);
        true -> true
    end.

```

With this code i need to display the content (false or true) of each array in Rooms till the number 59 is reached. However this creates a weird code which displays all of Rooms contents about 60 times (?). When I drop the if statement and only put in the recursive code it is working except for a exception error: Bad Argument.


So basically my question is how do I put a proper end to my ""for loop"".


Thanks in advance!


","This is a general misunderstanding of recursive loop definitions. What you are trying to check for is called the ""base condition"" or ""base case"". This is easiest to deal with by matching:



```
display(0, _) ->
    ok;
display(In, Rooms) ->
    Room = array:get(In, Rooms)
    io:format(""~w~n"", [Room]),
    display(In - 1, Rooms).

```

This is, however, rather unidiomatic. Instead of using a hand-made recursive function, something like a fold or map is more common.


Going a step beyond that, though, most folks would probably have chosen to represent the rooms as a set or list, and iterated over it using list operations. When hand-written the ""base case"" would be an empty list instead of a 0:



```
display([]) ->
    ok;
display([Room | Rooms]) ->
    io:format(""~w~n"", [Room]),
    display(Rooms).

```

Which would have been avoided in favor, once again, of a list operation like [foreach](http://erlang.org/doc/man/lists.html#foreach-2):



```
display(Rooms) ->
    lists:foreach(fun(Room) -> io:format(""~w~n"", [Room]) end, Rooms).

```

Some folks *really* dislike reading lambdas in-line this way. (In this case I find it readable, [but the larger they get the more likely the are to become genuinely distracting](http://zxq9.com/archives/1337).) An alternative representation of the exact same function:



```
display(Rooms) ->
    Display = fun(Room) -> io:format(""~w~n"", [Room]) end,
    lists:foreach(Display, Rooms).

```

Which might itself be passed up in favor of using a list comprehension as a shorthand for iteration:



```
_ = [io:format(""~w~n"", [Room]) | Room <- Rooms].

```

When *only* trying to get a side effect, though, I really think that `lists:foreach/2` is the best choice for semantic reasons.


I think part of the difficulty you are experiencing is that you have chosen to use a rather unusual structure as your base data for your first Erlang program that does anything (arrays are not used very often, and are not very idiomatic in functional languages). Try working with lists a bit first -- its not scary -- and some of the idioms and other code examples and general discussions about list processing and functional programming will make more sense.


**Wait! There's more...**


I didn't deal with the case where you have an irregular room layout. The assumption was always that everything was laid out in a nice even grid -- which is never the case when you get into the really interesting stuff (either because the map is irregular or because the topology is interesting).


The main difference here is that instead of simply carrying a list of `[Room]` where each `Room` value is a single value representing the Room's state, you would wrap the state value of the room in a tuple which also contained some extra data about that state such as its location or coordinates, name, etc. (You know, ""metadata"" -- which is such an overloaded, buzz-laden term today that I hate saying it.)


Let's say we need to maintain coordinates in a three-dimensional space in which the rooms reside, and that each room has a list of occupants. In the case of the array we would have divided the array by the dimensions of the layout. A 10\*10\*10 space would have an array index from 0 to 999, and each location would be found by an operation similar to



```
locate({X, Y, Z}) -> (1 * X) + (10 * Y) + (100 * Z).

```

and the value of each `Room` would be `[Occupant1, occupant2, ...]`.


It would be a real annoyance to define such an array and then mark arbitrarily large regions of it as ""unusable"" to give the impression of irregular layout, and then work around that trying to simulate a 3D universe.


Instead we could use a list (or something like a list) to represent the set of rooms, but the `Room` value would now be a tuple: `Room = {{X, Y, Z}, [Occupants]}`. You may have an additional element (or ten!), like the ""name"" of the room or some other status information or whatever, but the coordinates are the most certain real identity you're likely to get. To get the room status you would do the same as before, but mark what element you are looking at:



```
display(Rooms) ->
    Display =
        fun({ID, Occupants}) ->
            io:format(""ID ~p: Occupants ~p~n"", [ID, Occupants])
        end,
    lists:foreach(Display, Rooms).

```

To do anything more interesting than printing sequentially, you could replace the internals of `Display` with a function that uses the coordinates to plot the room on a chart, check for empty or full lists of `Occupants` (use pattern matching, don't do it procedurally!), or whatever else you might dream up.


"
"R dplyr::mutate - add all elements in a returned list
I have a function returning a list. I'm using `mutate` to add columns in a data frame that correspond to the output. The calculation is rather involved so I would prefer to only call the function once. I'm rather new to R and dplry and cannot figure out a more efficient way of doing this.


Here is a very simple example of what I am doing now.



```
library(dplyr)

testFun <- function(x,z)
{
  list(x2=x*x + z, x3=x*x*x + z)
}

have <- data.frame(x=seq(1:10),y=1,z=0)

want <- have %>%
        dplyr::mutate(x2=testFun(x,z)$x2,
                      x3=testFun(x,z)$x3)

```

How can I do this more efficiently?


","We can use `pmap`



```
library(purrr)
library(dplyr)
pmap_dfr(have %>% 
              select(x, z), testFun) %>%
   bind_cols(have, .)
#    x y z  x2   x3
#1   1 1 0   1    1
#2   2 1 0   4    8
#3   3 1 0   9   27
#4   4 1 0  16   64
#5   5 1 0  25  125
#6   6 1 0  36  216
#7   7 1 0  49  343
#8   8 1 0  64  512
#9   9 1 0  81  729
#10 10 1 0 100 1000

```



---


Or if we can change the function by quoting (`quote` or `quo`) it, this becomes more easier



```
testFun <- function(x,z){
  list(x2= quo(x*x + z), x3= quo(x*x*x + z))
 }

have %>% 
   mutate(!!! testFun(x, z))
#    x y z  x2   x3
#1   1 1 0   1    1
#2   2 1 0   4    8
#3   3 1 0   9   27
#4   4 1 0  16   64
#5   5 1 0  25  125
#6   6 1 0  36  216
#7   7 1 0  49  343
#8   8 1 0  64  512
#9   9 1 0  81  729
#10 10 1 0 100 1000

```

"
"Mount can't find device in /etc/fstab
It is my second day on Linux. I am having difficulties mounting an USB drive. It shows this error:



```
mount: can't find /dev/sdb1/mnt in /etc/fstab or /etc/mtab

```

How can I get rid of this error?


","## Why this error?


**You probably forgot to tell `mount` where to mount your drive.**


Linux uses device files (`/dev/sda`, `/dev/sdb1`, etc.). And unlike Windows drives (`C:`, `D:`, etc.), you cannot access them directly (`cd /dev/sdb1` will inevitably fail, telling you that it is not a directory but a file). If you want to open a drive with `mount`, you need to provide a **mountpoint**. A mountpoint is a directory wherein your USB drive will be opened and where you will be able to access your files.




---


## Solution


1. Create a directory that you will use as the mountpoint for your drive:



```
mkdir /mnt/mydrive

```
2. Mount your drive with this command:



```
mount /dev/sdb1 /mnt/mydrive

```

***Note:*** If you don't know your drive's device file, you can run `sudo fdisk -l` or `lsblk` to identify the partition you're looking for.
3. Now if you run `ls /mnt/mydrive`, it should list your drive's files.
4. When you're done, don't forget to **unmount** your USB drive before removing it from the computer:



```
umount /dev/sdb1

```




---


## More information about this error


`/etc/fstab` is a file in which you can associate a partition with a mountpoint, allowing you to run `mount <device>` instead of `mount <device> <mountpoint>`. **This is why you get this confusing error.**


*fstab* has many more uses like mounting a partition at boot time, etc. [More information about *fstab* on the Arch Linux wiki](https://wiki.archlinux.org/index.php/fstab)


"
"Getting task by name from taskqueue
How can I get a task by name?



```
from google.appengine.api import taskqueue
taskqueue.add(name='foobar', url='/some-handler', params={'foo': 'bar'}

task_queue = taskqueue.Queue('default')
task_queue.delete_tasks_by_name('foobar') # would work

# looking for a method like this:
foobar_task = task_queue.get_task_by_name('foobar')

```

It should be possible with the REST API (<https://developers.google.com/appengine/docs/python/taskqueue/rest/tasks/get>). But I would prefer something like `task_queue.get_task_by_name('foobar')`. Any ideas? Did I miss something?


","There is no guarantee that the task with this name exists - it may have been already executed. And even if you manage to `get` a task, it may be executed while you are trying to do something with it. So when you try to put it back, you have no idea if you are adding it for the first time or for the second time.


Because of this uncertainty, I can't see any use case where getting a task by name may be useful.


EDIT:


You can give a name to your task in order to ensure that a particular task only executes once. When you add a task with a name to a queue, App Engine will check if the task with such name already exists. If it does, the subsequent attempt will fail.


For example, you can have many instances running, and each instance may need to insert an entity in the Datastore. Your first option is to check if an entity already exists in a datastore. This is a relatively slow operation, and by the time you received your response (entity does not exist) and decide to insert it, another instance could have already inserted it. So you end up with two entities instead of one.


Your second option is to use tasks. Instead of inserting a new entity directly into a datastore, an instance creates a task to insert it, and it gives this task a name. If another instance tries to add a task with the same name, it will simply override the existing task. As a result, you are guaranteed that an entity will be inserted only once.


"
"How to parse HTML table with Powershell Core 7?
I have the following code:



```
    $html = New-Object -ComObject ""HTMLFile""
    $source = Get-Content -Path $FilePath -Raw
    try
    {
        $html.IHTMLDocument2_write($source) 2> $null
    }
    catch
    {
        $encoded = [Text.Encoding]::Unicode.GetBytes($source)
        $html.write($encoded)
    }
    $t = $html.getElementsByTagName(""table"") | Where-Object {
        $cells = $_.tBodies[0].rows[0].cells
        $cells[0].innerText -eq ""Name"" -and
        $cells[1].innerText -eq ""Description"" -and
        $cells[2].innerText -eq ""Default Value"" -and
        $cells[3].innerText -eq ""Release""
    }

```

The code works fine on Windows Powershell 5.1, but on Powershell Core 7 `$_.tBodies[0].rows` returns null. 


So, how does one access the rows of an HTML table in PS 7?


","
**PowerShell (Core)**, as of 7.3.4, **does *not* come with a built-in HTML parser** - and this **may never change**.


You must rely on a **third-party solution**, such as the **[`PowerHTML` module](https://www.powershellgallery.com/packages/PowerHTML)** that wraps the **[HTML Agility Pack](https://www.nuget.org/packages/HtmlAgilityPack/)**.


The **object model** works differently than the Internet Explorer-based one available in Windows PowerShell; it is *similar* to the XML DOM provided by the standard [`System.Xml.XmlDocument`](https://learn.microsoft.com/en-US/dotnet/api/System.Xml.XmlDocument) type (`[xml]`)[1]; see [the documentation](https://html-agility-pack.net/manipulation) and the sample code below.



```
# Install the module on demand
If (-not (Get-Module -ErrorAction Ignore -ListAvailable PowerHTML)) {
  Write-Verbose ""Installing PowerHTML module for the current user...""
  Install-Module PowerHTML -ErrorAction Stop
}
Import-Module -ErrorAction Stop PowerHTML

# Create a sample HTML file with a table with 2 columns.
Get-Item $HOME | Select-Object Name, Mode | ConvertTo-Html > sample.html

# Parse the HTML file into an HTML DOM.
$htmlDom = ConvertFrom-Html -Path sample.html

# Find a specific table by its column names, using an XPath
# query to iterate over all tables.
$table = $htmlDom.SelectNodes('//table') | Where-Object {
  $headerRow = $_.Element('tr') # or $tbl.Elements('tr')[0]
  # Filter by column names
  $headerRow.ChildNodes[0].InnerText -eq 'Name' -and 
    $headerRow.ChildNodes[1].InnerText -eq 'Mode'
}

# Print the table's HTML text.
$table.InnerHtml

# Extract the first data row's first column value.
# Note: @(...) is required around .Elements() for indexing to work.
@($table.Elements('tr'))[1].ChildNodes[0].InnerText

```



---


A **Windows-only alternative** is to use the **`HTMLFile` COM object**, as shown in [this answer](https://stackoverflow.com/a/71855426/45375), and as used in your own attempt - I'm unclear on why it didn't work in your specific case.




---


[1] Notably with respect to supporting XPath queries via the `.SelectSingleNode()` and `.SelectNodes()` methods, exposing child nodes via a `.ChildNodes` collection, and providing `.InnerHtml` / `.OuterHtml` / `.InnerText` properties. Instead of an *indexer* that supports child element names, methods `.Element(<name>)` and `.Elements(<name>)` are provided.


"
"How to edit files which are made by Laradock workspace
I've newly started to use **Laradock** to build my Laravel projects but I have a problem in editing the files such as Controllers, Models, etc which are made by the **php artisan** command in the Laradock workspace. The reason is the user in the workspace is a root and on the other side, I'm trying to edit the file in my editor by a common user. So every time I have to run the command **chmod -R 777 /newCreatedFile.php** to change the permission. So is there any solution to handle this problem?


By the way my OS is ubuntu 18.04


","In the [Laradock Getting Started guide](https://laradock.io/getting-started/), it explains how to get Laradock running as a specified user:



> 
> Note: You can add --user=laradock to have files created as your host’s user. Example:
> 
> 
> 
> ```
> docker-compose exec --user=laradock workspace bash
> 
> ```
> 
> 


I believe this should solve your issue, as you will no longer have the Docker user running these commands. Try it out!


Note: The core issue may just be that whatever user Laradock is running as is not creating files with group permissions that allows the host machine's user write capabilities, hence why the `--user` flag can be used. It may not actually be running as the `root` user itself.


"
"An analogue to dotted-pairs for pattern matching in Clojure
Scheme (and CL) has dotted-pairs where both elements of a `cons` cell are specified explicitly (e.g. `(1 . 2)`) rather than implicitly (e.g. `(1 2)` which is read as `(1 . (2 . nil))`).


I came across this [puzzle](http://funcall.blogspot.com/2012/01/small-puzzle.html) where dotted-pairs are used in pattern matching to capture the tail of a list in the object being matched, e.g.:



```
(pmatch '(foo . (? pvar)) '(foo bar baz))
;;      => ((pvar bar baz))

```

Here `'(foo . (? pvar))` is a pattern and `'(foo bar baz)` is the object matched against the pattern. `foo` in the pattern is a literal, whereas `(? pvar)` is a pattern variable which matches `(bar baz)` and binds the symbol `pvar` to that match. The `pmatch` function returns an association list of pattern variables and bound matches.


Had the pattern been `'(foo (? pvar))`, the match would fail, because `baz` would not match anything in the pattern.


I've implemented the puzzle in Clojure, and I pass all of JRM's test cases apart from the dotted-pair one. I'm trying to figure out how to possibly support the dotted-pair pattern as well.


Here's my current solution:



```
(defn pattern-variable? [pv]
  (when (seq? pv)
    (let [[qmark var] pv]
     (and (= (count pv) 2)
          (= qmark '?)
          (or (symbol? var)
              (keyword? var)))))

(defn pattern-variable [pv]
  (second pv))

(defn pmatch
  ([pat obj] (pmatch pat obj {}))
  ([pat obj binds]
     (cond (not (coll? pat))
           (when (= pat obj) binds)
           (pattern-variable? pat)
           (assoc binds (pattern-variable pat) obj)
           (seq? pat) (let [[pat-f & pat-r] pat]
                      (when (seq? obj)
                        (when-let [binds (pmatch pat-f (first obj) binds)]
                          (pmatch pat-r (next obj) binds))))
           :else nil)))

```

So how can I support patterns that match the rest of the object in Clojure without dotted-pairs?


","*(Edit: Added a slightly longer, but significantly clearer matcher impl + a demo. The original remains below the horizontal rule.)*


One solution would be to introduce a different notation to signify a variable to be matched against the tail of a seq, or ""variable after dot"". Another one would be to reserve `&` as a special symbol in patterns with the requirement that it may only be followed by a single pattern variable to be matched against the rest of the expression / object, which must be a seq. I'll explore the first approach below.


Here I took the liberty of changing the notation so that `~foo` is a regular occurrence of the variable `foo` and `~@foo` is a tail occurrence. (One could permit `~@`-matching against subsequences, perhaps matching the minimal initial fragment of a sequence, if any, such that the remainder can be matched against the rest of the pattern; I'll just say this is out of scope for this answer, though. ;-))


Note that these really are different occurrences of the same variable -- i.e. there is still only one variable type -- since no distinction is being made between bindings arising from `~`-occurrences and bindings arising from `~@`-occurrences.


Also note that the examples in the post you linked to do not test for attempts to rebind a previously bound variable (e.g. try `(pmatch '(~x ~x) '(foo bar))`, `(pmatch '((? x) (? x)) '(foo bar))` in the original syntax). The code below returns `nil` in such cases, as it does when the match fails for other reasons.


First, a demo:



```
user> (pmatch '(foo ~pvar1 ~pvar2 bar) '(foo 33 (xyzzy false) bar))
{pvar2 (xyzzy false), pvar1 33}
user> (pmatch '(~av ~@sv) '(foo bar baz))
{sv (bar baz), av foo}
user> (pmatch '(foo ~pvar1 ~pvar2 bar) '(foo 33 false bar))
{pvar2 false, pvar1 33}
user> (pmatch '(foo ~pvar bar) '(quux 33 bar))
nil
user> (pmatch '(a ~var1 (nested (c ~var2))) '(a b (nested (c d))))
{var2 d, var1 b}
user> (pmatch '(a b c) '(a b c))
{}
user> (pmatch '(foo ~pvar1 ~pvar2 bar) '(foo 33 (xyzzy false) bar))
{pvar2 (xyzzy false), pvar1 33}
user> (pmatch '(foo ~@pvar) '(foo bar baz))
{pvar (bar baz)}
user> (pmatch '(~? quux) '(foo quux))
{? foo}
user> (pmatch '~? '(foo quux))
{? (foo quux)}
user> (pmatch '(? ? ?) '(foo quux))
nil

```

Here's the matcher:



```
(defn var-type [pat]
  (when (seq? pat)
    (condp = (first pat)
      'clojure.core/unquote :atomic
      'clojure.core/unquote-splicing :sequential
      nil)))

(defn var-name [v]
  (when (var-type v)
    (second v)))

(defmulti pmatch*
  (fn [pat expr bs]
    (cond
      (= :atomic (var-type pat))        :atom
      (= :sequential (var-type pat))    nil
      (and (seq? pat) (seq? expr))      :walk
      (not (or (seq? pat) (seq? expr))) :exact
      :else                             nil)))

(defmethod pmatch* :exact [pat expr bs]
  (when (= pat expr) bs))

(defmethod pmatch* :atom [v expr bs]
  (if-let [[_ x] (find bs (var-name v))]
    (when (= x expr) bs)
    (assoc bs (var-name v) expr)))

(defmethod pmatch* :walk [pat expr bs]
  (if-let [[p] pat]
    (if (= :sequential (var-type p))
      (when (and (seq? expr) (not (next pat)))
        (if-let [[_ xs] (find bs (var-name p))]
          (when (= xs expr) bs)
          (assoc bs (var-name p) expr)))
      (when-let [[x] expr]
        (when-let [m (pmatch* p x bs)]
          (pmatch* (next pat) (next expr) m))))))

(defmethod pmatch* nil [& _] nil)

(defn pmatch
  ([pat expr] (pmatch pat expr {}))
  ([pat expr bs] (pmatch* pat expr bs)))

```



---


And here's the original monolithic version:



```
(defn pmatch
  ([pat expr] (pmatch pat expr {}))
  ([pat expr bs]
     (letfn [(atom-var? [pat]
               (and (seq? pat) (= 'clojure.core/unquote (first pat))))
             (seq-var? [pat]
               (and (seq? pat) (= 'clojure.core/unquote-splicing
                                  (first pat))))
             (v [var] (second var))
             (matcha [a e bs]
               (if-let [[_ x] (find bs (v a))]
                 (and (or (= x e) nil) bs)
                 (assoc bs (v a) e)))
             (matchs [s e bs]
               (when (seq? e)
                 (if-let [[_ xs] (find bs (v s))]
                   (or (= xs e) nil)
                   (assoc bs (v s) e))))]
       (when bs
         (cond
           (atom-var? pat)
           (matcha pat expr bs)

           (seq-var? pat)
           (matchs pat expr bs)

           (and (seq? pat) (seq? expr))
           (if-let [[p] pat]
             (if (seq-var? p)
               (matchs p expr bs)
               (when-let [[x] expr]
                 (when-let [m (pmatch p x bs)]
                   (recur (next pat) (next expr) m))))
             (when-not (first expr)
               bs))

           (not (or (seq? pat) (seq? expr)))
           (when (= pat expr)
             bs)

           :else nil)))))

```

"
"Can't extract data from Json
I can't extract data from json that I got from an api.
I tried for hours, tried all kinds of formats. Read Stackoverflow threads like [How do I extract data from JSON with PHP?](https://stackoverflow.com/questions/29308898/how-do-i-extract-data-from-json-with-php), but I can't see what I am doing wrong.


This is the code so far:



```
$api_results = '{""status"":""0000"",""data"":{""opening_price"":""6998000"",""closing_price"":""7270000"",""min_price"":""6750000"",""max_price"":""7997000"",""average_price"":""7188302.5804"",""units_traded"":""78484.9241002"",""volume_1day"":""78484.9241002"",""volume_7day"":""335611.84181738"",""buy_price"":""7268000"",""sell_price"":""7274000"",""date"":""1510563513031""}}';

$results = json_decode($api_results, true);

// Some variations I tried:
var_dump($results->status[1]);
var_dump($results->data[1]->opening_price);
var_dump($results[""data""][1][""opening_price""]);

```

End result: NULL NULL NULL


What am I doing wrong?


**Thanks for the answers! I will upvote the working ones. Seems I got confused in the formating!**


","
```
<?php

$api_results = '{""status"":""0000"",""data"":{""opening_price"":""6998000"",""closing_price"":""7270000"",""min_price"":""6750000"",""max_price"":""7997000"",""average_price"":""7188302.5804"",""units_traded"":""78484.9241002"",""volume_1day"":""78484.9241002"",""volume_7day"":""335611.84181738"",""buy_price"":""7268000"",""sell_price"":""7274000"",""date"":""1510563513031""}}';


$results = json_decode($api_results, true);
print_r($results['status']);
echo ""</br>"";
print_r($results['data']['opening_price']);

```

Try access your array that way.


The output is : 



```
0000
6998000

```

Keep an eye for the nested arrays. You need to access their parent array first in order to get their values.


"
"Routing doesn't work after Dockerizing Angular app
Am new to Angular and Docker. I have developed an Angular 7 application and trying to dockerize it. I did see many tutorials and was able to build the Docker image.
Below is my Dockerfile



```
FROM nginx:1.13.3-alpine
## Remove default nginx website
RUN rm -rf /usr/share/nginx/html/*
COPY ./ /usr/share/nginx/html
EXPOSE 80
CMD [""nginx"", ""-g"", ""daemon off;""]

```

And i run the docker container using command



```
docker run -d --name containerName -p 8082:80 imagename

```

My container did start and after looking in the browser <http://IP-address:8082/> I could see index html of my app. Apart from that none other url of my application like the login page (<http://IP-address:8082/login>) or dashboard (<http://IP-address:8082/dashboard>) works. I see 404 page not found issue. What else is missing to make sure routing works in my dockerized container?


Below is my default.conf file



```
server {
    listen       80;
    server_name  localhost;

    #charset koi8-r;
    #access_log  /var/log/nginx/host.access.log  main;

    location / {
        root   /usr/share/nginx/html;
        index  index.html index.htm;
    }

    #error_page  404              /404.html;

    # redirect server error pages to the static page /50x.html
    #
    error_page   500 502 503 504  /50x.html;
    location = /50x.html {
        root   /usr/share/nginx/html;
    }

    # proxy the PHP scripts to Apache listening on 127.0.0.1:80
    #
    #location ~ \.php$ {
    #    proxy_pass   http://127.0.0.1;
    #}

    # pass the PHP scripts to FastCGI server listening on 127.0.0.1:9000
    #
    #location ~ \.php$ {
    #    root           html;
    #    fastcgi_pass   127.0.0.1:9000;
    #    fastcgi_index  index.php;
    #    fastcgi_param  SCRIPT_FILENAME  /scripts$fastcgi_script_name;
    #    include        fastcgi_params;
    #}

    # deny access to .htaccess files, if Apache's document root
    # concurs with nginx's one
    #
    #location ~ /\.ht {
    #    deny  all;
    #}
}

```

And my docker version is Docker version 17.03.2-ce. 
Any help appreciated


","This is happening because the base `nginx` image will try to serve the requests from the `docroot`, so it will try to find a `login.html` when you send a request to `/login`.


To avoid this, you need nginx to serve the `index.html` **no matter the request**, and thus letting `angular` take care of the routes. 


To do so, you will need to change your `nginx.conf`, which is currently in the image, to include the following :


`try_files $uri $uri/ /index.html;`


Instead of the default being: 


`try_files $uri $uri/ =404;`


You can do so in many ways, but I guess the best approach is to have an extra command in your `Dockerfile`, that copies over the nginx configuration like so : 


`COPY nginx/default.conf /etc/nginx/conf.d/default.conf`


And have that `nginx/default.conf` file in your directory (containing the default nginx configuration) with the command specified above replaced.


**EDIT**


There is already images that does exactly that, so you can just use an image other than the official one, that is made specifically for that, and it should work fine: [example](https://hub.docker.com/r/trion/nginx-angular/)


"
"What is the difference between dynamic dispatch and late binding in C++?
I've recently read about the Dynamic Dispatch on [Wikipedia](http://en.wikipedia.org/wiki/Dynamic_dispatch) and couldn't understand the difference between dynamic dispatch and late binding in C++. 


When each one of the mechanisms is used?


The exact quote from Wikipedia:



> 
> Dynamic dispatch is different from late binding (also known as dynamic binding). In the context of selecting an operation, binding refers to the process of associating a name with an operation. Dispatching refers to choosing an implementation for the operation after you have decided which operation a name refers to. With dynamic dispatch, the name may be bound to a polymorphic operation at compile time, but the implementation not be chosen until runtime (this is how dynamic dispatch works in C++). However, late binding does imply dynamic dispatching since you cannot choose which implementation of a polymorphic operation to select until you have selected the operation that the name refers to.
> 
> 
> 


","A fairly decent answer to this is actually incorporated into a question on late vs. early binding on [programmers.stackexchange.com](https://softwareengineering.stackexchange.com/a/200123). 


In short, late binding refers to the *object*-side of an eval, dynamic dispatch refers to the functional-side. In late binding the *type* of a variable is the variant at runtime. In dynamic-dispatch, the function or subroutine being executed is the variant. 


In C++, we don't really have late binding because the type *is known* (not necessarily the end of the inheritance hierarchy, but at least a formal base class or interface). But we *do* have dynamic dispatch via virtual methods and polymorphism.


The best example I can offer for late-binding is the untyped ""object"" in Visual Basic. The runtime environment does all the late-binding heavy lifting for you.



```
Dim obj

- initialize object then..
obj.DoSomething()

```

The compiler will actually code the appropriate execution context for the runtime-engine to perform a named lookup of the method called `DoSomething`, and if discovered with the properly matching parameters, actually execute the underlying call. In reality, *something* about the type of the object is known (it inherits from `IDispatch` and supports `GetIDsOfNames()`, etc). but as far as the *language* is concerned the *type* of the variable is utterly unknown at compile time, and it has no idea if `DoSomething` is even a method for whatever `obj` actually *is* until runtime reaches the point of execution.


I won't bother dumping a C++ virtual interface et'al, as I'm confident you already know what they look like. I hope it is obvious that the C++ language simply can't do this. It is strongly-typed. It *can* (and does, obviously) do dynamic dispatch via the polymorphic virtual method feature. 


"
"How to get a valid font name on linux system which can be used in .Xresources config?
I'm interested in how can i get a valid font name to use in in .Xresources file. I've installed `ttf-ms-fonts` and want to switch my terminal to courier font. But setting it like this does't work



```
urxvt*font:     cour
urxvt*boldFont: courbd

```

How do i specify a valid name?


","First, you have to decide if you want to use core protocol or Xft for font rendering. As you see in another answer, `xfontsel` is the right tool to get a correct font name for core protocol. But it's unlikely to be what you want for truetype fonts (do you want antialiasing? Then Xft is your choice).


If urxvt is built with Xft support (check `urxvt --help 2>&1 | grep options` to be sure), you might want to give it font names prefixed by `xft:`



```
URxvt.font: xft:Courier New

```

Other options affecting font matching and rendering may be specified in xft font name:



```
URxvt.font: xft:Courier New:pixelsize=18:antialias=false

```

(search `man rxvt` for `xft:` for further details)


All available font names can be queried with `fc-list`. E.g. `fc-list|grep courbd.ttf` shows you the font name is Courier New and style is Bold (append `:style=Bold` to select it).


Use `fc-match ""Courier New""` to check which font is the best match for a given name from Xft's point of view.


Other applications may have their own conventions for X11 and Xft font names. E.g. the same `xft:` prefix is used by emacs; xterm uses `faceName` and `renderFont` resources to determine whether to use Xft and which font to request; `xedit` supports core protocol only. The mere fact that the application is configurable from X resources isn't enough to tell how the font names are interpreted.


"
"Difference between ""def setup"" and ""setup do"" in Minitest?
Are there any differences at all between calling `def setup` and `setup do` in Rails Minitests? I had been using `def setup` this whole time, but I suddenly found that my `setup` for a particular test file was not being called. When I changed it to `setup do`, it suddenly worked again (without changing anything else). But I find this very peculiar, and I'd rather stick to `def setup` for everything if possible, for consistency. Any advice is appreciated.



```
require 'test_helper'
require_relative '../../helpers/user_helper'

class FooTest < ActiveSupport::TestCase
  include UserHelper

  # This method doesn't get called as-is.
  # But it does get called if I change the below to `setup do`.
  def setup
    # create_three_users is a UserHelper method.
    create_three_users
    @test_user = User.first
  end


  test 'should abc' do
    # Trying to call @test_user here returned nil.
  end
end

```

","There was another test file with the class defined as `class FooTest < ActiveSupport::TestCase`. I imagine someone created it by copying the original `FooTest` file, and forgot to change the name.


In short, the other `FooTest`'s setup method had been getting called instead of this one. Coincidentally, the other `FooTest` had also been calling the same `create_three_users` in the setup, which was why I didn't realise this till I tried and failed to assign an instance variable.


I couldn't find much info on the actual difference between `def setup` and `setup do`, but one [blog](http://docs.komagata.org/5203) (you'll have to take my word because it's written in Japanese) writes that `setup do` calls the setup process for not only that class but also the its parent class, which might explain why my tests worked using `setup do` (maybe it called the `setup` for both `FooTest`s).


"
"Find last word of a string that has special characters
I am trying to add a `span` tag to the last word of a string. It works if the string has no special characters. I can't figure out the correct regex for it.



```
$string = ""Onun Mesajı"";
echo preg_replace(""~\W\w+\s*\S?$~"", ' <span>' . '\\0' . '</span>', $string);

```

Here is the Turkish character set : `ÇŞĞÜÖİçşğüöı`


","You need to use `/u` modifier to allow processing Unicode characters in the pattern and input string.



```
preg_replace('~\w+\s*$~u', '<span>$0</span>', $string); 
                       ^

```

[Full PHP demo](http://ideone.com/F0dmM6):



```
$string = ""Onun Mesajı"";
echo preg_replace(""~\w+\s*$~u"", '<span>$0</span>', $string);

```

Also, the regex you need is just `\w+\s*$`:


- `\w+` - 1 or more alphanumerics
- `\s*` - 0 or more whitespace (trailing)
- `$` - end of string


Since I removed the `\W` from the regex, there is no need to ""hardcode"" the leading space in the replacement string (removed, too).


"
"USB slots stop working suddenly from time to time
After re-installing Lubuntu 12.04 system on my laptop an older problem re-emerged after a few days and installation of different programs: without apparent reason the external mouse and sometimes other usb connected devices (including hdd) stop working. The hdd shows it has tension as it has a light there, and the external mouse flashes for a second when plugged.


[I have posted a different version of this problem before](https://askubuntu.com/q/178054/47206). I keep it for now as example of the two answers there. None of them works here.


**Logging out-in does nothing, restart does.**


The event seems entirely random, after reboot it will reappear after many days or weeks or, rarely, after a few hours.


- <http://pastebin.com/0qR8bhhX> in `var/log/syslog` after new occurrence (**with only external wired mouse and keyboard**)


What counts is at the end I guess:



```
Nov 24 14:06:55 cprq-HP-Compaq-nx8220-PY518EA-ABB kernel: [29953.822962] usb 3-1: USB disconnect, device number 3
Nov 24 14:06:57 cprq-HP-Compaq-nx8220-PY518EA-ABB kernel: [29955.069427] uhci_hcd 0000:00:1d.0: host controller process error, something bad happened!
Nov 24 14:06:57 cprq-HP-Compaq-nx8220-PY518EA-ABB kernel: [29955.069439] uhci_hcd 0000:00:1d.0: host controller halted, very bad!
Nov 24 14:06:57 cprq-HP-Compaq-nx8220-PY518EA-ABB kernel: [29955.069461] uhci_hcd 0000:00:1d.0: HC died; cleaning up
Nov 24 14:06:57 cprq-HP-Compaq-nx8220-PY518EA-ABB kernel: [29955.069492] usb 2-2: USB disconnect, device number 2

```

- I have noticed that **on most occasions only the external mouse and keyboard are affected, but not the external HDD. Or if it is, replugging it solves the problem**.
- I have a dual boot with WinXP: **in Windows this never happens, so it is not a hardware issue**
- **I have used Lubuntu Quantal 12.10 and the same problem happened there as well**. Upgrading to that would not be a solution
- On certain occasion **only restarting 2 or even 3 times** solved it.




---


**Using the same PC/hardware with Linux Mint 14 (Quantal) Xfce, the problem almost disappeared**(it happened *once* since then). I am not sure whether this 'solution' comes from using Xfce or Mint (I guess Mint 14 Nadia uses the same kernel as Lubuntu Quantal).


","## kernel vs hardware issues


If you have ruled out possible hardware issues such as voltage/current problems, failing USB ports/hubs then this is probably a kernel problem.


If you feel comfortable with possible non-booting issues/black screen issues on boot, you could try installing the quantal kernel available in the 12.04 repositories.


- [My computer boots to a black screen, what options do I have to fix it?](https://askubuntu.com/questions/162075/my-computer-boots-to-a-black-screen-what-options-do-i-have-to-fix-it)


Remember - most blackscreen issues are due to installation of proprietary graphics drivers. You will have most luck removing these first before upgrading your kernel.


To install the quantal kernel:



```
sudo apt-get install linux-generic-lts-quantal

```

Remember to do a full system backup to allow you to recover if the new kernel breaks more than it fixes.


- [Comparison of backup tools](https://askubuntu.com/questions/2596/comparison-of-backup-tools)


"
"Comparing multiple integers in IF statement, Java
I am brand new to coding. I'm currently using ""Java:How to program"" by Deitel, and one of the early exercises has me stumped.


It is asking me to use simple, chapter 1 if statements to compare 5 integers (input by the user) and display both the largest and the smallest. I am limited to the language used in the first chapter, so I made the assumption that I could list int values in the parenthetical after the if statement. Something like:



```
int firstInt;
int secondInt;
int thirdInt;
int fourthInt;
int fifthInt;

if (first Int> secondIt, thirdInt, fourthInt, fifthInt)

```

etc. but this is a syntax error. Does anyone know how to compare multiple values simultaneously? A google search proves to be no help since all the answers have some convoluted solution that I don't understand. 


If I try something like this:



```
     if (one > two)
     System.out.println(""%d is the largest number!"", one);

     if (one > three)

```

etc. for all 1 - 5, The issue I run into is when I continue with if statements 1>2,3,4,5 2>1,3,4,5 etc. then there will be 1 true for each integer regardless except for the smallest number, and it will display 4 outputs, when all I'm after is the larges and the smallest.


I'm stumped.


","The expression:



```
firstInt > secondInt, thirdInt, fourthInt, fifthInt

```

is not valid because Java is a computer language rather than a natural language (like English).


If you want to do it that way, what you need is the `and` conjunction, with multiple *complete* conditions (using `a,b,c,d,e` to minimise the code):



```
if ((a >= b) && (a >= c) && (a >= d) && (a >= e)) { // a >= b,c,d,e
    System.out.println (""Largest is: "" + a);
} else if ((b >= c) && (b >= d) && (b >= e)) {      // b >= c,d,e
    System.out.println (""Largest is: "" + b);
} else if ((c >= d) && (c >= e)) {                  // c >= d,e
    System.out.println (""Largest is: "" + c);
} else if (d >= e) {                                // d >= e
    System.out.println (""Largest is: "" + d);
} else {                                            // e > d
    System.out.println (""Largest is: "" + e);
}

```

keeping in mind that, once you decide `a` (or any of the others in later steps) *isn't* the largest, you never need to check it again, as one of the others *will* be larger than it.




---


Or you can use the facilities of the Java standard `Math` package, where `Math.max(a,b)` will give you the largest of the values from `a` and `b`:



```
int largest = Math.max(a,Math.max(b,Math.max(c,Math.max(d,e))));

```

Here, you define the largest of all five numbers as the maximum of the pair `{a, maximum-of-bcde}` then use a similar method to work out `maximum-of-bcde` as the maximum of the set `{b, maximum-of-cde}` and so on.


Or, in other words, choose the largest from `{d,e}` and call it `x`. Then choose the largest from `{c,x}` and call it `y`. Then the largest from `{b,y}` and call it `z`. Finally the largest from `{a,z}` and that's your largest value of the five.


Though, to be honest, if you're on chapter one of an introductory text, it probably hasn't covered much in the Java standard libraries yet.




---


But, if you're after a method where it's perhaps *easier* for a beginner to understand the flow, I don't think you can go past this one:



```
int maxOfFive (int a, int b, int c, int d, int e) {
    int largest = a;
    if (b > largest) largest = b;
    if (c > largest) largest = c;
    if (d > largest) largest = d;
    if (e > largest) largest = e;
    return largest;
}

```

This is basic selection of the largest value *as a person would do it,* scanning the list one by one and just remembering which value was the highest, then returning that.


Or, if you haven't learned about functions yet (and reverting to your original variables), just implement the same code *without* a function call:



```
int largest = firstInt;
if (secondInt > largest) largest = secondInt;
if (thirdInt  > largest) largest = thirdInt;
if (fourthInt > largest) largest = fourthInt;
if (fifthInt  > largest) largest = fifthInt;
System.out.println (""Largest is: "" + largest);

```

I believe that's probably only using stuff you've already indicated that you know about (variables, single-condition `if` statement, and output).


"
"How do I dispatch\_sync, dispatch\_async, dispatch\_after, etc in Swift 3, Swift 4, and beyond?
I have lots of code in Swift 2.x (or even 1.x) projects that looks like this:



```
// Move to a background thread to do some long running work
dispatch_async(dispatch_get_global_queue(DISPATCH_QUEUE_PRIORITY_DEFAULT, 0)) {
    let image = self.loadOrGenerateAnImage()
    // Bounce back to the main thread to update the UI
    dispatch_async(dispatch_get_main_queue()) {
        self.imageView.image = image
    }
}

```

Or stuff like this to delay execution:



```
dispatch_after(dispatch_time(DISPATCH_TIME_NOW, Int64(0.5 * Double(NSEC_PER_SEC))), dispatch_get_main_queue()) {
    print(""test"")
}

```

Or any of all kinds of other uses of the Grand Central Dispatch API...


Now that I've opened my project in Xcode 8 (beta) for Swift 3, I get all kinds of errors. Some of them offer to fix my code, but not all of the fixes produce working code. What do I do about this?


","Since the beginning, Swift has provided some facilities for making ObjC and C more Swifty, adding more with each version. Now, in Swift 3, the new [""import as member""](https://github.com/apple/swift-evolution/blob/master/proposals/0044-import-as-member.md) feature lets frameworks with certain styles of C API -- where you have a data type that works sort of like a class, and a bunch of global functions to work with it -- act more like Swift-native APIs. The data types import as Swift classes, their related global functions import as methods and properties on those classes, and some related things like sets of constants can become subtypes where appropriate.


In Xcode 8 / Swift 3 beta, Apple has applied this feature (along with a few others) to make the Dispatch framework much more Swifty. (And [Core Graphics](https://developer.apple.com/reference/coregraphics), too.) If you've been following the Swift open-source efforts, [this isn't news](https://github.com/apple/swift-evolution/blob/master/proposals/0088-libdispatch-for-swift3.md), but now is the first time it's part of Xcode.


**Your first step on moving any project to Swift 3** should be to open it in Xcode 8 and choose *Edit > Convert > To Current Swift Syntax...* in the menu. This will apply (with your review and approval) all of the changes at once needed for all the renamed APIs and other changes. (Often, a line of code is affected by more than one of these changes at once, so responding to error fix-its individually might not handle everything right.)


The result is that the common pattern for bouncing work to the background and back now looks like this:



```
// Move to a background thread to do some long running work
DispatchQueue.global(qos: .userInitiated).async {
    let image = self.loadOrGenerateAnImage()
    // Bounce back to the main thread to update the UI
    DispatchQueue.main.async {
        self.imageView.image = image
    }
}

```

Note we're using `.userInitiated` instead of one of the old `DISPATCH_QUEUE_PRIORITY` constants. Quality of Service (QoS) specifiers were introduced in OS X 10.10 / iOS 8.0, providing a clearer way for the system to prioritize work and deprecating the old priority specifiers. See Apple's [docs on background work and energy efficiency](https://developer.apple.com/library/ios/documentation/Performance/Conceptual/EnergyGuide-iOS/PrioritizeWorkWithQoS.html#//apple_ref/doc/uid/TP40015243-CH39-SW9) for details.


By the way, if you're keeping your own queues to organize work, the way to get one now looks like this (notice that `DispatchQueueAttributes` is an `OptionSet`, so you use collection-style literals to combine options):



```
class Foo { 
    let queue = DispatchQueue(label: ""com.example.my-serial-queue"",
                           attributes: [.serial, .qosUtility])
    func doStuff() {
        queue.async {
            print(""Hello World"")
        }
    }
}

```

Using `dispatch_after` to do work later? That's a method on queues, too, and it takes a `DispatchTime`, which has operators for various numeric types so you can just add whole or fractional seconds:



```
DispatchQueue.main.asyncAfter(deadline: .now() + 0.5) { // in half a second...
    print(""Are we there yet?"")
}

```

You can find your way around the new Dispatch API by opening its interface in Xcode 8 -- use Open Quickly to find the Dispatch module, or put a symbol (like `DispatchQueue`) in your Swift project/playground and command-click it, then brouse around the module from there. (You can find the Swift [Dispatch](https://developer.apple.com/reference/dispatch) API in Apple's spiffy new API Reference website and in-Xcode doc viewer, but it looks like the doc content from the C version hasn't moved into it just yet.)


See the [Migration Guide](https://swift.org/migration-guide/) for more tips.


"
"How to access wmi in python?
So I am trying to access the data from [here](http://openhardwaremonitor.org/wordpress/wp-content/uploads/2011/04/OpenHardwareMonitor-WMI.pdf)


in Python. As you can see, it uses wmi. I have tried to use wmi in python before but I am having trouble interpreting the data they are giving me. Please be patient with me as I am a noob to how wmi works. It says that the wmi data is stored in root/OpenHardwareMontor and that it uses two different wmi classes(Hardware and Sensor). But all this information is going over my head. 


could someone please give me some sample code to read some data from this? 


For example, the code to check cpu core 1 frequency. 


EDIT: i have sort of got it working. i run this code: 



```
for Temperature in c.sensor():
    print Temperature.identifier
    print Temperature.value

```

and i get this:



```
/hdd/0/load/0
37.6608924866
/intelcpu/0/temperature/1
53.0
/intelcpu/0/temperature/0
42.0
/ram/data/1
2.88324356079
/intelcpu/0/load/2
1.53846144676
/hdd/0/temperature/0
43.0
/intelcpu/0/load/0
2.30768918991
/intelcpu/0/clock/1
1463.29663086
/intelcpu/0/clock/0
133.02696228
/intelcpu/0/clock/2
1463.29663086
/ram/load/0
49.224521637
/ram/data/0
2.79517364502
/intelcpu/0/load/1
3.07692289352

```

how can i request **only** the value associated with the identifier /intelcpu/0/temperature/1 ignoring all other values?


","The most simple example to use WMI:



```
c = wmi.WMI()
wql = ""Select * From Win32_SerialPort""
for item in c.query(wql):
    print item

```

Output Example:



```
instance of Win32_SerialPort
{
    Availability = 2;
    Binary = TRUE;
    Caption = ""SpectrumAnalyzer1 (COM15)"";
    ConfigManagerErrorCode = 0;
    ConfigManagerUserConfig = FALSE;
    CreationClassName = ""Win32_SerialPort"";
    Description = ""SpectrumAnalyzer1"";
    DeviceID = ""COM15"";
    MaxBaudRate = 128000;
    MaximumInputBufferSize = 0;
    MaximumOutputBufferSize = 0;
    Name = ""SpectrumAnalyzer1 (COM15)"";
    OSAutoDiscovered = TRUE;
    PNPDeviceID = ""USB\\VID_10C4&PID_ED00\\1269376"";
    PowerManagementCapabilities = {1};
    PowerManagementSupported = FALSE;
    ProviderType = ""RS232 Serial Port"";
    SettableBaudRate = TRUE;
    SettableDataBits = TRUE;
    SettableFlowControl = TRUE;
    SettableParity = TRUE;
    SettableParityCheck = TRUE;
    SettableRLSD = TRUE;
    SettableStopBits = TRUE;
    Status = ""OK"";
    StatusInfo = 3;
    Supports16BitMode = FALSE;
    SupportsDTRDSR = TRUE;
    SupportsElapsedTimeouts = TRUE;
    SupportsIntTimeouts = TRUE;
    SupportsParityCheck = TRUE;
    SupportsRLSD = TRUE;
    SupportsRTSCTS = TRUE;
    SupportsSpecialCharacters = TRUE;
    SupportsXOnXOff = TRUE;
    SupportsXOnXOffSet = TRUE;
    SystemCreationClassName = ""Win32_ComputerSystem"";
    SystemName = ""......."";
};

```

You can access each `item` by:



```
myQuery = c.query(wql)
myQuery.Availability 

```

Output:



```
2

```

For more information, try the WMI [cookbook](http://timgolden.me.uk/python/wmi/cookbook.html).


Edit #1:


Using if statements and `in` you can do what you want.



```
for Temperature in c.sensor():
    if ""/intelcpu/0/temperature/1"" in Temperature.identifier:
        print Temperature.identifier
        print Temperature.value 

```

"
"Get current date time from server and convert it into local time in c#
Help: I have a server which is having time in GMT-07.00 hours. My local time is GMT+05.30 hours. I need to get current date and time from server and convert this date and time into my local time. I have tried many codes, but still have not found a successive way of doing this. Can somebody please help me out.



```
string zoneId = ""India Standard Time"";
TimeZoneInfo tzi = TimeZoneInfo.FindSystemTimeZoneById(zoneId);
DateTime result = TimeZoneInfo.ConvertTimeFromUtc(DateTime.UtcNow, tzi);
DateTime cu = result.ToUniversalTime();
DateTime cur = cu.ToLocalTime();

```

I have tried all the above methods, still I'm not getting the correct answer. Suppose my current local time is 09-Mar-2014 12:51:00 PM, then my server time would be 12.30 hours different from my local time, ie subtract 12 hours and 30 minutes from my local time to get my server time. I need to get my local time from the server time. How can it be acheived?? Please suggest me some solutions.. Thanks in advance for your answers


","no need to know server time zone. if the server time setting is correct you can try this :



```
DateTime serverTime = DateTime.Now; // gives you current Time in server timeZone
DateTime utcTime = serverTime.ToUniversalTime(); // convert it to Utc using timezone setting of server computer

TimeZoneInfo tzi = TimeZoneInfo.FindSystemTimeZoneById(""India Standard Time"");
DateTime localTime = TimeZoneInfo.ConvertTimeFromUtc(utcTime, tzi); // convert from utc to local

```

to check it locally , change your computer timezone to match your server. then run the code. I check and it's working fine.


***update:*** 


the first two lines can be mixed into one line as below . that has a better performance :



```
DateTime utcTime = DateTime.UtcNow;

```

"
"How can I get a list of available methods in a WebAPI web service?
I'm building a small test tool that should provide the user a list of web services (built using WebAPI). The user should be able to choose a service to test.
I'm using 



```
HttpClient client = new HttpClient();
client.BaseAddress = new Uri(""http://'localhost':51062/"");

// Add an Accept header for JSON format.

client.DefaultRequestHeaders.Accept.Add(
                new MediaTypeWithQualityHeaderValue(""application/json""));

```

and am looking for something like 



```
client.GetAllWebServices()

```

which would return a list of methods that the user can see. Meaning, the methods that he developed on the controller and wants to test.


","Michael was correct to mention [`ApiExplorer`](http://msdn.microsoft.com/en-us/library/system.web.http.description.apiexplorer(v=vs.118).aspx). This gives you details of all the WebApi methods for you. You just need to format it how you want the response.


Here is a simple example to get a list of all the methods with their parameters and return types. You can of course make this much more comprehensive - just browse the objects to find what you need:



```
using System.Collections.Generic;
using System.Linq;
using System.Web.Http;
using System.Web.Http.Description;

namespace WebApplication1.Controllers
{
    public class ApiMethodController : ApiController
    {
        public IEnumerable<HelpMethod> GetMethods()
        {
            // get the IApiExplorer registered automatically
            IApiExplorer ex = this.Configuration.Services.GetApiExplorer();

            // loop, convert and return all descriptions 
            return ex.ApiDescriptions
                // ignore self
                .Where(d => d.ActionDescriptor.ControllerDescriptor.ControllerName != ""ApiMethod"")
                .Select(d =>
                {
                    // convert to a serializable structure
                    return new HelpMethod
                    {
                        Parameters = d.ParameterDescriptions.Select(p => new HelpParameter
                        {
                            Name = p.Name,
                            Type = p.ParameterDescriptor.ParameterType.FullName,
                            IsOptional = p.ParameterDescriptor.IsOptional
                        }).ToArray(),
                        Method = d.HttpMethod.ToString(),
                        RelativePath = d.RelativePath,
                        ReturnType = d.ResponseDescription.DeclaredType == null ?
                            null : d.ResponseDescription.DeclaredType.ToString()
                    };
                });
        }
    }

    public class HelpMethod
    {
        public string Method { get; set; }
        public string RelativePath { get; set; }
        public string ReturnType { get; set; }
        public IEnumerable<HelpParameter> Parameters { get; set; }
    }

    public class HelpParameter
    {
        public string Name { get; set; }
        public string Type { get; set; }
        public bool IsOptional { get; set; }
    }
}

```

The nice thing is that it is a WebApi call itself, so you can use the `HttpClient` to call and process it using <http://www.localhost.com/api/ApiMethod/Methods>. Here's a sample JSON response:



```
[
    {
        ""Method"": ""GET"",
        ""RelativePath"": ""api/Account/{id}"",
        ""ReturnType"": ""WebApplication1.Models.Account"",
        ""Parameters"": [
            {
                ""Name"": ""id"",
                ""Type"": ""System.Int32"",
                ""IsOptional"": false
            }
        ]
    },
    {
        ""Method"": ""POST"",
        ""RelativePath"": ""api/Account"",
        ""ReturnType"": null,
        ""Parameters"": [
            {
                ""Name"": ""a"",
                ""Type"": ""WebApplication1.Models.Account"",
                ""IsOptional"": false
            }
        ]
    },
    {
        ""Method"": ""GET"",
        ""RelativePath"": ""api/Maths?i={i}&j={j}"",
        ""ReturnType"": ""System.Int32"",
        ""Parameters"": [
            {
                ""Name"": ""i"",
                ""Type"": ""System.Int32"",
                ""IsOptional"": false
            },
            {
                ""Name"": ""j"",
                ""Type"": ""System.Int32"",
                ""IsOptional"": false
            }
        ]
    }
]

```

**Going Forward**


Getting XML doc comments out isn't so clear cut, but there is a tutorial on [MSDN Blogs](http://blogs.msdn.com/b/yaohuang1/archive/2012/12/10/asp-net-web-api-help-page-part-3-advanced-help-page-customizations.aspx).


Also, there are other packages available which you can use, hook into, steal from, which do similar to what you need, for example


- [Swagger.Net](https://github.com/Swagger-Net/Swagger.Net)
- [RAML](http://raml.org/)


More details on these [in VS Mag](http://visualstudiomagazine.com/articles/2014/09/01/web-api-documentation-tools.aspx)


"
"Why does an imported function ""as"" another name keep its original \_\_name\_\_?
Here:



```
from os.path import exists as foo
print foo.__name__

```

we get: `'exists'`.
Why not `'foo'`? Which attribute would give `'foo'`?


","You can view `import foo as bar` as just an assignment. You would not expect a function to change its `__name__` attribute when you assign another name to the function.



```
>>> def foo(): pass
>>> 
>>> foo.__name__
'foo'
>>> bar = foo
>>> bar.__name__
'foo'

```


> 
> Thanks. What attribute of the variable `bar` would return the string `'bar'` then?
> 
> 
> 


There is no such attribute. Names (`bar`) refer to values (the function object) unidirectionally.


The `__name__` attribute of a function is set as the name the function was defined with using the  

`def ...` syntax. That's why you don't get a meaningful `__name__` attribute if you define an anonymous function and assign the name `foo` after it has been created.



```
>>> foo = lambda: None
>>> foo.__name__
'<lambda>'

```

"
"Soft Reference in the role of GC in java
Recently I come across this article which does the GC tuning the Jenkins which talks about this paramter : -XX:SoftRefLRUPolicyMSPerMB


<https://jenkins.io/blog/2016/11/21/gc-tuning/>


I understand it prevents OOM error, Because it clears the Soft reference object when it reaches to the threshold. 


1) What is this threshold (default = 1000ms in -XX:SoftRefLRUPolicyMSPerMB) means? What does this value denotes?


2) My jenkins seems to have 80% of the softreference (observed using HProf)


3) As suggested in the above article, If I reduce this -XX:SoftRefLRUPolicyMSPerMB flag to 10ms What will be consequence? 


NOTE : We use G1GC


Thanks,  

Harry


","1) [From Oracle](http://www.oracle.com/technetwork/java/hotspotfaq-138619.html):



> 
> Starting with 1.3.1, softly reachable objects will remain alive for
>  some amount of time after the last time they were referenced. The
>  default value is one second of lifetime per free megabyte in the heap.
>  This value can be adjusted using the -XX:SoftRefLRUPolicyMSPerMB flag,
>  which accepts integer values representing milliseconds. For example,
>  to change the value from one second to 2.5 seconds, use this flag:
> 
> 
> -XX:SoftRefLRUPolicyMSPerMB=2500
> 
> 
> 


3) I mean, it specifically says in the article you linked to. You'll possibly free up heap space by potentially sacrificing some performance. What more can we tell you? 



> 
> If Jenkins consumes excessive old generation memory, it may help to
>  make soft references easier to flush by reducing
>  `-XX:SoftRefLRUPolicyMSPerMB` from its default (1000) to something smaller (say 10-200). The catch is that SoftReferences are often used
>  for objects that are relatively expensive to load, such lazy-loaded
>  build records and pipeline FlowNode data
> 
> 
> 


"
"Order of reducer and saga
When dispatching an action is the order when it arrives to a reducer and a saga guaranteed?


Can I rely on that it 


1. first enters the reducer
2. then the saga?


Reducer:



```
 function reducer(state, action) {

    switch (action.type) {
       case 'MY_ACTION':
       // decorate action so that an epic doesn't have to take data from store
       action.ports = state.itemsModified;                 
       return state;
     }
    }

```

Saga:



```
export function* sagaUpdatePorts() {
    yield* ReduxSaga.takeEvery(actions.GRID_PORTS_ASYNC_UPDATE_PORTS, updatePorts);
}

function* updatePorts(action) {
    const {response, error} = yield SagaEffects.call(portsService.updatePorts, action.ports);
}

```

","Yes. The action hits the reducer first then Sagas.


This is [mentioned in the docs](https://redux-saga.js.org/docs/api/index.html#selectselector-args) under the `select()` effect API:



> 
> It's important to note that when an action is dispatched to the store, the middleware first forwards the action to the reducers and then notifies the Sagas. This means that when you query the Store's State, you get the State **after** the action has been applied. However, this behavior is only guaranteed if all subsequent middlewares call `next(action)` synchronously. If any subsequent middleware calls `next(action)` asynchronously (which is unusual but possible), then the sagas will get the state from **before** the action is applied. Therefore it is recommended to review the source of each subsequent middleware to ensure it calls `next(action)` synchronously, or else ensure that redux-saga is the last middleware in the call chain.
> 
> 
> 


"
"Index array out of bounds

```
if (testModetrue)
{
    try
    {
        Console.Write(""What number do you want the roll to be set to? (1-6)"");
        string diceString = Console.ReadLine();
        int diceCheck = int.Parse(diceString);
        if ((diceCheck >= minDiceValue) || (diceCheck <= maxDiceValue))
        {
            diceNo = int.Parse(diceString);
        }
        else if ((diceCheck <= minDiceValue) || (diceCheck >= maxDiceValue))
        {
            Console.WriteLine(""Please enter a number between 1-6."");
            break;
        }
    }
    catch (Exception)
    {
        Console.WriteLine(""An error has occured."");
        return;
    }
}

```

This code checks to see whether the answer given doesn't go past 6 or below 1, however whenever I run it, it does it anyway then it throws the out of array error, anybody help?


","
```
int diceCheck = int.Parse(diceString);
if ((diceCheck >= minDiceValue) || (diceCheck <= maxDiceValue))
{
    diceNo = int.Parse(diceString);
}

```

This conditional should be AND rather than OR. Also, since you're parsing the string before the conditional, you don't need to do it inside it, so you should change that part to:



```
int diceCheck = int.Parse(diceString);
if (diceCheck > maxDiceValue && diceCheck < minDiceValue)
{
   Console.Writeline(""Please write a number between 1 and 6"");
   break;
}

```

Your other if statement was also kind of redundant because you already have other variable (`dicecheck`) with the value, so remove it.


"
"How can I play video files?
I like to play video files, such as AVIs, through my C# program. Is it possible to play video files like that?


","You should be able to use the Media Player control to play media files. 


Example of playing audio from <http://msdn.microsoft.com/en-us/library/dd562692(VS.85).aspx>, you should be able to adapt it to video:



```
// [ C# ]
WMPLib.WindowsMediaPlayer Player;

private void PlayFile(String url)
{
    Player = new WMPLib.WindowsMediaPlayer();
    Player.PlayStateChange += 
        new WMPLib._WMPOCXEvents_PlayStateChangeEventHandler(Player_PlayStateChange);
    Player.MediaError += 
        new WMPLib._WMPOCXEvents_MediaErrorEventHandler(Player_MediaError);
    Player.URL = url;
    Player.controls.play();
}

private void Form1_Load(object sender, System.EventArgs e)
{
    // TODO  Insert a valid path in the line below.
    PlayFile(@""c:\myaudio.wma"");
}

private void Player_PlayStateChange(int NewState)
{
    if ((WMPLib.WMPPlayState)NewState == WMPLib.WMPPlayState.wmppsStopped)
    {
        this.Close();
    }
}

private void Player_MediaError(object pMediaObject)
{
    MessageBox.Show(""Cannot play media file."");
    this.Close();
}

```

There's a bit more information available on MSDN at <http://msdn.microsoft.com/en-us/library/dd564582(VS.85).aspx>


"
"Rails app, REMOTE\_USER attribute in IIS 8.5 with windows authentication
I have setup a IIS 8.5 server that runs a rails app using `HttpPlatformHandler` and Windows authentication. We have also tried using ARR as reverse proxy with same results.


This all works completely fine, users are authenticated using kerberos, and anonymous authentication is disabled, no problem there.
We can deny specific users and see that it has effect.


The problem is that there is no `REMOTE_USER` or `AUTH_USER` attribute anywhere in the http response headers. However the `HTTP_AUTHORIZATION` header is set and is eg ""Negotiate YIIHJAYGKw...."".


We haven't found a reasonable way to decode the `HTTP_AUTORIZATION` header either.


Any form of help would be much appreciated.


","To fix this, you need to install helicontech's ISAPI\_Rewrite 3 on the server running IIS. download url is here: [ISAPI\_Rewrite 3](http://www.helicontech.com/isapi_rewrite/download.html).


After that navigate to the install folder, the default is `C:\Program Files\Helicon\ISAPI_Rewrite3`


In there you will find a file called httpd.conf, open it up and add:



```
RewriteBase /
RewriteCond %{REQUEST_URI} ^/.*
RewriteHeader X-Remote-User: .* %{REMOTE_USER}

```

Now you will be able to get the remote username in your ruby code with:



```
request.env['HTTP_X_REMOTE_USER']

```

"
"Extending constants in Golang
Here's my requirement: I have 2 sets of constants. I want to form a third set which is nothing but a union of the above 2 sets. How do I achieve that? 



```
type CompOp byte

const (
    EQUAL CompOp = iota
    NOT_EQUAL
)

type LogOp byte

const (
    AND LogOp = iota
    OR
)

```

I want a third set, say Operator 



```
type Op {CompOp, LogOp}

```

or



```
var Op = CompOp + LogOp

```

But neither of the above 2 ways work. How do I achieve this requirement?


The above is important to me I'm trying to achieve this :



```
type logExpr struct { 
    expr Expression
    op LogOp 
}  

type compExpr struct { 
    expr Expression
    op CompOp 
} 

type filterExpr struct { 
    expr Expression
    op Op 
}

```

","`CompOp` and `LogOp` are not sets of the same type. They cannot be combined this way. If they could, they would conflict anyway, since both `EQUAL` and `AND` are both 0 (since they're the first `iota` in their blocks). You will need another design.


The most common design would be to combine all the operators into a single `const` block, and then provide functions like `IsCompare()` or `IsLogic()` to distinguish them if that's necessary. See `os.IsExist()` and `os.IsPermission()` as a template.


Here's one way that I might implement it. It wastes a few values for min/max, but it keeps the code very simple to read and easy to update.


[Playground](http://play.golang.org/p/jH14oETyDr)



```
const (
    // Comparison operators
    minComparison Op = iota
    EQUAL
    NOT_EQUAL
    maxComparison

    // Logic operators
    minLogic
    AND
    OR
    maxLogic
)

func IsComparsion(op Op) bool {
    return op >= minComparison && op <= maxComparison
}

func IsLogic(op Op) bool {
    return op >= minLogic && op <= maxLogic
}

```



---


But can you treat the different kinds of ops as types? Yeah, you can, and maybe it'll work better for you. For example, consider ([playground](http://play.golang.org/p/q5x8AjE_mS)):



```
type Op interface {
    isOp()
}

type CompOp byte

const (
    EQUAL CompOp = iota
    NOT_EQUAL
)

func (op CompOp) isOp() {}

type LogOp byte

const (
    AND LogOp = iota
    OR
)

func (op LogOp) isOp() {}

func doOpThingBasedOnValue(op Op) {
    switch op {
    case EQUAL:
        println(""passed equal"")
    case AND:
        println(""passed and"")
    }
}

func doOpThingBasedOnType(op Op) {
    switch op.(type) {
    case CompOp:
        println(""passed a comp"")
    case LogOp:
        println(""passed a logic"")
    }
}

```

So maybe that's even closer to what you're thinking of. Note that even though both `AND` and `EQUAL` are ""0"", as an interface, they are distinguishable, so we can switch on them like you'd want. Go is cool that way.


"
"What is the lifetime of a delegate created by a lambda in C#?
Lambdas are nice, as they offer [brevity and locality](https://stackoverflow.com/questions/5873603/whats-the-point-of-a-lambda-expression/5873685#5873685) and [an extra form of encapsulation](http://whathecode.wordpress.com/2011/06/05/improved-encapsulation-using-lambdas/). Instead of having to write functions which are only used once you can use a lambda.


While wondering how they worked, I intuitively figured they are **probably only created once**. This inspired me to create a solution which allows [to restrict the scope of a class member beyond private](https://stackoverflow.com/questions/4986023/restricting-scope-of-a-class-member-beyond-private/4986143#4986143) to one particular scope by using the lambda as an identifier of the scope it was created in.


This implementation works, although perhaps overkill (still researching it), proving my assumption to be correct.


A smaller example:



```
class SomeClass
{
    public void Bleh()
    {
        Action action = () => {};
    }

    public void CallBleh()
    {
        Bleh();  // `action` == {Method = {Void <SomeClass>b__0()}}
        Bleh();  // `action` still == {Method = {Void <SomeClass>b__0()}}
    }
}

```

Would the lambda ever return a new instance, or is it guaranteed to always be the same?


","Based on your question here and your comment to Jon's answer I think you are confusing multiple things. To make sure it is clear:


- The *method* that backs the delegate for a given lambda is always the same.
- The *method* that backs the delegate for ""the same"" lambda that appears lexically twice is *permitted* to be the same, but in practice is *not* the same in our implementation.
- The *delegate instance* that is created for a given lambda might or might not always be the same, depending on how smart the compiler is about caching it.


So if you have something like:



```
for(i = 0; i < 10; ++i)
    M( ()=>{} )

```

then every time M is called, you get the *same instance* of the delegate because the compiler is smart and generates



```
static void MyAction() {}
static Action DelegateCache = null;

...
for(i = 0; i < 10; ++i)
{
    if (C.DelegateCache == null) C.DelegateCache = new Action ( C.MyAction )
    M(C.DelegateCache);
}

```

If you have



```
for(i = 0; i < 10; ++i)
    M( ()=>{this.Bar();} )

```

then the compiler generates



```
void MyAction() { this.Bar(); }
...
for(i = 0; i < 10; ++i)
{
    M(new Action(this.MyAction));
}

```

You get a new delegate every time, with the same method.


The compiler is *permitted* to (but in fact does not at this time) generate



```
void MyAction() { this.Bar(); }
Action DelegateCache = null;
...
for(i = 0; i < 10; ++i)
{
    if (this.DelegateCache == null) this.DelegateCache = new Action ( this.MyAction )
    M(this.DelegateCache);
}

```

In that case you would always get the same delegate instance if possible, and every delegate would be backed by the same method.


If you have



```
Action a1 = ()=>{};
Action a2 = ()=>{};

```

Then in practice the compiler generates this as



```
static void MyAction1() {}
static void MyAction2() {}
static Action ActionCache1 = null;
static Action ActionCache2 = null;
...
if (ActionCache1 == null) ActionCache1 = new Action(MyAction1);
Action a1 = ActionCache1;
if (ActionCache2 == null) ActionCache2 = new Action(MyAction2);
Action a2 = ActionCache2;

```

However the compiler is *permitted* to detect that the two lambdas are identical and generate



```
static void MyAction1() {}
static Action ActionCache1 = null;
...
if (ActionCache1 == null) ActionCache1 = new Action(MyAction1);
Action a1 = ActionCache1;
Action a2 = ActionCache1;

```

Is that now clear?


"
"TTTAttributedLabel links are being styled but not clickable
I've been looking into a solution to getting clickable links working. I can get this working when using UITextView + NSAttributedString but it just doesn't autolayout properly when it's a UITableViewCell.


Now I've added the TTTAttributedLabel to my project and it styles the views just perfectly. The links also turn blue and are underlined.


However clicking them does nothing. I did implement TTTAttributedLabelDelegate on my controller, made the label in the storyboard implement MyLabel (Which just extends TTTAttributedLabel and has the delegate options since I want them to fire inside the same function). For now I've set the controller to be the delegate I was thinking it might not work pointing to itself.


But none of these functions get fired, I got breakpoints and logs in it.


I Implemented didSelectLinkWithUrl and didLongPressLinkWithUrl.



```
 func attributedLabel(label: TTTAttributedLabel!, didSelectLinkWithURL url: NSURL!) {
        Debug.log(""link clicked"")
    }
    func attributedLabel(label: TTTAttributedLabel!, didLongPressLinkWithURL url: NSURL!, atPoint point: CGPoint) {
        Debug.log(""link long clicked"")
    }

```

Outlet



```
@IBOutlet weak var content: MyLabel!

```

MyLabel


import UIKit
import TTTAttributedLabel



```
class MyLabel : TTTAttributedLabel, TTTAttributedLabelDelegate {

override func didMoveToSuperview() {
    if (self.delegate == nil) {
        self.delegate = self
    }
    self.enabledTextCheckingTypes = NSTextCheckingType.Link.rawValue
    self.userInteractionEnabled = true
}

func attributedLabel(label: TTTAttributedLabel!, didSelectLinkWithURL url: NSURL!) {
    Debug.log(""link clicked"")
}
func attributedLabel(label: TTTAttributedLabel!, didLongPressLinkWithURL url: NSURL!, atPoint point: CGPoint) {
    Debug.log(""link long clicked"")
}

```

Anyone know what I could be missing?


**Update**


I found out that just pasting in an url f/e <http://example.com> becomes active and is actually clickable and the didSelectLinkWithUrl becomes clickable, allthough I need an attributed string and it's based on a HTML String.


","[The implementation of `setAttributedText:`](https://github.com/TTTAttributedLabel/TTTAttributedLabel/blob/13d2d877707032dc4d4b888bb3832bac822bf9b0/TTTAttributedLabel/TTTAttributedLabel.m#L465) doesn't update the `linkModels` array, while [the implementation of `setText:`](https://github.com/TTTAttributedLabel/TTTAttributedLabel/blob/13d2d877707032dc4d4b888bb3832bac822bf9b0/TTTAttributedLabel/TTTAttributedLabel.m#L1129) does. I believe this is what causes your issue.


To resolve, set the label's `text` property instead of the `attributedText` property.


[The docs](https://github.com/TTTAttributedLabel/TTTAttributedLabel/) also include this warning:



> 
> `TTTAttributedLabel` can display both plain and attributed text: just pass an `NSString` or `NSAttributedString` to the `setText:` setter. **Never assign to the `attributedText` property.**
> 
> 
> 


The docs also show this example usage:



```
TTTAttributedLabel *attributedLabel = [[TTTAttributedLabel alloc] initWithFrame:CGRectZero];

NSAttributedString *attString = [[NSAttributedString alloc] initWithString:@""Tom Bombadil""
                                                                attributes:@{
        (id)kCTForegroundColorAttributeName : (id)[UIColor redColor].CGColor,
        NSFontAttributeName : [UIFont boldSystemFontOfSize:16],
        NSKernAttributeName : [NSNull null],
        (id)kTTTBackgroundFillColorAttributeName : (id)[UIColor greenColor].CGColor
}];

// The attributed string is directly set, without inheriting any other text
// properties of the label.
attributedLabel.text = attString;

```

"
"TFS Branches and Folders Best Practice
I have my TFS 2010 structure as follows:



```
ProjectCollectionA
       ..trunk
            ....Dev (this is a branch) 
                  ...Website1(these are folders)
                  ...Website2

            ....Release(this is a branch)
                  ...Website1(these are folders)
                  ...Website2

```

For whatever reason, I have dev and release as Branches and Website1, Website2 as Folders. My question is should I be having dev and release as folders and Website1 and Website2 as branches? Does that make more sense? What is the best practice here? Any advantages or disadvantages going one way or other? Thanks for the help


","There's a very comprehensive guide [here](https://vsardata.blob.core.windows.net/projects/TFS%20Version%20Control%20Part%201%20-%20Branching%20Strategies.pdf) from the ALM Rangers that explains the different branching practices in TFS.


If Website 1 and Website 2 are related applications within the same overall solution (not VS solution) or if there are any dependencies between the two Websites, then this (IMO) is the best approach. If they are two completely unrelated projects with isolated code bases, I would seperate it into two TFS Projects.


It's hard to decide on a correct branching pattern for someone without knowing the full ins and outs of thier business practices. What we do here, might not work for you (Actually, what we do we probably shouldn't). 


"
"Javascript (ECMA-6) class magic method \_\_call like PHP
This is my use case



```
getSomeFields(persons, fields){

    let personsWithSpecificFields = [];

    _.each(persons, (person) => {

        let personSpecificFields = {};

        _.each(fields, (field) => {
            // here im thinking to modify the field to match the method name 
            // ( if something like __call as in php is available)
            // e.g. field is first_name and i want to change it to getFirstName
            personSpecificFields[field] = person[field]();

        });

        personsWithSpecificFields.push(personSpecificFields);
    });

    return personsWithSpecificFields;
}

```

Here is my `person class`



```
import _ from 'lodash';

export default class Person{

    // not working..
    __noSuchMethod__(funcName, args){
        alert(funcName);
    }

    constructor( data ){
        this.init(data);
    }

    init(data) {
        _.each(data, (value, key) => {
           this[key] = value;
        });
    }
}

```

I have gone through [Monitor All JavaScript Object Properties (magic getters and setters)](https://stackoverflow.com/questions/6985582/monitor-all-javascript-object-properties-magic-getters-and-setters), tried to implement this [JavaScript getter for all properties](https://stackoverflow.com/questions/994143/javascript-getter-for-all-properties) but no luck.


I know i can do this by just writing another method which will transform my `first_name` to `getFirstName` and give it a shot. But is there any way to do this in `ECMA-6` way like in class.


thanks.


","You can use a [proxy](https://developer.mozilla.org/en-US/docs/Web/JavaScript/Reference/Global_Objects/Proxy) to detect access to a property that your object does not have, and deal with it -- this comes close to PHP's `__call`:



```
var person = new Person();
// Wrap the object in a proxy
var person = new Proxy(person, {
    get: function(person, field) {
        if (field in person) return person[field]; // normal case
        console.log(""Access to non-existent property '"" + field + ""'"");
        // Check some particular cases:
        if (field == 'first_name') return person.getFirstName;
        // ...
        // Or other cases:
        return function () {
            // This function will be executed when property is accessed as a function
        }
    }
});

```

You could even do this in the constructor of your class:



```
class Person {
    constructor(data) {
        this.init(data);
        return new Proxy(this, {
            get: function(person, field) {
                if (field in person) return person[field]; // normal case
                console.log(""Access to non-existent property '"" + field + ""'"");
                // Check some particular cases:
                if (field == 'first_name') return person.getFirstName;
                // ...
                // Or other cases:
                return function () {
                    // This function will be executed when property is accessed as a function
                    return 15; // example
                }
            }
        });
    }
    // other methods ...
    //
}

```

The nice thing about proxies is that the returned object is still considered an instance of the original class. With the above code, the following will be true:



```
new Person() instanceof Person

```

"
"Functions, Callable Objects, and how both are created in Python
I'm wondering about the more intricate differences between functions and callable objects. For example, if you do:



```
def foo1():
   return 2 + 4

class foo2:
   def __call__():
      return 2 + 4

import sys 
sys.getsizeof(foo1)  # returns 136
sys.getsizeof(foo2)  # returns 1016

```

There's clearly a big difference between functions and callable objects. However, I can't find a lot of documentation on what is going on behind the scenes. I know functions are first-class objects, but I also know that classes have a lot more going on than your regular functions. class foo2 is created with a metaclass, `type()`. 


My questions then, are these: 


1. When you create a function `def foo1():`, how does this differ from
the process of defining a class with a metaclass? Is there a version of `type()` but for functions, a metafunction?
2. Say someone wanted to write their own metafunction (the real reason behind this), would it be better to just use decorators, or maybe a metaclass that makes callable classes? What advantages would either offer (a metaclass that makes callable classes seems clunky)?
3. Is the only purpose behind having a callable object to have a function that can store info in it as well?


","Functions are also callable objects:



```
>>> foo1.__call__
<method-wrapper '__call__' of function object at 0x105bafd90>
>>> callable(foo1)
True

```

But a class needs to keep track of *more information*; it doesn't matter here that you gave it a `__call__` method. *Any* class is bigger than a function:



```
>>> import sys
>>> def foo1():
...    return 2 + 4
...
>>> class foo3:
...    pass
...
>>> sys.getsizeof(foo1)
136
>>> sys.getsizeof(foo3)
1056

```

A function object is a distinct object type:



```
>>> type(foo1)
<class 'function'>

```

and is reasonably compact because the meat is not actually in the function object but in other objects *referenced by* the function object:



```
>>> sys.getsizeof(foo1.__code__)
144
>>> sys.getsizeof(foo1.__dict__)
240

```

And that's it really; different types of objects have different sizes because they track different things or use composition to store stuff in other objects.


You can use the `type(foo1)` return value (or [`types.FunctionType`](https://docs.python.org/3/library/types.html#types.FunctionType), which is the same object) to produce new function objects if you so desire:



```
>>> import types
>>> types.FunctionType(foo1.__code__, globals(), 'somename')
<function foo1 at 0x105fbc510>

```

which is basically what the interpreter does whenever a `def function(..): ...` statement is being executed.


Use `__call__` to make custom classes callable when that makes sense to your API. The [`enum.Enum()` class is callable](https://docs.python.org/3/library/enum.html#functional-api), for example, specifically because using the call syntax gives you a syntax distinct from subscription, which was used for other purposes. And a [`xmlrpc.client.ServerProxy()` object](https://docs.python.org/3/library/xmlrpc.client.html#serverproxy-objects) produces method objects that are instances of `_Method`, because they proxy a remote call, not a local function.


"
"What is the difference between the various unistd.h under /usr/include in Linux?
Under the /usr/include directory in Linux i entered the command:
find -type f -name unistd.h which gave the following output:


./unistd.h
./linux/unistd.h
./asm-generic/unistd.h
./bits/unistd.h
./asm/unistd.h
./sys/unistd.h


my question is, what is the purpose of each unistd.h, since there is only one definiton of that file in the single unix specification ?


Thanks in advance.


","`linux/unistd.h` actually points to `asm/unistd.h`, which in turn points to either `asm/unistd_32.h` or `asm/unistd_64.h`, which is where system call numbers are defined and presented to user space depending on the system's architecture. These come from the kernel.


`bits/unistd.h` is a collection of macros that augment `unistd.h` (mostly stuff to help prevent buffer overflows), which is conditionally included via:



```
/* Define some macros helping to catch buffer overflows.  */
#if __USE_FORTIFY_LEVEL > 0 && defined __extern_always_inline
# include <bits/unistd.h>
#endif

```

In essence, the only POSIX required header is in fact, just `unistd.h`, the rest are either extensions, or definitions from the kernel. So, just including `unistd.h` is all you have to worry about doing, everything you need will be pulled in depending on your architecture and whatever build options you've selected. 


"
"JAXB location in file for unmarshalled objects
I have some objects being unmarshalled from an XML file by JAXB. Is it possible to have JAXB tell me or somehow find out where in the XML file (line and column) each object comes from?


This information is available at some point, because JAXB gives it to me during schema validation errors. But I would like to have it available for validated objects too.


","You could do this in JAXB by leveraging an `XMLStreamReader` and an `Unmarshaller.Listener`:


**Demo**



```
package forum383861;

import java.io.FileInputStream;
import java.util.HashMap;
import java.util.Map;

import javax.xml.bind.JAXBContext;
import javax.xml.bind.Unmarshaller;
import javax.xml.bind.Unmarshaller.Listener;
import javax.xml.stream.Location;
import javax.xml.stream.XMLInputFactory;
import javax.xml.stream.XMLStreamReader;

public class Demo {

    public static void main(String[] args) throws Exception {
        JAXBContext jc = JAXBContext.newInstance(Customer.class);


        XMLInputFactory xif = XMLInputFactory.newFactory();
        FileInputStream xml = new FileInputStream(""src/forum383861/input.xml"");
        XMLStreamReader xsr = xif.createXMLStreamReader(xml);

        Unmarshaller unmarshaller = jc.createUnmarshaller();
        LocationListener ll = new LocationListener(xsr);
        unmarshaller.setListener(ll);

        Customer customer = (Customer) unmarshaller.unmarshal(xsr);
        System.out.println(ll.getLocation(customer));
        System.out.println(ll.getLocation(customer.getAddress()));
    }

    private static class LocationListener extends Listener {

        private XMLStreamReader xsr;
        private Map<Object, Location> locations;

        public LocationListener(XMLStreamReader xsr) {
            this.xsr = xsr;
            this.locations = new HashMap<Object, Location>();
        }

        @Override
        public void beforeUnmarshal(Object target, Object parent) {
            locations.put(target, xsr.getLocation());
        }

        public Location getLocation(Object o) {
            return locations.get(o);
        }

    }

}

```

**input.xml**



```
<?xml version=""1.0"" encoding=""UTF-8""?>
<customer>
    <address/>
</customer>

```

**Output**



```
[row,col {unknown-source}]: [2,1]
[row,col {unknown-source}]: [3,5]

```

**Customer**



```
package forum383861;

import javax.xml.bind.annotation.XmlRootElement;

@XmlRootElement
public class Customer {

    private Address address;

    public Address getAddress() {
        return address;
    }

    public void setAddress(Address address) {
        this.address = address;
    }

}

```

**Address**



```
package forum383861;

public class Address {

}

```

**For More Information**


- <http://blog.bdoughan.com/2011/08/using-unmarshallerlistener-to-capture.html>


"
"Ray Tracing Calculations
I'm trying to do some calculations involving ray tracing, but am a little confused. Lets say I had an n-by-n image with N geometric primitives, l light sources, and k x k supersampling. How many ray intersections would I be calculating in the worst case? What if I added in reflection/refraction with depth d? 


","You would have to launch k x k rays for each of the n x n pixels of the image. For each of these rays you have to do collision testing, which, in a very simplified and inefficient way would translate to N comparisons (each depending on the complexity of the primitive). Now, if any of these rays hits an specular or translucid surface, you have to split the ray and call recursively. However, this time you don't do supersampling, so you just send one ray out in the reflect/refract direction. For depth d you would send that number of extra rays, one for each recursive call.


So, in total: k^2 x n^2 x d.


This is without counting the intersection calculations, which do not add any more rays but do add a lot of complexity.


There are, however, many simplifications. 


- Adaptive multisampling, reducing the k^2 factor.
- Pixel interpolations, reducing the n^2 factor.
- Use some space partition structure, like BSP and/or OctTree for collision testing.
- Use some heuristic to cut out the recursion.


"
"In a CQRS system, how should I show the user that their request has been received?
I'm trying to decouple some of the bits of our big-ball-of-mud architecture, and identified several boundaries that are obvious candidates for using CQRS to provide a more resilient and scalable solution.


Typical example: when a customer places an order, at the moment we block their thread whilst the order is submitted for payment, approved by the sales system, etc, etc.


This can all be handled asynchronously - allowing us to accept and queue orders whilst the payment processing system is unavailable, etc. - but I'm not sure how I should manage the UI data for the customer.


In other words - they place an order. Their order goes in a queue. If they log back into their account five seconds later and click ""review orders"" - what happens?


- If I draw it from the central repo (or from a cache that's updated based on that repo), then the user won't see their order and will probably try and place it again - or phone us and panic.
- If I draw it from a local database, then I have the overhead of maintaining *another* database of orders - which will need to be synchronised in a load-balanced environment, and seems to undermine a lot of the advantages of CQRS.


I want to do this in lots of places - and not all of them are actions as significant as confirming an order; in some cases it's as simple as a customer changing a phone number or something - so they're not all cases where I can just say ""thanks a lot, we'll send you a confirmation e-mail"" - because sending confirmation e-mails for every modification to a record strikes me as a little excessive.


Any patterns or solutions I should look at to help with this?


","- Something worth considering is a 'user' inbox: a place in your app the user can consult 'in-progress' commands at. You could also 'push' notifications back to the user's UI when he has already moved onto another screen, but still resides in your app. This might also be an option when the user logs back on.
- Another option could be faking the synchronous experience, i.e. wait around and do polling while in the background everything happens asynchronously. Granted, this might involve including timeouts as well, but I'd argue that those are embraced in today's synchronous processing as well.
- On top of all this, you may want to both inform and solicit feedback from your end users about how they experience your app and its behavior.
- Regardless what anybody tells you, if you want to handle this elegantly, it will take some effort on your part.


"
"My attempt to use a ""connection"" while trying to read in input causes R to freeze or crash
Sorry, but the terminology I use in the title may not be used correctly. Whenever I try to run this code, it seems like it is trying to run it but never completes the command. When I click the stop command sign (red), it doesn't do anything. I cannot close out of R. So why is this taking forever to run?



```
con <- file('stdin', open = 'r')

inputs <- readLines(con)

```

","When working in RStudio, you need to use `readLines(stdin())` rather than `readLines(file('stdin'))`, though you can use either if running R in the terminal.


However, there is also an issue from not specifying the number of lines of input since you are using RStudio. When reading input from stdin, `Ctrl`+`D` signals the end of input. However, if you are doing this from RStudio rather than from the terminal [`Ctrl`+`D` is unavailable](https://support.rstudio.com/hc/en-us/community/posts/200656347-Is-there-a-way-to-emulate-Ctrl-D-terminate-stdin-), so without specifying the lines of input there is no way to terminate the reading from stdin.


So, if you are running R from the terminal, your code will work, and you signal the end of input via `Ctrl`+`D`. If you must work from RStudio, you can still use `readLines(stdin())` [if you know the number of lines of input](https://stackoverflow.com/a/32936135/8386140); e.g.,



```
> readLines(stdin(), n=2)
Hello
World
[1] ""Hello"" ""World""

```

An alternate workaround is to use `scan()`, e.g.:



```
> scan(,'')
1: Hello
2: World
3: 
Read 2 items
[1] ""Hello"" ""World""

```

(On the third line I just pressed `Enter` to terminate input). The advantage there is that you don't need to know the number of lines of input beforehand.


"
"RestKit Dynamic nested mapping
I see that the restkit document is quite nice and has variety of examples on object modelling. There is also an example of nested mapping but I find my scenario little bit different than this. RestKit documentation provides the example mapping of the nested attribute with the following json format.


**Sample JSON structure from the RestKit Documentation :**



```
{
  ""blake"": {        
    ""email"": ""blake@restkit.org"",        
    ""favorite_animal"": ""Monkey""    
  },    
  ""sarah"": {
    ""email"": ""sarah@restkit.org"",   
    ""favorite_animal"": ""Cat""
  }
}

```

Suppose that my json is a bit different as this;


**My JSON structure :**



```
{ 
  ""id"" : 1,
  ""author"" : ""RestKit"",
  ""blake"": {        
    ""email"": ""blake@restkit.org"",        
    ""favorite_animal"": ""Monkey""    
  },    
  ""sarah"": {
    ""email"": ""sarah@restkit.org"",   
    ""favorite_animal"": ""Cat""
  }
}

```

I created two different managedobject model with the following attributes and to many relations.


Two different entities for my structure Product and creator to map the above JSON object.



```
Product                                           Creator 

identifier          <------------------- >>       name
author                                            email
                                                  favouriteAnimal

```

Now, my mapping would look like this for Product model would be;


This is how I map the Product entity, 
 [mapping mapKeyPath:""id"" toAttribute:""identifier""];
 [mapping mapKeyPath:""author"" toAttribute: ""author""];


But note here, mapping the nested dictionary attribute does not work for me.
// [mapping mapKeyOfNestedDictionaryToAttribute:@""creators""];


Now, in the authors class. 


I could not figure out the usual way to map the above JSON structure.


","If you have control over the web service, I would strongly recommend reorganizing your response data like this:



```
{
  product: 
  {
    id: 1,
    author: 'RestKit',
    creators: [
      {
        id: 1,
        name: 'Blake',
        email: '...',
        favorite_animal: 'Monkey'
      },
      {
        id: 2,
        name: 'Sarah',
        email: '...',
        favorite_animal: 'Cat'
      }
    ]
  }
}

```

Following this structure, you'd be able to use RestKit's nested mapping features, and the relationship would be correctly reflected in the deserialized objects received by the object loader delegate. RestKit relies on naming and structure standards to simplify the code required to achieve the task. Your example deviates from key-value coding standards, so RK doesn't provide an easy way to interact with your data format.


If you don't have access or you can't change it, I think you'll need to map known key-value pairs with a mapping and perform the remaining assignments with a custom evaluator. You'd need to assume the unknown keys are actually name values for associated creators and their associated values contain the attribute hash for each. Using that, you'd then reconstruct each object manually.


"
"WP-Query : check if the post content is empty
I'm trying to check if my post has content, in the loop. Currently, i added a condition in the loop : 



```
if ( $post->post_content ) 

```

and putted argument in the 



```
wp query ('posts_per_page' => 8).

```

I thought it worked, but actually, `WP query` go find last `8 posts`, and check the content of those `8 lasts`. So it renders `2` or `3` posts.


The thing I want is a way to show last `8` posts that has content.


See what I mean ?


I really appreciate some help :)


Best regards.


","This isn't possible with a standard WP query, and you'll have to leverage the use of `posts_where` before the WP\_Query is called.



```
function filter_where($where = ''){
    return $where .= ""AND trim(coalesce(post_content, '')) <>''"";
}

```

In the above, we're simply only selecting posts where the column `post_content` isn't empty.


Then add the filter.



```
add_filter('posts_where', 'filter_where');

```

Now perform the query.



```
$query = new WP_Query(array('post_type' => 'post', 'posts_per_page' => 8));

```

And then when you're done, remove the filter from the query so it doesn't interfere.



```
remove_filter('posts_where', 'filter_where');

```

"
"Memory allocation to docker containers after moving to WSL 2 in Windows
I recently updated my Docker environment to run on `WSL 2` on Windows.


For setting memory allocation limits on containers in previous versions, I had option in Docker Desktop GUI under `Settings->Resources->Advanced->Preferences` to adjust memory and CPU allocation.


After WSL 2 integration, I am not able to find that option.
[![enter image description here](https://i.stack.imgur.com/ZHGI3.png)](https://i.stack.imgur.com/ZHGI3.png)


I assume I should run everything through my Linux distro from now on, so this is the solution I was able to find:



```
docker run -d -p 8081:80 --memory=""256m"" container_name

```

I dont want to have to set a flag each time when running a container. Is there a way to permanently set the memory allocation?


","The Memory and CPU settings were removed for WSL2 integration. However, starting in Windows [Build 18945](https://learn.microsoft.com/en-us/windows/wsl/release-notes#build-18945), there is a workaround to limit WSL2 memory usage.


Create a `%UserProfile%\.wslconfig` file for configuring WSL2 settings:



```
[wsl2]
memory=6GB  # Any size you feel like (must be an integer!)
swap=0
localhostForwarding=true

```

Run `Get-Service LxssManager | Restart-Service` in an admin Powershell *(or reboot)* and verify that the `vmmem` usage in Task Manager drops off.


For the complete list of settings, please visit [Advanced settings configuration in WSL](https://learn.microsoft.com/en-us/windows/wsl/wsl-config).


"
"REGEX to find the first one or two capitalized words in a string
I am looking for a REGEX to find the first one or two capitalized words in a string. If the first two words is capitalized I want the first two words. A hyphen should be considered part of a word.


1. for `Madonna has a new album` I'm looking for `madonna`
2. for `Paul Young has no new album` I'm looking for `Paul Young`
3. for `Emmerson Lake-palmer is not here` I'm looking for `Emmerson Lake-palmer`


I have been using `^[A-Z]+.*?\b( [A-Z]+.*?\b){0,1}` which does great on the first two, but for the 3rd example I get `Emmerson Lake`, instead of `Emmerson Lake-palmer`. 


What REGEX can I use to find the first one or two capitalized words in the above examples?


","You may use



```
^[A-Z][-a-zA-Z]*(?:\s+[A-Z][-a-zA-Z]*)?

```

See the [regex demo](https://regex101.com/r/DiQUic/2)


Basically, use a character class `[-a-zA-Z]*` instead of a dot matching pattern to only match letters and a hyphen.


**Details**


- `^` - start of string
- `[A-Z]` - an uppercase ASCII letter
- `[-a-zA-Z]*` - zero or more ASCII letters / hyphens
- `(?:\s+[A-Z][-a-zA-Z]*)?` - an optional (1 or 0 due to `?` quantifier) sequence of:
	- `\s+` - 1+ whitespace
	- `[A-Z]` - an uppercase ASCII letter
	- `[-a-zA-Z]*` - zero or more ASCII letters / hyphens


A Unicode aware equivalent (for the regex flavors supporting Unicode property classes):



```
^\p{Lu}[-\p{L}]*(?:\s+\p{Lu}[-\p{L}]*)?

```

where `\p{L}` matches any letter and `\p{Lu}` matches any uppercase letter.


"
"is there a way to handle standard fields in this DU, in F#?
I have some quite ugly code here:



```
type AnalysisEvent =
    | ZoneStart             of DateTime * ConsolidationZone
    | ZoneEndExitHigh       of DateTime * ConsolidationZone
    | ZoneEndExitLow        of DateTime * ConsolidationZone
    | ZoneContraction       of DateTime * ConsolidationZone
    | ZoneExpansion         of DateTime * ConsolidationZone
    | FairPriceCrossHigh    of DateTime * FairPriceLine
    | FairPriceCrossLow     of DateTime * FairPriceLine
    | LiquidityLineCreated  of DateTime * LiquidityLine
    | LiquidityLineReached  of DateTime * LiquidityLine

    member this.GetTimestamp () =
        match this with
        | ZoneStart             (ts, _) -> ts
        | ZoneEndExitHigh       (ts, _) -> ts
        | ZoneEndExitLow        (ts, _) -> ts
        | ZoneContraction       (ts, _) -> ts
        | ZoneExpansion         (ts, _) -> ts
        | FairPriceCrossHigh    (ts, _) -> ts
        | FairPriceCrossLow     (ts, _) -> ts
        | LiquidityLineCreated  (ts, _) -> ts
        | LiquidityLineReached  (ts, _) -> ts

    member this.GetInterval () =
        match this with
        | ZoneStart             (_, x) -> x.Interval
        | ZoneEndExitHigh       (_, x) -> x.Interval
        | ZoneEndExitLow        (_, x) -> x.Interval
        | ZoneContraction       (_, x) -> x.Interval
        | ZoneExpansion         (_, x) -> x.Interval
        | FairPriceCrossHigh    (_, x) -> x.Interval
        | FairPriceCrossLow     (_, x) -> x.Interval
        | LiquidityLineCreated  (_, x) -> x.Interval
        | LiquidityLineReached  (_, x) -> x.Interval

```

And there are many common fields across all DU types that I need to extract like this; and they have, of course, a lot of different fields as well.


Is there a nicer way to extract the common fields? The DateTime is separated from the underlying type, but there are also common fields across all the types.


How could this be rewritten in a nicer way?


","As Fyodor suggests, this can be refactored as:



```
type AnalysisEventType =
    | ZoneStart             of ConsolidationZone
    | ZoneEndExitHigh       of ConsolidationZone
    | ZoneEndExitLow        of ConsolidationZone
    | ZoneContraction       of ConsolidationZone
    | ZoneExpansion         of ConsolidationZone
    | FairPriceCrossHigh    of FairPriceLine
    | FairPriceCrossLow     of FairPriceLine
    | LiquidityLineCreated  of LiquidityLine
    | LiquidityLineReached  of LiquidityLine

type AnalysisEvent =
    {
        EventType : AnalysisEventType
        Timestamp : DateTime
    }

    member this.GetTimestamp () =   // you probably don't even need this any more
        this.Timestamp

    member this.GetInterval () =
        match this.EventType with
        | ZoneStart             x
        | ZoneEndExitHigh       x
        | ZoneEndExitLow        x
        | ZoneContraction       x
        | ZoneExpansion         x -> x.Interval
        | FairPriceCrossHigh    x
        | FairPriceCrossLow     x -> x.Interval
        | LiquidityLineCreated  x
        | LiquidityLineReached  x -> x.Interval

```

And if you want to refactor even further, you could do something like this:



```
type ConsolidationZoneEventType =
    | Start
    | EndExitHigh
    | EndExitLow
    | Contraction
    | Expansion

type FairPriceEventType =
    | CrossHigh
    | CrossLow

type LiquidityLineEventType =
    | Created
    | Reached

type AnalysisEventType =
    | ConsolidationZoneEventType of ConsolidationZoneEventType * ConsolidationZone
    | FairPriceEventType of FairPriceEventType * FairPriceLine
    | LiquidityLineEventType of LiquidityLineEventType * LiquidityLine

    member this.GetInterval () =
        match this with
        | ConsolidationZoneEventType (_, x) -> x.Interval
        | FairPriceEventType (_, x) -> x.Interval
        | LiquidityLineEventType (_, x) -> x.Interval

type AnalysisEvent =
    {
        EventType : AnalysisEventType
        Timestamp : DateTime
    }

    member this.GetTimestamp () =
        this.Timestamp

    member this.GetInterval () =
        this.EventType.GetInterval ()

```

"
"Save an UIImagePicker Video
I have searched and searched but can not find a basic `UIImagepicker` video save code.
I have everything set up and just have a blank void.



```
- (void)imagePickerController:(UIImagePickerController *)
         picker didFinishPickingMediaWithInfo:(NSDictionary *)
              info{
}

```

I know it has something to do with the dictionary, but after that I'm lost, any help?


","Try this code for save the video on photo album.



```
    - (void)imagePickerController:(UIImagePickerController *)picker didFinishPickingMediaWithInfo:(NSDictionary *)info
    {
        [self dismissModalViewControllerAnimated:YES];

        //get the videoURL 
        NSString *tempFilePath = [[info objectForKey:UIImagePickerControllerMediaURL] path];

        if ( UIVideoAtPathIsCompatibleWithSavedPhotosAlbum(tempFilePath))
        {
                // Copy it to the camera roll.
                UISaveVideoAtPathToSavedPhotosAlbum(tempFilePath, self, @selector(video:didFinishSavingWithError:contextInfo:), tempFilePath);
            } 
    }

    - (void)imagePickerControllerDidCancel:(UIImagePickerController *)picker {
        [self dismissModalViewControllerAnimated:YES];
    }

    - (void) video: (NSString *) videoPath
    didFinishSavingWithError: (NSError *) error
       contextInfo: (void *) contextInfo {
        NSLog(@""Finished saving video with error: %@"", error);
    }

```

"
"How to get HTTP StatusCodes in ember-data
When I invoke 



```
 App.store.createRecord(App.User,  { name: this.get(""name"") });
 App.store.commit();

```

how do I know if its successful and how to wait for the asyn message?


","Very limited error handling was [recently added](https://github.com/emberjs/data/commit/36a70c9839c9807248856ce2e25959bf80d6b941) to DS.RESTAdapter in ember-data master.


When creating or updating records (with bulk commit disabled) and a status code between 400 and 599 is returned, the following will happen:


- A `422 Unprocessable Entity` will transition the record to the ""invalid"" state and will add any errors returned from the server to the record's `errors` property.


The adapter assumes the server will respond with JSON in the following format:




```
{
  errors: {
    name: [""can't be blank""],
    password: [""must be at least 8 characters"", ""must contain a number""]
  {
}

```

(The error messages themselves could be arrays of strings or just strings. ember-data doesn't currently care which.)


To detect this state:




```
record.get('isValid') === false

```
- All other status codes will transition the record to the ""error"" state.


To detect this state, use:




```
record.get('isError') === true

```


More cases may eventually be handled by ember-data out of the box, but for the moment if you need something specific, you'll have to extend `DS.RESTAdapter`, customizing its `didError` function to add it in yourself.


"
"pointer member of a class has to point to heap-allocated data?
The following code is to verify if I have a pointer member in a class, without making an overloaded assignment operator. When I make two instances equal,it will just make the pointers point to the same thing instead of copying the data. However, I got an error trying to compile before I could verify. It appears to me that it implies a pointer member in class has to be pointed to a heap-allocated data. Is it right? 



> 
> The error is: initializing pointer member 'a' with the stack address of parameter 'aa' [-Werror,-Wdangling-field]
> 
> 
> 


Second question is when do we need an overloaded ""=""operator? I guess would be when we have a pointer member in the class and we want whatever the pointer points to to be copied again instead of just making two pointer from the two classes point to same thing. Could somebody tell me this is correct. Thanks! 



```
class ClassA {    
  int* a;  
  int* b;  

  ClassA():a(NULL),b(NULL){};

  ClassA(int aa,int bb):a(&aa),b(&bb){};                                                                                                                                                                                     
};

int main(){ 

  ClassA test;
  ClassA subject(5,6);

  test = subject;
}

```

","### Explanation


**No**, there's nothing saying that he address contained in a pointer must be the location of an object with *dynamic storage duration*, but please be careful when initializing them with addresses of variables on the stack (having *automatic storage duration*).


  

As your compiler tells you; you are currently assigning the address of two *temporaries* to the pointers inside your class. This means that after the constructor is finished, and `aa` and `bb` are no longer alive, but you can still potentially refer to them through `this->a`, and `this->b`.


If you try to use `*(this->a)` and `*(this->b)` when their *pointed-to-object* in no longer available you are invoking *undefined behavior*, meaning that your application will most likely crash since you are trying to access memory which is no longer valid to use.




---


### When do I need `operator=`?


This is an often asked question on stackoverflow, and I'd recommend you to read through the below questions/answers:


- [**What is The Rule of Three?**](https://stackoverflow.com/q/4172722/1090079)
- [**Rule-of-Three becomes Rule-of-Five with C++11?**](https://stackoverflow.com/q/4782757/1090079)


"
"Strategy for designing a plugin API
I am working on an application for which a rich plugin API is crucial. What is the best process for creating an API that is intuitive and easy to learn?


","
> 
> What is the best process for creating an API that is intuitive and easy to learn?
> 
> 
> 


First, create an API that works.


Then, after using it for a while, get feedback on the parts that are not intuitive or easy to learn.


Third, rewrite the API to be more intuitive and easier to learn.


Then, after using the revised API for a while, get more feedback on the parts that are not intuitive or easy to learn. 


When you get to the 3.0 rewrite of the API to be more intuitive and easier to learn, you may actually have something that works.


This is very, very hard to do ""out of the box"".


It takes experience, and -- more importantly -- it takes a lot of back and forth between the producers and consumers of the API's. 


Until you know your users, you have a lot to learn about them and what amounts to ""intuitive"" and ""easy to learn"".


Most API's don't really strive for this. They strive for ""efficient"" or ""simple"" or something that's much easier to achieve.


Here's the point.


- ""Intuitive"" and ""easy-to-learn"" are very personal attributes. What is intuitive for one person is stupid to another. What is easy to learn for one person is baffling to another.
- You can't achieve ""Intuitive"" and ""easy-to-learn"" without knowing your users really, really well.
- You can't know your users really, really well until you interact with them by providing them an interface, which they use, and then you get detailed feedback on their experience.


Learning about your users takes a lot of back and forth between you and them.


You must make active steps to learn about your users so that you know what constitutes ""intuitive"" and ""easy-to-learn"". When your users argue amongst themselves, you'll learn a lot about what ""intuitive"" *really* means. [Hint: it means nothing.]


One easy way to take active steps to learning about the users is to cycle through several releases. There are other ways to learn about your users. 


"
"Extracting multiple values from a character field into a new data table
My data looks like the following:



```
ids <- c(1,2,3)
titles <- c(""Entry1"", ""Entry2"", ""Entry3"")
tags <- c(""<self-help><motivation>"", ""<programming><r><data.frame>"", ""<photography>"")
df <- data.frame(id = ids, title = titles, tags = tags)
df

```

Output:



```
   id  title                         tags
1   1 Entry1      <self-help><motivation>
2   2 Entry2 <programming><r><data.frame>
3   3 Entry3                <photography>

```

I'm struggling to extract out the tags into a new data frame. What I want is a second data frame that looks like this:



```
  id         tag
1  1   self-help
2  1  motivation
3  2 programming
4  2           r
5  2  data.frame
6  3 photography

```

I've managed to do *something* with lists using the following, but the resulting column seems to be nested somehow. My attempts to `unlist()` either have no effect or produce too many entries in the resulting vector.



```
df$tags_list <- lapply(df$tags, function(x)strsplit(gsub(""^.|.$"", """", x), ""><"")[[1]])

```

My question is: can someone please help me generate this new `data.frame` (or `data.table`) given `df` as shown here? Any words of wisdom on how to deal with nested lists in general are greatly appreciated as well.


","With `base R` version `3.2.0+` you could try:



```
lst <- strsplit(gsub(""^<|>$"", """", df$tags), split=""><"")
data.frame(id=rep(df$id, lengths(lst)), tags_list=unlist(lst))
#   id   tags_list
# 1  1   self-help
# 2  1  motivation
# 3  2 programming
# 4  2           r
# 5  2  data.frame
# 6  3 photography

```

With `gsub(""^<|>$""..)` we eliminate beginning and terminal brackets leaving the pattern `""><""` as word separators. Then we create the data frame with the ids and the separated list. 


**Update**


The function `lengths` was added in R 3.2.0 and does for lists what `length` similarly does for vectors. Before it was required to use `sapply(lst, length)`.


"
"Regexp to capture string between delimiters
In [this](https://stackoverflow.com/questions/1454913/regular-expression-to-find-a-string-included-between-two-characters-while-exclu/1454936#comment21122977_1454936) question a regex for capturing a string between delimiters is provided:


**Test:** `This is a test string [more or less]`


**Regexp:** `(?<=\[)(.*?)(?=\])`


**Returns:** `more or less`


What if the string to be captured also contains delimiters?


**Test 1:** `This is a test string [more [or] less]`


**Return 1:** `more [or] less`


**Test 2:** `This is a test string [more [or [and] or] less]`


**Return 2:** `more [or [and] or] less`


And multiple brackets?


**Test 3:** `This is a test string [more [or [and] or] less] and [less [or [and] or] more]`


**Return 3:** `more [or [and] or] less`, `less [or [and] or] more`


Which regex would do this? Or which small ruby/python script can do this?


","In javascript



```
var str = 'This is a test string [more [or [and] or] less]';    
str = str.match( /\[(.+)\]/ )[1];
// ""more [or [and] or] less""

```

If you omit the `?`, the `.+` will match greedily up to the last `]`.


In python



```
str = ""This is a test string [more [or [and] or] less]""
re.search( ""(?<=\[).+(?=\])"", str ).group()
// ""more [or [and] or] less""

```

**Update** for multiple nested brackets


In javascript



```
var matches = [],
    str = 'This is a test string [more [or [and] or] less] and [less [or [and] or] more] and [more]';

str.replace( /\[([^\]]*\[?[^\]]*\]?[^[]*)\]/g, function ( $0, $1 ) {
    $1 && matches.push( $1 );
});

console.log( matches );
// [ ""more [or [and] or] less"", ""less [or [and] or] more"", ""more"" ]

```

In python



```
import re
str = 'This is a test string [more [or [and] or] less] and [less [or [and] or] more] and [more]'

matches = re.findall( r'\[([^\]]*\[?[^\]]*\]?[^[]*)\]', str )

print matches
# [ 'more [or [and] or] less', 'less [or [and] or] more', 'more' ]

```

"
"Training custom SVM to use with HOGDescriptor in OpenCV
I'm trying to train my own detector for use with OpenCV::HOGDescriptor but I'm having trouble making the existing HOGDescriptor work with my newly trained SVM. 


I have calculated HOG features for positive and negative training images, labeled them and trained the SVM using CvSVM. The parameters I have used are:



```
    CvSVMParams params;
    params.svm_type =CvSVM::EPS_SVR;
    params.kernel_type = CvSVM::LINEAR;
    params.C = 0.01;
    params.p = 0.5;

```

Then I calculate Primal Form of the support vectors so that I only get one vector instead of many and set the calculated support vector using HOGDescriptor.setSVMDetector(vector);


[This is Primal Form](http://en.wikipedia.org/wiki/Support_vector_machine#Primal_form)


When I use CvSVM.predict() I am able to correctly classify objects with the SVM, but HOGDescriptor.detect() or detectMultiScale() always returns a lot of positive matches and does not give accurate predictions. 


CvSVM.predict() uses the original support vectors for classification so there might be something wrong with the way I'm calculating primal form. 


Is there anyone who has trained their own detector who can point me in the right direction? 


","I wrote a child class of CvSVM to extract primal form after a linear svm is trained. Positive samples are labeled 1 and negative samples are labeled -1. It is strange that I have to put negative sign in front of alphas and leaving the sign of rho unchanged in order to get correct results from HogDescriptor. 


LinearSVM.h



```
#ifndef LINEAR_SVM_H_
#define LINEAR_SVM_H_
#include <opencv2/core/core.hpp>
#include <opencv2/ml/ml.hpp>

class LinearSVM: public CvSVM {
public:
  void getSupportVector(std::vector<float>& support_vector) const;
};  

#endif /* LINEAR_SVM_H_ */

```

LinearSVM.cc



```
#include ""linear_svm.h""    
void LinearSVM::getSupportVector(std::vector<float>& support_vector) const {

    int sv_count = get_support_vector_count();
    const CvSVMDecisionFunc* df = decision_func;
    const double* alphas = df[0].alpha;
    double rho = df[0].rho;
    int var_count = get_var_count();
    support_vector.resize(var_count, 0);
    for (unsigned int r = 0; r < (unsigned)sv_count; r++) {
      float myalpha = alphas[r];
      const float* v = get_support_vector(r);
      for (int j = 0; j < var_count; j++,v++) {
        support_vector[j] += (-myalpha) * (*v);
      }
    }
    support_vector.push_back(rho);
}

```

"
"VB.NET For each exception on custom controls
in VB.NET i have 2 custom controls, one is a TextBox and second one is a ComboBox.
These have custom values like Bool \_IsHidden and are added on runtime to a form.


Now, at some point in the code I want to check if the \_IsHidden is set to True or False and display that information. Since the user can edit this values when creating the control these are not set on creation.


So what I tried is:


(all of this is on MDI Forms)



```
For Each frm as CustomForm in Main.MdiChildren
If frm.MyName = calledBy Then 'this part is just to know which form called the form to create the object
For Each cntrl as CustomTextBox in frm.Controls
'DO Something
Next
End if
Next

```

Now.. if the first control is a custom ComboBox it thorws an error since it sees that it does not match the custom TextBox control.. 


how do i get around this? By my understanding it should just go through all of the controls on the said form and just check those who match CustomTextBox control ?


Thank you


","`For Each x As T In collection` does *not* filter your collection items to those of type `T`. It tries to convert *every* item in `collection` to `T` and throws an exception if that fails.


Thus, you have the following options:


1. Do the check yourself, for example, using the code provided by [RB](https://stackoverflow.com/a/8443091/87698).
2. Alternatively, you could filter your list *first*, and then loop through the items. Here, LINQ can help:



```
For Each cntrl In frm.Controls.OfType(Of CustomTextBox)()
    ... ' Do this for all CustomTextBoxes
Next

For Each cntrl In frm.Controls.OfType(Of CustomComboBox)()
    ... ' Do this for all CustomComboBoxes
Next

```

You don't need the `As CustomTextBox` clause here, since [`frm.Controls.OfType(Of CustomTextBox)()`](http://msdn.microsoft.com/en-us/library/bb360913.aspx) returns an `IEnumerable(Of CustomTextBox)`, so `For Each` can infer by itself that `cntrl` must be of type `CustomTextBox`.


"
"Run tests concurrently
I would like to run several tests concurrently using asyncio (/curio/trio) and pytest, but I couldn't find any information on that. Do I need to schedule them myself? And if I do, is there a way to have a nice output that separates (sub-)tests cases?


Here is a small toy example for which I'm trying this:



```
import pytest
import time
import asyncio

pytestmark = pytest.mark.asyncio
expected_duration = 1
accepted_error = 0.1

async def test_sleep():
    start = time.time()
    time.sleep(expected_duration)
    duration = time.time() - start
    assert abs(duration-expected_duration) < accepted_error

async def test_async_sleep():
    start = time.time()
    await asyncio.sleep(expected_duration)
    duration = time.time() - start
    assert abs(duration-expected_duration) < accepted_error

```

","Unfortunately, the way pytest works internally, you can't really run multiple tests at the same time under the same call to `trio.run`/`asyncio.run`/`curio.run`. (This is also good in some ways – it prevents state from leaking between tests, and with trio at least it lets you configure trio differently for different tests, e.g. setting one test to use the [autojump clock](https://pytest-trio.readthedocs.io/en/latest/quickstart.html#trio-s-magic-autojump-clock) while another test doesn't.)


Definitely the simplest option is to use pytest-xdist to run tests in separate threads. You can still use async internally inside each test – all these async libraries support running different loops in different threads.


If you really need to use async concurrency, then I think you'll have to write a single pytest test function, and then inside that function do your own scheduling and concurrency. If you do it this way, then from pytest's perspective there's only one test, so it won't be easy to get nice per-test output. I guess you could try using [pytest-subtests](https://pypi.org/project/pytest-subtests/)?


"
"How can I combine hue and style groups in a Seaborn legend?
I'm doing a Seaborn lineplot for longitudinal data which is grouped by ""Subscale"" using `hue` and by ""Item"" using `style`.


Here is my code (I hope this is understandable also without data):



```
ax = sns.lineplot(data = df, x = 'Week', y = 'Value', style = 'Item', hue = 'Subscale', palette = 'colorblind', markers = True)
plt.legend(bbox_to_anchor = (1.03, 1.02), fontsize = 10)

```

Whicht gets me this plot:
[![enter image description here](https://i.stack.imgur.com/qGnm9.png)](https://i.stack.imgur.com/qGnm9.png)


What I want is to **combine the sublegends** so that it only shows the legend for ""Item"" but the items are colored according to ""Subscale"", kind of like this:
[![enter image description here](https://i.stack.imgur.com/hI5Y3.png)](https://i.stack.imgur.com/hI5Y3.png)


I've failed to create this, so if any of you could help, this would be highly appreciated! Thanks :)


","If I understand correctly, all items of a certain type have the same subscale. So, you already have (or can create) a dictionary that maps an item type to the corresponding subscale.


Seaborn creates following labels for the legend:


- 'Subscale' for a subtitle
- each of the subscales
- 'Item' for a second subtitle
- each of the items


Each label corresponds to a ""handle"" identifying how the marker looks like.


The following code:


- finds the index of `'Item'` to be able to split the arrays of labels and handles
- extracts the colors of the 'Subscale's
- applies these colors to the item markers
- only uses the items for the legend



```
import matplotlib.pyplot as plt
import seaborn as sns
import pandas as pd
import numpy as np

# first create some random data similar to the description
np.random.seed(123)
items = ['01', '04', '05', '06', '07', '10', '11', '13']
N = len(items)
subscale_dict = {'01': 'A', '04': 'C', '05': 'C', '06': 'C', '07': 'A', '10': 'B', '11': 'B', '13': 'A'}

df = pd.DataFrame({'Week': np.tile(np.arange(7), N),
                   'Value': np.random.rand(7 * N),
                   'Item': np.repeat(items, 7)})
df['Subscale'] = df['Item'].apply(lambda i: subscale_dict[i])
df['Subscale'] = pd.Categorical(df['Subscale'])  # creates a fixed order

# create the line plot as before
ax = sns.lineplot(data=df, x='Week', y='Value', style='Item', hue='Subscale', palette='colorblind', markers=True)

# create a dictionary mapping the subscales to their color
handles, labels = ax.get_legend_handles_labels()
index_item_title = labels.index('Item')
color_dict = {label: handle.get_color()
              for handle, label in zip(handles[1:index_item_title], labels[1:index_item_title])}

# loop through the items, assign color via the subscale of the item idem
for handle, label in zip(handles[index_item_title + 1:], labels[index_item_title + 1:]):
    handle.set_color(color_dict[subscale_dict[ label]])

# create a legend only using the items
ax.legend(handles[index_item_title + 1:], labels[index_item_title + 1:], title='Item',
          bbox_to_anchor=(1.03, 1.02), fontsize=10)

plt.tight_layout()
plt.show()

```

[![sns.lineplot with custom legend](https://i.stack.imgur.com/JsQqS.png)](https://i.stack.imgur.com/JsQqS.png)


"
"ggplot2: Is there a way to overlay a single plot to all facets in a ggplot
I would like to use ggplot and faceting to construct a series of density plots grouped by a factor. Additionally, I would like to a layer another density plot on each of the facets that is not subject to the constraints imposed by the facet. 


For example, the faceted plot would look like this: 



```
require(ggplot2)
ggplot(diamonds, aes(price)) + facet_grid(.~clarity) + geom_density()

```

and then I would like to have the following single density plot layered on top of each of the facets:



```
ggplot(diamonds, aes(price)) + geom_density()

```

Furthermore, is ggplot with faceting the best way to do this, or is there a preferred method? 


","One way to achieve this would be to make new data frame `diamonds2` that contains just column `price` and then two `geom_density()` calls - one which will use original `diamonds` and second that uses `diamonds2`. As in `diamonds2` there will be no column `clarity` all values will be used in all facets.



```
diamonds2<-diamonds[""price""]
ggplot(diamonds, aes(price)) + geom_density()+facet_grid(.~clarity) + 
     geom_density(data=diamonds2,aes(price),colour=""blue"")

```

![enter image description here](https://i.stack.imgur.com/Yt8qA.png)


**UPDATE** - as suggested by @BrianDiggs the same result can be achieved without making new data frame but transforming it inside the `geom_density()`.



```
ggplot(diamonds, aes(price)) + geom_density()+facet_grid(.~clarity) +
     geom_density(data=transform(diamonds, clarity=NULL),aes(price),colour=""blue"")

```

Another approach would be to plot data without faceting. Add two calls to `geom_density()` - in one add `aes(color=clarity)` to have density lines in different colors for each level of `clarity` and leave empty second `geom_density()` - that will add overall black density line.



```
ggplot(diamonds,aes(price))+geom_density(aes(color=clarity))+geom_density()

```

![enter image description here](https://i.stack.imgur.com/bjYSj.png)


"
"Visual Studio - ""The environment IronPython|2.7-32 appears to be incorrectly configured or missing""
Steps to the Reproduce:


1. Open Visual Studio 2017 (Pro on Windows 10 64-bit)
2. File > New Project > IronPython Application
3. Run the default program: `print('Hello world')`


When I run it, I get the following error:



> 
> The environment ""IronPython|2.7-32"" appears to be incorrectly configured or missing.  
> 
>  You may need to install it or create a virtual environment
> 
> 
> 


[![Python Error](https://i.stack.imgur.com/hGrcx.png)](https://i.stack.imgur.com/hGrcx.png)


I went to VS Installer, Individual Components tab, Compilers section, and checked on the Python 2.7 & 3.6 selections, but I'm still getting the problem


","You need to set up the IronPython 2.7 environment\*.  

\*Currently you cannot do it through the Visual Studio Installer (there isn't an IronPython Package option to install).


1. You need to download IronPython and install it separately from [Github](https://github.com/IronLanguages/ironpython2/releases/).
2. After it's installed, you need to [select the correct python environment within your VS project](https://learn.microsoft.com/en-us/visualstudio/python/managing-python-environments-in-visual-studio?view=vs-2019#the-python-environments-window).
3. If the IronPython 2.7 environment is not properly set then you have to do it [manually](https://learn.microsoft.com/en-us/visualstudio/python/managing-python-environments-in-visual-studio?view=vs-2019#manually-identify-an-existing-environment)


Hopefully, Microsoft will patch this in the future.


"
"UILexicon in Objective-C
How do you use UILexicon in Objective-C? I find the documentation Apple provides is extremely unhelpful.


What does it do? Does it return a dictionary or proper spellings of words? Or do I provide a word like ""hellllo"" and it matches it with the proper spelling ""Hello"" and returns that as a string?


Any help would be appreciated.


`requestSupplementaryLexiconWithCompletion:` 




---


Here's my error report, but obviously I'll have errors because I'm completely guessing how to use the function, no clue what goes inside the block statement (because the docs (at the time) don't say! (Beta 4 docs)) Hahahah!
![enter image description here](https://i.stack.imgur.com/OMFui.png)


","I've never used this feature, but a quick web search for ""UILexicon"" landed me in Apple's documentation; reading and following links from there filled in the picture pretty quick.


[*App Extension Programming Guide*](https://developer.apple.com/library/prerelease/ios/documentation/General/Conceptual/ExtensibilityPG/Keyboard.html#//apple_ref/doc/uid/TP40014214-CH16-SW1) has a quick explanation of what lexicons are for:



> 
> Every custom keyboard (independent of the value of its `RequestsOpenAccess` key) has access to a basic autocorrection lexicon through the [`UILexicon`](https://developer.apple.com/library/prerelease/ios/documentation/UIKit/Reference/UILexicon_Class/index.html#//apple_ref/occ/instp/UILexicon/entries) class. Make use of this class, along with a lexicon of your own design, to provide suggestions and autocorrections as users are entering text.
> 
> 
> 


Clicking the [`UILexicon`](https://developer.apple.com/library/prerelease/ios/documentation/UIKit/Reference/UILexicon_Class/index.html#//apple_ref/occ/instp/UILexicon/entries) link on that page took me to the reference doc for that class, which explains that it's a read-only list of Apple-provided term pairs. Each of its entries is a [`UILexiconEntry`](https://developer.apple.com/library/prerelease/ios/documentation/UIKit/Reference/UILexiconEntry_Class/index.html#//apple_ref/occ/cl/UILexiconEntry) object -- the docs for that class say it provides a `userInput` (what the user typed, e.g. ""ipad"") and a `documentText` (what to substitute for it, e.g. ""iPad""). Since those classes are read-only, it follows that they're probably not a way for you to provide your own autocorrection pairs -- as stated in the docs, they're for supplementing whatever autocorrection system you implement.




---


At this point, I don't even have to look at the doc for [`requestSupplementaryLexiconWithCompletion:`](https://developer.apple.com/library/prerelease/ios/documentation/UIKit/Reference/UIInputViewController_Class/index.html#//apple_ref/occ/instm/UIInputViewController/requestSupplementaryLexiconWithCompletion:) to get a good idea how to use it: just the declaration tells me: 


- It's a method on `UIInputViewController`, the class I'd have to subclass to create a custom keyboard. Somewhere in that subclass I should probably call it on `self`.
- Its return type is `void`, so I can't get a lexicon by assigning the result of a `requestSupplementaryLexiconWithCompletion` call to to a variable.
- It calls the block I provide, passing me a `UILexicon` object as a parameter to that block.
- It's got words like ""request"" and ""completionHander"" in it, so it'll probably do something asynchronous that takes awhile, and call that block when it's done.


So, I'm guessing that if I were writing a custom keyboard, I'd call this method early on (in `viewDidLoad`, perhaps) and stash the `UILexicon` it provides so I can refer to it later when the user is typing. Something like this:



```
@property UILexicon *lexicon;

- (void)viewDidLoad {
    [super viewDidLoad];
    [self requestSupplementaryLexiconWithCompletion:^(UILexicon *lexicon){
        self.lexicon = lexicon;
    }];
}

```

Because it's unclear how long `requestSupplementaryLexiconWithCompletion` will take to complete, any place where I'm using `self.lexicon` I should check to see if it's `nil`.




---


Back in the *App Extension Programming Guide*, it lists ""Autocorrection and suggestion"" under ""Keyboard Features That iOS Users Expect"", right before saying: 



> 
> You can decide whether or not to implement such features; there is no dedicated API for any of the features just listed
> 
> 
> 


So it sounds like autocorrection is something you have to do yourself, with your own UI that's part of the view presented by your [`UIInputViewController`](https://developer.apple.com/library/prerelease/ios/documentation/UIKit/Reference/UIInputViewController_Class/index.html) subclass. The [API Quick Start for Custom Keyboards](https://developer.apple.com/library/prerelease/ios/documentation/General/Conceptual/ExtensibilityPG/Keyboard.html#//apple_ref/doc/uid/TP40014214-CH16-SW11) section in the programming guide seems to hint at how you'd do that: use [`documentContextBeforeInput`](https://developer.apple.com/library/prerelease/ios/documentation/UIKit/Reference/UITextDocumentProxy_Protocol/index.html#//apple_ref/occ/intfp/UITextDocumentProxy/documentContextBeforeInput) to see what the user has recently typed, [`deleteBackward`](https://developer.apple.com/library/prerelease/ios/documentation/UIKit/Reference/UIKeyInput_Protocol/index.html#//apple_ref/occ/intfm/UIKeyInput/deleteBackward) to get rid of it, and [`insertText:`](https://developer.apple.com/library/prerelease/ios/documentation/UIKit/Reference/UIKeyInput_Protocol/index.html#//apple_ref/occ/intfm/UIKeyInput/insertText:) to insert a correction.


"
"Icon file names iOS 7
How do I have to name the Icon files for Xcode 5?
It gives always errors that de app is not on the top level, it is really frustrating.
Can someone give me the filenames that you have to use for every resolution?


","icon file name and size for ios 7 compatible app as below (iPhone)


Icon.png - 57\*57


Icon@2x.png - 114\*114


Icon-Small.png - 29\*29


Icon-Small@2x.png - 58\*58


Icon-40.png - 40\*40


Icon-40@2x.png - 80\*80


Icon-60@2x.png - 120\*120


Icon-76.png - 76 x 76


Icon-76@2x.png 152 x 152


(EDIT: corrected names of 120 & 40 sizes, and added 76 + 152)


"
"What is the purpose of the ""get"" and ""set"" properties in C#
I saw some get set method to set values. Can anyone tell me the purpose of this?



```
public string HTTP_USER_NAME
{
      get 
      {
            return UserName; 
      }
      set 
      {
            UserName = value; 
      }
}

public string HTTP_USER_PASSWORD
{
      get 
      {
            return UserPwd; 
      }
      set 
      {
            UserPwd = value; 
      }
}

```

Actually why use these things. For global access, or is there some other reason for this type of thing?


","They are just accessors and mutators. That's how properties are implemented in C#


In C# 3 you can use auto-implemented properties like this:



```
public int MyProperty { get; set; }

```

This code is automatically translated by the compiler to code similar to the one you posted, with this code is easier to declare properties and they are ideal if you don't want to implement custom logic inside the `set` or `get` methods, you can even use a different accessor for the `set` method making the property immutable



```
public int MyProperty { get; private set; }

```

In the previous sample the `MyProperty` will be read only outside the class where it was declared, the only way to mutate it is by exposing a method to do it or just through the constructor of the class. This is useful when you want to control and make explicit the state change of your entity


When you want to add some logic to the properties then you need to write the properties manually implementing the `get` and `set` methods just like you posted:


Example implementing custom logic



```
private int myProperty;
public int MyProperty
{
   get
   {
       return this.myProperty;
   }
   set
   {
       if(this.myProperty <=5)
          throw new ArgumentOutOfRangeException(""bad user"");
       this.myProperty = value;
   }
}

```

"
"XML parsing error with Python Script with pythonanywhere (but not on local machine)
I'm running a flask app with python, part of which uses XML data retrieved from a third-party API. I use minidom to parse the XML within the python script.


**Relevant python code:**



```
from xml.dom import minidom
import requests

usa_xml = requests.get(URL_HERE)
usa_parsed = minidom.parseString(usa_xml.content)

```

The script goes on to locate and display values from the XML. Running the python script on my local machine, everything works as it should. Having put a repository up on **pythonanywhere**, the parsing fails with the same XML data.


**Error traceback:**



```
Traceback (most recent call last):
  File ""/usr/local/lib/python3.4/dist-packages/flask/app.py"", line 1817, in wsgi_app
    response = self.full_dispatch_request()
  File ""/usr/local/lib/python3.4/dist-packages/flask/app.py"", line 1477, in full_dispatch_request
    rv = self.handle_user_exception(e)
  File ""/usr/local/lib/python3.4/dist-packages/flask/app.py"", line 1381, in handle_user_exception
    reraise(exc_type, exc_value, tb)
  File ""/usr/local/lib/python3.4/dist-packages/flask/_compat.py"", line 33, in reraise
    raise value
  File ""/usr/local/lib/python3.4/dist-packages/flask/app.py"", line 1475, in full_dispatch_request
    rv = self.dispatch_request()
  File ""/usr/local/lib/python3.4/dist-packages/flask/app.py"", line 1461, in dispatch_request
    return self.view_functions[rule.endpoint](**req.view_args)
  File ""/home/jshorty/OwlWire/owlwire.py"", line 65, in select
    usa_parsed = minidom.parseString(usa_xml.content)
  File ""/usr/lib/python3.4/xml/dom/minidom.py"", line 1970, in parseString
    return expatbuilder.parseString(string)
  File ""/usr/lib/python3.4/xml/dom/expatbuilder.py"", line 925, in parseString
    return builder.parseString(string)
  File ""/usr/lib/python3.4/xml/dom/expatbuilder.py"", line 223, in parseString
    parser.Parse(string, True)
xml.parsers.expat.ExpatError: no element found: line 29, column 7    

```

Since it's failing with the same XML that would work otherwise, it doesn't seem like this is an issue with the XML itself. I'm stumped at where to start looking for the problem- I'm using all the default modules pre-installed on pythonanywhere, could this be an issue of different versions of minidom?


**The error is always at line 29, column 7, so here is a link to one instance of the XML I'm accessing:** <http://ebird.org/ws1.1/data/obs/region_spp/recent?rtype=country&r=US&sci=surnia%20ulula&back=30&maxResults=1&includeProvisional=true>


","My guess is that you're using a free account. Free accounts on PythonAnywhere have restricted Internet access, you can only access sites that are on the whitelist:


<https://www.pythonanywhere.com/wiki/403ForbiddenError>


<https://www.pythonanywhere.com/whitelist/>


You'll see that if you adjust your code to do a:



```
usa_xml = requests.get(URL_HERE)
print(usa_xml)

```

You'll probably see a



```
<Response [403]> 

```

403 being forbidden.


We (the PythonAnywhere team) are usually happy to add sites with a public API to the whitelist. ebird.com looks fine, I'll see if I can get that added. For anyone else with a similar request, don't hesitate to get in touch with us if you see a 403!


"
"ef-core load collection property of nested tph inherited member
Given the following class structure



```
 public class Parent 
    {
        public Guid Id { get; 
        public List<BaseChild> Children { get; set; }
    }

 public abstract class BaseChild
    {
        public int Id { get; set; }
        public string ChildName { get; set; }
    }

public class NormalChild : BaseChild
    {
         public DateTime BirthDate { get; set; }
    }

public class RichChild : BaseChild
    {
        public List<OffshoreAccount> OffshoreAccounts { get; set; }
    }

public class OffshoreAccount 
    {
        public string AccountNumber { get; set; }
        public AccountInfo AccountInfo { get; set; }
    }

```

What is the best way to query parent data to include information about the children's offshore accounts?. I came up with the solution below, using ef-core's explicit loading, but it just doesn't feel right. Is there a more elegant solution?



```
var parent = Context.Set<Parent>()
    .Where(o => o.Id == Guid.Parse(parentId))
    .Include(o => o.Children)
    .SingleOrDefault();

foreach (var child in parent.Children.OfType<RichChild>())
    {
        Context.Entry<RichChild>(child).Collection(f => f.OffshoreAccounts).Load();
        foreach (var account in child.OffshoreAccounts)
            {
                 Context.Entry<OffshoreAccount>(account).Reference(f => f.AccountInfo).Load();
            }
     }

```

","**Update (EF Core 2.1+):**


Starting with v2.1, EF Core native supports [Include on derived types](https://learn.microsoft.com/en-us/ef/core/querying/related-data#include-on-derived-types) through C# cast or `as` operators.


e.g



```
.Include(e => e.Children)
    .ThenInclude(e => ((RichChild)e).OffshoreAccounts)
        .ThenInclude(e => e.AccountInfo)

```

or



```
.Include(e => e.Children)
    .ThenInclude(e => (e as RichChild).OffshoreAccounts)
        .ThenInclude(e => e.AccountInfo)

```

The documentation claims that the `string` overload of `Include` coudld also be used, e.g. according to it



```
.Include(e => ""Children.OffshoreAccounts.AccountInfo"")

```

should also work, but it doesn't (checked up to v3.1.4).


**Original:**


Currently there is no way to accomplish that in the parent query, but the explicit loading can be improved by using a combination of `Entry`, `Collection`, `Query`, `Include` / `ThenInclude` and `Load` calls:



```
var parent = Context.Set<Parent>()
    .Where(o => o.Id == Guid.Parse(parentId))
    .Include(o => o.Children)
    .SingleOrDefault();

Context.Entry(parent).Collection(e => e.Children)
    .Query().OfType<RichChild>()
    .Include(e => e.OffshoreAccounts)
        .ThenInclude(e => e.AccountInfo)
    .Load();

```

"
"Can the \_\_VIEWSTATE and \_\_EVENTVALIDATION be use for research in order to breach a web app?
I'm learning ASP.NET now and I am a bit confused by the \_\_VIEWSTATE and \_\_EVENTVALIDATION.


1. Would it be possible to read the value those two items to learn about the internals of the app and possibly manipulate it. For eg. people write that the \_\_VIEWSTATE contains information about the properties of elements that are not sent back through the POSTBACK, like a label for example. Wouldn't it be possible then to manipulate the value of labels in an app to make it display wrong information?
2. Would it be possible to change the value of \_\_VIEWSTATE with a much larger value so that when it is posted back to the server it adds serious overhead to decompressing and/or decripting the information and thus basically making a DDOS?


","1. Yes it is possible to read the values of the viewstate. It is base64 encoded which does not mean it is encrypted so to read its values all you need to do is convert it from base64 to UTF-8 and you will be able to read its contents. [Try it out here for yourself](http://lachlankeown.blogspot.com/2008/05/online-viewstate-viewer-decoder.html) Each control is listed and several of their properties. In regards to manipulating the contents, this is possible but difficult as the contents are validated before being processed on the server end.
2. Yes its possible, if your site is targeted for an attack and a huge number of large requests were sent with large viewstates then it will have a corresponding effect on the server.


Take a look at the following:


- [Denial of service attacks](http://www.nikhilk.net/WebResourcesDenialOfServiceAttack.aspx)
- [Viewstate info](http://wiki.asp.net/page.aspx/58/viewstate/)
- [Truly understanding viewstate](http://weblogs.asp.net/infinitiesloop/archive/2006/08/03/Truly-Understanding-Viewstate.aspx)


"
"What is the theorical foundation for scikit-learn dummy classifier?
By the [documentation](http://scikit-learn.org/stable/modules/generated/sklearn.dummy.DummyClassifier.html) I read that a dummy classifier can be used to test it against a classification algorithm.



> 
> This classifier is useful as a simple baseline to compare with other
>  (real) classifiers. Do not use it for real problems.
> 
> 
> 


What does the dummy classifier do when it uses the stratified aproach. I know that the docummentation says that:



> 
> generates predictions by respecting the training set’s class
>  distribution.
> 
> 
> 


Could anybody give me a more theorical explanation of why this is a proof for the performance of the classifier?.


","The dummy classifier gives you a measure of ""baseline"" performance--i.e. the success rate one should expect to achieve even if simply guessing. 


Suppose you wish to determine whether a given object possesses or does not possess a certain property. If you have analyzed a large number of those objects and have found that 90% contain the target property, then guessing that every future instance of the object possesses the target property gives you a 90% likelihood of guessing correctly. Structuring your guesses this way is equivalent to using the `most_frequent` method in the documentation you cite.


Because many machine learning tasks attempt to increase the success rate of (e.g.) classification tasks, evaluating the baseline success rate can afford a floor value for the minimal value one's classifier should out-perform. In the hypothetical discussed above, you would want your classifier to get more than 90% accuracy, because 90% is the success rate available to even ""dummy"" classifiers. 


If one trains a dummy classifier with the `stratified` parameter using the data discussed above, that classifier will predict that there is a 90% probability that each object it encounters possesses the target property. This is different from training a dummy classifier with the `most_frequent` parameter, as the latter would guess that *all* future objects possess the target property. Here's some code to illustrate:



```
from sklearn.dummy import DummyClassifier
import numpy as np

two_dimensional_values = []
class_labels           = []

for i in xrange(90):
    two_dimensional_values.append( [1,1] )
    class_labels.append(1)

for i in xrange(10):
    two_dimensional_values.append( [0,0] )
    class_labels.append(0)

#now 90% of the training data contains the target property
X = np.array( two_dimensional_values )
y = np.array( class_labels )

#train a dummy classifier to make predictions based on the most_frequent class value
dummy_classifier = DummyClassifier(strategy=""most_frequent"")
dummy_classifier.fit( X,y )

#this produces 100 predictions that say ""1""
for i in two_dimensional_values:
    print( dummy_classifier.predict( [i]) )

#train a dummy classifier to make predictions based on the class values
new_dummy_classifier = DummyClassifier(strategy=""stratified"")
new_dummy_classifier.fit( X,y )

#this produces roughly 90 guesses that say ""1"" and roughly 10 guesses that say ""0""
for i in two_dimensional_values:
    print( new_dummy_classifier.predict( [i]) )

```

"
"Saving any file to in the database, just convert it to a byte array?
Is converting a file to a byte array the best way to save ANY file format to disk or database var binary column?


So if someone wants to save a .gif or .doc/.docx or .pdf file, can I just convert it to a bytearray UFT8 and save it to the db as a stream of bytes?


","Since it's not mentioned what database you mean I'm assuming SQL Server. Below solution works for both 2005 and 2008.


You have to create table with `VARBINARY(MAX)` as one of the columns. In my example I've created Table `Raporty` with column `RaportPlik` being `VARBINARY(MAX)` column.


**Method to put `file` into database from `drive`**:



```
public static void databaseFilePut(string varFilePath) {
    byte[] file;
    using (var stream = new FileStream(varFilePath, FileMode.Open, FileAccess.Read)) {
        using (var reader = new BinaryReader(stream)) {
            file = reader.ReadBytes((int) stream.Length);       
        }          
    }
    using (var varConnection = Locale.sqlConnectOneTime(Locale.sqlDataConnectionDetails))
    using (var sqlWrite = new SqlCommand(""INSERT INTO Raporty (RaportPlik) Values(@File)"", varConnection)) {
        sqlWrite.Parameters.Add(""@File"", SqlDbType.VarBinary, file.Length).Value = file;
        sqlWrite.ExecuteNonQuery();
    }
}

```

**This method is to get `file` from database and save it on `drive`**:



```
public static void databaseFileRead(string varID, string varPathToNewLocation) {
    using (var varConnection = Locale.sqlConnectOneTime(Locale.sqlDataConnectionDetails))
    using (var sqlQuery = new SqlCommand(@""SELECT [RaportPlik] FROM [dbo].[Raporty] WHERE [RaportID] = @varID"", varConnection)) {
        sqlQuery.Parameters.AddWithValue(""@varID"", varID);
        using (var sqlQueryResult = sqlQuery.ExecuteReader())
            if (sqlQueryResult != null) {
                sqlQueryResult.Read();
                var blob = new Byte[(sqlQueryResult.GetBytes(0, 0, null, 0, int.MaxValue))];
                sqlQueryResult.GetBytes(0, 0, blob, 0, blob.Length);
                using (var fs = new FileStream(varPathToNewLocation, FileMode.Create, FileAccess.Write)) 
                    fs.Write(blob, 0, blob.Length);
            }
    }
}

```

**This method is to get `file` from database and put it as `MemoryStream`**:



```
public static MemoryStream databaseFileRead(string varID) {
    MemoryStream memoryStream = new MemoryStream();
    using (var varConnection = Locale.sqlConnectOneTime(Locale.sqlDataConnectionDetails))
    using (var sqlQuery = new SqlCommand(@""SELECT [RaportPlik] FROM [dbo].[Raporty] WHERE [RaportID] = @varID"", varConnection)) {
        sqlQuery.Parameters.AddWithValue(""@varID"", varID);
        using (var sqlQueryResult = sqlQuery.ExecuteReader())
            if (sqlQueryResult != null) {
                sqlQueryResult.Read();
                var blob = new Byte[(sqlQueryResult.GetBytes(0, 0, null, 0, int.MaxValue))];
                sqlQueryResult.GetBytes(0, 0, blob, 0, blob.Length);
                //using (var fs = new MemoryStream(memoryStream, FileMode.Create, FileAccess.Write)) {
                memoryStream.Write(blob, 0, blob.Length);
                //}
            }
    }
    return memoryStream;
}

```

**This method is to put `MemoryStream` into database:**



```
public static int databaseFilePut(MemoryStream fileToPut) {
        int varID = 0;
        byte[] file = fileToPut.ToArray();
        const string preparedCommand = @""
                    INSERT INTO [dbo].[Raporty]
                               ([RaportPlik])
                         VALUES
                               (@File)
                        SELECT [RaportID] FROM [dbo].[Raporty]
            WHERE [RaportID] = SCOPE_IDENTITY()
                    "";
        using (var varConnection = Locale.sqlConnectOneTime(Locale.sqlDataConnectionDetails))
        using (var sqlWrite = new SqlCommand(preparedCommand, varConnection)) {
            sqlWrite.Parameters.Add(""@File"", SqlDbType.VarBinary, file.Length).Value = file;

            using (var sqlWriteQuery = sqlWrite.ExecuteReader())
                while (sqlWriteQuery != null && sqlWriteQuery.Read()) {
                    varID = sqlWriteQuery[""RaportID""] is int ? (int) sqlWriteQuery[""RaportID""] : 0;
                }
        }
        return varID;
    }

```

"
"Dynamically Running Mocha Tests
I'm trying to run a series of tests dynamically. I have the following setup but it doesn't seem to run and I'm not getting any errors:



```
import Mocha from 'mocha';
const Test = Mocha.Test;
const Suite = Mocha.Suite;
const mocha = new Mocha();
for (let s in tests) {
  let suite = Suite.create(mocha.suite, s);
  tests[s].forEach((test) => {
    console.log('add test', test.name)
    suite.addTest(new Test(test.name), () => {
      expect(1+1).to.equal(2);
    });
  });
}
mocha.run();

```

The `tests` I'm running look like this:



```
{ todo: 
  [ { name: 'POST /todos',
      should: 'create a new todo',
      method: 'POST',
      endpoint: '/todos',
      body: [Object] } ] }

```

(though at this point my test is just trying to check a basic expect)


Based on the console.logs the iteration seems fine and it appears to be adding the tests, so I'm confident in the flow of operations, I just can't get any execution or errors.


","You have to pass the test function to the `Test` constructor, not to `suite.addTest`. So change your code to add your tests like this:



```
suite.addTest(new Test(test.name, () => {
    expect(1+1).to.equal(2);
}));

```

Here is the entire code I'm running, adapted from your question:



```
import Mocha from 'mocha';
import { expect } from 'chai';
const Test = Mocha.Test;
const Suite = Mocha.Suite;
const mocha = new Mocha();

var tests = { todo:
  [ { name: 'POST /todos',
      should: 'create a new todo',
      method: 'POST',
      endpoint: '/todos',
      body: [Object] } ] };

for (let s in tests) {
  let suite = Suite.create(mocha.suite, s);
  tests[s].forEach((test) => {
      console.log('add test', test.name);
      suite.addTest(new Test(test.name, () => {
          expect(1+1).to.equal(2);
      }));
  });
}
mocha.run();

```

When I run the above with `node_modules/.bin/babel-node test.es6`, I get the output:



```
  todo
    ✓ POST /todos


  1 passing (5ms)

```

"
"How to display readme.md file on profile in Github?
I have created a `README.md` file in my `username.github.io` repository but it doesn't appear on my profile page:


[![enter image description here](https://i.stack.imgur.com/HTr1Y.png)](https://i.stack.imgur.com/HTr1Y.png)


","[Profile READMEs](https://docs.github.com/en/account-and-profile/setting-up-and-managing-your-github-profile/customizing-your-profile/managing-your-profile-readme) must be [located in a public repository named after your GitHub user name](https://docs.github.com/en/account-and-profile/setting-up-and-managing-your-github-profile/customizing-your-profile/managing-your-profile-readme#prerequisites). This is different from your GitHub Pages User site.


Assuming your GitHub username is `username`:


- Your GitHub Pages User Site should be in a repository named `username.github.io`. You already have one of these.
- The repository for your profile `README.md` should *just* be named `username`. Create a new repository with that name and add your `README.md` there.


"
"How ro re-render everything in React Native
My app has it's theme (colors and styles) delivered via a REST API. Whenever a theme is changed on the server the app gets notified via socket communication so it can get the new theme via REST.


The theme is fetched in a Redux Saga and the colors are saved to both the Redux store (to enable persisting to disk) as well as a ""global"" singleton object which I can reach from everywhere, i.e. even non connected components.


When a new theme arrives I want to re-render every element in the app. I have managed to re-render all store-connected components in a hacky way - by injecting a dummy property which changes when new a theme is changed.


Is there a way to `forceUpdate` the whole app?


Problem areas:


- dumb components (not aware of theme changes)
- really smart components (aware of the store, but knows not to re-render in vain)


Here is an example of a hacky force-re-render of react-navigation:



```
myAppContainer.render() {

  <AppRouter
    screenProps={this.props.styling}
  />
}

```

I.e. my AppContainer gets `this.props.styling` via `mapStateToProps`. The wrapped `StackNavigator` gets this via `screenProps` which forces the navigator to re-render when the store gets new styling data.


I don't want to continue on this hacky path. Instead I'd like a `forceUpdateEverything()` function which I could call from my `AppContainer`. Is there such a thing?




---


**Edit to comment the answer from @bennygenel:**


I think what Benny describes is essentially what I did. A change in this.props.styling triggers a re-render. 


But I'd have to implement this into all components and ""hack"" react-navigation via its screenProps as I described. This is what I'd hoped to avoid, but I guess I'll have to take the long way.. 


I'm using immutable.js, so sending in the complete style instead of just the theme name is no big deal as it's only a pointer which is speedy to compare with it's former value to look for changes.


I didn't get forceUpdate() to work other than on the component it's called on. I assume it doesn't get propagated recursively through all children.


","I think the simplest and most performance way of doing this rather then holding all styling object in your state and redux you can just hold some sort of theme identifier. This way when any changes happen for the theme that your components use can be applied.


In react there is a method called [`forceUpdate`](https://reactjs.org/docs/react-component.html#forceupdate). Although it sounds like the thing you are looking for, its not a really good practice to use it.


*From the react docs;*



> 
> By default, when your component’s state or props change, your
>  component will re-render. If your `render()` method depends on some
>  other data, you can tell React that the component needs re-rendering
>  by calling `forceUpdate()`.
> 
> 
> Calling `forceUpdate()` will cause render() to be called on the
>  component, skipping `shouldComponentUpdate()`. This will trigger the
>  normal lifecycle methods for child components, including the
>  `shouldComponentUpdate()` method of each child. React will still only
>  update the DOM if the markup changes.
> 
> 
> **Normally you should try to avoid all uses of `forceUpdate()` and only read from this.props and this.state in `render()`.**
> 
> 
> 


I think what you should do is to create a theme logic in your app and change the theme accordingly.


**Example**



```
// styles.js
const styles = {
  button: {
    blue: {
      backgroundColor: 'blue'
    },
    red: {
      backgroundColor: 'red'
    },
    green: {
      backgroundColor: 'green'
    }    
  }
};

export default styles;

// Dumb component example
import styles from './styles';

const Button = (props) => {
  return <Button {...props} style={[props.style, styles.button[props.theme]]} />
};

// Complex component example
import styles from './styles';

export default Button extends React.Component {
  constructor(props) {
    super(props);
  }
  render() {
    const {style, theme} = this.props;
    return <Button {...this.props} style={[style, styles.button[theme]]} />
  }
}

// usage in another component
import React, { Component } from 'react';
import { Text, View, StyleSheet } from 'react-native';
import Button from './components/Button'; // Custom button component
import styles from './styles';

export default App extends React.Component {
  constructor(props) {
    super(props);
  }
  render() {
    return(
      <View style={styles.container}>
        <Text style={styles.paragraph}>
          Change code in the editor and watch it change on your phone!
          Save to get a shareable url. You get a new url each time you save.
        </Text>
        <Button theme=""blue"" {/* other Button component props */} />
      </View>
    );
  }
}

```

In my example I use styles defined as hard-coded object. You can add your API call and socket communication logic to it. 


In example rather than passing the complete style object from one component/screen to another I can just simply pass the theme name and any change on `theme` prop will force the component to fetch new style object from `styles` constant.


Theme doesn't need to be set for every single object. You can use a higher property in your object.


**Example**



```
const styles = {
  id5248698745: {
    button: {
      backgroundColor: 'green'
    },
    label: {
      color: 'yellow',
      fontSize: 18
    },
    paragraph: {
      color: 'red',
      fontSize: 19
    },
    headings: {
      color: 'wihte',
      fontSize: 24
    }
  }
};

// ...
render() {
    const {style, theme} = this.props;
    return <Button {...this.props} style={[style, styles[theme].button]} />
}

// ...
// <Button theme={this.props.themeId} {/* other Button component props */} />

```

"
"R programming: plyr how to count values from a column with ddply
I would like to summarize the pass/fail status for my data as below. In other words, I would like to tell the number of pass and fail cases for each product/type.



```
library(ggplot2)
library(plyr)
product=c(""p1"",""p1"",""p1"",""p1"",""p1"",""p1"",""p1"",""p1"",""p1"",""p1"",""p1"",""p1"",""p2"",""p2"",""p2"",""p2"",""p2"",""p2"",""p2"",""p2"",""p2"",""p2"",""p2"",""p2"")
type=c(""t1"",""t1"",""t1"",""t1"",""t1"",""t1"",""t2"",""t2"",""t2"",""t2"",""t2"",""t2"",""t1"",""t1"",""t1"",""t1"",""t1"",""t1"",""t2"",""t2"",""t2"",""t2"",""t2"",""t2"")
skew=c(""s1"",""s1"",""s1"",""s2"",""s2"",""s2"",""s1"",""s1"",""s1"",""s2"",""s2"",""s2"",""s1"",""s1"",""s1"",""s2"",""s2"",""s2"",""s1"",""s1"",""s1"",""s2"",""s2"",""s2"")
color=c(""c1"",""c2"",""c3"",""c1"",""c2"",""c3"",""c1"",""c2"",""c3"",""c1"",""c2"",""c3"",""c1"",""c2"",""c3"",""c1"",""c2"",""c3"",""c1"",""c2"",""c3"",""c1"",""c2"",""c3"")
result=c(""pass"",""pass"",""fail"",""pass"",""pass"",""pass"",""fail"",""pass"",""fail"",""pass"",""fail"",""pass"",""fail"",""pass"",""fail"",""pass"",""pass"",""pass"",""pass"",""fail"",""fail"",""pass"",""pass"",""fail"")
df = data.frame(product, type, skew, color, result)

```

The following cmd returns the total number of pass+fail cases but I want separate columns for pass and fail



```
dfSummary <- ddply(df, c(""product"", ""type""), summarise, N=length(result))

```

Result is:



```
        product type N
 1      p1      t1   6
 2      p1      t2   6
 3      p2      t1   6
 4      p2      t2   6

```

The desireable result would be 



```
         product type Pass Fail
 1       p1      t1   5    1
 2       p1      t2   3    3
 3       p2      t1   4    2
 4       p2      t2   3    3

```

I have attempted somthing like: 



```
 dfSummary <- ddply(df, c(""product"", ""type""), summarise, Pass=length(df$product[df$result==""pass""]), Fail=length(df$product[df$result==""fail""]) )

```

but obviously it’s wrong since the results are the grand totatl for fail and pass.


Thanks in advance for your advice !
Regards,
Riad.


","Try:



```
dfSummary <- ddply(df, c(""product"", ""type""), summarise, 
                   Pass=sum(result==""pass""), Fail=sum(result==""fail"") )

```

Which gives me result:



```
  product type Pass Fail
1      p1   t1    5    1
2      p1   t2    3    3
3      p2   t1    4    2
4      p2   t2    3    3

```

Explanation:


1. You are giving the data set, `df` to the `ddply` function.
2. `ddply` is splitting on the variables, ""product"" and ""type""
	- This results in `length(unique(product)) * length(unique(type))` pieces (i.e. subsets of the data `df`) split on every combination of the two variables.
3. With each of the pieces, `ddply` applies some function that you provide. In this case, you count the number of `result==""pass""` and `result==""fail""` there are.
4. Now `ddply` is left with some results for each piece, namely the variables you split on (product and type) and the results you requested (Pass and Fail).
5. It combines all of the pieces together and returns it


"
"Algorithm for counting common group memberships with big data
I need to write a program that counts the number of times two users are in the same group. The users are given by username and groups by id.
For example, with the input (stored in a text-file):



```
john 32
john 21
jim 21
jim 32
bob 32

```

I want the result:



```
john-jim 2 
john-bob 1
jim-bob 1

```

This sounds trivial. But the problem is: I have 1,8 million groups and 300,000 users. And a lot of memberships (I'm expecting at least an average of 50 per user, possibly much more). This means a HUGE amount of data and processing.


I've written 5 different programs doing this, none of which has been able to cut the amount of data: It was too slow as an PostgreSQL query. Too memory consuming running in a Map in Java working memory (first heap space, after optimization I got the rare ""GC overhead limit exceeded""). Too slow to write continuously to database from Java (even when optimized using batch-queries). Growing increasingly desperate, I've tried some more exotic things, like writing all the pairs to an array, then sorting them (O(n log (n))) and then counting them peu à peu. But it was still way too much data to store in memory.


Any ideas on an algorithm for doing this? Or is it impossible?


","
An RDBMS is specialized in operations like sorting. Doing this outside the DB will hardly ever come even close in performance. Do it with SQL!


This would do the job (simplified in update):



```
SELECT t1.usr || '-' || t2.usr, count(*) AS ct
FROM   usr_grp t1
JOIN   usr_grp t2 USING (grp_id) 
WHERE  t2.usr > t1.usr   -- prevent dupes and get sorted pair
GROUP  BY t1.usr, t2.usr;

```

Depending on how many overlaps you have, this potentially produces a HUGE amount of rows, as you said. So this is never going to be fast.


*Raises the question*: What's the purpose of producing millions of rows that nobody can process? Are you sure, the operation makes sense to begin with?


To make it *faster*, you could ..


- **Upgrade!** [PostgreSQL 8.4 is rather outdated by now](http://www.postgresql.org/support/versioning/). In particular, PostgreSQL 9.2 had its focus on big data. You can expect *much* better performance for a job like this.  

And *nobody* should be running 8.4.0. For security reasons alone, but you are missing out on lot of bug-fixes, too. The current point-release is 8.4.17. I quote the linked web-site:



> 
> *We always recommend that all users run the latest available minor
>  release for whatever major version is in use.*
> 
> 
> 


- use an **`integer`** as surrogate key for users, so you deal with integers only in `usr_grp`. Makes table and indexes smaller and processing faster. If the n:m table (`usr_grp`) has a much bigger cardinality than the table `usr`, this should be faster, even if it means additional joins.




```
SELECT u1.usr  || '-' || u2.usr, count(*) AS ct
FROM   usr_grp t1
JOIN   usr_grp t2 USING (grp_id) 
JOIN   usr u1 ON t1.usr_id = u1.usr_id
JOIN   usr u2 ON t2.usr_id = u2.usr_id
WHERE  t2.usr_id > t1.usr_id
GROUP  BY u1.usr_id, u2.usr_id;

```

- Create a **multi-column index** (if you don't have it yet).  

`grp_id` must come first. [Why does this matter?](https://dba.stackexchange.com/questions/27481/is-a-composite-index-also-good-for-queries-on-the-first-field/27493#27493)




```
    CREATE INDEX usr_grp_gu_idx ON usr_grp(grp_id, usr_id);

```

- Put **a lot of RAM** into your machine and increase the settings for [`work_mem`](http://www.postgresql.org/docs/current/interactive/runtime-config-resource.html#GUC-WORK-MEM) and [`shared_buffers`](http://www.postgresql.org/docs/current/interactive/runtime-config-resource.html#GUC-SHARED-BUFFERS).


### Test case


I took the numbers [@OldCurmudgeon reported](https://stackoverflow.com/a/15831671/939860) for his test case and created a comparable test case in PostgreSQL.


[-> **SQLfiddle** demo.](http://www.sqlfiddle.com/#!12/0b72c/2)


~ **250 ms** in this public test database.  

The result is not ordered (no `ORDER BY`) since this hasn't been specified.  

As compared to **2.5 minutes**, [reported below](https://stackoverflow.com/a/15831671/939860). Factor 600.


"
"Play Framework: Dependency Inject Action Builder
since Play Framework 2.4 there is the possibility to use dependency injection (with Guice).


Before I used objects (for example `AuthenticationService`) in my ActionBuilders:



```
object AuthenticatedAction extends ActionBuilder[AuthenticatedRequest] {
  override def invokeBlock[A](request: Request[A], block: (AuthenticatedRequest[A]) => Future[Result]): Future[Result] = {
    ...
    AuthenticationService.authenticate (...)
    ...
  }
}

```

Now `AuthenticationService` is not an object anymore, but a class. How can I still use the `AuthenticationService` in my `ActionBuilder`? 


","Define your action builders inside a trait with the authentication service as an abstract field. Then mix them into your controllers, into which you inject the service. For example:



```
trait MyActionBuilders {
  // the abstract dependency
  def authService: AuthenticationService

  def AuthenticatedAction = new ActionBuilder[AuthenticatedRequest] {
    override def invokeBlock[A](request: Request[A], block(AuthenticatedRequest[A]) => Future[Result]): Future[Result] = {
      authService.authenticate(...)
      ...
    }
  }
}

```

and the controller:



```
@Singleton
class MyController @Inject()(authService: AuthenticationService) extends Controller with MyActionBuilders {    
  def myAction(...) = AuthenticatedAction { implicit request =>
    Ok(""authenticated!"")
  }
}

```

"
"I can't bind events to elements



```
const color = ['red', 'orange', 'yellow', 'green', 'blue', 'navy', 'purple'];
let a = [];
for (let i = 0; i < color.length; i++) {
  a[i] = document.createElement(""input"");
  a[i].type = 'button';
  a[i].id = 'b' + (i + 1);
  a[i].value = color[i];
  a[i].addEventListener('click', function() {
    alert('color');
  })
  document.body.appendChild(a[i]);
  document.body.innerHTML += ""<br>""
  console.log(a[0].innerHTML);
}
```






It seems that the listener is not getting bound despite the `addEventListener`. What is the problem?


","The problem is that, when concatenating with the `innerHTML` of a container (for example, with your `document.body.innerHTML += ""<br>""`), the container will be *emptied* and then re-parsed with the new HTML string. If you previously attached a listener to an element in the container, that listener will not be in the HTML string, so it will not transfer over to the *new* element in the same position.





```
const div1 = document.querySelector('#somediv');
document.body.innerHTML += '';
const div2 = document.querySelector('#somediv');

console.log(div1 === div2);
// False, the container's contents were re-parsed, the new div is different!
```


```
<div id=""somediv""></div>
```






Either append your `br` using the same `appendChild` method you're using for the `a[i]`:





```
const color = ['red', 'orange', 'yellow', 'green', 'blue', 'navy', 'purple'];
let a = [];
for (let i = 0; i < color.length; i++) {
  a[i] = document.createElement(""input"");
  a[i].type = 'button';
  a[i].id = 'b' + (i + 1);
  a[i].value = color[i];
  a[i].addEventListener('click', function() {
    alert('color');
  })
  document.body.appendChild(a[i]);
  document.body.appendChild(document.createElement('br'));
}
```






Or use `insertAdjacentHTML` instead, which can act similarly to `.innerHTML +=`, but unlike `.innerHTML +=`, does *not* re-create all elements in the container:





```
const color = ['red', 'orange', 'yellow', 'green', 'blue', 'navy', 'purple'];
let a = [];
for (let i = 0; i < color.length; i++) {
  a[i] = document.createElement(""input"");
  a[i].type = 'button';
  a[i].id = 'b' + (i + 1);
  a[i].value = color[i];
  a[i].addEventListener('click', function() {
    alert('color');
  })
  document.body.appendChild(a[i]);
  document.body.insertAdjacentHTML('beforeend', '<br>');
}
```






"
"How to format XSD documentation?
**The Problem:**


I have a xsd schema with included documentation, like:



```
<xs:element name=""Tag"" type=""XTag"">
  <xs:annotation>
    <xs:documentation>
      Do like this:
       - foo
       - bar
    </xs:documentation>
  </xs:annotation>
</xs:element>

```

When I view the documentation in Eclipse, the white-spaces are truncated to a single space, so I end up with:



```
Do like this: - foo - bar

```

This naturally limits readability when documentation is more than a little note. In javadoc for instance, one can one html tags to format the documentation


**The question:**


Is there any way to format the documentation, at least just adding newlines? 


**Some details:**


I'm doing everything in Eclipse. The XSD i write is added to `preferences -> XML -> XML catalog`, so I can get content assist and view documentation in the XML editor.


It's an internal tool and the only foreseeable place the documentation will be viewed is through Eclipse in the way described above. So if it works in eclipse, it's good enough :)


","Embed your documentation into `<![CDATA[ ]]>` and use html expressions. Sample below should be rendered correctly.



```
<xs:annotation>
    <xs:documentation>
    <![CDATA[
Lorem ipsum dolor sit amet, consectetur adipiscing elit, sed do eiusmod tempor incididunt ut labore et dolore magna aliqua.<br/>
Ut enim ad minim veniam, quis nostrud exercitation ullamco laboris nisi ut aliquip ex ea commodo consequat. 
<p><b>Duis aute</b> irure dolor in reprehenderit in voluptate velit esse cillum dolore eu fugiat nulla pariatur. </p>
<br/>
<b>Excepteur sint occaecat cupidatat non proident,<br/>
<ul>
    <li>sunt in culpa qui officia</li> 
    <li>deserunt mollit anim</li>
    <li>id est laborum.</li>
</ul>
    ]]>
    </xs:documentation>
</xs:annotation>

```

Tested with Spring Tool Suite v3.6.4 (Eclipse 4.4.2)


"
"How to use Vue Router from Vuex state?
In my components I've been using:


`this.$router.push({ name: 'home', params: { id: this.searchText }});`


To change route. I've now moved a method into my Vuex actions, and of course `this.$router` no longer works. Nor does `Vue.router`. So, how do I call router methods from the Vuex state, please?


","I'm assuming [vuex-router-sync](https://github.com/vuejs/vuex-router-sync) won't help here as you need the router instance.


Therefore although this doesn't feel ideal you could set the instance as a global within webpack, i.e.



```
global.router = new VueRouter({
  routes
})

const app = new Vue({
  router
  ...

```

now you should be able to: `router.push({ name: 'home', params: { id: 1234 }})` from anywhere within your app




---


As an alternative if you don't like the idea of the above you could return a Promise from your action. Then if the action completes successfully I assume it calls a mutation or something and you can `resolve` the Promise. However if it fails and whatever condition the redirect needs is hit you `reject` the Promise.


This way you can move the routers redirect into a component that simply catches the rejected Promise and fires the vue-router push, i.e.



```
# vuex
actions: {
  foo: ({ commit }, payload) =>
    new Promise((resolve, reject) => {
      if (payload.title) {
        commit('updateTitle', payload.title)
        resolve()
      } else {
        reject()
      }
    })

# component
methods: {
  updateFoo () {
    this.$store.dispatch('foo', {})
      .then(response => { // success })
      .catch(response => {
        // fail
        this.$router.push({ name: 'home', params: { id: 1234 }})
      })

```

"
"Use stopPropagation() in $(document).on(...)
I would add some ""Li"" tags in my page **dynamically**.


**Each ""Li"" have some codes like this:**



```
<li id=""file-1000"">
    <a target=""_blank"" href=""pack.jpg"" style=""display: block;"" class=""pirobox_gall first last"">
       <img alt="""" src=""pack_thumb.jpg"" class=""attached"">
       <i style=""display: none;"" class=""icon-close2"" id=""delIcon-1000""></i>
    </a>
</li>

```

**Result:**


![enter image description here](https://i.stack.imgur.com/H1VPO.jpg)




---


Events:


**1) I bind some events to the ""a"" and ""i"" tags like this:**



```
$(document)
    .on('mouseenter',
        '#file-1000',
        function () {
            $('#delIcon-1000').fadeIn(200);
        })
    .on('mouseleave',
        '#file-1000',
        function () {
            $('#delIcon-1000').fadeOut(200);
        })
    .on('click',
        '#delIcon-1000',
        function () {
            alert('Close btn is clicked');
        });

```

**2) An other hand, the ""a"" tag have other assigned events on its click event (by a jQuery image slide plugin: piroBox)**




---


So when I click on the image, I see a nice popup of my image (as a large picture) but when I click the Delete button I see the large image + ""Close btn is clicked"" message"" together.


**I want prevent ""a"" tag click event (parent event) when I click the Delete button (when I click it's child).**


I know using `stopPropagation()` is a solution but it does not work here and I don't know how can I use it in the `$(document).on(....)` method.




---


**Edit:**
I tried some answers but they failed and the result is like this:


![enter image description here](https://i.stack.imgur.com/UK3R1.jpg)


**Above Black Frame is >>** a popup box for showing larger picture (I want prevent opening this frame)


","### edit


# ***[`jsFiddle Demo`](http://jsfiddle.net/YZ7UW/1/)***


It was brought to my attention that the deprecated approach first listed will in fact not work here. That is 100% correct, sorry for missing that. In fact, that scenario seems rather complex! How can you, from the outside, manage to prevent a click action which takes place prior to yours which is also attached to another element? Well, I am not sure entirely what the ""best practice"" is. I came up with this approach though.


Iterate from the element up the dom tree looking for click events. Collect those click events. Iterate through all of the handlers collected, and then overload them to screen for your element if the event was based on your element.


*Initial Setup*



```
$(""#file-1000"").click(function(e){
 console.log(e);
 alert(""some other click""); 
});

$(document)
.on('mouseenter',
    $('#file-1000').selector,
    function () {
        $('#delIcon-1000').fadeIn(200);
    })
.on('mouseleave',
    $('#file-1000').selector,
    function () {
        $('#delIcon-1000').fadeOut(200);
    })
.on('click',
    $('#delIcon-1000').selector,
    function (e) {
        alert('Close btn is clicked');
        e.preventDefault();
    });

```

*Implementation of iteration as described above*



```
function ProtectClick(element){
 var handlers = [];
 var elCopy = element;
 do{
  var handler = jQuery._data( elCopy, ""events"" );
  if( handler != void 0 && handler.click != void 0){
   handlers.push({el: elCopy, handle: handler.click});
  }
  elCopy = elCopy.parentNode;
 }while( elCopy );

 for( var h in handlers ){
  (function(i){
   for( var cur in handlers[i].handle ){
    (function(n){
     var handlerCopy = handlers[i].handle[n].handler;
     var $sourced;
     if( handlers[i].handle[n].selector ){
      $sourced = $(handlers[i].handle[n].selector);
     }else{
      $sourced = $(handlers[i].el);
     }
     handlers[i].handle[n].handler = function(e){
      if( $sourced.is(element) || element != e.target )handlerCopy(e);
     };       
    })(cur);
   }
  })(h);
 }
}

```

*Call*



```
ProtectClick($('#delIcon-1000')[0]);

```



---


# *deprecated*


***[`jsFiddle Demo`](http://jsfiddle.net/arVYB/)***


I believe what you were looking for is `preventDefault`



```
$(document)
.on('mouseenter',
    '#file-1000',
    function () {
        $('#delIcon-1000').fadeIn(200);
    })
.on('mouseleave',
    '#file-1000',
    function () {
        $('#delIcon-1000').fadeOut(200);
    })
.on('click',
    '#delIcon-1000',
    function (e) {
        alert('Close btn is clicked');
        e.preventDefault();
    });

```

"
"What is the recommended approach to style a SlickGrid?
I am just starting to use SlickGrid and amazed by its quality. However, when it comes to styling, I did not find any docs or examples recommending an overall styling approach. There are options and APIs scattered in various places, but it's very difficult to extract a strategy out of those. Also the grid leverages jQuery UI themes. Unfortunately those are interfering with what I am trying to achieve. We have picked up jQuery UI only for the calendar widget along with the ui-darkness theme. This theme works perfectly fine for the calendar widget, but the grid needs to override every aspect of it.


Here's a jsFiddle that shows the look I am trying to achieve: <http://jsfiddle.net/nareshbhatia/3q6RD/>. Just for illustration, it uses a regular HTML table. However I would like to achieve the exact same styling using SlickGrid. The CSS in this jsFiddle is essentially the requirement I have from my visual designer, e.g.



```
#positions-table th {
    background-color: #505050;
    color: #eeeeee;
    text-shadow: none;
    font-size: 13px;
    height: 40px;
    line-height: 40px;
}

```

**Edit:** I also created a jsFiddle with a starter SlickGrid implementation: <http://jsfiddle.net/nareshbhatia/vJshY/>. As you can see, the ui-darkness theme has completely taken over!


","Within your second/last jsFiddle you can modify the CSS to have this code



```
.slick-header-column.ui-state-default {
    background:none ;
    background-color: #505050 ;
    color: #eeeeee;  
    border: none;  
    padding: 0;
    text-shadow: none;
    font-family: Arial, Verdana, Helvetica, sans-serif;
    font-size: 13px;
    height: 40px;
    line-height: 40px;    
}

.slick-row.ui-widget-content, .slick-cell {
    background: none;
    background-color: #eeeeee;
    color: #666666;
    border: none;
    border-bottom: solid 1px #ffffff;
    font-size: 14px;
    height: 60px;
    line-height: 60px;
    padding: 0 5px;
}

```

"
"Active Choices Reactive Reference Parameter in jenkins pipeline
I'm using the Active Choices Reactive Reference Parameter plugin in a dsl job here the code 



```
 parameters {
                  activeChoiceParam('choice1') {
                      description('select your choice')
                      choiceType('RADIO')
                      groovyScript {
                          script(""return['aaa','bbb']"")
                          fallbackScript('return [""error""]')
                      }
                  }
                  activeChoiceReactiveParam('choice2') {
                      description('select your choice')
                      choiceType('RADIO')
                      groovyScript {
                          script(""if(choice1.equals(""aaa"")){return ['a', 'b']} else {return ['aaaaaa','fffffff']}"")
                          fallbackScript('return [""error""]')
                      }
                      referencedParameter('choice1')
                  }

```

and it work's fine what i want now is how to use the activeChoiceReactiveParam in a jenkinsFile for pipeline job i did that :



```
properties(
    [
            [
                    $class              : 'ParametersDefinitionProperty',
                    parameterDefinitions: [
                            [
                                    $class     : 'ChoiceParameterDefinition',
                                    choices    : 'aaa\nbbb',
                                    description: 'select your choice : ',
                                    name       : 'choice1'
                            ]

```

but how can i add the choice2 parameter !!!


","Instead of using the **Active Choices Reactive Reference Parameter** i did below and it's working fine !!!



```
node('slave') {
    def choice1
    def choice2

    stage ('Select'){
        choice1 = input( id: 'userInput', message: 'Select your choice', parameters: [ [\$class: 'ChoiceParameterDefinition', choices: 'aa\nbb', description: '', name: ''] ])
        if(choice1.equals(""aa"")){
            choice2 = input( id: 'userInput', message: 'Select your choice', parameters: [ [\$class: 'ChoiceParameterDefinition', choices: 'yy\nww', description: '', name: ''] ])
        }else{
            choice2 = input( id: 'userInput', message: 'Select your choice', parameters: [ [\$class: 'ChoiceParameterDefinition', choices: 'gg\nkk', description: '', name: ''] ])
        }
    }
}

```

"
"Infinite Auto Scroll ListView with Scroll Speed Controlled
I have been working on a `ListView`idea where it keeps scrolling automatically with no user interaction and that is absolutely doable using the android APIs for instance ***smoothScrollToPositionFromTop***.


I have implemented `ListView BaseAdapter` where it load items forever (almost) to get a non stopping self repeated `ListView`.


What I want to achieve here is to keep my`ListView`scrolling forever with certain speed (slow) to make items clear and readable while scrolling down, I not sure yet if `ListView` is my best choice here.


below is a snippet of what I am trying to do. the result is good somehow but it's not smooth enough, I can feel the ListView flickers.


I need to improve **smoothness, efficiency and control the speed**



```
new Thread(new Runnable() {

    @Override
    public void run() {
        int listViewSize = mListView.getAdapter().getCount();

        for (int index = 0; index < listViewSize ; index++) {
            mListView.smoothScrollToPositionFromTop(mListViewA.getLastVisiblePosition() + 100, 0, 6000);
            try {
                // it helps scrolling to stay smooth as possible (by experiment)
                Thread.sleep(60);
            } catch (InterruptedException e) {

            }
        }
    }
}).start();

```

","I suggest, thath your adapter implemented in effective way.
so this code is just scrolls listview


you need to try another values of variables



```
final long totalScrollTime = Long.MAX_VALUE; //total scroll time. I think that 300 000 000 years is close enouth to infinity. if not enought you can restart timer in onFinish()

final int scrollPeriod = 20; // every 20 ms scoll will happened. smaller values for smoother

final int heightToScroll = 20; // will be scrolled to 20 px every time. smaller values for smoother scrolling

listView.post(new Runnable() {
                        @Override
                        public void run() {
                                new CountDownTimer(totalScrollTime, scrollPeriod ) {
                                    public void onTick(long millisUntilFinished) {
                                        listView.scrollBy(0, heightToScroll);
                                    }

                                public void onFinish() {
                                    //you can add code for restarting timer here
                                }
                            }.start();
                        }
                    });

```

"
"What happens when Git pull or push to different branch being in some other branch
I am familiar with GIT and using the same for versioning of my project.


I am having few queries which I would like to know. I googled but couldn't get the good answer.


So the thing is I am having **master, module1, feature1** branches.



```
master
 ---------- 
         | module1
         ----------
                  | feature1
                  ------------

```

module1 is branched from master and feature1 is branched from module1.


**Query 1** : Being in feature1 branch what if I work on few changes and commit and push the same to module1.



```
git add .
git commit -m ""Changes of feature1""
git push origin module1 //Being in feature1 branch

```

What happens here to the code of feature1 to module1 branch and how module1 branch takes it.


**My Understaing** : As per my understanding the feature1 changes will be pushed to module1 branch. And later I realised that I should have pushed to feature1 branch then I will push the same to feature1 branch and will checkout to module1 branch and revert the code which I had pushed recently.


**Query 2** : What if being in the feature1 branch and I pull the code of module1 in this branch by the following



```
git pull origin module1 //Being in feature1 branch

```

**My Understanding** : The changes of the module1 code will get merged to my feature1 branch and is same as the following in commands



```
git checkout moduel1
git pull origin module1
git checkout feature1
git merge module1

```

If there are any conflicts will be shown. Which I need to resolve.


Can anyone help me whether my understanding is correct or not. If not please help me to understand this concept properly. Thanks in advance.


","You have, I think, a few broad misconceptions about branch names and the use of `git pull`. Let me split this into several parts, and give you this executive summary overview to start with:


- `push`'s counterpart is not `pull`, it's `fetch`;
- `git pull` just runs `git fetch` followed by a second Git command, usually `git merge`, and I believe it's best for new Git users to avoid `git pull`, using instead the two separate commands;
- while `push` and `fetch` use names for transferring hash IDs, it's the *commits*, identified by hash IDs, that matter; and
- for `git merge` or `git rebase`, your current branch *does* matter. During push or fetch, the current branch does not matter, but if you use `git pull`, that runs `git merge` or `git rebase`, and *now* the current branch matters.


Now, let's dive into the details.


### A branch name is only a pointer to a (single) commit


Git is all about *commits*. Git would, in a sense, ""like it"" if us mere humans did not need branch names, and just talked about commits by hash ID all the time. I might ask you if you are using commit `95ec6b1b3393eb6e26da40c565520a8db9796e9f`, and you would say ""yes"" or ""no, but I have that one"" or ""no, and I have not heard of that one yet"".


You mention that:



> 
> module1 is branched from master and feature1 is branched from module1.
> 
> 
> 


but in Git's eyes, branches don't branch *from* another *branch*. Instead, each commit links to a previous, or *parent*, commit. You drew this:



> 
> 
> ```
> master
>  ---------- 
>          | module1
>          ----------
>                   | feature1
>                   ------------
> 
> ```
> 
> 


which suggests to me that you think of commits as belonging to (or being ""on"") *only one branch*. That's not how Git sees them, though. Instead, most commits are on *many branches at the same time*. For instance, consider a graph that we might draw like this:



```
          o--o--o   <-- br1
         /
...--o--o--o   <-- master
         \
          o--o   <-- br2

```

where each round `o` represents a commit. All the commits down the middle row are on `master`, but *most* of those commits are *also* on `br1` and `br2`. The *last* (newest and right-most) commit on `master` is *only* on `master`; the rest are on the other branches as well.


This is all because, in Git, a branch *name* like `master` points only to *one* commit. The commit to which the name points is the one furthest to the right, if we draw the commit graph this way, left (earlier) to right (later). Git calls this the *tip* commit, or sometimes the *head* (note lowercase here) of the branch. To find the rest of the commits that you—or Git—can reach from this tip commit, Git will look up the commit by its big ugly hash ID, such as `95ec6b1...`. Again, it's the *hash ID* that lets Git find the commit. The *name* just lets Git find this hash ID!


The commit itself stores a parent commit hash ID, so Git will look up the parent commit by *its* hash ID, finding the commit one step back. That commit has another parent ID, and so on. Walking backwards through this sequence of parent hash IDs, one commit at a time, from later to earlier, is how we see `git log`, for instance.


If you run:



```
git checkout br1

```

and then do some work and then run:



```
git add -a && git commit

```

and make a *new* commit—let's draw this as `*` instead of `o` so that we can see it—here's what happens to the branch name `br1`:



```
          o--o--o--*   <-- br1 (HEAD)
         /
...--o--o--o   <-- master
         \
          o--o   <-- br2

```

We draw this with the `(HEAD)` (note: all-uppercase) added to remember which name we gave to `git checkout`. The *new* commit `*` goes in the graph, and points backwards to whichever commit *was* the tip of `br1`. Meanwhile Git *changes* the name `br1` so that it points to the new commit we just made. This is how branches grow, in Git: we add new commits to the graph, and Git updates the *name* that `HEAD` is attached-to.


Of course, it's no accident that if we start at the tips of both `br1` and `master` and work backwards, we'll eventually come back to a single meeting-point commit. But this confluence of commits doesn't arise from the *names*. It doesn't matter *how we choose to start from* the tip commit of `br1` and the tip commit of `master`; what matters is *these two particular commits, and then each commit we find along the way*.


The branch names, in other words, get us started in the commit graph. ***It's the commit graph that matters most.*** The names just serve as starting points!


### The opposite of `push` is `fetch`


Everything above is about working within a single Git repository. But when we do work with Git, we work with more than one repository. In particular, we have *our* repository, where we do our own work, but then we often need to talk with another Git repository stored on some other machine, like those provided by GitHub, or by an employer or a friend or colleague.


This means that we want to *share* commits. As we saw above, Git is all about the commits and the commit graph—but we mere humans need *names* to get us started. This is where `git fetch` and its counterpart, `git push`, come in. Running either command connects our Git to some other Git.


Our Git has all the commits we have, some of which may be commits that we made. Some of our Git's commits might be commits we got from elsewhere. Similarly, their Git has all the commits they have, some of which are the same as commits we have. Some may be different commits. But in any case, *all* of these commits are identified by their unique hash IDs. These IDs are the same in *every* Git, all around the world, if they have the same commits. If they don't have our commits (because we just made them), our new commits have IDs that are different from every commit ID they have! (This seems like magic, but it's just cryptographic mathematics—some of which is similar to what's behind Bitcoin, for instance, though Git uses a weaker set of hashes.)


In the end, this means that each of these two Gits can tell which commits one of us has that the other doesn't, just by looking at these hash IDs. That's how our Git can give their commits we have that they don't—`git push`—or their Git can give our Git commits they have that we don't: `git fetch`.


Once they have sent commit objects (and other related Git objects needed to make those commits complete), the two Gits then need to set up *names* for any new *tip* commits. This is the second place that branch names start to matter.


The fetch direction is simpler. Your Git calls up some other Git. Usually we use the name `origin` to stand for some URL, where the other Git is listening for calls, so we run `git fetch origin`. Your Git calls up that Git, and asks it: *What are your branch tip names? What hash IDs are those?* Their Git tells your Git about its branch tips and hash IDs, and your Git either says: *Ah, I have that hash ID* or *Hmm, I don't have that hash ID, please send me that commit, and by the way what's its parent hash ID because maybe I need that one too*, and so on.


Eventually, your Git has all the commit and other hashed objects that they suggested. Now your Git takes their branch names, like their `master`, and *saves those names*. But you have your own branches. Your Git cannot save those names as *your* branches. It *renames* all those names. Their branches, like `master`, become your Git's *remote-tracking names*, like `origin/master`. Note that your Git is simply using your short-hand name, `origin`, plus a slash, in front of their branch names.


Once `git fetch` finishes, your Git now remembers where *their* Git's branch tips were, using *your* `origin/*` remote-tracking names. You have their commits, plus any required associated stuff so that you can check out those commits and get the files that go with them, but any *new* commits are only going to be *found* via these remote-tracking names. If they have, as branch tip commits, some of your *older* commits, you might already have other ways to find them.


The counterpart to `git fetch` is `git push`, but it's not completely symmetric. To do a `git push` to `origin`, for instance, you have your Git call up their Git as before, but this time, you want to *send* them things. The first part of this sending operation is to hand over any commits that you have, that they don't, that they will need. You identify these commits by having your `git push` take some extra arguments:



```
git push origin module1:module1

```

Note that I've put in the same name twice here. The name on the left, the `module1:`, is to find the *specific commit hash* that you want to send to them. That's the *tip commit* of your `module1`. The name on the right, the `:module1` part, is the *name you want them to use*. (These names don't have to be the same! But it gets tricky if you use different names on each side, so avoid that if possible.)


When you run `git fetch origin`, you generally want to pick up *everything* they have that you don't. That's quite safe, because whatever tip commit they have as their `master`, your Git will call your `origin/master`. Whatever tip commit they have as their `module1`, your Git will call that your `origin/module1`. Your remote-tracking names are your own private entries, all reserved for that one remote named `origin`, and it's harmless, or even a good thing, to just get them all up to date right away.1


But `git push` doesn't work this way. You send them a commit hash ID, and then ask them to set their *branch*, `master` or `module1` or `feature1`, to that hash ID. They have your Git send the commit object (and as many parent commits and other objects they need, all identified by hash IDs) if they don't already have that commit-ID, and then they evaluate for themselves *whether they will let you set their branch name*. Sometimes they will, and your push succeeds; sometimes they find a rule that says ""don't allow this"", and your push fails.


Note that their rules are up to them! You send a request (""please set your `module1` to `a9fc312...`""); they can look at their current `module1` branch hash ID, and the commit you've sent if it's new, and *choose* whether or not to take the request. You can use `--force` to send it as a command, but even if you do that, they can still choose whether to obey the command. All you can do is make a request (or forceful command) and see whether they accept it. But there's a standard rule that most Gits use most of the time: a request is allowed if the only thing it does is *add new commits*.


Look what happened when you added a commit to `br1` above. The existing commits, which were all *reachable from* the original tip commit, were still reachable from the new tip commit. The new commit ""grew the branch"". It did not change any of the existing—no new commit *can* ever change any existing commit—but the new commit's parent was the old tip commit. If we start at the new tip and work backwards, as Git does, we arrive at the old tip.


The same rules work well with `git push`: if the request you send to the other Git will keep all its existing commits reachable from the new branch tip, the other Git will probably allow the request.


Note that we've been using `git push origin module1:module1`, but you suggested that we run:



> 
> 
> ```
> git push origin module1 //Being in feature1 branch
> 
> ```
> 
> 


It doesn't matter which branch we're *in* (or `on`, as `git status` would say `on branch feature1`). What matters here is the two `module1:module1` parts to the `git push` command. That tells our Git: use *our* name `module1` to find the commit to push, and ask their Git to set *their* `module1` name.


We didn't give two parts. We only gave one part. For `git push`, if you only give one part, Git just assumes you mean ""use that same name twice"". So `git push origin module1` means `git push origin module1:module1`, after all.


### `git pull` is `git fetch` followed by a second Git command


You then asked about this:



> 
> 
> ```
> git pull origin module1 //Being in feature1 branch
> 
> ```
> 
> 


What `git pull` does is simple to describe:


1. It runs `git fetch` with most of the arguments you pass it:



```
git fetch origin module1

```
2. If that works, it runs a second Git command, which you pick in advance. The default command to run is `git merge`. The exact arguments to this second command depend a little bit on what came in during the `git fetch`, but you still have to pick `git merge` vs `git rebase` *before* you get to see what came in.


We mentioned before that we normally just run `git fetch origin`, which brings over all the names `origin` has, and your Git then renames. If you run:



```
git fetch origin module1

```

this simply *limits* the fetch: it finds out what they have as their `module1`, brings over the commit(s) by ID if necessary, and then sets your `origin/module1`.2 It ignores all their other names. The `module1` here is a lot like in `git push`, except that if you don't use two names—if you don't write `module1:module1`—your Git will just update your remote-tracking name. (And you should rarely if ever use two names separated by a colon here. It does work and can be used for some purposes, but you will need to learn more details.)


During the `git fetch`, it doesn't matter which branch you have checked out. But the *second* Git command is either `git merge` or `git rebase`, and for these two commands, it *does* matter which branch you have checked out.


The second command that `git pull` runs, runs *without changing the current branch*. Assuming the second branch is `git merge`, the arguments are:


- the branch tip hash ID3 obtained from the remote Git, and
- the message ""merge branch '*name*' of *url*""


This is a lot like running `git merge origin/*name*`, though the message is a bit different.


What `git merge` does with this is itself a bit complicated, since `git merge` sometimes does nothing, sometimes does a *fast-forward* instead of merging, sometimes does a real merge, and sometimes has merge conflicts that make the merge stop in the middle and get help from you. I'll leave all of this to other answers.


Let's summarize here:


- Your `git push` and `git fetch` operations always transfer *whole commits*. They never work on individual files: they copy *commits* from one Git repository to another. The process of copying commits normally involves setting some name(s) in the receiving Git, so as to remember any adjusted tip commits. For the most part, these don't care about your current branch (though `git push` can be told to push the current branch by default: see `push.default`).
- Your `git pull` simply runs two separate Git commands. If you're not already very familiar with both of those commands, I suggest you run each command separately instead.
- Your `git merge` command is the most complicated one, and should be a separate question. If you don't think you ran a `git merge` command, see `git pull` above.




---


1The one drawback to getting *everything* is that it could take a long time, if there is a lot of ""everything"" and your net connection is slow. On the other hand, if you get everything today, you will have almost nothing to get tomorrow, so the *next* `git fetch` will go fast. If you selectively get just a little today, maybe tomorrow's `git fetch` will be slow.


2This assumes your Git version is at least 1.8.4. In Git before 1.8.4, `git fetch origin module1` brings over their hash ID, but then throws away the name, failing to update your own `origin/module1`.


3If you run `git pull origin name1 name2`, Git passes multiple hash IDs to `git merge`, which produces what Git calls an *octopus merge*. This is almost never what you want, so avoid that. If you avoid `git pull` you won't make this particular mistake!


"
"Gradle's Maven Publish plugin not publishing POM or proper version to Maven Local
I am trying to get the Gradle [Maven Publish Plugin](https://docs.gradle.org/current/userguide/publishing_maven.html) to publish a snapshot version of my Java library to my local Maven repo such that:


1. The version of the jar is `1.0.0.SNAPSHOT-<timestamp>`, where `<timestamp>` is the current system time in millis (similar to something *like* `System.currentTimeInMillis()`); and
2. I log to STDOUT/console the full name of the jar being published, including the version above; and
3. A properly-formatted pom.xml is published to Maven local alongside the jar, so that any other Gradle/Maven projects can ""pull it down"" locally and fetch its transitive dependencies properly


My best attempt so far:



```
plugins {
    id 'java-library'
    id 'maven-publish'
}

dependencies {
    compile(
        'org.hibernate:hibernate-core:5.0.12.Final'
        ,'com.fasterxml.jackson.core:jackson-core:2.8.10'
        ,'com.fasterxml.jackson.core:jackson-databind:2.8.10'
        ,'com.fasterxml.jackson.core:jackson-annotations:2.8.0'
    )

    testCompile(
        'junit:junit:4.12'
    )
}

repositories {
    jcenter()
    mavenCentral()
}

sourceCompatibility = 1.8
targetCompatibility = 1.8
group 'com.me'

jar {
    baseName = 'my-lib'
    version = '1.0.0-SNAPSHOT'
}

publishing {
    publications {
        mavenJava(MavenPublication) {
            from components.java
        }
    }
}

```

However, with this setup, when I run `./gradlew publishToMavenLocal`:


- I do see the jar being deployed to `~/.m2/repository/com/me/my-lib/` but without a pom.xml and no `1.0.0.SNAPSHOT` version appended to it
- I don't even know how/where I would append the timestamp onto the version
- I don't even know how/where I would do a `println(...)` to report the full name of the jar being published


**Any ideas?**


","Regarding **#3**, To install your artifact to a local repository you do not need the `maven-publish` plugin, rather the `maven` plugin


See [The Maven plugin documentation, specifically the Tasks section and the Installing to the local repository section](https://docs.gradle.org/current/userguide/maven_plugin.html) with it, you can run `gradle clean build install`


It works for me with a build.gradle file as simple as this



```
version '1.0-SNAPSHOT'

apply plugin: 'java'
apply plugin: 'maven'

```

Note, if you need to publish something other then the default generated jar then you need to change the `archives` configuration




---


Regarding **#1** appending the timestamp, move the version line outside the jar clause and change it from



```
version = '1.0.0-SNAPSHOT'

```

to



```
version = ""1.0-SNAPSHOT-${System.currentTimeMillis()}""

```

This is using Groovy GString (AKA string interpolation - **note the change from single quotes to double quotes**) to append the current time in millis to the version




---


Last but not least, regarding **#2** printing the jar full name append the following to the build.gradle file



```
install.doLast {
        println jar.archiveName
}

```

Essentially we're appending to the install task (the one executed in the top of my answer) a println of the jar configuration's archiveName (see [here](https://docs.gradle.org/current/dsl/org.gradle.api.tasks.bundling.Jar.html) if you want something else)




---


So all in all my build.gradle file looks like this:



```
group 'com.boazj'
version ""1.0-SNAPSHOT-${System.currentTimeMillis()}""

apply plugin: 'java'
apply plugin: 'maven'

install.doLast {
        println jar.archiveName
}

```

"
"How can I convert strings like ""\u5c0f\u738b\u5b50\u003a\u6c49\u6cd5\u82f1\u5bf9\u7167"" to Chinese characters
I am now working on a small tool to request and decode a webpage, on which the Chinese characters are stored as string like 



```
\u5c0f\u738b\u5b50\u003a\u6c49\u6cd5\u82f1\u5bf9\u7167 

```

in the source code, something of unicode. I want to convert it to Chinese characters.


I can make it through this website <http://rishida.net/tools/conversion/>. But How can I make it using python?


","Those are Unicode codepoints *already*. They represent Chinese characters, but using escape codes that are easier on the developer:



```
>>> print u'\u5c0f\u738b\u5b50\u003a\u6c49\u6cd5\u82f1\u5bf9\u7167'
小王子:汉法英对照

```

You do not have to do *anything* to convert those; the `\uxxxx` escape form is simply another way to express the same codepoint. See [String Literals](http://docs.python.org/2/reference/lexical_analysis.html#string-literals):



> 
> `\uxxxx`  
> 
>  Character with 16-bit hex value xxxx (Unicode only)  
> 
> `\Uxxxxxxxx`  
> 
>  Character with 32-bit hex value xxxxxxxx (Unicode only)
> 
> 
> 


Python interprets those escape codes when reading the source code to construct the unicode value.


If the source of the data is not from Python source code but from the web, you have JSON data instead, which uses the same escape format:



```
>>> import json
>>> print json.loads('""\u5c0f\u738b\u5b50\u003a\u6c49\u6cd5\u82f1\u5bf9\u7167""')
小王子:汉法英对照

```

Note that the value then needs to be part of a larger string, one that *at least* includes quotes to mark this a string.


Also note that the JSON string escape format differs from Python's when it comes to non-BMP (supplementary) codepoints; JSON treats those like UTF-16 does, by [creating a surrogate pair](https://en.wikipedia.org/wiki/UTF-16#U.2B10000_to_U.2B10FFFF) and use two `\uxxxx` sequences for such a codepoint. In Python you'd use a `\Uhhhhhhhh` 32-bit hex value.


"
"Where is d3.svg.diagonal()?
I was trying to execute the code of collapsible-tree as mentioned [here](https://bl.ocks.org/mbostock/4339083). But it seems the diagonal method is not applicable in v4 (I may be wrong).


For: 



```
var diagonal = d3.svg.diagonal()

```

I get this error:



> 
> TypeError: Cannot read property 'diagonal' of undefined
> 
> 
> 


What is the equivalent in v4? Looking at d3.js API reference didn't give me any clue.


","D3 version 4.9.0 introduced [link shapes](https://github.com/d3/d3-shape/blob/master/README.md#links), which have the same functionality of the old `d3.svg.diagonal` in D3 v3.


According to the [API](https://github.com/d3/d3-shape/blob/master/README.md#links):



> 
> The link shape generates a smooth cubic Bézier curve from a source point to a target point. The tangents of the curve at the start and end are either vertical, horizontal or radial.
> 
> 
> 


There are three methods:


- [d3.linkHorizontal()](https://github.com/d3/d3-shape/blob/master/README.md#linkHorizontal)
- [d3.linkVertical()](https://github.com/d3/d3-shape/blob/master/README.md#linkVertical)
- [d3.linkRadial()](https://github.com/d3/d3-shape/blob/master/README.md#linkRadial)


So, for a collapsible tree like that one you linked, you define the path `d` attribute as:



```
.attr(""d"", d3.linkHorizontal()
    .x(function(d) { return d.y; })
    .y(function(d) { return d.x; }));

```

**Demo**:


Suppose you have an object with `source` and `target`, each one with `x` and `y` properties:



```
var data = {
  source: {
    x: 20,
    y: 10
  },
  target: {
    x: 280,
    y: 100
  }
};

```

First, you create the link generator:



```
var link = d3.linkHorizontal()
  .x(function(d) {
    return d.x;
  })
  .y(function(d) {
    return d.y;
  });

```

And then you can draw the path just by passing that data to the link generator:



```
.attr(""d"", link(data))

```

Here is the demo:





```
var svg = d3.select(""svg"");

var data = {
  source: {
    x: 20,
    y: 10
  },
  target: {
    x: 280,
    y: 100
  }
};

var link = d3.linkHorizontal()
  .x(function(d) {
    return d.x;
  })
  .y(function(d) {
    return d.y;
  });

svg.append(""path"")
  .attr(""d"", link(data))
  .style(""fill"", ""none"")
  .style(""stroke"", ""darkslateblue"")
  .style(""stroke-width"", ""4px"");
```


```
<script src=""https://d3js.org/d3.v5.min.js""></script>
<svg></svg>
```






"
"What is changelist in Git/IntelliJ and how to get rid of it?
If I am commiting code in IntelliJ with Git, I see the following dialog:


[![enter image description here](https://i.stack.imgur.com/ZUimX.png)](https://i.stack.imgur.com/ZUimX.png)


As you see, it has `Change list` set to ""Recovering from unsuccessfull..."". This is my title, created long time ago and I don't know how to use it.


So, I would like to remove this title from the dialog.


But if I change it to ""Default"", then I get empty window:


[![enter image description here](https://i.stack.imgur.com/RzdGN.png)](https://i.stack.imgur.com/RzdGN.png)


What all these things mean and how to get rid of old change list title, still being able to commit code?


","
> 
> What all these things mean and how to get rid of old change list
>  title, still being able to commit code?
> 
> 
> 


First off, you should move your changes from the old changelist to the default one. In order to do that, go to `View | Tool Windows | Version Control` or press `Alt` + `9` in windows:


[![](https://i.stack.imgur.com/esM72.png)](https://i.stack.imgur.com/esM72.png)
*Version Control Panel*  

Then select the old changelist title and drag it to the default one or use **Move to Another Changelist** button. After you moved changes to the default changelist, just right click on the old one and select **Delete** option. 



> 
> What is changelist in Git/IntelliJ and how to get rid of it?
> 
> 
> 


Checkout the IntelliJ's documentation on [Changelist](https://www.jetbrains.com/help/idea/2016.1/changelist.html) and [Managing Changelist](https://www.jetbrains.com/help/idea/2016.1/managing-changelists.html).


"
"How to profile a C++ function at assembly level?
I have a function that is the bottleneck of my program. It requires no access to memory and requires only calculation. It is the inner loop and called many times so any small gains to this function is big wins for my program. 


I come from a background in optimizing SPU code on the PS3 where you take a SPU program and run it through a pipeline analyzer where you can put each assembly statement in its own column and you minimize the amount of cycles the function takes. Then you overlay loops so you can minimized pipeline dependencies even more. With that program and a list of all the cycles each assembly instruction takes I could optimize much better then the compiler ever could.


On a different platform it had events I could register (cache misses, cycles, etc.) and I could run the function and track CPU events. That was pretty nice as well.


Now I'm doing a hobby project on Windows using Visual Studio C++ 2010 w/ a Core i7 Intel processor. I don't have the money to justify paying the large cost of VTune.


**My question:**


*How do I profile a function at the assembly level for an Intel processor on Windows?*


I want to compile, view disassembly, get performance metrics, adjust my code and repeat.


","There are some great free tools available, mainly AMD's CodeAnalyst (from my experiences on my i7 vs my phenom II, its a bit handicapped on the Intel processor cause it doesn't have access to the direct hardware specific counters, though that might have been bad config). 


However, a lesser know tool is the [Intel Architecture Code Analyser](http://software.intel.com/en-us/articles/intel-architecture-code-analyzer/) (which is free like CodeAnalyst), which is similar to the spu tool you described, as it details latency, throughput and port pressure (basically the request dispatches to the ALU's, MMU and the like) line by line for your programs assembly. Stan Melax gave a nice [talk](http://www.gdcvault.com/play/1014645) on it and x86 optimization at this years GDC, under the title ""hotspots, flops and uops: to-the-metal cpu optimization"". 


Intel also has a few more tools in the same vein as IACA, avaibale under the performance tuning section of [their experimental/what-if code site](http://software.intel.com/en-us/whatif/), such as PTU, which is (or was) an experimental evolution of VTune, from what I can see, its free.


Its also a good idea to have read the intel optimization manual before diving into this.


EDIT: as Ben pointed out, the timings might not be correct for older processors, but that can be easily made up for using [Agner Fog's Optimization manuals](http://www.agner.org/optimize/), which also contain many other gems.


"
"Understanding interpn function of scipy.interpolate
I am trying to understand the working of the function `scipy.interpolate`. I created a small setup but it is giving errors. Here is what I did:



```
import numpy as np
import scipy.interpolate
import matplotlib.pyplot as plt

x = np.arange(10)
y = np.arange(10)
gx, gy = np.meshgrid(x,y)
v = np.ones((10,10))


sample_at = np.random.random((10,10))

interpolated = scipy.interpolate.interpn((gx, gy), v, sample_at)


print(interpolated.shape)

```

This gives an error:



```
    Traceback (most recent call last):
File ""test.py"", line 13, in <module>
    interpolated = scipy.interpolate.interpn((gx, gy), v, sample_at)
File ""/home/lib/python3.5/site-packages/scipy/interpolate/interpolate.py"", line 2624, in interpn
    ""1-dimensional"" % i)
ValueError: The points in dimension 0 must be 1-dimensional

```

What is happening here? 


","You are making a wrong assumption wrt the structure of the grid requested by `interpn`



> 
> `ValueError: The points in dimension 0 must be 1-dimensional`
> 
> 
> 


This message, a little bit obscure, references the elements of the grid and tells you that they must be 1-dimensional, or in other words you must call `interpn` not with `(gx, gy)` as first argument but with `(x, y)`.



```
scipy.interpolate.interpn((x, y), v, sample_at)

```

But there is another wrong assumption in your code, because if you use the call above, with your definition of `sample_at`, you'll get a different error



> 
> `ValueError: The requested sample points xi have dimension 10, but this RegularGridInterpolator has dimension 2`
> 
> 
> 


that hopefully has a clearer meaning: if your grid has two dimensions (*x* and *y*, that is) all your points on which `interpn` has to interpolate must be 2-dimensional as well... In other words, in your example the *LAST* dimension of the grid must be `2`


Putting it all together (I'll use a location matrix `sample_at` of 5 rows by 3 columns of 2-D points in this example) we have



```
In [69]: import numpy as np 
    ...: import scipy.interpolate 
    ...: import matplotlib.pyplot as plt 
    ...:  
    ...: x = np.arange(10) 
    ...: y = np.arange(10) 
    ...: v = np.ones((10,10)) 
    ...: sample_at = 8*np.random.random((30)).reshape(5, 3, 2) 
    ...:  
    ...: scipy.interpolate.interpn((x, y), v, sample_at)                                                       
Out[69]: 
array([[1., 1., 1.],
       [1., 1., 1.],
       [1., 1., 1.],
       [1., 1., 1.],
       [1., 1., 1.]])

```

PS: read the source code1, the comments in it are much better than the error messages...




---


(1) for me the source code is located at



```
~/lib/miniconda3/lib/python3.7/site-packages/scipy/interpolate/interpolate.py

```

"
"Shortest decimal representations of rational numbers
I want to to represent two numbers in `A/B` format,wWhere `A` and `B` are any integers. I want the decimal result in shortest format. 


Examples:


- The shortest decimal representation of `3/28` is `""0.10(714285)""`.
- Decimal representation of `1/7` is `""0.(142857)""`


How can I do this with C#?


","Hopefully, someone can convert this to C# code:



```
def get_decimal(a, b):
    result = str(a / b) + '.'
    r = a % b
    remainders = []
    while r not in remainders:
        remainders.append(r)
        result += str(10 * r / b)
        r = 10 * r % b
    return result

```

Some tests:



```
>>> get_decimal(1, 7)
'0.142857'
>>> get_decimal(10, 7)
'1.428571'
>>> get_decimal(3, 28)
'0.10714285'
>>> get_decimal(99099167, 990000000)
'0.100100168'

```

If you need the parentheses to indicate the repeating set of digits, then something like this:



```
def get_decimal(a, b):
    result = str(a / b) + '.'
    r = a % b
    remainders = []
    digits = []
    while r not in remainders:
        remainders.append(r)
        digits.append(str(10 * r / b))
        r = 10 * r % b
    start = remainders.index(r)
    result += ''.join(digits[0:start]) + '(' + ''.join(digits[start:]) + ')'
    return result

```

Some tests:



```
>>> get_decimal(1, 7)
'0.(142857)'
>>> get_decimal(10, 7)
'1.(428571)'
>>> get_decimal(3, 28)
'0.10(714285)'
>>> get_decimal(99099167, 990000000)
'0.1001001(68)'

```

"
"Allow access to Apache on both port 80 and 443 in Ubuntu 16.04
I've just installed UFW 0.35 on Ubuntu 16.04:



```
root@localhost:/etc# ufw --version
ufw 0.35
Copyright 2008-2015 Canonical Ltd.

```

and



```
root@localhost:/etc# ufw app list
Available applications:
  OpenSSH

```

I would like to allow access to Apache on both port 80 and 443, with the command 



```
$ ufw allow ""Apache Full""

```

but I got an error



```
ERROR: Could not find a profile matching 'Apache Full'

```

","You are likely receiving that error because there has not been a profile created for 'Apache Full'. You can see which profiles exist on your system by checking the directory:



```
/etc/ufw/applications.d/

```

To create a profile known as 'Apache Full' create a file in the above directory using the following syntax (from the man page):



```
     [Apache Full]
     title=<title>
     description=<description>
     ports=80/tcp,443/tcp

```

Next, you will update ufw app:



```
ufw app update ""Apache Full""

```

Now you should be able to run the command from your question:



```
ufw allow ""Apache Full""

```

To confirm that your profile is included in ufw's rules run:



```
ufw status

```

"
"After upgrading to xcode 9, cordova app won't build, error 70, requires provisioning profile
Yesterday we upgraded from xcode 8.3.2 to version 9. And now our enterprise distribution apache cordova ios app refuses to build.



```
2017-09-21 07:37:16.787 xcodebuild[70400:217569] [MT] IDEDistribution: -[IDEDistributionLogging _createLoggingBundleAtPath:]: Created bundle at path '/var/folders/wj/yj3cfvh954gbc_btlhcrcx7nk7t4dj/T/App Name_2017-09-21_07-37-16.786.xcdistributionlogs'.
2017-09-21 07:37:16.938 xcodebuild[70400:217569] [MT] IDEDistribution: Step failed: <IDEDistributionSigningAssetsStep: 0x7ff756bbdf70>: Error Domain=IDEDistributionSigningAssetStepErrorDomain Code=0 ""Locating signing assets failed."" UserInfo={NSLocalizedDescription=Locating signing assets failed., IDEDistributionSigningAssetStepUnderlyingErrors=(
    ""Error Domain=IDEProvisioningErrorDomain Code=9 \""\""App Name.app\"" requires a provisioning profile.\"" UserInfo={NSLocalizedDescription=\""App Name.app\"" requires a provisioning profile., NSLocalizedRecoverySuggestion=Add a profile to the \""provisioningProfiles\"" dictionary in your Export Options property list.}""
)}
error: exportArchive: ""App Name.app"" requires a provisioning profile.

Error Domain=IDEProvisioningErrorDomain Code=9 """"App Name.app"" requires a provisioning profile."" UserInfo={NSLocalizedDescription=""App Name.app"" requires a provisioning profile., NSLocalizedRecoverySuggestion=Add a profile to the ""provisioningProfiles"" dictionary in your Export Options property list.}

** EXPORT FAILED **

Error: Error code 70 for command: xcodebuild with args: -exportArchive,-archivePath,App Name.xcarchive,-exportOptionsPlist,/Users/Shared/Workspace/github/AppName/platforms/ios/exportOptions.plist,-exportPath,/Users/Shared/Workspace/github/AppName/platforms/ios/build/device

```

(I replaced the apps name with ""App Name"" for this question)


I have verified that all the certs and provisioning profiles are set in xcode. And the build.json is set. I'm not really a Mac guy and I'm stumped on what to do next.


-Edit, Downgrading xcode to 8.3.3 fixed the problem. Not an ideal solution but not much I can do.


","If you specify your provisioning profile explicitly, like me. Like this in your Cordova build.json:



```
""ios"": {
    ""debug"": {
        ""codeSignIdentitiy"": ""iPhone Developer"",
        ""developmentTeam"":""MYTEAMID"",
        ""packageType"": ""developer"",
        ""iCloudContainerEnvironment"": ""Development""
    },
    ""release"": {
        ""codeSignIdentitiy"": ""iPhone Distribution"",
        ""developmentTeam"":""MYTEAMID"",
        ""provisioningProfile"": ""XXXXXXXX-XXXX-XXXX-XXXX-XXXXXXXXXXXX"",
        ""packageType"": ""ad-hoc"",
        ""iCloudContainerEnvironment"": ""Production""
    }
}

```

**Please Note**
iCloudContainerEnvironment = Production/Development is only required if you use push notifications


You need to explicitly set manual signing and provide the provisioning keys in your ExportOptions.plist that is generated by Cordova. Unfortunately Cordova is not currently generating all of the required keys.


Here is what it needs to look like, at least for me:



```
<?xml version=""1.0"" encoding=""UTF-8""?>
<!DOCTYPE plist PUBLIC ""-//Apple//DTD PLIST 1.0//EN"" ""http://www.apple.com/DTDs/PropertyList-1.0.dtd"">
<plist version=""1.0"">
<dict>
  <key>compileBitcode</key>
  <false/>
  <key>method</key>
  <string>ad-hoc</string>
  <key>iCloudContainerEnvironment</key >
  <string>Production</string>
  <key>provisioningProfiles</key>
  <dict>
    <key>my.bundle.idenifier</key>
    <string>My Provisioning Profile Name</string>
  </dict>
  <key>signingCertificate</key>
  <string>iPhone Distribution</string>
  <key>signingStyle</key>
  <string>manual</string>
  <key>stripSwiftSymbols</key>
  <true/>
  <key>teamID</key>
  <string>YOURTEAMID</string>
  <key>thinning</key>
  <string>&lt;none&gt;</string>
</dict>
</plist>

```

The file Cordova generates @ cordova/app/platforms/ios/exportOptions.plist looks like:



```
<?xml version=""1.0"" encoding=""UTF-8""?>
<!DOCTYPE plist PUBLIC ""-//Apple//DTD PLIST 1.0//EN"" ""http://www.apple.com/DTDs/PropertyList-1.0.dtd"">
<plist version=""1.0"">
  <dict>
    <key>compileBitcode</key>
    <false/>
    <key>method</key>
    <string>development</string>
    <key>teamID</key>
    <string>MYTEAMID</string>
  </dict>
</plist>

```

notice it is missing the important bits that Xcode 9 requires. 


I generated the correct file by archiving the build manually, then exporting it which also creates the exportOptions.plist that I now use as reference. 


After digging deeper, I found that this cannot be fixed after running ""Cordova add platform iOS"", because it is generated during the build phase dynamically. I decided to fork the Cordova-ios repo and submit a pull request. You may use my fork directly, or wait until the pull request is merged.


**Pull Request**
<https://github.com/apache/cordova-ios/pull/338/commits>


**Fork**
<https://github.com/jrryhrtn/cordova-ios>


**Usage notes from pull request**


See example below, please note that the provisioning profile can be the name or UUID of the profile. Name is preferred for maintenence, as UUID changes each time to regenerate the profile.



```
{
""ios"": {
    ""debug"": {
        ""codeSignIdentity"": ""iPhone Developer"",
        ""developmentTeam"":""YOURTEAMID"",
        ""provisioningProfile"": ""provisioning profile name or UUID"",
        ""packageType"": ""development""
    },
    ""release"": {
        ""codeSignIdentity"": ""iPhone Distribution"",
        ""developmentTeam"":""YOURTEAMID"",
        ""provisioningProfile"": ""provisioning profile name or UUID"",
        ""packageType"": ""ad-hoc""
    }
}
}

```

I plan to manually patch until the/a fix is merged into the next Cordova release. You will have to regenerate your iOS platform after the patch via ""Cordova platform rm iOS"" and then ""Cordova platform add ~/forks/cordova-ios"". ~/forks/cordova-ios my local path, use the path on your local machine where you downloaded the forked Cordova-ios repo.


**Update**


cordova-ios 4.5.2 has been officially released! Upgrade by running the following commands: ""cordova platform rm ios"", and then ""cordova platform add ios@4.5.2""


Cheers!


"
"Visualize arbitrary tree in Racket using tree-layout
How to visualise arbitrary tree?


for example:
 `(define T1 '(and (or x1 x2)(or x3 x4 x5)))`


or one generated with:



```
(define functions '(not if and or))
(define terminals '(A0 A1 A2 D0 D1))
(define functions&terminals (append terminals functions ))

(define (pick-one list)
  (list-ref list (random (length list))))

(define arities '((if . 3)(and . 2)(or . 2)(not . 1)))

(define (terminal? symbol)
  (find (lambda (x)(eq? x symbol)) terminals))

(define (function? symbol)
  (find (lambda (x)(eq? x symbol)) functions))

(define (arity non-terminal)
  (let ((arity (find (lambda (x)(eq? non-terminal (car x))) arities)))
    (if arity
        (cdr arity)
        0)))

(define (numbers n)
  (if (= n 0) 
      '()
      (cons n (numbers (- n 1)))))

(define (gen-tree) 
  (let ((node (pick-one functions&terminals))) 
    (if (terminal? node) 
        node 
        (cons node (map (lambda (x) (gen-tree)) (numbers (arity 
node)))))))

> (gen-tree)
'(or (if A1 (and A1 (not (if D1 (and A0 A0) (or A0 A0)))) (or A0 A0)) D0)

```

Racket seems to have:
<https://docs.racket-lang.org/pict/Tree_Layout.html>


is it enough to visualise trees of functions with the name of the function and params in the circle?


","You can do something like this to visualize arbitrarily sized trees:



```
(require pict
         pict/tree-layout)

(define (draw tree)
  (define (viz tree)
    (cond
      ((null? tree) #f)
      ((not (pair? tree))
       (tree-layout #:pict (cc-superimpose
                            (disk 30 #:color ""white"")
                            (text (symbol->string tree)))))
      ((not (pair? (car tree)))
       (apply tree-layout (map viz (cdr tree))
              #:pict (cc-superimpose
                      (disk 30 #:color ""white"")
                      (text (symbol->string (car tree))))))))
  (if (null? tree)
      #f
      (naive-layered (viz tree))))

```

For example, using the lists you provided:



```
(define t1 '(and (or x1 x2) (or x3 x4 x5)))
(define t2 '(or (if A1 (and A1 (not (if D1 (and A0 A0) (or A0 A0)))) (or A0 A0)) D0))

```

[![enter image description here](https://i.stack.imgur.com/eGXLW.png)](https://i.stack.imgur.com/eGXLW.png)


"
"Firebase queryEqualToValue with childKey
I am trying to query data in my Firebase database using:



```
queryEqual(toValue: Any?, childKey: String?)

```

My database structure:



```
Schools {
    testUID1 {
        schoolNameLC: test school 1
    }
}

```

My query is:



```
databaseReference.child(""Schools"").queryEqual(toValue: ""test school 1"", childKey: ""schoolNameLC"").observe(.value) { (snap) in
        print(snap)
    }

```

This query prints out `null` and I can't quite get it to work. Because of the way my app is set up, I don't know that the key `schoolNameLC` has a value `testSchool1` under the parent key of `testUID1`. All I want to do is search through the `Schools` in my database and return anything with a `schoolNameLC` value of `test school 1`.


","The two-parameter `queryEqualToValue:childKey:` (and its brethren `queryStartingAtValue:childKey:` and `queryEndingAtValue:childKey:`) are unfortunately some of the most misunderstood methods in the Firebase Database API.


To order by a child and then filter on a value of that child, you have to call `queryOrderedByChild().queryEqualToValue()`. So as @ElCaptainV2.0 said:



```
databaseReference.child(""Schools"")
    .queryOrderedByChild(""scho‌​olNameLC"")
    .queryEqua‌​lToValue(""test school 1"")

```

You only should use the `childKey:` parameter if you **also** want to start at a specific key in all nodes that have the same matching `test school 1` value. This overload is really mostly useful when you're trying to paginate/endless scroll results.


"
"XmlDocument CreateElement without xmlns under a prefixed element
I'm trying to compose a SOAP request to ebay FindingAPI web service by using C# XmlDocument class in the following code:



```
XmlDocument doc = new XmlDocument();
XmlElement root = (XmlElement)doc.AppendChild(doc.CreateElement(""soap"", ""Envelope"", ""http://www.w3.org/2003/05/soap-envelope""));
root.SetAttribute(""xmlns"", ""http://www.ebay.com/marketplace/search/v1/services"");
XmlElement header = (XmlElement)root.AppendChild(doc.CreateElement(""soap"", ""Header"", ""http://www.w3.org/2003/05/soap-envelope""));
XmlElement body = (XmlElement)root.AppendChild(doc.CreateElement(""soap"", ""Body"", ""http://www.w3.org/2003/05/soap-envelope""));
XmlElement request = (XmlElement)body.AppendChild(doc.CreateElement(""findItemsByKeywordsRequest""));
XmlElement param = (XmlElement)request.AppendChild(doc.CreateElement(""keywords""));
param.InnerText = ""harry potter phoenix"";

```

And, the XML output of above code is: 



```
<soap:Envelope xmlns=""http://www.ebay.com/marketplace/search/v1/services"" xmlns:soap=""http://www.w3.org/2003/05/soap-envelope"">
    <soap:Header />
    <soap:Body>
        <findItemsByKeywordsRequest xmlns="""">
            <keywords>harry potter phoenix</keywords>
        </findItemsByKeywordsRequest>
    </soap:Body>
</soap:Envelope>

```

However, this XML can't be recognized by the server because of the extra xmlns="""" attribute in the findItemsByKeywordsRequest element. The desired XML output should be as below:



```
<soap:Envelope xmlns:soap=""http://www.w3.org/2003/05/soap-envelope"" xmlns=""http://www.ebay.com/marketplace/search/v1/services"">
    <soap:Header/>
    <soap:Body>
        <findItemsByKeywordsRequest>
            <keywords>harry potter phoenix</keywords>
        </findItemsByKeywordsRequest>
    </soap:Body>
</soap:Envelope>

```

Does anyone know what is the problem of my code and please give me some hints. Thanks!


","Because your document has default namespace declared in the most outer element you have to repeat that namespace on every child element to avoid adding additional empty one.


Change `request` and `param` elements declaration to contain `""http://www.ebay.com/marketplace/search/v1/services""` namespace



```
XmlElement request = (XmlElement)body.AppendChild(doc.CreateElement(""findItemsByKeywordsRequest"", ""http://www.ebay.com/marketplace/search/v1/services""));
XmlElement param = (XmlElement)request.AppendChild(doc.CreateElement(""keywords"", ""http://www.ebay.com/marketplace/search/v1/services""));

```

With these changes your code produces following XML:



```
<soap:Envelope xmlns=""http://www.ebay.com/marketplace/search/v1/services"" xmlns:soap=""http://www.w3.org/2003/05/soap-envelope"">
    <soap:Header />
    <soap:Body>
        <findItemsByKeywordsRequest>
            <keywords>harry potter phoenix</keywords>
        </findItemsByKeywordsRequest>
    </soap:Body>
</soap:Envelope>

```

"
"Is it possible to pause and resume buffering of AVPlayer?
Is it possible to pause and resume buffering of AVPlayer?


","*Yes it is possible to some extent!*


You can use the playerItem's `preferredForwardBufferDuration` property to decide how much duration from the current playing time the player should prefetch. But sadly, this is only available from iOS version 10.


*Macro to check system version:*



```
#define SYSTEM_VERSION_GREATER_THAN_OR_EQUAL_TO(v)  ([[[UIDevice currentDevice] systemVersion] compare:v options:NSNumericSearch] != NSOrderedAscending)

```

You can now set how much duration you want to prefetch (in seconds).



```
if (SYSTEM_VERSION_GREATER_THAN_OR_EQUAL_TO(@""10.0"")) {
    NSTimeInterval interval = 1; // set to  0 for default duration.
    _player.currentItem.preferredForwardBufferDuration = interval;
    _player.automaticallyWaitsToMinimizeStalling = YES;

}

```

`automaticallyWaitsToMinimizeStalling` is another property which enables the `autoplay` or `autowait` feature in avplayer. Because the player is likely to stall frequently if the `preferredForwardBufferDuration` is set to a smaller duration.


You can also use the playeritem's `canUseNetworkResourcesForLiveStreamingWhilePaused` property to set if the player should continue or pause buffering when the player is in pause state. This is available from iOS 9 onwards.



```
if (SYSTEM_VERSION_GREATER_THAN_OR_EQUAL_TO(@""9.0"")) {
_player.currentItem.canUseNetworkResourcesForLiveStreamingWhilePaused = NO;
}

```


> 
> The condition check for the system version is very important. You can
>  use the macro mentioned above to do this. Otherwise the app will
>  crash.
> 
> 
> 


***UPDATE - swift:***



```
    if #available(iOS 10.0, *) {
            player.currentItem?.preferredForwardBufferDuration = TimeInterval(1)
            player.automaticallyWaitsToMinimizeStalling = true;
    }

```

"
"Where is the struct task\_struct definition in the 2.6.30.5 Linux Kernel?
In the version 2.6.15 kernel, I got that I can rewrite the `task_struct` in the file (include/linux/sched.h)，like:



```
struct task_struct {  
    unsigned did_exec:1;  
    pid_t pid;  
    pid_t tgid;  
    ...
    char hide;
}  

```

But, unfortunately, when I upgraded to the version 2.6.30.5, I looked through the same file, I just find a declaration of the `task_struct`, like:



```
struct task_struct;

```

And I have no idea which file I should refer to for the purpose of specifying my own `task_struct`? Can someone help me?


","Use `grep` or any other search tool to look for the definition:



```
grep -r '^struct task_struct ' include

```

Or search online at [LXR](http://lxr.linux.no/linux/):
<http://lxr.linux.no/linux+v2.6.30.5/+search?search=task_struct>


The structure is still defined in `include/linux/sched.h`. There's a [forward declaration](http://lxr.linux.no/linux+v2.6.30.5/include/linux/sched.h#L246) which is used in mutually recursive type definitions, and the [definition](http://lxr.linux.no/linux+v2.6.30.5/include/linux/sched.h#L1117) is further down.


"
"if command in python

```
if ""aa"" or ""bb"" or ""cc"" or ""dd"" or ""ee"" or ""ff"" in attrs[""show""]:
    self.xx = xxxx

```

I have a code like this, to check if attrs[""show""] contains either of these strings, then assign some value to self.xx


But is this ""IF"" command is correct? Because from my results, it seems like this if command is always true(which is impossible)


","Try the following:



```
if any(s in attrs[""show""] for s in (""aa"", ""bb"", ""cc"", ""dd"", ""ee"", ""ff"")):
    self.xx = xxxx

```

Your current if statement will always evaluate to `True`, because instead of checking if each string is in `attrs[""show""]` you are checking if `""aa""` is True, or if `""bb""` is True, and on and on. Since `""aa""` will always evaluate as True in a Boolean context, you will always enter the `if` clause.


Instead, use the [`any()`](http://docs.python.org/library/functions.html#any) function as in my example, which accepts an iterable and returns True if any of its elements are True. Similar to a bunch of `or`'s as in your code, this will short-circuit as soon as it finds an element that is True. In this example, the above code is equivalent to the following (but much more concise!):



```
if (""aa"" in attrs[""show""] or ""bb"" in attrs[""show""] or ""cc"" in attrs[""show""] or
    ""dd"" in attrs[""show""] or ""ee"" in attrs[""show""] or ""ff"" in attrs[""show""]):
    self.xx = xxxx

```

"
"Model selection: can I compare the AIC from models of count data between linear and poisson models?
I am modeling count data (with offset / exposure parameter). My modeling strategy is use of a Poisson model and a negative binomial regression model. I compare model AICs, which are about -760 for my case. I selected the model with the best AIC. Finally, I was interested how a linear model (GLM with Gaussian distribution and identity link) would perform in comparisson to the count models. The linear model obviously does not fulfil the distributional assumptions as well as the count models. However, now I find model AICs of arround -150 suggesting much better fit. 


Should I use the linear model based on AIC despite the fact that basic distributional assumptions are not met? Are the AICs comparable?


I used a GLMM function for fitting: <http://glmmadmb.r-forge.r-project.org/> (my data are clustered).


","You cannot use likelihood-based statistics like AIC to compare across models with different likelihood functions - the underlying formulas are different. In linear regression, the likelihood function is the normal density function, in Poisson regression it is the Poisson function. That will account for the differences in the AIC probably more than any differences in fit.


Before you decide to even use a linear model, you need to make sure that the residuals from the model are normally distributed (you can proxy that by looking at the distribution of the outcome variable, though keep in mind it isn't the same). If they are not normally distributed, or close enough for the eye, then you can't use a normal regression model to do any hypothesis testing.


Assuming that it is approximately normal, I would take a two broad approaches to choose the model to report.


1) Predicted outcomes. Estimate the predicted outcomes of each model and compare. Does the linear model have better predictive ability? You may want to do this in a cross-validation framework, where you ""train"" your model on part of your data and use the rest for prediction.


2) Intuitive interpretation of coefficients. Poisson coefficients can be complicated to understand - they are not the change in number of *y* but rather a proportional change. Depending on your context this may be more or less useful. Sometimes it is worth sacrificing fit if your model can be more easily interpreted by the end-user - for example, some researchers are willing to avoid the complexity of logit and probit models for the easier-to-interpret coefficients in a linear probability model, even though the LPM has tons of setbacks. Think about who your audience is, what is your context, what is your research question, etc., as you make these decisions.


EDIT: I forgot to add [this](http://www.ncbi.nlm.nih.gov/pmc/articles/PMC3238139/) paper, which gives a good comparison across a range of different count models and may be helpful.


"
"Control the application from control center in iOS
I am using AVAudioPlayer in my iOS application and the following functionalities have been given.


**1. Play
2. Pause
3. Next
4. Previous**


My requirement is, I want to give these functionalities from control center. Is it possible?


","Control Center uses same API as lock screen. The audio controls appear on lock screen when a properly configured `AVAudioSession` is active. In addition to that you need to use `MPNowPlayingInfoCenter` API to set correct media information.



```
MPMediaItemArtwork *artwork = [[MPMediaItemArtwork alloc]initWithImage:albumImage];

[MPNowPlayingInfoCenter defaultCenter].nowPlayingInfo = @{ 
    MPMediaItemPropertyTitle : aSong.songTitle,
    MPMediaItemPropertyArtist : aSong.artistName,
    MPMediaItemPropertyArtwork : artwork,
    MPNowPlayingInfoPropertyPlaybackRate : 1.0f
};

```

See the questions below for more information:


- [Why do the Lock Screen audio controls disappear when I pause AVAudioPlayer?](https://stackoverflow.com/questions/21892151/why-do-the-lock-screen-audio-controls-disappear-when-i-pause-avaudioplayer)
- [AVAudioPlayer on Lock Screen](https://stackoverflow.com/questions/20363058/avaudioplayer-on-lock-screen)


"
"Great Circle Distance between two ""P = (x, y, z)"" points on a unit sphere
I need to calculate distance between two points, using their coordinates (*x, y, z*) on a unit sphere. 


Both [Haversine and Great Arc Distance](https://math.stackexchange.com/questions/231221/great-arc-distance-between-two-points-on-a-unit-sphere/231225) use *φ* and *λ* (longitude and latitude). How do i write those Formulas for *x, y, z*?


","Generalizing to a sphere of radius `R`, so that the physical dimensions of expressions and variables is always apparent, we draw the following figure where you can see, on the plane that contains the three points, the points *A* and *B* on the surface of the sphere and the point *O*, the centre of the sphere, and also the arc of minimum length that connects *A* and *B*.


[![enter image description here](https://i.stack.imgur.com/W8OIL.png)](https://i.stack.imgur.com/W8OIL.png)


With reference to the figure and the text in it, you can compute the distance between points *A* and *B* like this:


1. Compute the distance `Δ = math.sqrt((x_B-x_A)**2+(y_B-y_A)**2+(z_B-z_A)**2)`
2. Compute the angle underlying one half of Δ, `φ = math.asin((Δ/2/R))`
3. the distance on the great circle is `gc_dist = 2*phi*R` .


(you may want to use `R = 1`).


"
"dplyr: passing a grouped tibble to a custom function
(The following scenario simplifies my actual situation)  

My data comes from villages, and I would like to summarize an outcome variable by a village variable.



```
> data
   village     A     Z      Y 
     <chr> <int> <int>   <dbl> 
 1       a     1     1   500     
 2       a     1     1   400     
 3       a     1     0   800  
 4       b     1     0   300  
 5       b     1     1   700  

```

For example, I would like to calculate the mean of `Y` only using `Z==z` by villages. In this case, I want to have (500 + 400)/2 = 450 for village ""a"" and 700 for village ""b"". 


Please note that the actual situation is more complicated and I cannot directly use [this answer](https://stackoverflow.com/questions/50925734/dplyr-use-a-custom-function-in-summarize-after-group-by/50926037?noredirect=1#comment88854293_50926037), but **the point is I need to pass a grouped tibble and a global variable (z) to my function**.



```
z <- 1 # z takes 0 or 1
data %>%
    group_by(village) %>% # grouping by village
    summarize(Y_village = Y_hat_village(., z)) # pass a part of tibble and a global variable

Y_hat_village <- function(data_village, z){
    # This function takes a part of tibble (`data_village`) and a variable `z`
    # Calculate the mean for a specific z in a village
    data_z <- data_village %>% filter(Z==get(""z""))
    return(mean(data_z$Y))
}

```

However, I found `.` passes entire tibble and the code above returns the same values for all groups.


","There are a couple things you can simplify. One is in your function: since you're passing in a value `z` to the function, you don't need to use `get(""z"")`. You have a `z` in the global environment that you pass in; or, more safely, assign your z value to a variable with some other name so you don't run into scoping issues, and pass that in to the function. In this case, I'm calling it `z_val`.



```
library(tidyverse)

z_val <- 1

Y_hat_village2 <- function(data, z) {
  data_z <- data %>% filter(Z == z)
  return(mean(data_z$Y))
}

```

You can make the function call on each group using `do`, which will get you a list-column, and then unnesting that column. Again note that I'm passing in the variable `z_val` to the argument `z`.



```
df %>%
  group_by(village) %>%
  do(y_hat = Y_hat_village2(., z = z_val)) %>%
  unnest()
#> # A tibble: 2 x 2
#>   village y_hat
#>   <chr>   <dbl>
#> 1 a         450
#> 2 b         700

```

However, `do` is being deprecated in favor of `purrr::map`, which I am still having trouble getting the hang of. In this case, you can group and nest, which gives a column of data frames called `data`, then map over that column and again supply `z = z_val`. When you unnest the `y_hat` column, you still have the original data as a nested column, since you wanted access to the rest of the columns still.



```
df %>%
  group_by(village) %>%
  nest() %>%
  mutate(y_hat = map(data, ~Y_hat_village2(., z = z_val))) %>%
  unnest(y_hat)
#> # A tibble: 2 x 3
#>   village data             y_hat
#>   <chr>   <list>           <dbl>
#> 1 a       <tibble [3 × 3]>   450
#> 2 b       <tibble [2 × 3]>   700

```

Just to check that everything works okay, I also passed in `z = 0` to check for 1. scoping issues, and 2. that other values of z work.



```
df %>%
  group_by(village) %>%
  nest() %>%
  mutate(y_hat = map(data, ~Y_hat_village2(., z = 0))) %>%
  unnest(y_hat)
#> # A tibble: 2 x 3
#>   village data             y_hat
#>   <chr>   <list>           <dbl>
#> 1 a       <tibble [3 × 3]>   800
#> 2 b       <tibble [2 × 3]>   300

```

"
"What does DateTimeStyles.RoundtripKind enumeration mean?
I was reading this answer [here](https://stackoverflow.com/questions/39508178/xml-delete-node-according-to-timestamp-c-sharp) where I came across this enumeration value `DateTimeStyles.RoundtripKind`. I'm trying to understand it now. I looked into MSDN docs [here](https://msdn.microsoft.com/en-us/library/91hfhz89%28v=vs.110%29.aspx?f=255&MSPPError=-2147217396) which says:



> 
> The DateTimeKind field of a date is preserved when a DateTime object
> is converted to a string using the ""o"" or ""r"" standard format
> specifier, and the string is then converted back to a DateTime object.
> 
> 
> 


The timestamp pointed by `input` variable in the post I referred in the previous paragraph is as below:



```
<timestamp time='2016-09-16T13:45:30'>

```

I ran her code and it works. Now it has become bit messy to correlate all the information I have:


1. The above time stamp contains some identifier `T`
2. The MSDN documentation talks about `o` and `r` format specifiers which it doesn't tell what it is?
3. If you go into dig more details on `DateTimeKind` enumeration on the MSDN link I've quoted above it says nothing about `o` and `r` format specifiers. [Here](https://msdn.microsoft.com/en-us/library/shx7s921(v=vs.110).aspx) is the link which says:



```
Member Name   |  Description
--------------------------------------------------------------------------------

Local         |  The time represented is local time.

```



---



```
Unspecified   |  The time represented is not specified as either local time or Coordinated Universal Time (UTC).

```



---



```
Utc           |      The time represented is UTC.

```


So can someone help me understand `DateTimeStyles.RoundtripKind` enumeration and how it works?


","So I was finally able to understand this and sharing the same information here if it can be helpful for others too:


First part is conversion of C# **DateTime** object into string. There are many format specifiers to do that but for us **r** and **o** format specifiers are of concern to us with regards to **DateTimeStyles.RoundtripKind**. You can see all date time format specifiers [here](https://msdn.microsoft.com/en-us/library/az4se3k1%28v=vs.110%29.aspx?f=255&MSPPError=-2147217396). See what happens when we do the conversion in code using these format specifiers:



```
//r corresponds to RFC 1123 format (GMT date time format)
var gmtDateTimeString = DateTime.Now.ToString(""r""); //gives Fri, 23 Sep 2016 15:39:21 GMT 

//o corresponds to ISO 8601 (Local date time format)
var localDateTimeString = DateTime.Now.ToString(""o""); //gives 2016-09-23T15:39:21.8899216+05:30

```

You can clearly see that string date time being output has the information embedded inside it which suggests:


- **Fri, 23 Sep 2016 15:39:21 GMT** is of **DateTimeKind.Utc** (**GMT** text is present)
- **2016-09-23T15:39:21.8899216+05:30** represents a date time of **DateTimeKind.Local** (**T** character is present as per **ISO 8601** standard)


Now comes the second part. If I've to convert the date time strings stored in `gmtDateTimeString` and `localDateTimeString` back to a date time object then we need to parse them. So with the help of **DateTimeStyles.RoundtripKind** enumeration value passed to **DateTime.Parse** API you actually signify that time zone information is already baked in the string and API parses the date time appropriately using that information.


Normally, when date time data is transferred over the wire in XML format then **ISO 8601** format is used which I saw in the post which I referred before posting the question in this thread. So while parsing such a date time string obtained from an XML document it was appropriate to use the **DateTimeStyles.RoundtripKind** to get the right date time value as per the time-zone information present in the string.


"
"What are the benefits of Apache Kafka's native binary TCP protocol over it's restful API?
As per Apache Kafka's documentation, Kafka uses binary TCP protocol in it's native API's communication but they have also provided URL based restful API for the languages which don't support Apache Kafka's native API. I was just wondering if there is any benefit of native binary TCP protocol (supported in native API) over restful URL based communication with broker node? And I was also thinking that will restful API still maintain `only once` property?


Edit:
The restful API guide is here: <https://www.confluent.io/blog/a-comprehensive-open-source-rest-proxy-for-kafka> which explains how to produce and consume Kafka's message over restful API


","There is no REST API included in Apache Kafka for producing or consuming messages as with the native Kafka protocol client implemented in Java.


There is a REST APIs in Apache Kafka for configuring Kafka Connect.


There are a number of third party REST Proxy implementations (such as the Confluent Kafka REST Proxy) which allow pub/sub over a REST interface but these are separate open source projects outside of Apache Kafka.


If you mean to ask what are the advantages to use the native Kafka Java Producer/Consumer API rather than these third party REST/HTTP Proxy implementation then these are some reasons:


One benefit is greater parallelism. A Kafka client will typically open up TCP connections to multiple brokers in the cluster and send or fetch data in parallel across multiple partitions of the same topic.


Another benefit is better network utilization as HTTP headers can add a lot of size to otherwise small messages while Kafka’s wire protocol is a compact binary protocol.


Kafka clients handle load balancing, failover, and cluster expansion or contraction automatically while REST clients typically require a third party load balancer to achieve the same functionality.


Kafka client can send their own authentication credentials for access control and bandwidth throttling (quotas) while all REST clients look to the kafka cluster as one Kafka client and therefore have common ACL privileges.


Kafka client libraries buffer and batch messages together into smaller numbers of Kafka produce or fetch requests while HTTP can only batch data if the programmer thought to publish them as a single batch.


The native Kafka protocol supports more than just what the producer/consumer api exposes. There is also an Admin API for creating topics, and modifying topic configurations. These functions are not (yet) exposed through the most popular REST Proxy implementations.


"
"How can I display a 2D binary matrix as a black & white plot?
I have a 2D binary matrix that I want to display as a black and white plot. For example, let's say I have a 4-by-4 matrix as follows:



```
1 1 0 1
0 0 1 0
1 1 0 1
1 0 0 0

```

How can this be plotted as a black and white matrix? Some of my input binary matrices are of size 100-by-9, so I would ideally need a solution that generalizes to different sized matrices.


","If you want to make a crossword-type plot [as shown here](https://www.mathworks.com/help/matlab/ref/pcolor.html#bua4ysn) (with grid lines and black and white squares) you can use the [`imagesc`](https://www.mathworks.com/help/matlab/ref/imagesc.html) function, a [gray colormap](https://www.mathworks.com/help/matlab/ref/colormap.html#input_argument_d0e131488), and modify the [axes properties](https://www.mathworks.com/help/matlab/ref/axes-properties.html) like so:



```
mat = [1 1 0 1; 0 0 1 0; 1 1 0 1; 1 0 0 0];  % Your sample matrix
[r, c] = size(mat);                          % Get the matrix size
imagesc((1:c)+0.5, (1:r)+0.5, mat);          % Plot the image
colormap(gray);                              % Use a gray colormap
axis equal                                   % Make axes grid sizes equal
set(gca, 'XTick', 1:(c+1), 'YTick', 1:(r+1), ...  % Change some axes properties
         'XLim', [1 c+1], 'YLim', [1 r+1], ...
         'GridLineStyle', '-', 'XGrid', 'on', 'YGrid', 'on');

```

And here's the image you should get:


[![enter image description here](https://i.stack.imgur.com/o0lPl.png)](https://i.stack.imgur.com/o0lPl.png)


"
"What are all those partition table type in gparted?
I was just curious about what are those available partition table types in gparted useful for. Is there any reason to use them instead of MBR or GPT ?


- atari
- aix
- amiga
- bsd
- dvh
- mac
- pc98
- sun


","### What are those available partition table types in `gparted` useful for?


The options you mention (other than Atari) are covered below (and I guess Atari is there to support Atari partition tables).



> 
> The options correspond to the various partitioning systems supported
>  in [`libparted`](https://unix.stackexchange.com/users/86440/stephen-kitt); there's not [much documentation](https://www.gnu.org/software/parted/manual/html_node/mklabel.html#mklabel), but looking
>  at [the source code](http://git.savannah.gnu.org/cgit/parted.git/tree/libparted/labels):
> 
> 
> - `aix` provides support for the volumes used in IBM’s AIX (which introduced what we now know as LVM);
> - `amiga` provides support for the Amiga’s RDB partitioning scheme;
> - `bsd` provides support for BSD disk labels;
> - `dvh` provides support for SGI disk volume headers;
> - `gpt` provides support for GUID partition tables;
> - `mac` provides support for old (pre-GPT) Apple partition tables;
> - `msdos` provides support for DOS-style MBR partition tables;
> - `pc98` provides support for [PC-98](http://people.freebsd.org/~kato/pc98.html) partition tables;
> - `sun` provides support for Sun’s partitioning scheme;
> - `loop` provides support for raw disk access (loopback-style) — I’m not sure about the uses for this one.
> 
> 
> As you can see, the majority of these are for older systems, and you
>  probably won’t need to create a partition table of any type other than
>  `gpt` or `msdos`.
> 
> 
> For a new disk, **I recommend `gpt`**: it allows more partitions, it
>  can be booted even in pre-UEFI systems (using `grub`), and supports
>  disks larger than 2 TiB (up to 8 ZiB for 512-byte sector disks).
>  Actually, if you don’t need to boot from the disk, I’d recommend not
>  using a partitioning scheme at all and simply adding the whole disk to
>  `mdadm`, LVM, or a zpool, depending on whether you use LVM (on top of
>  `mdadm` or not) or ZFS.
> 
> 
> 


Source [What are the differences between the various partition tables?](https://unix.stackexchange.com/a/289401), answer by [Stephen Kitt](https://unix.stackexchange.com/users/86440/stephen-kitt)


"
"Check if column exists, if not, add in MySQL through PHP
I'm using the following code to check if the column exists, and if it doesn't, add it:



```
mysql_query(""select $column from $table"") or mysql_query(""alter table $table add $column varchar (20)"");

```

But there's no change in the database. Any suggestions why?


P.S. The database is connected.


","You can change your SQL to...



```
//THIS IS BETTER BUT DONT USE THIS
$qry = ""ALTER IGNORE TABLE {$table} ADD {$column} VARCHAR(20);""

```

Instead use PHP PDO or MySQLi with prepared statements.
Instead of that legacy horror with concatinated unescaped strings.


**MySQLi Solution:**



```
$mysqli = new mysqli($cfg->host, $cfg->user, $cfg->password, $cfg->db);

if ($mysqli->connect_errno) {
    echo 'Connect failed: ', $mysqli->connect_error, '"" }';
    exit();
}

if ($stmt = $mysqli->prepare(""ALTER IGNORE TABLE ? ADD ? VARCHAR(20);"")) {
    $stmt->bind_param(""ss"", $table, $column);
    $stmt->execute();
    $stmt->close();
}
$mysqli->close();

```

"
"How to avoid memory leaks in Mule Applications?
Are there some special things that must be considered to avoid memory leaks in [Mule Applications](https://www.mulesoft.com/de/resources/esb/what-mule-esb)?


How can we avoid memory leaks in Mule Applications?


For example; Do we actually have to remove flow variables? What must be done done explicitly by the developers of the Mule Applications and what is done (automatically) by the [Mule Runtime](https://www.mulesoft.com/de/platform/mule) and the [JVM GC](https://stackoverflow.com/questions/3798424/what-is-the-garbage-collector-in-java)?


","A good way to get to the memory leak suspects would be to take a heap dump (of all the nodes) right after you start seeing a decline in memory reclaim post-major GC. There are multiple tools available that help analyze the memory leaks. 


There is [a great blog post](https://dzone.com/articles/enduring-black-fridays-with-mulesoft-apis) in the topic. This has summarized some memory-leak related issues like the following findings for example:


**Finding**: *Pooled memory manager generally grabs 10% of JVM heap and lives with it without releasing.* 
Fix: Switch the Grizzly Memory Manager implementation HeapMemoryManager. Note that HeapMemoryManager is the default implementation and is recommended by Grizzly for performance; albeit, Mule treats PoolMemoryManager implementation as the default.


Wrapper.conf changes:



```
wrapper.java.additional.<XX>=-Dorg.glassfish.grizzly.DEFAULT_MEMORY_MANAGER=org.glassfish.grizzly.memory.HeapMemoryManager

```

**Finding**: *Async logging was being used widely and associated Log4J was observed to be holding a lot of JVM memory.* The default setting of 256\*1024 slots was apparently too high. Since this RingBuffer does not grow or shrink, a high fixed size with each slot allocated as a separate object (RingBufferLogEvent), each holding a log event, could occupy a considerable amount of memory.


**Fix**: *Reduce the Log4J RingBuffer size to 128 in wrapper.conf or log4j2.xml*



```
wrapper.java.additional.<XX>=-DAsyncLoggerConfig.RingBufferSize=128

```

Or, in log4j2.xml:



```
<AsyncLogger name=""DebugLog"" level=""info"" includeLocation=""true"" ringBufferSize=""128"">

```

Memory leak due to default HazelCast implementation used for aggregator components (splitter-Aggregator pattern).


**Finding**: Heap analysis pointed memory being held up by default HazelCast objectstore implementation used in splitter-aggregator components used in specific flows. It appeared as if the store was not getting expired appropriately.


**Fix**: Custom Object store implementation (subclass of PartitionedInMemoryObjectStore) was written and TTL (TimeToLive) for entries explicitly defined.



```
@Override 
public void expire(int entryTTL, int maxEntries, String partitionName) throws ObjectStoreException
{
    super.expire(entryTTL, maxEntries, partitionName);
    if (getPrivatePartitionSize(partitionName) == 0) {
disposePartition(partitionName);
    }
}

```

**reference : <https://dzone.com/articles/enduring-black-fridays-with-mulesoft-apis>**


"
"How to add custom validator to paper-input?
Need to add a custom validator which does some complex validation based on the values of other fields in the html. 


Tried adding custom validator function as an attribute to the paper-input element but it **wont get called** at all. 



```
    <dom-module id='custom-ele'>
            <paper-input is=""iron-input"" id='input_1' label='Label_1' validator='validate_1'></paper-input>
            <paper-input is=""iron-input"" id='input_2' label='Label_2' validator='validate_2'></paper-input>
            ...
    </dom-module>
    <script>
    Polymer({

        is: 'custom-ele',

        validate_1: function() {
            //validation code
        },

        validate_2: function() {
           //validation code 
        }

    });
    </script>

```

","The validator has to implement `IronValidatorBehavior` (see [docs](https://elements.polymer-project.org/elements/iron-validator-behavior?active=Polymer.IronValidatorBehavior)). Also, if you want to validate automatically, you need to set the `auto-validate` attribute. To achieve your goal you could create a custom validator for each type of validation that you want to use. Alternatively, you could create a generic custom validator and set the validate function upon initialisation. Here's an example.



```
Polymer({

    is: 'custom-validator',

    behaviors: [
        Polymer.IronValidatorBehavior
    ]
});

<dom-module id='validation-element'>
    <template>
        <custom-validator id=""valid1"" validator-name=""validator1""></custom-validator>
        <custom-validator id=""valid2"" validator-name=""validator2""></custom-validator>
        <paper-input auto-validate id='input_1' label='Label_1' validator='validator1'></paper-input>
        <paper-input auto-validate id='input_2' label='Label_2' validator='validator2'></paper-input>
    </template>
</dom-module>
<script>

    Polymer({

        is: 'validation-element',

        validate1: function(value) {
            //validation code
            console.log(""validate1"");
            return value.length > 3;
        },

        validate2: function(value) {
            //validation code
            console.log(""validate2"");
            return value.length > 5;
        },

        ready: function() {
            this.$.valid1.validate = this.validate1.bind(this);
            this.$.valid2.validate = this.validate2.bind(this);
        }

    });

</script>

```

You can find a working example in this [plunk](http://plnkr.co/edit/ageC9c?p=preview).


"
"What is the difference between ""Import-Bundle"" and ""Require-Bundle""?
What is the difference between spring source dm server specific **Import-Bundle** and OSGi's **Require-Bundle**?   

I am confused whether to use Import-Bundle or Require-Bundle in my project.


","Import-Bundle is similar to Require-Bundle, it creates a complete dependency on the other bundle, including that bundle's dependencies. This *transitivity* is bad because you have no idea what you depend, creating the infamous ""[big ball of mud](http://www.laputan.org/mud/)"" problem we're so familiar with in Object oriented programming.


In OO, we've found a solution to this entanglement by using interfaces, they separate implementation from specification. OSGi is built around a similar albeit of an higher order concept of *service contracts*. These contracts (interfaces, permissions, helper classes) are stored in a package. In contract based programming you depend on the contracts, not the implementations. Ergo, an OSGi bundle should depend on packages since they represent the contracts.



```
 Import-Package               <=> interface
 Import-Bundle/Require-Bundle <=> implementation class

```

Import-Bundle is NOT OSGi, it is a proprietary Spring extension. It is a cleaner form for Require-Bundle; the uncleanliness was necessary to support some Eclipse use cases. The OSGi decided not to adopt this header since the Require-Bundle/Import-Bundle is fundamentally broken if you want to build systems from components.


"
"How does the predict function of StatsModels interact with roc\_auc\_score of scikit-learn?
I am trying to understand the `predict` function in Python statsmodels for a Logit model. Its documentation is [here](https://www.statsmodels.org/stable/generated/statsmodels.discrete.discrete_model.LogitResults.predict.html).


When I build a Logit Model and use `predict`, it returns values from 0 to 1 as opposed to 0 or 1. Now I read this saying these are probabilities and we need a threshold. [Python statsmodel.api logistic regression (Logit)](https://stackoverflow.com/questions/26528019/python-statsmodel-api-logistic-regression-logit)


Now, I want to produce AUC numbers and I use `roc_auc_score` from sklearn ([docs](https://scikit-learn.org/stable/modules/generated/sklearn.metrics.roc_auc_score.html)).


Here is when I start getting confused.


1. When I put in the raw predicted values (probabilities) from my Logit model into the `roc_auc_score` as the second argument `y_score`, I get a reasonable AUC value of around 80%. How does the `roc_auc_score` function know which of my probabilities equate to 1 and which equate to 0? Nowhere was I given an opportunity to set the threshold.
2. When I manually convert my probabilities into 0 or 1 using a threshold of 0.5, I get an AUC of around 50%. Why would this happen?


Here's some code:



```
m1_result = m1.fit(disp = False)

roc_auc_score(y, m1_result.predict(X1))

AUC: 0.80

roc_auc_score(y, [1 if X >=0.5 else 0 for X in m1_result.predict(X1)])

AUC: 0.50

```

Why is this the case?


","Your 2nd way of calculating the AUC is wrong; by definition, AUC needs *probabilities*, and not hard class predictions 0/1 generated after thresholding, as you do here. So, your AUC is 0.80.


You don't set a threshold yourself in AUC calculation; roughly speaking, as I have explained [elsewhere](https://stackoverflow.com/questions/47104129/getting-a-low-roc-auc-score-but-a-high-accuracy/47111246#47111246), the AUC measures the performance of a binary classifier *averaged across all possible decision thresholds*.


It would be overkill to explain again here the rationale and details of AUC calculation; instead, these other SE threads (and the links therein) will help you get the idea:


- [In Classification, what is the difference between the test accuracy and the AUC score?](https://stackoverflow.com/questions/60905517/in-classification-what-is-the-difference-between-the-test-accuracy-and-the-auc)
- [Advantages of AUC vs standard accuracy](https://datascience.stackexchange.com/questions/806/advantages-of-auc-vs-standard-accuracy)
- [Getting a low ROC AUC score but a high accuracy](https://stackoverflow.com/questions/47104129/getting-a-low-roc-auc-score-but-a-high-accuracy)
- [Comparing AUC, log loss and accuracy scores between models](https://stackoverflow.com/questions/58610117/comparing-auc-log-loss-and-accuracy-scores-between-models)


"
"jQuery Deferred vs jqXHR
I use:



```
var jqXHR = $.ajax(settings);
jqXHR.success(function(result){});
jqXHR.error(function(result){});
jqXHR.complete(function(result){});

```

But version 1.5 has added the deferred object.


Q: In general, when do you use success, error and complete methods vs. the new hotness of deferred then, done and fail?


","For `$.ajax()` and family `.success` is merely a synonym for Deferred's `.done`, and likewise `.error` is a synonym for `.fail`.


So in fact the examples you show are *already* deferred methods, but with different names.


`.complete` is *mostly* a synonym for the new jQuery 1.6 `.always`, and you can get the same effect using `$.then(cb, cb)`, which will cause `cb` to be invoked whether the AJAX call succeeds or not. I believe there are minor differences in which parameters are passed to the ""fail"" callbacks between the `.complete`, `.always` and `$.then` variants.


I personally prefer to use the Deferred version of those named functions, because then you don't need to worry about whether your deferred objects are jqXHRs or not. Only `jqXHRs` have `.success`, `.error`, and `.complete`, but *every* Deferred (including jqXHRs) has `.done`, `.fail` and `.always`.


**EDIT** it seems the jQuery devs agree with me - they've [announced](http://api.jquery.com/jQuery.ajax/#jqXHR) that `.success`, `.error` and `.complete` will be deprecated in jQuery 1.8


"
"How to fix ""Whitelabel Error Page"" on spring framework - GET request is not working
I'm trying to make a ""Hello World"" controller with spring boot and the GET request is not working. What can i do to fix this?


I used spring initializer at <https://start.spring.io/> and in the Dependencies I chose web


I've tried to use diffrent annotations like @GetMapping



```
package com.example.hello;

import org.springframework.web.bind.annotation.RequestMapping;
import org.springframework.web.bind.annotation.RestController;

@RestController
public class HelloController {

    @RequestMapping(""/"")
    public String helloWorld() {
        return ""Hello World"";
    }
}


```


```
<?xml version=""1.0"" encoding=""UTF-8""?>
<project xmlns=""http://maven.apache.org/POM/4.0.0"" xmlns:xsi=""http://www.w3.org/2001/XMLSchema-instance""
    xsi:schemaLocation=""http://maven.apache.org/POM/4.0.0 http://maven.apache.org/xsd/maven-4.0.0.xsd"">
    <modelVersion>4.0.0</modelVersion>
    <parent>
        <groupId>org.springframework.boot</groupId>
        <artifactId>spring-boot-starter-parent</artifactId>
        <version>2.1.3.RELEASE</version>
        <relativePath/> <!-- lookup parent from repository -->
    </parent>
    <groupId>com.example</groupId>
    <artifactId>demo</artifactId>
    <version>0.0.1-SNAPSHOT</version>
    <name>demo</name>
    <description>Demo project for Spring Boot</description>

    <properties>
        <java.version>1.8</java.version>
    </properties>

    <dependencies>
        <dependency>
            <groupId>org.springframework.boot</groupId>
            <artifactId>spring-boot-starter-web</artifactId>
        </dependency>

        <dependency>
            <groupId>org.springframework.boot</groupId>
            <artifactId>spring-boot-starter-test</artifactId>
            <scope>test</scope>
        </dependency>
    </dependencies>

    <build>
        <plugins>
            <plugin>
                <groupId>org.springframework.boot</groupId>
                <artifactId>spring-boot-maven-plugin</artifactId>
            </plugin>
        </plugins>
    </build>

</project>


```


```
package com.example.demo;

import org.springframework.boot.SpringApplication;
import org.springframework.boot.autoconfigure.SpringBootApplication;

@SpringBootApplication
public class DemoApplication {

    public static void main(String[] args) {
        SpringApplication.run(DemoApplication.class, args);
    }

}


```

I want the output to be ""Hello World"" but the actual output is ""Whitelabel Error Page""


","The issue is related to your project structure, by default Spring Boot will scan the components below your main application class.


In your case your main is located at `package com.example.demo;`, and your controller is located at `package com.example.hello;`.


I would recommend you to keep your controller in a new package below the `com.example.demo` package, like `com.example.demo.controller` as per structure mentioned in the [documentation](https://docs.spring.io/spring-boot/docs/current/reference/html/using-boot-structuring-your-code.html).


If you really want to use the `com.example.hello` package, you may need to add the `@ComponentScan(""com.example.hello"")`, which becomes hard to mantain in the future.


More info can be found [here](https://docs.spring.io/spring-boot/docs/current/reference/html/using-boot-using-springbootapplication-annotation.html).


"
"Button listener on a user control
Here's the problem:


I have a 'home button' with a listener



```
    private void Item_Tap(object sender, TappedRoutedEventArgs e)
    {
        if (this.Frame != null)
        {
            this.Frame.Navigate(typeof(Home));
        }
    }

```

If I built a User Control, using the 'home button' inside it, how can I accomplish the same navigation?


\*\* I'm near to the solution\*\*.
This is the user control:



```
public sealed partial class TopBarControl : UserControl
    {
        public Frame Frame { get; set; }
        public object Parameter { get; set; }
        public TopBarControl(Frame frame)
        {
            this.InitializeComponent();
            this.Frame = frame;
            var localSettings = Windows.Storage.ApplicationData.Current.LocalSettings;
            var dateAndTime = DateTime.Now;
            Date.Text = dateAndTime.Date.ToString(""MM/dd/yyyy"");
            User.Text = localSettings.Values[""userName""].ToString();
            //SectionTitle.Text = Parameter.ToString();
        }

        private void GoHome_Tap(object sender, TappedRoutedEventArgs e)
        {
            if (this.Frame != null)
            {
                this.Frame.Navigate(typeof(Main_Menu));
            }
        }
    }

```

on the xaml user control I have:



```
<Button Grid.Column=""0"" HorizontalAlignment=""Right"" BorderThickness=""0"">
    <Image Source=""/Assets/home_icon.png"" Tapped=""GoHome_Tap""/>
</Button>

```

Page associated to the user control



```
public WantedVehicles()
        {
            this.InitializeComponent();
        }

        private TopBarControl topBarControl;
        protected override void OnNavigatedTo(NavigationEventArgs e)
        {
            Tit.Text = e.Parameter.ToString();
            topBarControl.Frame = this.Frame;
            topBarControl.Parameter = e.Parameter;
        }

 protected override void LoadState(Object navigationParameter, Dictionary<String, Object> pageState)
        {
        }
protected override void SaveState(Dictionary<String, Object> pageState)
        {
        }

```

But the button doesn't work yet.


","I understand your problem. Here's a simple way to create a home button that you can add anywhere in your application and it will navigate home. No need to cascade events or anything like that. Just create a custom control. **This is different than a user control**. This is just sub-classing the button control. Far more clean an approach. 


Three easy steps.


Note: My custom control is going to be called HomeButton and my home page class is going to be called HomePage. Other than that, this should all be out-of-the-box for you.


First, define how your button will look:



```
<!-- custom home button -->
<Style TargetType=""local:HomeButton"" BasedOn=""{StaticResource BackButtonStyle}"">
    <Setter Property=""Template"">
        <Setter.Value>
            <ControlTemplate TargetType=""Button"">
                <Grid Height=""48"" Width=""48"">
                    <Ellipse Stroke=""White"" StrokeThickness=""2"" />
                    <TextBlock Text=""&#xE10F;"" FontSize=""20"" 
                               VerticalAlignment=""Center"" HorizontalAlignment=""Center"" />
                </Grid>
            </ControlTemplate>
        </Setter.Value>
    </Setter>
</Style>

```

Second, define your custom control like this:



```
public sealed class HomeButton : Button
{
    public HomeButton()
    {
        this.DefaultStyleKey = this.GetType();
        this.Click += HomeButton_Click;
    }

    void HomeButton_Click(object sender, RoutedEventArgs e)
    {
        var _Frame = Window.Current.Content as Frame;
        var _Type = typeof(HomePage);
        _Frame.Navigate(_Type);
    }
}

```

Third, and finally, add your home button to your app!



```
<local:HomeButton HorizontalAlignment=""Left"" VerticalAlignment=""Top"" Margin=""40"" />

```

Here's how it looks:


![enter image description here](https://i.stack.imgur.com/dzVG1.jpg)


It even works! :) If you want to style the button more, it will be easy to steal the styling of the BackButtonStyle in your /Common/StandardStyles.xaml file. Then you can have the same View States and all. Anyway, this will get you where you want.


So, this is home button that solves the problem you described in your question. It even solves it in a way that is less problematic than the solution you assumed you would get - since manually bubbling events across contexts can cause nightmares.


I hope this helps.


"
"Frame buffer object support
Do most OpenGL 2.0 and 2.1 graphics cards that still are in use support frame buffer objects (through the GL\_ARB\_framebuffer\_object or GL\_EXT\_framebuffer\_object extensions)?


","In my experience, they do.


Among nVidia, GPUs at least as far back as the GeForce FX 5xxx (which support OpenGL 2.0) have FBO support, and I suspect even older cards do.


Among ATI GPUs old enough to only support OpenGL 2.0, I have seen such GPUs as the HD 2400 and the X1300, and they all support FBOs.


Among Intel GPUs, I think that it is mainly the HD Graphics families that have OpenGL 2.0 support at all, and all the HD Graphics GPUs I've seen have FBO support. I have also seen some other GPUs with 2.0 and FBO support, including some versions of the 965, and something called the ""Eaglelake"". I'm not sure why only some 965s support OpenGL 2.0, though. It could be a driver issue.


I have, on the other hand, not yet found any 2.0-compatible GPUs that do *not* support FBOs.


I hope this purely empirical answer helps somewhat.


"
"C++: how to find max\_element using boost::range?
I am trying to return an iterator to the largest element in a filtered range. Here is what I have so far:



```
#include <boost/lambda/lambda.hpp>
#include <boost/range/adaptors.hpp>
#include <boost/range/algorithm.hpp>
#include <vector>
#include <iostream>

using namespace boost::adaptors;
using namespace boost::lambda;
using namespace std;

int main ()
{
  vector<double> x = {100, 150, 200, 110};
  auto it = boost::max_element(x | indexed(0) | filtered(_1>100)); /* problem here */
  cout << it.index() << endl;

  return 0;
}

```

I expected the code to print out the index in the vector x which has the largest element (ie 2), but unfortunately it does not compile (Linux 64bit, GCC 4.7.2), the problem being in the line indicated above. The first compilation error I get from the compiler (amongst others) is the following:


/boost/tuple/detail/tuple\_basic.hpp:396:36: error: assignment of read-only member ‘boost::tuples::cons::head’


Any ideas what I am doing wrong? Or how else I can achieve what I am trying to do? Thanks in advance!


EDIT:


Changing the problematic line to :



```
auto it = boost::max_element<boost::return_found>(x | sliced(1,4) |   filtered(boost::function<bool(double)>(_1>100)));

```

seems to return the iterator to the largest element. However, is there a way to check that the iterator is within the range? Comparing it with boost::end(x) gives me an error. The only thing I can think of is to return 



```
auto another_range = boost::max_element<boost::return_found_end>(x | sliced(1,4) |   filtered(boost::function<bool(double)>(_1>100)));

```

and check if boost::empty(another\_range). Is this the only option? Thanks.


","The specific error you've encountered appears because boost lambdas are not CopyAssignable. Here's a simpler way to achieve the same message:



```
auto f1 = _1 > 100;
auto f2 = f1;
f2 = f1; // same error

```

If you provide a CopyAssignable functor to `filtered`, boost.phoenix (which you should be using anyway, boost.lambda is on the road to deprecation in favor of phoenix), a hand-written struct, or the old faithful `std::bind2nd(std::greater<double>(), 100)`, this line compiles with clang++:


bind2nd demo: <http://liveworkspace.org/code/2xKZIf>


phoenix demo: <http://liveworkspace.org/code/18425g>


It fails with gcc due to some boost.concept check, which is probably a bug, but it's a moot point because the result of `filtered` is `boost::filtered_range`, whose iterators don't have the `.index()` member function.


EDIT in response to comment:
comparing iterator into filtered\_range with the iterator into the original vector wouldn't work. However, since you used vector, and since it's still accessible, you can compare addresses, since neither `indexed` nor `filtered` make copies



```
#include <vector>
#include <iostream>
#include <cassert>
#include <boost/range/adaptors.hpp>
#include <boost/range/algorithm.hpp>
#include <boost/phoenix.hpp>

using namespace boost::adaptors;
using namespace boost::phoenix::placeholders;

int main ()
{
    std::vector<double> x = {100, 150, 200, 110};
    auto it = boost::max_element( x | indexed(0) | filtered(arg1 < 110) );
    assert(&x[0] <= &*it && &*it < &x[0] + x.size());
    std::cout << ""Element "" << *it << "" is at index "" << &*it - &x[0] << '\n';
}

```

demo <http://liveworkspace.org/code/1zBIJ9>


Or, for a more general solution, you could transform your vector into a vector of pairs (when boost gets zip adaptor, it could be neatly zipped with counting\_range), and carry the original sequence index along with the value through all the transformations.


"
"Get Number of Rows from a Select statement
I have this:



```
$dbh = new PDO(""odbc:Driver={Microsoft Access Driver (*.mdb, *.accdb)};Dbq=$mdbFilename"", $username, $password);

$sql = ""SELECT * FROM this_table"";

$stmt = $dbh->query($sql);

//num of rows?

```

How do I get the number of rows returned from that SELECT statement?


Thanks all


","I have found a solution, using fetchAll and then using count on this array - which is what MySQL does anyway internally, a bit inefficient but it works for me.



```
$q = $db->query(""SELECT ..."");
$rows = $q->fetchAll();
$rowCount = count($rows);

```

From another question [Chad](https://stackoverflow.com/users/41665/chad-birch) provided this insight:



> 
> It seems as though the only reason
>  this was possible with MySQL is
>  because it internally fetched all the
>  result rows and buffered them, to be
>  able to give you this information. See
>  mysql\_unbuffered\_query(). If you use
>  that function instead of
>  mysql\_query(), the mysql\_num\_rows()
>  function will not work. If you really
>  need to know the number of rows while
>  using PDO, you can fetch all of the
>  rows from PDO into an array and then
>  use count().
> 
> 
> 


Hope this is useful to someone.


"
"Search for text files where two different words exist (any order, any line)
I'm looking for a way to search files where two word instances exist in the same file. I've been using the following to perform my searches up to this point:



```
find . -exec grep -l ""FIND ME"" {} \;

```

The problem I'm running into is that if there isn't exactly one space that between ""FIND"" and ""ME"", the search result does not yield the file. How do I adapt the former search string where both words ""FIND"" and ""ME exist in a file as opposed to ""FIND ME""?


I'm using AIX.


","With GNU tools:



```
find . -type f  -exec grep -lZ FIND {} + | xargs -r0 grep -l ME

```

You can do standardly:



```
find . -type f -exec grep -q FIND {} \; -exec grep -l ME {} \;

```

But that would run up to two `grep`s per file. To avoid running that many `grep`s and still be portable while still allowing any character in file names, you could do:



```
convert_to_xargs() {
  sed ""s/[[:blank:]\""\']/\\\\&/g"" | awk '
    {
      if (NR > 1) {
        printf ""%s"", line
        if (!index($0, ""//"")) printf ""\\""
        print """"
      }
      line = $0
    }'
    END { print line }'
}

export LC_ALL=C
find .//. -type f |
  convert_to_xargs |
  xargs grep -l FIND |
  convert_to_xargs |
  xargs grep -l ME

```

The idea being to convert the output of `find` into a format suitable for xargs (that expects a blank (SPC/TAB/NL in the `C` locale, YMMV in other locales) separated list of words where single, double quotes and backslashes can escape blanks and each other).


Generally you can't post-process the output of `find -print`, because it separates the file names with a newline character and doesn't escape the newline characters that are found in file names. For instance if we see:



```
./a
./b

```

We've got no way to know whether it's one file called `b` in a directory called `a<NL>.` or if it's the two files `a` and `b` in the current directory.


By using `.//.`, because `//` cannot appear otherwise in a file path as output by `find` (because there's no such thing as a directory with an empty name and `/` is not allowed in a file name), we know that if we see a line that contains `//`, then that's the first line of a new filename. So we can use that `awk` command to escape all newline characters but those that precede those lines.


If we take the example above, `find` would output in the first case (one file):



```
.//a
./b

```

Which awk escapes to:



```
.//a\
./b

```

So that `xargs` sees it as one argument. And in the second case (two files):



```
.//a
.//b

```

Which `awk` would leave as is, so `xargs` sees two arguments.


You need the `LC_ALL=C` so `sed`, `awk` (and some implementations of `xargs`) work for arbitrary sequences of bytes (even though that don't form valid characters in the user's locale), to simplify the *blank* definition to just SPC and TAB and to avoid problems with different interpretations of characters whose encoding contains the encoding of backslash by the different utilities.


"
"NuGet TypeScript tooling:Why the added version of typescript compiler is still not available?
**Context**


I am using VS 2017.3. Just created an ASP MVC project, then added two nuget packages, 


- Microsoft.TypeScript.Compiler v2.5.3
- Microsoft.TypeScript.MSBuild v2.5.3


When trying to configure the typescript build, I see that TypeScript version 2.5 is not available. 


**Question**


What am I missing? I thought that adding those nuget packages will the project (and build) independent from the development machine state of installation...


**What I've tried so far**


- Unloaded, reloaded the project
- Exit VS,
- restart VS Build the project


[![enter image description here](https://i.stack.imgur.com/0y8I9.png)](https://i.stack.imgur.com/0y8I9.png)


[![enter image description here](https://i.stack.imgur.com/Oa38n.png)](https://i.stack.imgur.com/Oa38n.png)


","
> 
> Why the added version of typescript compiler is still not available?
> 
> 
> 


You need download and install the typescripts SDK version 2.5.3 from the [download center](https://www.microsoft.com/en-us/download/details.aspx?id=55258) and restarted Visual Studio.


The NuGet package of `Microsoft.TypeScript.MSBuild` is used for **providing the TypeScript MSBuild task, and matching compiler version**.


[![enter image description here](https://i.stack.imgur.com/xSrxl.png)](https://i.stack.imgur.com/xSrxl.png)


Starting with Visual Studio 2017, we could have multiple TypeScript versions installed and choose specific versions for each of your projects, after you install `Microsoft.TypeScript.MSBuild` package, you can choose specific versions for your project, for example, I have install several versions of the Typescripts SDKs on my machine:


[![enter image description here](https://i.stack.imgur.com/MT5k9.png)](https://i.stack.imgur.com/MT5k9.png)


Then I install that package to my ASP MVC project with the version **2.3**, restart the Visual Studio, re-open project, on the Typescripts build tab, you will notice **the matching version of Typescript 2.3 is selected by default**:


[![enter image description here](https://i.stack.imgur.com/TkUNS.png)](https://i.stack.imgur.com/TkUNS.png)


Of course, you can manually select other version of Typescript.


"
"How to correctly undo a bad commit in P4V (Perforce)?
Ok, so I am using Perforce P4V to back up my work for a Unity project. One of my teammates checked in some changes to the metafiles which broke everything. No problem though right? That's the whole point of using P4. We can just revert that. Only... revert didn't work?


The behavior I am seeing is
File A was changed in changelist 1
File B was changed in changelist 2,
File C And A were changed in changelist 3


Let's say Changelist 3 contains the bad change
I clicked on changelist 2 in my history, then clicked get revision, and checked the Force Operation box.
changelist 2 being the last known good state what I expected to happen was to have all of my files restored to the state I was in when changelist 2 was submitted.


Instead, file C was reverted, but File A was not. It's like, since file A didn't change in changelist 2 it didn't bother to get that version.


So I am in a state where all of the unity metafiles are maimed and all prefab references are broken.


When that didn't work I tried using get a revision to go back to the most current state. Then using Backout. That similarly didn't work, metafiles still maimed. I then tried selecting the last known good state and rolling the entire project folder back to that state. Again, didn't work. But then again, I may have maimed my project so badly at that point that nothing would have worked.


The only way I have found that appears to correctly be reverting the files and restoring the broken links is manually selecting each file or folder and reverting it to the last good commit, which is different for each file/folder since they were added and changed in different commits.


What I don't understand is why the force get revision didn't do that on its own. Or what the ""correct"" way to undo a bad commit is.


I even tried deleting the entire assets folder then using get revision force to pull an entirely new copy from the server using the last known good commit. This appeared to work perfectly once, but when I tried to repeat it to verify my results it went back to losing all of the meta file links. The only dependable way of getting back into a good state appears to be manually force getting each file and folder to the individual last known good commit.


I have consigned myself to having to manually fix my blunder this time, but I'd really appreciate help to know how to do this the right way for the future.


","Use the `p4 undo` command.



```
p4 undo @BADCHANGELIST
p4 submit

```

That's all there is to it!


(There is a similar operation in P4V called ""Back Out"" -- I have encountered difficulties using this and prefer using the command line because it's easier, but YMMV.)


Note that `p4 undo` and `p4 revert` are entirely different commands! Many people get confused about this because `git` calls its ""undo"" command `git revert`, but a `p4 revert` is not the same as a `git revert` (it's more like `git reset --hard`). You can't use `p4 revert` to undo something that's already submitted; it's only for reverting your pending work to the latest submitted state from the depot.


I'll see if I can explain what went wrong with what you attempted:



> 
> Let's say Changelist 3 contains the bad change I clicked on changelist 2 in my history, then clicked get revision, and checked the Force Operation box. changelist 2 being the last known good state what I expected to happen was to have all of my files restored to the state I was in when changelist 2 was submitted.
> 
> 
> Instead, file C was reverted, but File A was not. It's like, since file A didn't change in changelist 2 it didn't bother to get that version.
> 
> 
> 


IIRC when you click on a changelist and use it to sync in P4V, it defaults to doing `p4 sync @CHANGE,CHANGE` rather than `p4 sync @CHANGE` -- as you've observed, this is very unhelpful if you're trying to ""roll back"" to an earlier state across the entire workspace. You can probably fix this by clicking checkboxes in the sync dialog somewhere. I believe if you click on a changelist in the ""Folder History"" view it uses the path (`p4 sync path/...@CHANGE`) which should work better.


Once you've managed to sync the workspace to the older state, if you want to make that state permanent, you need to open the files for `edit` (""check out"" in P4V), `sync` to schedule a resolve (""get latest"" in P4V), `resolve` (which is also ""resolve"" in P4V, make sure you accept `your` workspace version (P4V might call this ""accept target"", I'm not sure), and then `submit`. Again, I recommend just using `p4 undo` so you can skip all of those manual steps.


"
"ASP.NET Vertically wrapping a list generated by a repeater
I've got my repeater generating a list of links that need to appear in a certain order. Meaning I need my list to appear like so



```
-Item1 -Item4
-Item2 -Item5
-Item3

```

Every solution I've found involves knowing whats going to be in your list and setting classes where the list should break. My issue is that it could be anywhere from 1 to 18 items.


So my question is, is there a good, simple way to vertically wrap a list that's being dynamically generated using an ASP.NET repeater control?


","Here's an idea. This assumes that the number of items to display vertically before wrapping is a fixed number, or at least a number that can be determined by the server (see the VerticalWrapLimit constant). I.e. it is not related to the height of the users's browser.



```
<script runat=""server"">
    void Page_Load(object sender, EventArgs e)
    {
        rptVerticalWrap.DataSource = new[] {
            new { ID = 1, Name = ""ABC"" },
            new { ID = 2, Name = ""BCD"" },
            new { ID = 3, Name = ""CDE"" },
            new { ID = 4, Name = ""DEF"" },
            new { ID = 5, Name = ""EFG"" },
            new { ID = 6, Name = ""FGH"" },
            new { ID = 7, Name = ""GHI"" },
            new { ID = 8, Name = ""HIJ"" },
        };
        rptVerticalWrap.DataBind(); 
    }

    int count;
    const int VerticalWrapLimit = 5;
</script>

<!DOCTYPE html PUBLIC ""-//W3C//DTD XHTML 1.0 Transitional//EN"" ""http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"">

<html xmlns=""http://www.w3.org/1999/xhtml"" >
<head runat=""server"">
    <title></title>
</head>
<body>
    <form id=""form1"" runat=""server"">
    <div>

    <% count = 0; %>
    <div style=""float:left;"">
            <asp:Repeater ID=""rptVerticalWrap"" runat=""server"">
                <ItemTemplate>
                    <%
                        if (count % VerticalWrapLimit == 0 && count > 0)
                        {
                            %></div><div style=""float:left;""><%                     
                        }
                        count++;
                    %>
                    <div>
                        <asp:Label runat=""server"" Text=""ID: "" /><asp:TextBox runat=""server"" Text='<%# Eval(""ID"") %>' />
                        <asp:Label runat=""server"" Text=""Name: "" /><asp:TextBox runat=""server"" Text='<%# Eval(""Name"") %>' />
                    </div>
                </ItemTemplate>
            </asp:Repeater>
    </div>
    </div>
    </form>
</body>
</html>

```

For this sort of problem, I think what you need to do is identify what you want the resulting HTML to look like first, then write the ASP code to generate that.


In this case, I wrapped each vertical column of data in a with style=""float:left;"". I used a class scoped ""count"" variable in order to identify when I needed to close off the current div and start a new one. The count variable has to be a class-level variable because the contents of the Repeater's ItemTemplate has a different scope than the rest of the page.


"
"How a maven dependecy can be defined without version tag?
I have a pom which works without defining dependency version in pom works fine and another one without dependency version which does not work. 


The one which works:



```
<project>
    <parent>
        <artifactId>artifact-parent</artifactId>
        <groupId>group-parent</groupId>
        <version>1.0-SNAPSHOT</version>
    </parent>
    <dependencies>
        <dependency>
            <groupId>group-a</groupId>
            <artifactId>artifact-a</artifactId>
        </dependency>
        <dependency>
            <groupId>group-a</groupId>
            <artifactId>artifact-b</artifactId>
        </dependency>
    </dependencies>
</project>

```

This one which does not work:



```
<project>
    <dependencies>
        <dependency>
            <groupId>group-a</groupId>
            <artifactId>artifact-a</artifactId>
        </dependency>
        <dependency>
            <groupId>group-a</groupId>
            <artifactId>artifact-b</artifactId>
        </dependency>
    </dependencies>
</project>

```

The only thing differ in these two seems to be:



```
<parent>
    <artifactId>artifact-parent</artifactId>
    <groupId>group-parent</groupId>
    <version>1.0-SNAPSHOT</version>
</parent>

```

Second one does not work it seems fine to me but my question is why the first one works ?


Quoted from [maven pom reference](https://maven.apache.org/pom.html#Dependencies) :



> 
> This trinity represents the coordinate of a specific project in time,
>  demarcating it as a dependency of this project.
> 
> 
> 


So my question is how the first one is working?


","The main thing to notice here is:



```
<parent>
    <artifactId>artifact-parent</artifactId>
    <groupId>group-parent</groupId>
    <version>1.0-SNAPSHOT</version>
</parent>

```

Version of the dependency looks like to be defined in the parent pom. This can be something like this:



```
<project>
    <dependencyManagement>
        <dependencies>
            <dependency>
                <groupId>group-a</groupId>
                <artifactId>artifact-a</artifactId>
                <version>1.0</version>
            </dependency>
            <dependency>
                <groupId>group-a</groupId>
                <artifactId>artifact-b</artifactId>
                <version>1.0</version>
            </dependency>
        </dependencies>
    </dependencyManagement>
</project>

```

Again quoting from the [doc](https://maven.apache.org/guides/introduction/introduction-to-dependency-mechanism.html) :



> 
> `This is because the minimal set of information for matching a
>  dependency reference against a dependencyManagement section is
>  actually {groupId, artifactId, type, classifier}`.
> 
> 
> 


Here we did not need to define the `{type, classifier}` as it is same as default value which is as follows :



```
<type>jar</type>
<classifier><!-- no value --></classifier>

```

If this value differ form the default, you need to define it explicitly in both parent pom and child pom.


"
"Variables lack correlation, but have pattern
Below is the graph of two variables, X and Y, each representing count data. N=348. Note the scales of the axes:  

![http://i.imgur.com/tNGyTX5.jpg](https://i.stack.imgur.com/ih1yU.jpg)


Y is very approximately lognormal, but X has no decent fit (including Poisson, negative binomial, lognormal and gamma of the log transform).  

Spearman coefficient between X and Y is close to 0, and p-value to reject no correlation is very high.


From the plot, there appears to be no combinations of extreme values of both x and y.


When I log transform both X and Y, the following plot results:  

![enter image description here](https://i.stack.imgur.com/DVVHU.jpg)  

Clearly any appearance of pattern has disappeared.


My questions are:


- Why is there a lack of combinations of ""extreme"" values on the linear scale, but not on the log scale?
- Is there any significance to the lack of combination of extreme values on the linear scale, and is there anyway to investigate further?


The purpose of this study is exploratory.


","Consider flipping your questions around.


Begin with uncorrelated data - I generated this data randomly, so these variables are independent; my y is normal and my x is log(1+X1) where X1 is a mixture of several geometric distributions chosen to give a roughly similar appearance to your plot:


![enter image description here](https://i.stack.imgur.com/VpOLB.png)


The y variable is symmetric and the x-variable is mildly skew, but critically, neither of these variables is very long-tailed.


You then get many cases of relatively large values of X and Y together, because the probability that either is above its midrange (the center of the plot) is high, so the probability that both will be is also reasonably high (say, somewhere around 0.15-0.25), the product in this case of 0.5 for the y-variable and something a bit less that 0.5 for the x-variable.


What happens if you exponentiate two such variables, which are simply independent, shortish-tailed variables:


![enter image description here](https://i.stack.imgur.com/ENZIm.png)


You get something broadly similar to your first plot. It makes both variables - though still independent - strongly right skew (long tailed to the right), ... and that's where the appearance comes from.


Why does it look ""L"" shaped? Simply because high values of the X and Y variables are both relatively rare, and the combination of the two (due to their independence) rarer still. Because for each variable, almost all the other variable's values are far below its midrange, an extreme X or Y is likely to be associated with values of the other variable below/left-of the middle of the plot.


e.g. if 5% of each distribution is above the midrange, then about 0.25% of values (i.e. not quite 1 on average) will be in the upper right quadrant.


You see pretty much the same phenomenon with any two sufficiently right-skew variables that are independent, and in many that are close to independent. Here are two independent random variables (the absolute values of t-distributed random variables with 1.5 d.f. and different scales):


![enter image description here](https://i.stack.imgur.com/jzF5J.png)


"
"What is target in Python's sklearn coef\_ output?
When I do ridge regression using sklearn in Python, the coef\_ output gives me a 2D array. According to the [documentation](http://scikit-learn.org/stable/modules/generated/sklearn.linear_model.Ridge.html) it is (n\_targets, n\_features). 


I understand that features are my coefficients. However, I am not sure what targets are. What is this?


","The targets are the values you want to predict. The ridge regression can in fact predict more values for each instance, not only one. The `coef_` contain the coefficients for the prediction of each of the targets. It is also the same as if you trained a model to predict each of the targets separately.


Let's have a look at a simple example. I will use `LinearRegression` instead of `Ridge`, as `Ridge` shrinks the values of the coefficients and make it harder to understand.


First, we create some random data:



```
X = np.random.uniform(size=100).reshape(50, 2)
y = np.dot(X, [[1, 2, 3], [3, 4, 5]])

```

The first three instances in `X` are: 



```
[[ 0.70335619  0.42612165]
 [ 0.2959883   0.10571314]
 [ 0.33868804  0.07351525]]

```

The targets `y` for these instances are 



```
[[ 1.98172114  3.11119897  4.24067681]
 [ 0.61312771  1.01482915  1.41653058]
 [ 0.55923378  0.97143708  1.38364037]]

```

Notice, that `y[0] = x[0]+3*x[1]`, `y[1] = 2*x[0] + 4*x[1]` and `y[2] = 3*x[0] + 5*x[1]` (that's how we created the data with the matrix multiplication).


If we now fit the linear regression model



```
clf = linear_model.LinearRegression()
clf.fit(X, y) 

```

the `coef_`s are:



```
[[ 1.  3.]
 [ 2.  4.]
 [ 3.  5.]]

```

This exactly matches the equations we used to create the data.


"
"Compute the mean of two columns in a dataframe
I have a dataframe storing different values. Sample:



```
a$open  a$high  a$low   a$close

1.08648 1.08707 1.08476 1.08551
1.08552 1.08623 1.08426 1.08542
1.08542 1.08572 1.08453 1.08465
1.08468 1.08566 1.08402 1.08554
1.08552 1.08565 1.08436 1.08464
1.08463 1.08543 1.08452 1.08475
1.08475 1.08504 1.08427 1.08436
1.08433 1.08438 1.08275 1.08285
1.08275 1.08353 1.08275 1.08325
1.08325 1.08431 1.08315 1.08378
1.08379 1.08383 1.08275 1.08294
1.08292 1.08338 1.08271 1.08325

```

What I want to do, is creating a new column `a$mean` storing the mean of `a$high` and `a$low` for each row.


Here is how I achieved that:



```
highlowmean <- function(highs, lows){
  m <- vector(mode=""numeric"", length=0)
  for (i in 1:length(highs)){
    m[i] <- mean(highs[i], lows[i])
  }
  return(m)
}

a$mean <- highlowmean(a$high, a$low)

```

However I'm a bit new into R and in functionnal languages in general, so I'm pretty sure that there is a more efficient/simple way to achieve that.


How to achieve that the smartest way?


","We can use `rowMeans`



```
 a$mean <- rowMeans(a[,c('high', 'low')], na.rm=TRUE)

```

NOTE: If there are NA values, it is better to use `rowMeans`


For example



```
 a <- data.frame(High= c(NA, 3, 2), low= c(3, NA, 0))
 rowMeans(a, na.rm=TRUE)    
 #[1] 3 3 1

```

and using `+`



```
 a1 <- replace(a, is.na(a), 0)
 (a1[1] + a1[2])/2
#  High
#1  1.5
#2  1.5
#3  1.0

```

NOTE: This is no way trying to tarnish the other answer. It works in most cases and is fast as well.


"
"How to resize a tiff image with multiple channels?
I have a tiff image of size 21 X 513 X 513 where (513, 513) is the height and width of the image containing 21 channels. How can I resize this image to 21 X 500 X 375?


I am trying to use PILLOW to do so. But can't figure out if I am doing something wrong.



```
>>> from PIL import Image
>>> from tifffile import imread
>>> img = Image.open('new.tif')
>>> img

    <PIL.TiffImagePlugin.TiffImageFile image mode=F size=513x513 at 0x7FB0C8E5B940>

>>> resized_img = img.resize((500, 375), Image.ANTIALIAS)
>>> resized_img

    <PIL.Image.Image image mode=F size=500x375 at 0x7FB0C8E5B908>

>>> resized_img.save('temp.tif')

>>> img = imread('temp.tif')
>>> img.shape
  (500, 375)

```

The channel information is lost here. 


","Try using `tifffile` and `scikit-image`:



```
from tifffile import imread, imwrite
from skimage.transform import resize

data = imread('2009_003961_SEG.tif')
resized_data = resize(data, (375, 500, 21))
imwrite('multi-channel_resized.tif', resized_data, planarconfig='CONTIG')

```

The file `2009_003961_SEG.tif` linked in [comment98601187\_55975161](https://stackoverflow.com/questions/55974717/how-to-resize-a-tiff-image-with-multiple-channels/55975161#comment98601187_55975161) is not a multi-channel 513x513x21 image. Instead the file contains 513 images of size 513x21. The [`tifffile`](https://pypi.org/project/tifffile/) library will read the series of images in the file and return it as a numpy array of shape 513x513x21.


To resize the numpy array to 375x500x21, use [`skimage.transform.resize`](https://scikit-image.org/docs/stable/api/skimage.transform.html#skimage.transform.resize) (or `scipy.ndimage.zoom`). It might be faster to resize the 21 channels separately.


To write a TIFF file containing a single multi-channel image of size 375x500x21 using `tifffile`, specify the `planarconfig` parameter. Not many libraries or apps can handle such files.


"
"Using coefficient of variation in aggregate
I have a data frame with 50000 rows and 200 columns. There are duplicate rows in the data and I want to aggregate the data by choosing the row with maximum coefficient of variation among the duplicates using aggregate function in R. With aggregate I can use ""mean"", ""sum"" by default but not coefficient variation.


For example



```
aggregate(data, as.columnname, FUN=mean)

```

Works fine.


I have a custom function for calculating coefficient of variation but not sure how to use it with aggregate.



```
co.var <- function(x)
(
 100*sd(x)/mean(x)
)

```

I have tried



```
aggregate(data, as.columnname, function (x) max (co.var (x, data[index (x),])

```

but it is giving an error as object x is not found.


","Assuming that I understand your problem, I would suggest using `tapply()` instead of `aggregate()` (see `?tapply` for more info). However, a minimal working example would be very helpful.



```
co.var <- function(x) ( 100*sd(x)/mean(x) )

## Data with multiple repeated measurements.
## There are three things (ID 1, 2, 3) that 
## are measured two times, twice each (val1 and val2)
myDF<-data.frame(ID=c(1,2,3,1,2,3),val1=c(20,10,5,25,7,2),
  val2=c(19,9,4,24,4,1))

## Calculate coefficient of variation for each measurement set
myDF$coVar<-apply(myDF[,c(""val1"",""val2"")],1,co.var)

## Use tapply() instead of aggregate
mySel<-tapply(seq_len(nrow(myDF)),myDF$ID,function(x){
  curSub<-myDF[x,]
  return(x[which(curSub$coVar==max(curSub$coVar))])
})

## The mySel vector is then the vector of rows that correspond to the
## maximum coefficient of variation for each ID
myDF[mySel,]

```

**EDIT:**


There are faster ways, one of which is below. However, with a 40000 by 100 dataset, the above code only took between 16 and 20 seconds on my machine.



```
# Create a big dataset

myDF <- data.frame(val1 = c(20, 10, 5, 25, 7, 2),
  val2 = c(19, 9, 4, 24, 4, 1))
myDF <- myDF[sample(seq_len(nrow(myDF)), 40000, replace = TRUE), ]
myDF <- cbind(myDF, rep(myDF, 49))
myDF$ID <- sample.int(nrow(myDF)/5, nrow(myDF), replace = TRUE)

# Define a new function to work (slightly) better with large datasets

co.var.df <- function(x) ( 100*apply(x,1,sd)/rowMeans(x) )

# Create two datasets to benchmark the two methods
# (A second method proved slower than the third, hence the naming)

myDF.firstMethod <- myDF
myDF.thirdMethod <- myDF

```

Time the original method



```
startTime <- Sys.time()
myDF.firstMethod$coVar <- apply(myDF.firstMethod[,
  grep(""val"", names(myDF.firstMethod))], 1, co.var)
mySel <- tapply(seq_len(nrow(myDF.firstMethod)),
  myDF.firstMethod$ID, function(x) {
    curSub <- myDF.firstMethod[x, ]
    return(x[which(curSub$coVar == max(curSub$coVar))])
}, simplify = FALSE)
endTime <- Sys.time()

R> endTime-startTime
Time difference of 17.87806 secs

```

Time second method



```
startTime3 <- Sys.time()
coVar3<-co.var.df(myDF.thirdMethod[,
  grep(""val"",names(myDF.thirdMethod))])
mySel3 <- tapply(seq_along(coVar3),
  myDF[, ""ID""], function(x) {
    return(x[which(coVar3[x] == max(coVar3[x]))])
}, simplify = FALSE)
endTime3 <- Sys.time()

R> endTime3-startTime3
Time difference of 2.024207 secs

```

And check to see that we get the same results:



```
R> all.equal(mySel,mySel3)
[1] TRUE

```

There is an additional change from the original post, in that the edited code considers that there may be more than one row with the highest CV for a given ID. Therefore, to get the results from the edited code, you must `unlist` the `mySel` or `mySel3` objects:



```
myDF.firstMethod[unlist(mySel),]

myDF.thirdMethod[unlist(mySel3),]

```

"
"How to set up an SSH config-file for beginners
First I am fairly new to SSH. 


From [this question](https://stackoverflow.com/questions/56285972/can-you-and-is-it-advisable-to-use-the-same-ssh-key-for-github-and-gitlab-gitbuc?noredirect=1#comment99185285_56285972) I have seen the need and benefit of setting up an SSH config file. While I was doing my research I noticed that there is a lot to know on SSH and I also found out that I have been using SSH keys and not the SSH server. I have been using the keys to push my code to my hosted repos. 


So now my questions are(I am using windows 10):


1. Can I set up an SSH config file without a tool like openSSH, if so how do I do it?
2. Where in my computer is this config file stored?


","**Generally, in Windows machine, the SSH config file stored in the following location:** `/c/Users/PC_USER_NAME/.ssh/`


Just follow the steps in below (if you're using the Git Bash):


1. Go to the .ssh directory `/c/Users/PC_USER_NAME/.ssh/`, click right mouse button and choose ""Git Bash Here""
2. Create a file named ""config"" with the following command:



```
touch config

```

3. Now open the config file with the command:



```
nano config

```

4. Now write the following lines inside the config file


Let's assume you've created two files named `id_rsa_hub` for Github and `id_rsa_lab` for GitLab



```
# GITHUB
Host github.com
   HostName github.com
   PreferredAuthentications publickey
   IdentityFile ~/.ssh/id_rsa_hub

# GITLAB
Host gitlab.com
   HostName gitlab.com
   PreferredAuthentications publickey
   IdentityFile ~/.ssh/id_rsa_lab

```

"
"Overriding a WPF control
I have a WPF control ParentWPFControl from a third party that I would like to inherit from (let's call the child class ChildWPFControl). In the process, I plan to override some of the back-end logic and parts of the front end styles. I can do the former just fine but I have problems doing the latter.


I attempt to use a xaml <-> xaml.cs structure for the child country, but that appears to be not allowed with the following warning from VS:



```
Partial declarations of 'ChildWPFControl' must not specify different base classes

```

Now, I suppose I can write a ResourceDictionary XAML and define the front end there, but that becomes a problem if I want to add event handlers to the XAML (at least I couldn't find a way to do that)


Another alternative I have is to define the override template directly in the objects that use the ChildWPFControl but that makes the design less modular.


A final alternative I can think of is to make a xaml <-> xaml.cs pair that is a XAML style container and then force the ChildWPFControl to use the ControlTemplate defined within through the back end event handler.


Anyway, what I am looking for is an elegant and modular solution for my problem. Any advice would be welcomed.


Thanks


","There are a couple of steps necessary to completely override a WPF Control. Some are necessary some are optional depending on your needs. I will explain the two important ones for you:


**Creating a new default style**


Every WPF control has somewhere a default style which contains it visual representation and override properties. Now if you derive from control WPF still thinks you want to use this default style, to change that you change the DefaultStyle in a static constructor like this



```
class MyButton : Button
{
    static MyButton()
    {
        DefaultStyleKeyProperty.OverrideMetadata(typeof(MyButton), new FrameworkPropertyMetadata(typeof(MyButton)));
    }
}

```

Now if you use MyButton WPF tries to find a Style for MyButton, not for Button. OverridesDefaultStyle is a property in a style which might also be handy at some points. Usually these default styles should be placed in a theme related xaml.


**Event Handlers when overriding classes**


It is correct in a `ControlTemplate` or `Style` you can't use the syntactic sugar of using event like `Click=""OnClick""`. The point is, that the visual representation is decoupled from the logic part as much as possible. There are other ways though to overcome this, using the OnApplyTemplate method. By overriding this you ask the template ""Give me this control"" and then you just add your events there.



```
override OnApplyTemplate()
{
    var innerChild = Template.FindName(""PART_InnerChild"", this) as MyInnerControl;
    if(innerChild != null)
        innerChild.SomeEvent += OnSomeEvent;
}

```

Note: The name of these controls usually begin with a PART\_ by convention, this can be seen in WPF basic controls aswell. Its a nice way to tell the designers ""Without this control, the logic part might break"". There is also the attribute [TemplatePart](http://msdn.microsoft.com/en-us/library/system.windows.templatepartattribute.aspx) but it is not really important, WPF doesn't care about it. AFAIK Expression blend does some with it, personally i use it to tell other people what kind of inner controls are absolutely necessary to make this control work.


**Personal advice**


Deriving from a class is usually the last step we do when trying to customize controls. Because a lot of work is necessary to fully make it work and it can be limiting in reusability, we try to avoid it, for example a good alternatives are besides template overriding and styling; [attached behaviors](http://www.codeproject.com/Articles/28959/Introduction-to-Attached-Behaviors-in-WPF).


Lastly,
The whole subject is covered in a [nice MSDN](http://msdn.microsoft.com/en-us/library/ms745025.aspx) article.


Hope that helps


"
"slime-connect not working
I am an Emacs and Common Lisp novice. I have successfully installed SLIME in my Emacs, but I am unable to run it. `Alt` + `x` `slime-connect` gives the following options, accepting which gives the error pasted below the options.



```
Host: 127.0.0.1
Port: 4005

Connecting to Swank on port 4005...
make client process failed: connection refused, :name, SLIME Lisp, :buffer, nil,
                                                :host, 127.0.0.1, :service 4005,
                                                :nowait, nil

```

How do I get this to work?


","1. to fix the whitespace - escape it with double backward slash.
2. When you do `slime-connect` it will expect a Lisp image (Lisp runtime) to be up and listening on the server and port you connect to. Probably you didn't mean it.
3. `slime` will first launch `inferior-lisp-program`, and then connect to it. This is what you want to do, actually.


<https://github.com/quicklisp/quicklisp-slime-helper> this is something you might want to try (when you are done with the setup, or instead). Or just keep it for the time you feel more comfortable with more setup. The program in the link is a helper, that works through quicklisp - something you might want to get familiar with, a kind of package manager for Lisp code, and SLIME you run in Emacs. Possibly, if you will find that it is easier for you to 


- first install the Lisp you want.
- install quicklisp. (follow the instructions here: <http://www.quicklisp.org/beta/> with the exception that for Windows you might either use Cygwin, or instead of curl, just right-click on the link and do the ""save target as"" voodoo magic :). If you don't use SBCL, then --load may not be an option for your lisp. Replace it with something like: start your Lisp, type into REPL: `(load ""quicklisp.lisp"")`, provided your current working directory is the one where you saved the quicklisp.lisp file. To print out the current directory, if not sure: `(directory ""."")`.
- Read the README for quicklisp-slime-helper and follow the instructions in it. Basically, there's only one thing you want to configure after you install it via quicklisp, it is:


(load ""c:/full-path-to/quicklisp/slime-helper.el"")


in the .emacs file.


If you aren't familiar with \*nix path names, tilde means the current user's home directory, aka `$HOME`, you probably set it up as an environment variable when you installed Emacs. Note that `quicklisp-slime-helper` and `slime` packages are mutually exclusive (it's the same package, just wrapped differently), so if you have something like:



```
(add-to-list 'load-path ""/path/to/slime"")
(require 'slime)

```

Then, if you want to use the `quicklisp-slime-helper`, comment that away, and use only the configuration for one of them.


PS. Re' 



```
(add-hook 'inferior-lisp-mode-hook (lambda () (inferior-slime-mode t)))

```

This looks redundant, why do you need that / what is that? Where did you get `inferior-slime-mode`? That's something that doesn't look familiar to me. What happens if you position the point (caret) on it and hit `C-h f`?


"
"Responsive menu - reducing margin on resize
I'm trying to set up my horizontal menu so that it will 'squeeze' my navigation items together until the space between them has been used up, keeping an even space between each one as it reduces.


Once it gets to the point that the items are too close together my media query will kick in to re-arrange the menu.


I'm having trouble creating the 'squeezing' behaviour though. As soon as the window size is reduced, rather than squeezing my nav items up it's pushing them onto two lines.


Is there any way to get around this behaviour with just HTML and CSS?


Here's what i'm doing so far:



```
<nav id=""menu"">
    <ul>
        <li><a href=""#"">Menu item1</a></li>
        <li><a href=""#"">Menu item2</a></li>
        <li><a href=""#"">Menu item3</a></li>        
        <li><a href=""#"">Menu item4</a></li>
        <li class=""last""><a href=""#"">Menu item5</a></li>
    </ul>
</nav>

```

And the CSS:



```
#menu ul, #menu li { 
    margin:0; 
    padding:0; 
    list-style-type:none; 
}

#menu li { 
    float:left; 
    margin-right:4%; 
}

```

I've also put this into a fiddle so you can see what I mean - <http://jsfiddle.net/q3gjg/>. I don't want the items pushed onto two lines until there is no horizontal space left between them.


(I know this could be done using a series of media queries that reduced the margin between the items in steps, but this would result in a lot of media queries and doesn't seem very elegant).


Thank you


","You need a different approach. Don't use margins in this instance, use width and percentage.


DEMO <http://jsfiddle.net/kevinPHPkevin/q3gjg/1/>



```
#menu li { 
    float:left; 
    display: block;
    width: 20%;
    text-align: center;
}

```

**EDITED**


You state you do not wish to use media queries due to the amount you would need, but in this instance you would only need 2


DEMO <http://jsfiddle.net/kevinPHPkevin/q3gjg/3/>



```
@media (max-width: 440px) {
   #menu li { margin-right:2%;} 
}
@media (max-width: 400px) {
   #menu li { margin-right:0;} 
}

```

"
"Scraping dynamic content through Selenium?
I'm trying to scrap dynamic content from a [Blog](http://justgivemechocolateandnobodygetshurt.blogspot.com/) through Selenium but it always returns un rendered JavaScript.


To test this behavior I tried to wait till iframe loads completely and printed it's content which prints fine but again when I move back to parent frame it just displays un rendered JavaScript.


I'm looking for something in which I'm able to print completely rendered HTML content



```
from selenium import webdriver
from selenium.webdriver.support.ui import WebDriverWait
from selenium.webdriver.common.by import By
from selenium.webdriver.support import expected_conditions

driver = webdriver.Chrome(""path to chrome driver"")   
driver.get('http://justgivemechocolateandnobodygetshurt.blogspot.com/')

WebDriverWait(driver, 40).until(expected_conditions.frame_to_be_available_and_switch_to_it((By.ID, ""navbar-iframe"")))

# Rendered iframe HTML is printed.
content = driver.page_source
print content.encode(""utf-8"")

# When I switch back to parent frame it again prints non rendered JavaScript.
driver.switch_to.parent_frame()
content = driver.page_source
print content.encode(""utf-8"")

```

","The problem is - [the `.page_source` works only in the current context](https://w3c.github.io/webdriver/webdriver-spec.html#getting-page-source). There is that [""current top-level browsing context""](https://w3c.github.io/webdriver/webdriver-spec.html#dfn-current-top-level-browsing-context) notation..Meaning, if you would call it on a default content - you would not get the inner HTML of the child `iframe`elements - for that you would have to switch into the context of a `frame` and call `.page_source`.


In other words, to get the very complete HTML of the page including the page source of the iframes, you would have to switch into the iframe contexts one by one and get the sources separately.


See also:


- [Command Contexts](https://w3c.github.io/webdriver/webdriver-spec.html#h-command-contexts)
- [Switch To Frame](https://w3c.github.io/webdriver/webdriver-spec.html#switch-to-frame)




---


**Old answer:**


I would [wait](http://selenium-python.readthedocs.org/waits.html#explicit-waits) for at least one blog entry to be loaded before getting the `page_source`:



```
from selenium.webdriver.common.by import By
from selenium.webdriver.support.ui import WebDriverWait
from selenium.webdriver.support import expected_conditions as EC

wait = WebDriverWait(driver, 40)
wait.until(EC.visibility_of_element_located((By.CSS_SELECTOR, "".entry-content"")))

print(driver.page_source)

```

"
"Oracle RAC and sequences
I have various database applications that use sequences, I´m migrating these applications to Oracle RAC from 10g without RAC to 11g with RAC. I need ordered sequences and gaps are tolerated.


I'm thinking in cache sequences with order, I don´t know what are the effect in performance. Do you think this is a good option? What are your experience with sequences and RAC?


Thanks,


","Exactly what do you mean by ""ordered"" in this context?


By default, each node in the cluster has a separate cache of sequence numbers. So node 1 may be handing out values 1-100 while node 2 is handing out values 101-200. The values returned from a single node are sequential, but session A on node 1 may get a value of 15 while session B on node 2 gets a value of 107 so the values returned across sessions appear out of order.


If you specify that the sequence has to be ordered, you're basically defeating the purpose of the sequence cache because Oracle now has to communicate among nodes every time you request a new sequence value. That has the potential to create a decent amount of performance overhead. If you're using the sequence as a sort of timestamp, that overhead may be necessary but it's not generally desirable.


The overhead difference in practical terms is going to be highly application dependent-- it will be unmeasurably small for some applications and a significant problem for others. The number of RAC nodes, the speed of the interconnect, and how much interconnect traffic there is will also contribute. And since this is primarily a scalability issue, the practical effect is going to limit how well your application scales up which is inherently non-linear. Doubling the transaction volume your application handles is going to far more than double the overhead.


If you specify NOCACHE, the choice of ORDER or NOORDER is basically irrelevent. If you specify ORDER, the choice of CACHE or NOCACHE is basically irrelevent. So CACHE NOORDER is by far the most efficient, the other three are relatively interchangable. They are all going to involve inter-node coordination and network traffic every time you request a sequence value which is, obviously, a potential bottleneck. 


It would generally be preferrable to add a TIMESTAMP column to the table to store the actual timestamp rather than relying on the sequence to provide a timestamp order.


"
"How to Implement Error Handling
Even though I've programmed on a professional level for some years I still do not fully understand error handling. Although my applications work fine, the error handling isn't implemented at a professional level and is a mix and match of a number of techniques.


There is no structure behind my error handling. I'd like to learn and understand how it's implemented at a professional level. This is one area where I lack knowledge.


When should I use an exceptions and when should I return a success status, to be checked in the logic flow? Is it OK to mix exception and returning a status? 


I code in C# mainly.


","1. Use exceptions for exceptional things, the things you can't reasonably expect to encounter too often, things which indicate that something goes wrong. For example, if the network is down, it is an exceptional thing for a web server. If the database is unavailable, it means that something is wrong. If the configuration file is missing, it probably means that the user messed up with it.
2. Don't use exceptions to handle incorrect code. In order to check the correctness of the code, you should use either the assertions, or, in .NET Framework 4 and later, Code contracts (which replace assertions and have additional, particularly valuable features).
3. Don't use exceptions in non-exceptional cases. The fact that the user, when asked to enter a number, entered ""dog"" is not so exceptional to deserve an exception.
4. Be careful when choosing the types of exceptions. Create your own types when needed. Carefully chose the inheritance, keeping in mind that catching parents will catch the children as well. Never `throw Exception`.
5. Don't use return codes for errors. Error codes are easily masked, ignored, forgotten. If there is a error, either handle it, or propagate it to the upper stack.
6. In cases where a method is expected to return a error and the error is not exceptional, use enums, never error numbers. Example:



```
// Note that the operation fails pretty often, since it deals with the servers which are
// frequently unavailable, and the ones which send garbage instead of the actual data.
private LoadOperationResult LoadProductsFromWeb()
{
    ...
}

```

The meaning of `LoadOperationResult.ServerUnavailable`, `LoadOperationResult.ParsingError`, etc. is much more explicit than, say, remembering that code 12 means that the server is down, and code 13 — that the data cannot be parsed.
7. Use error codes when they refer to the common ones, known by every developer who works in the specific domain. For example, don't reinvent an enum value for HTTP 404 Not Found or HTTP 500 Internal Server Error.
8. Beware of booleans. Sooner or later, you will want to know not only whether a specific method succeeded or failed, but why. Exceptions and enums are much more powerful for that.
9. Don't catch every exception (unless you're at the very top of the stack). If you catch an exception, you should be ready to handle it. Catching everything is showing that you don't care if your code runs correctly. This may solve the ""I don't want to search right now how to fix this"", but will hurt you sooner or later.
10. In C#, never rethrow exceptions like this:



```
catch (SomeException ex)
{
    ...
    throw ex;
}

```

because you're breaking the stack. Do this instead:



```
catch (SomeException)
{
    ...
    throw;
}

```
11. Make an effort when writing exception messages. How many times I've seen something like `throw Exception(""wrong data"")` or `throw Exception(""shouldn't call this method in this context"")`. Other developers, including yourself six months later, would have no idea what data is wrong and why or why shouldn't we call some method in a context, nor which context precisely.
12. Don't show exception messages to the user. They are not expected for ordinary people, and often are even unreadable for developers themselves.
13. Don't localize exception messages. Searching the documentation for a localized message is exhausting and pointless: every message should be in English and English only.
14. Don't focus exclusively on exceptions and errors: logs are also extremely important.
15. In .NET, don't forget to include exceptions in XML documentation of the method:



```
/// <exception cref=""MyException"">Description of the exception</exception>

```

Including exceptions in XML documentation makes things much easier for the person who is using the library. There is nothing more annoying than trying to guess which exception could be possibly thrown by a method and why.


In this sense¹, [Java exception handling](http://www.ntu.edu.sg/home/ehchua/programming/java/J5a_ExceptionAssert.html) provides a stricter, better approach. It forces you to either deal with exceptions potentially thrown by the called methods, or declare in your own method that it can throw the exceptions you don't handle, making things particularly transparent.




---


¹ This being said, I find Java distinction between exceptions and errors pretty useless and confusing, given that the language has checked and unchecked exceptions. Luckily, .NET Framework has only exceptions, and no errors.


"
"Best practice when adding whitespace in JSX
I understand how (and [why](https://stackoverflow.com/questions/32686728/how-to-avoid-jsx-component-from-condensing-when-react-js-rendering-it)) to add a whitespace in JSX, but I am wondering what's best practice or if any makes any real difference?


**Wrap both elements in a span**



```
<div className=""top-element-formatting"">
  <span>Hello </span>
  <span className=""second-word-formatting"">World!</span>
</div>

```

**Add them on one line**



```
  <div className=""top-element-formatting"">
    Hello <span className=""second-word-formatting"">World!</span>
  </div>

```

**Add space with JS**



```
<div className=""top-element-formatting"">
    Hello {"" ""}
    <span className=""second-word-formatting"">World!</span>
</div>

```

","Because `&nbsp` causes you to have non-breaking spaces, you should only use it where necessary. In most cases, this will have unintended side effects.


Older versions of React, I believe all those before v14, would automatically insert `<span> </span>` when you had a newline inside of a tag.


While they no longer do this, that's a safe way to handle this in your own code. Unless you have styling that specifically targets `span` (bad practice in general), then this is the safest route.


Per your example, you can put them on a single line together as it's pretty short. In longer-line scenarios, this is how you should probably do it:



```
  <div className=""top-element-formatting"">
    Hello <span className=""second-word-formatting"">World!</span>
    <span> </span>
    So much more text in this box that it really needs to be on another line.
  </div>

```

This method is also safe against auto-trimming text editors.


The other method is using `{' '}` which doesn't insert random HTML tags. This could be more useful when styling, highlighting elements, and removes clutter from the DOM. If you don't need backwards compatibility with React v14 or earlier, this should be your preferred method.



```
  <div className=""top-element-formatting"">
    Hello <span className=""second-word-formatting"">World!</span>
    {' '}
    So much more text in this box that it really needs to be on another line.
  </div>

```

"
"What is the Minimum SDK and which one should I choose
First off, I know very little about android development, I am just getting started.


What is the Minimum SDK choice that you get when creating a project in android studio? Is there a downside to using an older one? And if I follow a tutorial is it essential that I use the same one so I can follow along?


Thanks.


","
> 
> What is the Minimum SDK choice that you get when creating a project in android studio?
> 
> 
> 


That is the oldest version of Android that you are willing to support. It is expressed in terms of an API level. You can see common API levels in the [Android dashboards](http://developer.android.com/about/dashboards/index.html), and the documentation will point out in many places where things need such-and-so an API level to work.



> 
> Is there a downside to using an older one?
> 
> 
> 


Less stuff in Android will be supported. In your case, since you are following a tutorial, choosing a lower `minSdkVersion` may cause some more complaints from your IDE, saying that such-and-so is not available on your chosen `minSdkVersion`.



> 
> And if I follow a tutorial is it essential that I use the same one so I can follow along?
> 
> 
> 


IMHO, that depends on your overall programming experience. If you are a veteran developer, and you want to play around with a lower `minSdkVersion`, go ahead, bearing in mind that the tutorial code might not run on that API level. If you are fairly new to programming overall, stick with what the tutorial tells you to do. If your concern is that your test device is not new enough for the tutorial, find a different tutorial, find a different device, or use the emulator instead of a device for testing this tutorial.


"
"Call back in Python
Could some explain how call back methods work, and if possible, give me an example in Python? So as far as I understand them, they are methods which are provided by the user of an API, to the API, so that the user doesn't have to wait till that particular API function completes. So does the user program continue executing, and once the callback method is called by the API, return to point in the program where the callback method was provided? How does a callback method essentially affect the 'flow' of a program?


Sorry if I'm being vague here.


","Callbacks are just user-supplied hooks. They allow you to specify what function to call in case of certain events. `re.sub` has a callback, but it sounds like you are dealing with a GUI, so I'll give a GUI example:


Here is a very simple example of a callback:



```
from Tkinter import *

master = Tk()

def my_callback():
    print('Running my_callback')

b = Button(master, text=""OK"", command=my_callback)
b.pack()

mainloop()

```

When you press the `OK` button, the program prints ""Running my\_callback"".


If you play around with this code:



```
from Tkinter import *
import time

master = Tk()

def my_callback():
    print('Starting my_callback')
    time.sleep(5)
    print('Ending my_callback')    

def my_callback2():
    print('Starting my_callback2')
    time.sleep(5)
    print('Ending my_callback2')    

b = Button(master, text=""OK"", command=my_callback)
b.pack()
b = Button(master, text=""OK2"", command=my_callback2)
b.pack()

mainloop()

```

you'll see that pressing either button blocks the GUI from responding until the callback finishes. So the ""user **does** have to wait till that particular API function completes"".


"
"Meteor device detection android or ios?
I have an meteor app that is deployed for both ios and android device and i want certain code to run on only ios device and not on android. I know that I can detect device using meteor device-detection package like



```
Meteor.Device.isPhone()

```

But is there any possible way can know if its an android or iOS device.


EDIT: I have created bundle using meteor cordova. 


","Here's a global helper that should do the trick as far as detecting iOS:



```
Template.registerHelper('isIOS',() => {
  return ( navigator.userAgent.match(/(iPad|iPhone|iPod)/g) ? true : false );
});

```

And another for Android:



```
Template.registerHelper('isAndroid',() => {
  return navigator.userAgent.toLowerCase().indexOf(""android"") > -1;
});

```

To use anywhere in client js:



```
Blaze._globalHelpers.isIOS()
Blaze._globalHelpers.isAndroid()

```

And of course, to use in html template markup:



```
{{#if isIOS}}...{{/if}}
{{#if isAndroid}}...{{/if}}

```

"
"Return null or something nullable from lambda in transformation mehtod WebFlux
In a Spring WebFlux chain I use a map operation which sometimes may return null and I get a warning :  

`Return null or something nullable from lambda in transformation mehtod`.  

I believe when data is null it doesn't actually map input to `null` but it will raise an exception instead.  

What is the best way to handle the scenario ?


Map method which is nullable :



```
public Pojo parseJson(String json) {
    try {
        // parse 
        return pojo;
    } catch (Exception e) {
        e.printStackTrace();
    }
    return null;
}

```

My reactive chain :



```
public Mono<Pojo> query(long id) {

    Integer value = idMapper.getValue(id);

    if (value != null) {
        return repo.query(value)

                        Parse Method
                             |
                             v
                .map(this::parse);
    }
    return null;
}

```

","When working in a functional/reactive world you should try to avoid all null checks and try to never return null from any methods.


Instead return `Optional<T>` when there is a risk for a null return, and return `Mono.error` when there is an error in a function that returns `Mono`. Or return `Mono.empty` if you just want to skip returning something.


By using optional you can rewrite the code to something much cleaner.



```
public Optional<Pojo> parseJson(String json) {
    // Return an optional if there is risk for a null.
    return Optional.ofNullable(new Pojo());
}

private final IdMapper idMapper = new IdMapper();
private final Repo repo = new Repo();

public Mono<Pojo> query(long id) {

    // By returning an optional from the idMapper 
    // you can chain on and avoid null checks.
    return idMapper.getValue(id).map(integer -> repo.query(integer)
            .map(s -> parseJson(s).map(Mono::just)
                    .orElse(Mono.empty())))
            .orElse(Mono.error(() -> new NotFoundException(""No Pojo with ID could be found"")));
}

class Pojo {
    // Some pojo class
}

class Repo {
    public Mono<String> query(long id) {
        return Mono.just(""Foo"");
    }
}

class IdMapper {

    public Optional<Integer> getValue(long id) {
        // Here return an Optional instead of null!
        return Optional.of(1);
    }
}

```

Here i return Optionals and a make some decisions to either return a Mono.empty or a Mono.error depending on what happens.


"
"Nodejs setImmediate() function Realtime usecase and example
Can some some one please explain the realtime usecase of setImmediate() function in nodejs.I've gone through many blogs but every where they gave console.log example with setImmediate.


","`setImmediate()` is useful when you want to defer running some code or calling some callback until AFTER the current event has been fully processed and control returned back to the event loop.


There are a number of different reasons why you might want to do that:


1. You want other (already pending) events to have a chance to get processed before you run some code.
2. You want to call a callback, but you want to call it asynchronously so that the callers code that comes after gets to run before you call the callback. There are places in the nodejs library where it does this to guarantee that a callback is always called asynchronously, even if the result is known synchronously. This creates programming consistency for the caller so that the callback is not called synchronously sometimes and asynchronously sometimes which can lead to subtle bugs.
3. When you're trying to not block the event loop for too long. You may run a chunk of code, then call `setImmediate()` and run the next chunk of code when the `setImmediate()` callback gets called and so on. This allows the processing of other events that arrive in the event loop in between your chunks of processing.


Set this article for an easy-to-read summary of `setImmediate()`, `setTimeout()` and `process.nextTick()`:


[Scheduling Execution in Node.js - Differences between setTimeout, setImmediate, and process.nextTick.](https://rclayton.silvrback.com/scheduling-execution-in-node-js)


Here are some examples from the nodejs and Express source code:


[Socket: Ensure that write is dispatched asynchronously](https://github.com/nodejs/node/blob/f1ae7ea343020f608fdc1ca77d9cdfe2c093ac72/lib/internal/js_stream_socket.js#L144)


[HTTP: Make sure parser is not deleted until stack has unwound](https://github.com/nodejs/node/blob/f1ae7ea343020f608fdc1ca77d9cdfe2c093ac72/lib/_http_common.js#L191)


[Express calls callback in layer routing only after stack has unwound](https://github.com/expressjs/express/blob/f3fa758af9664526ee18c58764590e48f559e6ae/lib/router/index.js#L196)


[Express calls callback `sendFile()` callback upon abort, only after stack has unwound](https://github.com/expressjs/express/blob/0a48e18056865364b2461b2ece7ccb2d1075d3c9/lib/response.js#L1065)


"
"Comparing to a hardcoded string with quotes in C++
I'm writing a c++ function to generate XML using TinyXML. I'd like to verify that a (relatively small) tree produced by my function gets turned into a string identical to a reference string.



```
// intended XML:
<element type=""correct"" />

```

It seems like the easiest way to do this comparison is to hardcode the reference string into the code:



```
//assignment
std::string intended = ""<element type=\""correct\"" />"";

```

However, the backslashes to escape the quotes prevent the comparison from succeeding.



```
#include <tinyxml.h>
#include <string.h>
TiXmlElement test = TiXmlElement(""element"");
test.SetAttribute(""type"", ""correct"");
TiXmlPrinter printer;
test.Accept(&printer);
ASSERT_EQ(intended, printer.CStr()); // gtests macro

```

output:



```
Value of: printer.CStr()
Actual: ""<element type=""correct"" />""
Expected: intended
Which is: ""<element type=\""correct\"" />""

```

","On the [googletest Primer](https://github.com/google/googletest/blob/1f87a0970dc0880c954ea4066fd7b72fad6fb089/googletest/docs/Primer.md) page I read that `ASSERT_EQ()` compares pointers. (which are only equal if they point to the same memory location). If you want to compare C strings, you should use `ASSERT_STREQ()`.



```
ASSERT_STREQ(intended, printer.CStr());

```

If you want to compare `std::string` objects, you can do it this way1:



```
ASSERT_EQ(intended, printer.Str());

```



---


1 Contributed by [johnfo](https://stackoverflow.com/users/2336991/johnfo) through a [comment](https://stackoverflow.com/questions/22364502/comparing-to-a-hardcoded-string-with-quotes-in-c/22364916#comment110366910_22364916).


"
"emacs23 / elisp: how to properly autoload this library?
I am upgrading to emacs23. I find that my emacs.el loads much more slowly. 


It's my own fault really... I have a lot of stuff in there. 


So I am also trying to autoload everything possible that is currently ""required"" by my emacs.el. 


I have a module that exposes 12 entry points - interactive functions I can call. 


Is the correct approach to have 12 calls to `autoload` in order to insure that the module is loaded regardless of which function I call? Are there any problems with this approach? Will it present performance issues?


If not *that* approach, then what? 


","What you really want is to get the autoloads generated for you automatically, so that your .emacs file remains pristine. Most packages have the `;;;###autoload` lines in them already, and if not, you can easily add them.


To manage this, you can put all the packages in a directory, say `~/emacs/lisp`, and in there have a file named `update-auto-loads.el` which contains:



```
;; put this path into the load-path automatically
;;;###autoload
(progn
  (setq load-path (cons (file-name-directory load-file-name) load-path)))

;;;###autoload
(defun update-autoloads-in-package-area (&optional file)
  ""Update autoloads for files in the diretory containing this file.""
  (interactive)
  (let ((base (file-truename
       (file-name-directory
        (symbol-file 'update-autoloads-in-package-area 'defun)))))
(require 'autoload)         ;ironic, i know
(let ((generated-autoload-file (concat base ""loaddefs.el"")))
  (when (not (file-exists-p generated-autoload-file))
    (with-current-buffer (find-file-noselect generated-autoload-file)
      (insert "";;"") ;; create the file with non-zero size to appease autoload
      (save-buffer)))
  (cd base)
  (if file
      (update-file-autoloads file)
    (update-autoloads-from-directories base)))))

;;;###autoload
(defun update-autoloads-for-file-in-package-area (file)
  (interactive ""f"")
  (update-autoloads-in-package-area file))

```

If you add `'update-autoloads-in-package-area` to your `kill-emacs-hook`, then the `loaddefs.el` will automatically be updated every time you exit Emacs.


And, to tie it all together, add this to your `.emacs`:



```
(load-file ""~/emacs/lisp/loaddefs.el"")

```

Now, when you download a new package, just save it in the `~/emacs/lisp` directory, update the loaddefs via `M-x update-autoloads-in-package-area` (or exit emacs), and it'll be available the next time you run Emacs. No more changes to your `.emacs` to load things.


See this question for other alternatives to speeding up Emacs startup: [How can I make Emacs start-up faster?](https://stackoverflow.com/questions/778716/how-can-i-make-emacs-start-up-faster)


"
"How to add external JS scripts to VueJS Components?
I've to use two external scripts for the payment gateways.


Right now both are put in the `index.html` file.


However, I don't want to load these files at the beginning itself.


The payment gateway is needed only in when user open a specific component (`using router-view`).


Is there anyway to achieve this?


Thanks.


","A simple and effective way to solve this, it's by adding your external script into the vue `mounted()` of your component. I will illustrate you with the [Google Recaptcha](https://www.google.com/recaptcha/intro/) script:



```
<template>
   .... your HTML
</template>

<script>
  export default {
    data: () => ({
      ......data of your component
    }),
    mounted() {
      let recaptchaScript = document.createElement('script')
      recaptchaScript.setAttribute('src', 'https://www.google.com/recaptcha/api.js')
      document.head.appendChild(recaptchaScript)
    },
    methods: {
      ......methods of your component
    }
  }
</script>

```

Source: <https://medium.com/@lassiuosukainen/how-to-include-a-script-tag-on-a-vue-component-fe10940af9e8>


"
"Comparing PhoneGap and HTML5 APIs for geolocation and online status
I'm developing an mobile apps using PhoneGap 2.7. I require the geolocation and network status APIs only. I found that those APIs are available from HTML5 directly using navigator.gelocation and navigator.onLine. So, should I use the PhoneGap APIs? What're their difference?


","As you can read on the [docs](http://docs.phonegap.com/en/2.9.0/cordova_geolocation_geolocation.md.html#Geolocation)



> 
> This API is based on the W3C Geolocation API Specification, and only executes on devices that don't already provide an implementation.
> 
> 
> 


it is used to provide backward compatibility if the device does not natively support the method.




---


`navigator.geolocation` is a standard which was introduced with HTML5. If you try `navigator.geolocation` on browsers which don't support this new standard you will receive `undefined`. PhoneGap now helps you and simulates `navigator.geolocation` for such (old) browsers. The api is [exactly the same](http://dev.w3.org/geo/api/spec-source.html) as in the W3C specification. If the browser supports the HTML5 geolocation api PhoneGap simply pipes it through. If not PhoneGap makes sure you can still use it (backward compatibility for unsupported devices). Request permissions on the device to use the geolocation service (like described on the PhoneGap docs) and stop worrying, PhoneGap's doing the rest.


See [caniuse](http://caniuse.com/#feat=geolocation) for browsers supporting the geolocation standard.


"
"Array indexOf() vs includes() perfomance depending on browser and needle position
Is there any advantage to Array.prototype.includes() over Array.prototype.indexOf() depending on browsers (Chrome, Firefox) and needle item position (at the begging, middle, ending of the array)?


[Array.prototype.includes vs. Array.prototype.indexOf](https://stackoverflow.com/questions/35370222/array-prototype-includes-vs-array-prototype-indexof)
There is no browser specific information, there is no position in the array specific information, and I don't ask about `NaN` value.


","I made a test using array with 10 000 numeric values, here is results:


Chrome:


- beginning
	- includes (22,043,904 ops/sec)
	- indexOf (136,512,737 ops/sec)
- middle
	- includes (8,361 ops/sec)
	- indexOf (31,296 ops/sec)
- ending
	- includes (4,018 ops/sec)
	- indexOf (95,221 ops/sec)


Firefox:


- beginning
	- includes (34,087,623 ops/sec)
	- indexOf (33,196,839 ops/sec)
- middle
	- includes (84,880 ops/sec)
	- indexOf (86,612 ops/sec)
- ending
	- includes (25,253 ops/sec)
	- indexOf (14,994 ops/sec)


So, `indexOf()` in Chrome works much faster than `includes()` in all positions.


In Firefox both `indexOf()` and `includes()` works almost similar.


"
"BizTalk receive pipeline failing due to disk space
I'm doing some performance testing on a BizTalk 2010 installation and one of the tests involves consuming a multi-gigabyte file. I'm receiving an error which is pretty straightforward, but it's missing some key information and I'm not able to figure it out so far based on the documentation.


The error is that the receive pipeline failed; ""Reason: there is not enough space on the disk.""


Hmmm. Which disk? It could be the partition on which windows is installed, the partition on which BizTalk is installed, the database server, etc. They all seem to have enough space to accomodate the file, but it's going to get much larger as it's transformed to XML, so it could probably be any of them.


I'm looking for suggestions on tracking down the source of the issue so I can get my tests back up and running.


","It's probably your temp folder that is causing your disk to run full.
When large files are getting processed (especially in mapping or disassemblers), the virtualstream will write parts of the message to disk in order to work in a streaming way and prevent memory leaks.  

It is a best practice (especially on production environments) to move the temp folder (the %temp% variable) to another disk than the c:\ drive. This is where BizTalk is storing all the large message fragments.


From the MSDN help:



> 
> By default, documents that are buffered to the file system during mapping are written to the %temp% directory of the BizTalk Server computer. Change the setting for the %temp% environment variable to a non-system disk to improve performance when buffering large messages to the file system during mapping.
> 
> 
> 


More information can be found here :<http://msdn.microsoft.com/en-us/library/aa560481.aspx>


"
"Java Regex Pattern Matcher.matches() before Matcher.find() Strange behaviour
I tried this example just interchanging two lines it gives different outputs why



```
String inputString = ""username@gmail.com""; 
String pattern=""([a-z]+@)([a-z]+)(\\.[a-z]+)"";
Pattern p = Pattern.compile(pattern); 
Matcher m = p.matcher(inputString); 

```

///changes happens here



```
if(m.find()) 
{
   String resultString = m.replaceAll(""$1xxxx$3""); 
   System.out.println(resultString); 
}

System.out.println(m.matches());//line to be changed

```

output :


username@xxxx.com


true



```
System.out.println(m.matches());//line changed     
if(m.find()) 
{
   String resultString = m.replaceAll(""$1xxxx$3""); 
   System.out.println(resultString); 
}

```

output :
true


","Extracted from [`Matcher.find`](http://docs.oracle.com/javase/7/docs/api/java/util/regex/Matcher.html#find%28%29) documentation



> 
> find
> 
> 
> public boolean find() 
> 
> 
> Attempts to find the next subsequence of the
>  input sequence that matches the pattern. This method starts at the
>  beginning of this matcher's region, **or, if a previous invocation of
>  the method was successful and the matcher has not since been reset, at
>  the first character not matched by the previous match**.
> 
> 
> If the match succeeds then more information can be obtained via the
>  start, end, and group methods.
> 
> 
> Returns: true if, and only if, a subsequence of the input sequence
>  matches this matcher's pattern
> 
> 
> 


So, since you called `Matcher.matches` which attempts to match the whole String, and you did not reset the matcher, it tried to find starting after the first match. As there is only one match, it does not find anything.


"
"AWS CloudWatch agent shows ""cwoc\_status"": ""stopped""
I have installed AWS CloudWatch agent on my EC2 instance and configured with the `amazon-cloudwatch-agent-config-wizard`. The AWS credentials are stored in `.aws/credentials` and picked up by the wizard during configuration.


Now I am checking the status of the agent, but I do not understand what `""cwoc_status"": ""stopped""` refers to?


Here the output of the agent status:



```
ubuntu@ip-address:~$ sudo /opt/aws/amazon-cloudwatch-agent/bin/amazon-cloudwatch-agent-ctl -m ec2 -a status
{
  ""status"": ""running"",
  ""starttime"": ""2021-02-09T05:37:12+00:00"",
  ""configstatus"": ""configured"",
  ""cwoc_status"": ""stopped"",
  ""cwoc_starttime"": """",
  ""cwoc_configstatus"": ""not configured"",
  ""version"": ""1.247347.3b250378""
}


```

","I finally discovered that `cwoc` refers to [AWS-OpenTelemetry Collector](https://aws.amazon.com/blogs/opensource/monitoring-application-health-and-performance-with-aws-distro-for-opentelemetry/), which I think has been installed by default by `amazon-cloudwatch-agent.deb` that I used to install the CWAgent on my EC2.
I found it by doing the following grep



```
ubuntu@ip-address:/opt/aws/amazon-cloudwatch-agent$ grep -r cwoc

```

which found the string in the file `/opt/aws/amazon-cloudwatch-agent/bin/amazon-cloudwatch-agent-ctl` where I found that reference to the folder `/opt/aws/amazon-cloudwatch-agent/cwagent-otel-collector`.


Googling `otel-collector` I found the OpenTelemetry Collector.


"
"sudo ls not showing hidden files on Linux
Today I discovered that `sudo ls` shows hidden files (that is, those that have names starting with `.`) on OS X. This surprised me so much that I [asked a question](https://superuser.com/q/931554/216502) about this behaviour, which I still find somewhat strange and unexpected.


[Turns out](https://superuser.com/a/931622/216502), this behaviour goes back to [2BSD in 1979](https://github.com/dspinellis/unix-history-repo/blob/BSD-2-Snapshot-Development/src/ls.c#L105).


Given that, now I’d like to ask the following question. Why doesn’t `ls` on Linux behave this way?  

Was there a period of time when `ls` on some other kinds of \*nixes had this behaviour? Are there any documents, commit messages, emails explaining who and why decided that this feature should not be copied at all or that it should be dropped if it was copied initially?


","The POSIX standard says:



> 
> ""Filenames beginning with a ( '.' ) and any associated information shall not be written out unless explicitly referenced, the -A or -a option is supplied, or an implementation-defined condition causes them to be written.""
> 
> 
> 


Being root is evidently not considered a condition which causes hidden files to be written by the GNU Coreutils implementation of `ls` that is commonly packaged in Linux distros.


There are good reasons not to have the behavior of programs influenced by global variables, like which user ID is in effect. A script developed as non-root will change behavior when run as root.


The hiding of files that begin with dot is not a security mechanism; it shouldn't be connected to security contexts. It conceals things that we normally don't want to see, like the `.git` directory among your `.c` source files or whatever. If you have read access to another user's directory, you can list their hidden files. *The dot hides items whose presence is expected and uninteresting, not whose presence is intended to be secret.*


Dotted directory entries other than `..` and `.` have no special operating system status; just `ls` treats them specially.


I just tried Solaris 10; its `ls` also has no such behavior. It is not a universal ""Unixism"", which explains why the POSIX requirement is worded that way.


"
"Understanding ldd output
How does `ldd` knows it's depending on `libc.so.6` ,not `libc.so.5` or `libc.so.7`?



```
libc.so.6 => /lib64/libc.so.6 (0x00000034f4000000)
/lib64/ld-linux-x86-64.so.2 (0x00000034f3c00000)

```

","It is recorded inside application binary itself (specified at compile time, more exactly at link step, done with `ld`):



```
$ readelf -d /bin/echo

Dynamic section at offset 0x5f1c contains 21 entries:
  Tag        Type                         Name/Value
 0x00000001 (NEEDED)                     Shared library: [libc.so.6]
...

```

(there are some additional columns for how elf does store information in dynamic section. but you can see that libc.so.6 is hardcoded with `.6` suffix because of [SONAME](http://en.wikipedia.org/wiki/Soname))


or even without any knowledge of ELF file format:



```
$ strings /bin/echo |grep libc.so
libc.so.6

```

To find, how does linker find a library (it is done at final step of compilation), use `gcc` option `-Wl,--verbose` (this asks gcc to pass option `--verbose` to `ld`):



```
$ gcc a.c -Wl,--verbose

...
attempt to open /usr/lib/gcc/i686-pc-linux-gnu/4.4.4/libc.so failed
attempt to open /usr/lib/gcc/i686-pc-linux-gnu/4.4.4/libc.a failed
attempt to open /usr/lib/gcc/i686-pc-linux-gnu/4.4.4/libc.so failed
attempt to open /usr/lib/gcc/i686-pc-linux-gnu/4.4.4/libc.a failed
attempt to open /usr/lib/gcc/i686-pc-linux-gnu/4.4.4/../../../libc.so succeeded
opened script file /usr/lib/gcc/i686-pc-linux-gnu/4.4.4/../../../libc.so
opened script file /usr/lib/gcc/i686-pc-linux-gnu/4.4.4/../../../libc.so
attempt to open /lib/libc.so.6 succeeded
/lib/libc.so.6

```

Linker doesn't know anything about `.digit` suffix, it just iterate over all library search directories trying to open `libLIBNAME.so` and `libLIBNAME.a`, where LIBNAME is a string after `-l` option. ( `-lc` option is added by default).


First success is `/usr/lib/libc.so` which itself is not a library, but a linker script (text file). Here is content from typical `libc.so` script:



```
$ cat /usr/lib/libc.so
/* GNU ld script
   Use the shared library, but some functions are only in
   the static library, so try that secondarily.  */
OUTPUT_FORMAT(elf32-i386)
GROUP ( /lib/libc.so.6 /usr/lib/libc_nonshared.a  AS_NEEDED ( /lib/ld-linux.so.2 ) )

```

So, script `/usr/lib/libc.so` is found earlier than actual library, and this script says, what file will be linked, `libc.so.6` in this case.


In more common case, `lib___.so` is symlink to some version like `lib___.so.3.4.5` and there is SONAME field filled in `lib___.so.3.4.5` which says to `ld` link not to `lib___.so` but to `lib___.so.3.4` which is another symlink to `lib___.so.3.4.5`. The `.3.4` name will be recorded in NEEDED field of binary.


"
"Are sockaddr\_in and sockaddr\_in6 still using sin\_len and sin6\_len?
So, basically the title says it all. I've been porting my unix socket C code to Windows, and apparently those structures do not have sin\_len or sin6\_len in windows. 


I'm using a union between sockaddr\_storage, sockaddr\_in and sockaddr\_in6 everywhere, and just using the correct member according to ss\_family. It would make sense that the socket library could just deduce the size according to the family, so the length field would indeed be redundant.


If I comment out the code that sets the length field, everything still works on OSX and linux, but that may be just an illusion, so I decided to ask here.
Is that variable deprecated, somehow? Can I safely stop using it, and rely on the socket implementation to use the family variable?


","The sin\_len field is not required by the POSIX specification.


Here's the relevant information from **Unix Network Programming - Vol 1. The Sockets Networking API, 3rd Edition**:


**3.2 Socket Address Structures**



> 
> The length member, **sin\_len**, was added with 4.3BSD-Reno, when support
> for the OSI protocols was added (Figure 1.15). Before this release,
> the first member was **sin\_family**, which was historically an unsigned
> short. Not all vendors support a length field for socket address
> structures and the POSIX specification does not require this member.
> 
> 
> 


Further Steven's provides the motivation behind the field:



> 
> Having a length field simplifies the handling of variable-length socket address structures.
> 
> 
> Even if the length field is present, we need never set it and need never examine it, unless we are dealing with routing sockets (Chapter 18). It is used within the kernel by the routines that deal with socket address structures from various protocol families (e.g, the routing table code).
> 
> 
> The four socket functions that pass a socket address structure from the process to the kernel, **bind, connect, sendto, and sendmsg**, all go through the **sockargs** function in a Berkeley-derived implementation (p. 452 of TCPv2). This function copies the socket address structure from the process and explicitly sets its **sin\_len** member to the size of the structure that was passed as an argument to these four functions. The five socket functions that pass a socket address structure from the kernel to the process, **accept, recvfrom, recvmsg, getpeername, and getsockname**, all set the **sin\_len** member before returning to the process.
> 
> 
> Unfortunately, there is normally no simple compile-time test to determine whether an implementation defines a length field for its socket address structures...We will see in Figure 3.4 that IPv6 implementations are required to define **SIN6\_LEN** if the socket address structures have a length field. Some IPv4 implementations provide the length field of the socket address structure to the application based on a compile-time option (e.g., **\_SOCKADDR\_LEN**).
> 
> 
> 


You'll want to evaluate the usage of sin\_len in your code. If it's just initializing it to 0, you can remove the code. If you're reading the value from the result of **accept, recvfrom, recvmsg, getpeername, or getsockname**, you'll unfortunately need some platform specific compile switches or switch to using the separate address length variable


"
"Hash Table Confusion - How much space is needed for Hash Table with a good (eg. Cryptographic) Hash Function?
I am learning about Hash Tables, Hash Maps etc. I have just implemented a Hash Table in C, with operations: `insert(HTable, key)`, `delete(HTable, key)`, `initialize(HTable)` and `search(HTable, key)`.


I would like to ask something. Since in a (proper) Hash Table the computed hashed indexes could be very large, doesn't this mean that the space consumed will be like `INT_MAX` (which is still O(n) of course), or more? I mean given the input element that we want to store in a hash table (ie insert it in), the insert() function would call the hash function which would then compute the hashed index for the element to go in. Thus it would use the hash function to find this index.


When we use the hash function to operate on the element, the hashed index could become very large. With a proper, for example cryptographic hash function, this index could become huge (they are using prime numbers with 300 digits - Diffie Hellman public key cryptography etc.), right? I know that in normal hash functions (such as the trivial ones beginners use to learn) we apply mod operation in order for the element to fit within the hash table's bounds, but in doing so, maybe we limit the hash function potential?


So to uniquely map an element to the hash table we must use a HUGE Hash Table. How are these cryptographic hash tables implemented? They must be completely secure, right? Even the Stack Overflow tag on ""cryptographichashfunction"" says that it is extremely unlikely to find two inputs that will map to the same element (as such the possibility of collisions is tiny). Wouldn't this require though a HUGE array to be stored in memory (or to disk)? Therefore, the memory consumption would be huge. 


Of course, the time complexity is not a problem. We just see the start address of the hash table / array add it with the index and just go that place in memory to get the value (O(1) - search principle of Hash Table).


Am i wrong somewhere? Is there something i'm missing? I hope i made myself clear. So to conclude, i would like confirmation on this. Does a good hash function require a huge array (Hash Table) and as such a very large amount of memory to be properly implemented? Is so much space justified, or is there something i don't quite get? Thanks.


","In general cryptographic hash values are *not* used for hash tables. Instead a fast hash is used. Of that hash value only as many bits may be used to tweak the size of the table. If multiple key values map to the same index then the values are stored in a separate structure, possibly with additional information to choose between the two.


It is not required that the hash output is unique; the hash function output would be too large and the table required would certainly not fit in memory. Besides that, cryptographic hashes are generally quite slow.


Cryptographic hash functions are usually build from operations also used in symmetric block ciphers. That means mixing and bitwise operators used in a large amount of rounds. Modular arithmetic, as used for e.g. RSA are commonly **not** used.


All in all, the main thing is that the index generated doesn't need to be unique. Usually if one hash leeds to multiple values they are stored in a list or set where the key can be compared by value.


"
"Entity Framework unique on child class
I am confused about how this kind of situation should be handled in Entity Framework.


I have a character base class:



```
public class Character
{
    [Key]
    public Guid Id { get; set; }
    public Realm Realm { get; set; }
    public Guid RealmId { get; set; }
    public string Name { get; set; }
}

```

Character has name and is bound to realm.


Character controlled by a player is definitely a character. It is bound to a player/user:



```
public class PlayerCharacter : Character
{        
    public User User { get; set; }
    public Guid UserId { get; set; }
}

```

NPC is also a character: it does not need to have any properties for a [MCVE](https://stackoverflow.com/help/mcve)



```
public class Npc : Character { }

```

Now, obviously, you would not want to have to characters on the same realm with the same name, but you can have as many `Orcs` as you need on a single realm, so we would need to define unique constraint on `PlayerCharacter`, for example via `.HasAlternateKey`:



```
public DbSet<PlayerCharacter> Characters { get; set; }
public DbSet<Npc> Npcs { get; set; }

protected override void OnModelCreating(ModelBuilder modelBuilder)
{
    base.OnModelCreating(modelBuilder);

    modelBuilder.Entity<PlayerCharacter>()
        .HasAlternateKey(""Name"", ""RealmId"");
}

```

Now we add migration:



```
$ dotnet ef migrations add 5
System.InvalidOperationException: A key cannot be configured on 
'PlayerCharacter' because it is a derived type. The key must be 
configured on the root type 'Character'.

```

Ok, how about a unique index?



```
modelBuilder.Entity<PlayerCharacter>()
    .HasIndex(""Name"", ""RealmId"")
    .IsUnique();

$ dotnet ef migrations add 5
Done. To undo this action, use 'ef migrations remove'

```

It worked! But wait:



```
migrationBuilder.CreateTable(
    name: ""Character"",
    columns: table => new
    {
        Id = table.Column<Guid>(nullable: false),
        Discriminator = table.Column<string>(nullable: false),
        Name = table.Column<string>(nullable: true),
        RealmId = table.Column<Guid>(nullable: false),
        // other props...
    },
    constraints: table =>
    {
        table.PrimaryKey(""PK_Character"", x => x.Id);
        // foreign keys...
    });

migrationBuilder.CreateIndex(
    name: ""IX_Character_Name_RealmId"",
    table: ""Character"",
    columns: new[] { ""Name"", ""RealmId"" },
    unique: true);

```

Now both player characters and npcs are unique by name! How can I set unique only in derived class?


","The only way you can do it currently (EF Core 2.0.1) is to use filtered unique index based on [*Discriminator* column](https://learn.microsoft.com/en-us/ef/core/modeling/relational/inheritance) value.


Unfortunately the [`HasFilter`](https://learn.microsoft.com/en-us/dotnet/api/microsoft.entityframeworkcore.relationalindexbuilderextensions.hasfilter?view=efcore-2.0#Microsoft_EntityFrameworkCore_RelationalIndexBuilderExtensions_HasFilter_Microsoft_EntityFrameworkCore_Metadata_Builders_IndexBuilder_System_String_) fluent API is not db agnostic. Here is the configuration for SqlServer:



```
modelBuilder.Entity<PlayerCharacter>()
    .HasIndex(""Name"", ""RealmId"")
    .IsUnique()
    .HasFilter($""Discriminator = '{nameof(PlayerCharacter)}'""); // <--

```

"
"Junit5 with IntelliJ and Gradle
## Trying to migrate my project to java8 + Junit5 using `IntelliJ 2017.2`


I have added `junit-jupiter-api` version `5.0.0-M6` 


and `junit-platform-launcher` version `1.0.0-M6`


Project structure is a default maven convention `src/test/java`


Found a couple articles about this but none of them did solve my issue.


**It runs nicely in a console, I presume this is something to do with the IntelliJ default JUnit Runner, or I am missing some dependencies?**




---


When I Run a single test class all works fine but when I select the directory and `Run all 'Tests' in Java` like I used to do then I encounter few errors.



```
WARNING: TestEngine with ID 'junit-jupiter' failed to discover tests
java.lang.NoSuchMethodError: org.junit.platform.engine.TestDescriptor.pruneTree()V

Aug 02, 2017 2:44:56 PM org.junit.platform.launcher.core.DefaultLauncher handleThrowable
WARNING: TestEngine with ID 'junit-vintage' failed to discover tests
java.lang.NoSuchMethodError: org.junit.platform.engine.EngineDiscoveryRequest.getDiscoveryFiltersByType(Ljava/lang/Class;)Ljava/util/List;

Aug 02, 2017 2:44:56 PM org.junit.platform.launcher.core.DefaultLauncher handleThrowable
WARNING: TestEngine with ID 'junit-jupiter' failed to discover tests
java.lang.NoSuchMethodError: org.junit.platform.engine.TestDescriptor.pruneTree()V

Aug 02, 2017 2:44:56 PM org.junit.platform.launcher.core.DefaultLauncher handleThrowable
WARNING: TestEngine with ID 'junit-vintage' failed to discover tests
java.lang.NoSuchMethodError: org.junit.platform.engine.EngineDiscoveryRequest.getDiscoveryFiltersByType(Ljava/lang/Class;)Ljava/util/List;

```

**Note**: I have not migrated any tests yet, all are Junit 4 syntax.


","Adding specific dependencies solve the problem. 


**NOTE: UPDATE INTELLIJ ABOVE 2017.2.0 AS THERE WAS A BUG WITH THE JUNIT LAUNCHER**


**OXYGEN** if you using eclipse.




---


Below dependency enables Junit5 parametrized tests which can be used instead of a **DataProvider**.



```
""org.junit.jupiter:junit-jupiter-params:5.0.0""
//for JUnit5 parametrized tests.

```



---


Junit5 **API**.



```
""org.junit.jupiter:junit-jupiter-api:5.0.0""
//JUnit5 API

```



---


Needed if you want to run legacy **JUnit4** tests without changing the syntax and imports.



```
""org.junit.vintage:junit-vintage-engine:4:12.0""
//for legacy JUnit4 tests

```

**EDIT: 07/2018 Match the version of the vintage runner to the jupiter version**




---


Needed if you want to run **JUnit5** tests with new syntax and imports.



```
""org.junit.jupiter:junit-jupiter-engine:5.0.0""
//for JUnit5 tests

```


> 
> java.lang.NoSuchMethodError: org.junit.platform.engine.EngineDiscoveryRequest.getDiscoveryFiltersByType(Ljava/lang/Class;)Ljava/util/List;
> 
> 
> 




---


**Launcher**.



```
""org.junit.platform:junit-platform-launcher:1.0.0
//to handle default launcher

```


> 
> Exception in thread ""main"" java.lang.NoSuchMethodError: org.junit.platform.commons.util.ReflectionUtils.getDefaultClassLoader()Ljava/lang/ClassLoader;
> 
> 
> 




---


Additional info how to install [JUnit5](http://junit.org/junit5/docs/current/user-guide/)




---


**Since version 4.6 for Gradle, there is no need for plugins anymore
Gradle supports Junit5 natively just do:**
And the version of the vintage runner is now same as the JUnit 5 version.



```
dependencies {

    testImplementation ""org.junit.jupiter:junit-jupiter-params:$junitVersion""
    testImplementation ""org.junit.jupiter:junit-jupiter-api:$junitVersion""

    testRuntimeOnly ""org.junit.vintage:junit-vintage-engine:$junitVersion""
    testRuntimeOnly ""org.junit.jupiter:junit-jupiter-engine:$junitVersion""
}

test {  
    useJUnitPlatform {
        includeEngines 'junit-jupiter', 'junit-vintage'
    }
}

```

"
"Convert cURL command to ansible
I have two cURL commands which I'm trying to convert to Ansible. They look like this:



```
curl -v -ik --header ""Content-Type: application/json"" --header ""X-Application-Username: my_username"" --header ""X-Application-Password: my_password"" --request PUT --data '{""enabled"":""False"",""persisted"" : true,""concurrentExecution"" : false, ""type"" : ""cron"",""schedule"" : ""0/10 * * * * ?"",""invokeService"" : ""provisioner"",""invokeContext"" : { ""action"" : ""someAction"",""source"" : ""path/to/resource""}}' ""https://localhost:8443/path/to/url/that/I/need""


curl -v -ik --header ""Content-Type: application/json"" --header ""X-Application-Username: my_username"" --header ""X-Application-Password: my_password"" --request PUT --data '{""enabled"":""True"",""persisted"" : true,""concurrentExecution"" : false, ""type"" : ""cron"",""schedule"" : ""0/10 * * * * ?"",""invokeService"" : ""provisioner"",""invokeContext"" : { ""action"" : ""someAction"",""source"" : ""path/to/resource""}}' ""https://localhost:8443/path/to/url/that/I/need""

```

I've tried to convert to ansible, but it isn't working. What I've come up with so far is this:



```
- name: Run cURL commands
  hosts: role_app_server[1]
  vars:
    endpoint: ""https://localhost:8443/path/to/url/that/I/need""
    cron_schedule: ""0/10 * * * * ?""
    invocation_context:
      action: ""someAction""
      source: ""path/to/resource""
  tasks:
    - name: First task
      uri:
        url: ""{{ endpoint }}""
        headers:
          Content-Type: ""application/json""
          X-Application-Username: ""my_username""
          X-Application-Password: ""my_password""
        method: PUT
        body:
          enabled: ""False""
          persisted: ""true""
          concurrentExecution: ""false""
          type: ""cron""
          schedule: ""{{ cron_schedule }}""
          invokeService: ""provisioner""
          invokeContext: ""{{ invocation_context | to_json }}""
        body_format: json
        validate_certs: no
    - name: 2nd task
      uri:
        url: ""{{ endpoint }}""
        headers:
          Content-Type: ""application/json""
          X-Application-Username: ""my_username""
          X-Application-Password: ""my_password""
        method: PUT
        body:
          enabled: ""True""
          persisted: ""true""
          concurrentExecution: ""false""
          type: ""cron""
          schedule: ""{{ cron_schedule }}""
          invokeService: ""provisioner""
          invokeContext: ""{{ invocation_context | to_json }}""
        validate_certs: no
        body_format: json

```

Can anyone spot what I'm doing wrong?


","We can diagnose this problem using a debugging endpoint like <https://httpbin.org>, which you can use to see exactly what data your requests are sending.


Using the following playbook, we make the request first with `curl` and then with the `uri` module, in each case storing the respond from `httpbin` so that we can compare them:



```
---
- name: Run cURL commands
  hosts: localhost
  gather_facts: false
  vars:
    endpoint: ""https://httpbin.org/put""
    cron_schedule: ""0/10 * * * * ?""
    invocation_context:
      action: ""someAction""
      source: ""path/to/resource""
  tasks:

    - name: First task (curl)
      command: >-
        curl -k -o output-curl1.json 
        --header ""Content-Type: application/json""
        --header ""X-Application-Username: my_username""
        --header ""X-Application-Password: my_password""
        --request PUT
        --data '{""enabled"":""False"",""persisted"" : true,""concurrentExecution"" : false, ""type"" : ""cron"",""schedule"" : ""0/10 * * * * ?"",""invokeService"" : ""provisioner"",""invokeContext"" : { ""action"" : ""someAction"",""source"" : ""path/to/resource""}}'
        ""{{ endpoint }}""

    - name: First task (uri)
      uri:
        url: ""{{ endpoint }}""
        headers:
          Content-Type: ""application/json""
          X-Application-Username: ""my_username""
          X-Application-Password: ""my_password""
        method: PUT
        body:
          enabled: ""False""
          persisted: ""true""
          concurrentExecution: ""false""
          type: ""cron""
          schedule: ""{{ cron_schedule }}""
          invokeService: ""provisioner""
          invokeContext: ""{{ invocation_context | to_json }}""
        body_format: json
        validate_certs: no
        return_content: true
      register: output1

    - copy:
        content: ""{{ output1.content }}""
        dest: ./output-task1.json

```

You can just manually inspect the results, but it's probably easier if we combine `jq` and `diff` to highlight the differences:



```
$ diff -u <(jq -Sr .data output-curl1.json | jq -S .) <(jq -Sr .data output-task1.json | jq -S .)
--- /dev/fd/63  2019-04-07 18:00:46.597729568 -0400
+++ /dev/fd/62  2019-04-07 18:00:46.599729606 -0400
@@ -1,12 +1,9 @@
 {
-  ""concurrentExecution"": false,
+  ""concurrentExecution"": ""false"",
   ""enabled"": ""False"",
-  ""invokeContext"": {
-    ""action"": ""someAction"",
-    ""source"": ""path/to/resource""
-  },
+  ""invokeContext"": ""{\""action\"": \""someAction\"", \""source\"": \""path/to/resource\""}"",
   ""invokeService"": ""provisioner"",
-  ""persisted"": true,
+  ""persisted"": ""true"",
   ""schedule"": ""0/10 * * * * ?"",
   ""type"": ""cron""
 }

```

This highlights a few differences. The biggest one appears to be the content of the `invokeContext` attribute. When using `curl`, the value of that attribute is a JSON object:



```
$ jq -r .data output-curl1.json | jq .invokeContext
{
  ""action"": ""someAction"",
  ""source"": ""path/to/resource""
}

```

But when using the `uri` module, the value of `invokeContext` is a string:



```
$ jq -r .data output-task1.json | jq .invokeContext
""{\""action\"": \""someAction\"", \""source\"": \""path/to/resource\""}""

```

This is happening because you're passing the value of your `invocation_context` variable through the `to_json` filter:



```
invokeContext: ""{{ invocation_context | to_json }}""

```

That means when the `body` data structure in your task is serialized into JSON, the `invokeContext` -- which you have *already* converted into a JSON string -- gets double-converted. You want this:



```
invokeContext: ""{{ invocation_context }}""

```

You also have some boolean vs. string conflicts.


When using `curl` you're setting the `persisted` attribute to boolean `true`, but in your Ansible task you're setting it to the string value `""true""`. Instead of:



```
persisted: ""true""

```

You want:



```
persisted: true

```

And lastly, you have the same issue with the `concurrentExecution` attribute, which should be:



```
concurrentExecution: false

```

With all these changes, the first task becomes:



```
- name: First task (uri)
  uri:
    url: ""{{ endpoint }}""
    headers:
      Content-Type: ""application/json""
      X-Application-Username: ""my_username""
      X-Application-Password: ""my_password""
    method: PUT
    body:
      enabled: ""False""
      persisted: true
      concurrentExecution: false
      type: ""cron""
      schedule: ""{{ cron_schedule }}""
      invokeService: ""provisioner""
      invokeContext: ""{{ invocation_context }}""
    body_format: json
    validate_certs: no
    return_content: true
  register: output1

```

And if we repeat our earlier `diff` command, we see that the data send by `curl` and by the `uri` module is identical:



```
$ diff -u <(jq -Sr .data output-curl1.json | jq -S .) <(jq -Sr .data output-task1.json | jq -S .)
$ # no output from the previous command

```

"
"how do i test cookie expiry in rails rspec
There is a lot of confusion about setting cookies in rspec
<http://relishapp.com/rspec/rspec-rails/v/2-6/dir/controller-specs/file/cookies>


in your controller, normally you can write



```
cookies['transaction_code'] = { :expires => 300.seconds.from_now, :value => c }

```

but in rspec i can only write



```
request.cookies['transaction_code'] = transaction_code

```

if i say 



```
request.cookies['transaction_code'] = { :expires => 300.seconds.from_now, :value => c }

```

i get the hash back as value of cookies['transaction\_code'] in my controller.


Now my question is: how do i set/test cookie expiry then in an rspec controller test example? 


UPDATE: On seconds thought: What i mean is: how do i test if the controller is reacts to an expired cookie as expected, but in fact an expired cookie is just like no cookie if i trust cookie implementation, which i should do, so after all maybe my question makes no sense. If this is the case, i need to test if (another) controller action sets an expiring cookie correctly, but how do i do it if cookies['transaction\_code'] in the test only gives the value back?


","Browsers [do not send cookie attributes back to the server](http://en.wikipedia.org/wiki/HTTP_cookie#Cookie_attributes). This is why you can only send the key-value pair to the action.


Since you can assume that Rails, Rack and browsers do the right thing with the arguments, all you really need to test is the arguments your code is passing to the `CookieJar`. 


To test that expiry is being set properly *in the controller setting the cookie*, you could stub out the `#cookies` method and make sure the right settings are being passed to it. 



```
# app/controllers/widget_controller.rb
...
def index
    cookies[:expiring_cookie] = { :value   => 'All that we see or seem...', 
                                  :expires => 1.hour.from_now }
end
...

# spec/controllers/widget_controller_spec.rb
...
it ""sets the cookie"" do
  get :index
  response.cookies['expiring_cookie'].should eq('All that we see or seem...')
                                               # is but a dream within a dream.
                                               #                - Edgar Allan Poe
end

it ""sets the cookie expiration"" do
  stub_cookie_jar = HashWithIndifferentAccess.new
  controller.stub(:cookies) { stub_cookie_jar }

  get :index
  expiring_cookie = stub_cookie_jar['expiring_cookie']
  expiring_cookie[:expires].to_i.should be_within(1).of(1.hour.from_now.to_i)
end
...

```

Testing much more than this is boiling the ocean. At some point, you have to assume that the stack you are sitting on (e.g., Rails, Rack, the web server, TCP/IP, OS, web browsers, etc) works properly and focus on the code you control.


"
"How do I get to the root cause of high Deferred Procedure Calls?
I've got a Dual core processor, and one of the two is constantly at 100%. Looking in ProcessExplorer shows me that it's Deferred Procedure Calls. Reading around the net seems to give me tons of different answers.


Is it possible to set out a couple of steps to trying to narrow down what the problem might be in my case?


**Update 1:** FWIW, the problem persists even in Safe Mode.


**Update 2:** I unplugged everything I could from the back of the PC, and that bought me 40% more free processor. I also downloaded the [RATTV3 tool](http://www.microsoft.com/whdc/DevTools/tools/RATT.mspx), but for some reason on my machine it's not giving me a driver-by-driver breakdown. There's a good description of both DPCLatencyChecker and RATTV3 [here](http://forum.cockos.com/showthread.php?t=30896).


**Update 3:**, [LatencyMon](http://www.resplendence.com) (see my answer below) tells me it's `nvstor32.sys` - which is NVidia's SATA driver - with times of about 5300 µs. 


**Update 4:** The plot thickens, while pondering whether to try booting of a recovery disk (to see if it's really drivers, and not a hardware problem), I noticed that the DVD/CD player wasn't working (i.e. Not even opening the door when I press the button). Given that the machine just got back from having the motherboard replaced, I figured maybe they'd forgotten to plug it in. I opened the box, everything seemed ok, but I unplugged and plugged it back in again. On reboot, everything was fine - no more DPC (highest now 300µs)!


**Update 5:** On following day, problem back, CD player not working again, even the cursor in the password textbox is blinking in slow motion... Tried unplugging everything I could think of, and on second reboot, worked again (as at Update2). Next time I'm going to try unplugging the CD player completely...


**Update 6:** Just noticed System event log has `nvstor32.sys` giving an error saying `Parity error detected in \Device\RaidPort0`, then a warning about sending a reinitialisation. Now just have to work out which one `RaidPort0` is... (note, I've got no RAID set-up, it's just a bog standard Acer). Oh, and my Avast setup apparently got killed when I did a System Rollback (or whatever it's called), because it won't start (RPC error), won't uninstall (setiface error has occurred). 


**Update 7:** Finally got time to reboot with DVD unplugged. No more DPC problems! (lots of page faults though, but that's for later). Next step: work out if it's the cable, or the DVD player.


**Update 8:** Borrowed a SATA cable, booted with it, no problems. CD/DVD player works, no DPC problems with `nvstor32.sys`, no processors blocked. Happy end... almost: I've still got problems with Avast, apparent DPC problems with `storport.sys` at start up (maybe normal for USB?), and lots of hard page faults. But those will be the subject of other questions.


**Postscript:** I recently started having the same problem, and using the same method, managed to track it down to a USB stick (the one I was using for ReadyBoost) being shot. 


","Here is the story of how I found the cause of my high DPC latency.




---


My system was experiencing clicks and pops during sound playback. I knew this meant that something in kernel mode was hogging the CPU. My first thought was to poke around [Process Explorer](https://docs.microsoft.com/en-us/sysinternals/downloads/process-explorer), and see if anything looked out of place. The only thing that caught my attention was an excessive amount of time spent performing *Deferred Procedure Calls* (DPCs):


[![Screenshot of Process Explorer showing high DPC time](https://i.stack.imgur.com/JiSwP.png)](https://i.stack.imgur.com/JiSwP.png)


I knew that DPCs are code being run inside a driver; the challenge was to figure out *which* driver. I turned to [DPC Latency Checker](https://web.archive.org/web/20201015115944/https://www.thesycon.de/eng/latency_check.shtml), which showed me just how bad the latency was:


![screenshot of DPC Latency Checker](https://i.stack.imgur.com/TvPXg.png)


*DPC Latency Checker* suggests going through devices in **Device Manager**, and disabling non-essential hardware one-by-one (e.g. network card, sound card), hoping to isolate the buggy driver. (If you disable a device, and the DPC latency suddenly drops: you've found your culprit!)


![screenshot of disabling devices](https://i.stack.imgur.com/fs1Jw.png)


Unfortunately, after disabling everything I possibly could (while still being able to *use* the computer — don't disable your hard drive, video card, mouse, or USB hub the mouse is plugged into!), the latency was still high. Next I turned to the the [Windows Performance Toolkit](https://docs.microsoft.com/en-us/) (part of the [Windows SDK](https://www.microsoft.com/en-us/download/details.aspx?id=8279)), and an excellent blog post by Peter Weiland, [""Measuring DPC Time""](https://docs.microsoft.com/en-us/archive/blogs/peterwie/measuring-dpc-time). After installing the Windows Performance Toolkit:


[![Screenshot of Windows SDK installer, with Windows Performance Toolkit being selected](https://i.stack.imgur.com/QhHKD.png)](https://i.stack.imgur.com/QhHKD.png)


I opened an elevated command prompt and ran:



```
>xperf -on Latency

```


> 
> **Note**: The `Latency` *group* is a
> predefined set of events that can be
> traced from the *Kernel Group*
> provider:
> 
> 
> 
> ```
> >xperf -providers kg
>    Base           : PROC_THREAD+LOADER+DISK_IO+HARD_FAULTS+PROFILE+MEMINFO
>    Diag           : PROC_THREAD+LOADER+DISK_IO+HARD_FAULTS+DPC+INTERRUPT+CSWITCH+PERF_COUNTER+COMPACT_CSWITCH
>    DiagEasy       : PROC_THREAD+LOADER+DISK_IO+HARD_FAULTS+DPC+INTERRUPT+CSWITCH+PERF_COUNTER
>    Latency        : PROC_THREAD+LOADER+DISK_IO+HARD_FAULTS+DPC+INTERRUPT+CSWITCH+PROFILE
>    ...
> 
> ```
> 
> In this case `Latency` corresponds to
> the Kernel Flags:
> 
> 
> - **PROC\_THREAD** Process and Thread create/delete
> - **LOADER** Kernel and user mode Image Load/Unload events
> - **PROFILE** CPU Sample profile
> - **CSWITCH** Context Switch
> - **DPC** DPC Events
> - **INTERRUPT** Interrupt events
> - **DISK\_IO** Disk I/O
> - **HARD\_FAULTS** Hard Page Faults
> 
> 
> 


After letting that run for a minute, I stopped the trace, and had it save to a file:



```
C:\Users\Ian\Desktop\xperf -d thingy1.etl

```

And then I viewed the results of the trace with the command:



```
C:\Users\Ian\Desktop\xperf thingy1.etl

```

This loads the graphical **Windows Performance Analyzer**. Right clicking on the **DPC CPU Usage** graph, I selected **Summary Table**. This shows a breakdown of time spent in DPCs by driver:


![screenshot of XPerf output](https://i.stack.imgur.com/pOBlG.png)


Right away I can see one driver (`tsvp.sys`) taking an average of 2.8ms per DPC execution, which is an order of magnitude slower than any other driver:


![screenshot](https://i.stack.imgur.com/YH7g9.png)


Googling `tsvp.sys` gave me the answer: [CommView](https://www.tamos.com/products/commview/), which I had recently installed.


The question now is how to disable this driver. Using [AutoRuns](https://docs.microsoft.com/en-us/sysinternals/downloads/autoruns), I can see that it's installed as a driver service:


![screenshot of autoruns](https://i.stack.imgur.com/ajUoa.png)


Using Device Manager, I can disable the service that hosts this driver. First you have to **Show hidden devices**, then expand the `Non-Plug and Play Drivers` node:


![screenshot of device manager](https://i.stack.imgur.com/w9IfJ.png)


Finally I could stop the driver service, and I changed it's startup mode from `System` (meaning the driver is an essential part of Windows, and Windows cannot boot without it), to `Demand` (meaning I can start the driver when I want to):


![screenshot of device manager](https://i.stack.imgur.com/iaBF0.png)


Stopping the driver service immediately fixed my DPC latency:


![screenshot](https://i.stack.imgur.com/hgcnq.png)


I may or may not completely uninstall CommView, but for now I've solved the Case of the High DPC Latency.




---


**Update**: Starting with Windows 8 [you can no longer see **Non-Plug and Play Drivers** in Device Manager](https://docs.microsoft.com/en-us/windows-hardware/drivers/install/viewing-hidden-devices):



> 
> Note Starting with Windows 8 and Windows Server 2012, the Plug-and-Play Manager no longer creates device represetnations for non-PnP (legacy) devices. Thus there are no such devices to view in the Device Manager.
> To include hidden devices in Device Manager display, click View and select Show Hidden Devices.
> 
> 
> 


Microsoft took away the feature and replace it with nothing. Good job.


In typical nerd rage, [some unhelpful answers](https://community.osr.com/discussion/261449):


- *Device manager never showed non pnp drivers*
- *Why do you need this?*


Fortunately, NirSoft has created a replacement. [ServiWin](https://www.nirsoft.net/utils/serviwin.html) lets you see, stop, and start all services (even the ones Microsoft decided administrators should not be allowed to see):


[![screenshot of ServiWin](https://i.stack.imgur.com/fiIjr.png)](https://i.stack.imgur.com/fiIjr.png)


"
"How to remove global ""use strict"" added by babel
I'm using function form of ""use strict"" and don't want global form which Babel adds after transpilation. The problem is I'm using some libraries that aren't using ""use strict"" mode and it might throw error after scripts are concatenated


","### Babel 5


You'd blacklist `""useStrict""`. For instance here's an example in a Gruntfile:



```
babel: {
    options: {
        blacklist: [""useStrict""],
        // ...
    },
    // ...
}

```

### Babel 6


Since Babel 6 is [fully opt-in for plugins now](http://babeljs.io/blog/2015/10/29/6.0.0#opt-in-plugins), instead of blacklisting `useStrict`, you just don't include the [`strict-mode` plugin](http://babeljs.io/docs/plugins/transform-strict-mode/). If you're using a preset that includes it, I think you'll have to create your own that includes all the others, but not that one.


"
"Node.js JSON.stringify() causing "" in output. Can't parse with Jquery
I am using Node.js (with Express.js) to pass a JSON data object from the server to the client view.


When I render the JSON object directly to the view I get the JSON object shown on the page as expected (this WORKS):



```
pageprovider.findAllTag( function(error, pages){
    res.send(pages);
})

```

And my output looks like this (much bigger, many nested obj)



```
{""green"":{""title"":""green"",""pagesContaining"": """"}}

```

When I try to pass it to my Jade View like this:



```
pageprovider.findAllTag( function(error, tagsJSONObj){
        //res.send(pages);

    pageprovider.findAll( function(error, pages){
        res.render('search_tags.jade', { locals: {
            title: 'Search by Tags',
            'pages': pages,
            tagsJSON: JSON.stringify(tagsJSONObj) //pass the tags data as a JSON obj
            }
        });
    }) //pageprovider.findAll
}) //pageprovider.findAllTag

```



---


**The problem**  

When I pass 'tagsJSON' to the view, the output includes the html entities:



```
var obj = jQuery.parseJSON( ""{&quot;name&quot;: 'value'}"");

```

JQuery throws an error because it doesn't like '""'. How can I get Node to give me the proper quote, or get jQuery to accept this format?


Any thoughts?


","It's because when you call 



```
    res.render('search_tags.jade', { locals: {
        title: 'Search by Tags',
        'pages': pages,
        tagsJSON: JSON.stringify(tagsJSONObj) //pass the tags data as a JSON obj
        }
    });

```

`search_tags.jade` is meant to output HTML, so it encodes your quotes. You should use a renderer that doesn't HTML escape, or at least change your view so that your params aren't HTML encoded


If you don't want something in the output escaped, use `!{tagsJSON}` within the view. However, when outputting JSON, there's no need for a view. you can just take your object, call `JSON.stringify`. I don't use JADE so I'm not sure if there is a way to create view that can just call `JSON.stringify()`, but that's what I've done in JSP, velocity, ASP, PHP and Code Igniter (not using `JSON.stringify`, instead it uses a JSON tool for the given language)


"
"How to call function in angularjs factory
this is my factory and i want to call getData in saveData.Here is my code



```
.factory('dataSyncOperation', function($q,$http){
return {
    getData:function(){
        var q = $q.defer();
         var config = {
                    headers : {
                        'Content-Type': 'application/json'
                    }
                }
         $http.get(api+'/sync/').then(function(response){
            q.resolve(response);
        },function(error){
            q.reject();
        })
        return q.promise;

    },

    saveData:function(){

    }

}

```

});
How can i use the promise returned by getData into saveData.


","You can always do, something like this -



```
saveData:function(){
  this.getData().then(function(response){ // you can handle getData promise here
     // on success 
  }, function(reject){
     // on failure
  });
}

```

inside your `saveData` method, let me know if this is what something you are looking for.


Working Example - <http://plnkr.co/edit/y8WZQT8SvOAWpKj8Jgxs?p=preview>


Code -



```
// Code goes here

var myApp = angular.module('myApp', []);

myApp.controller('mainCtrl', function($scope, testService){
  testService.saveData().then(function(res){
    $scope.test = res.data;  
  });
})

myApp.factory('testService', function($q, $http){
  return {
      getData:function(){
        var q = $q.defer();
        $http.get('data.json').then(function(response){
          q.resolve(response);
        }, function(error){
            q.reject();
        })
        return q.promise;
      },
      saveData:function(){
        return this.getData();
      }
  }
})

```

"
"Couchdb Mango Performance vs Map Reduce Views
I've just noticed that in [the release notes](http://docs.couchdb.org/en/latest/whatsnew/2.0.html) of Couchdb 2.0, it is mentionned that Mango queries are recommended for new applications. It is also mentionned that apparently Mango indexes are from 2x to x10 faster than javascript queries which really surprised me, as such I have a number of questions :


- Are Map/Reduce views being phased out ? I'm expecting the answer to be no since it seems to me that Mango does not cover all the use cases of Map/Reduce (the easiest example being Reduce itself), and the flexibility of this querying style seems to be more limited too. But m prefer to ask because of the recommendation :



> 
> We recommend all new apps start using Mango as a default.
> 
> 
> 


- We know that Map/Reduce views rely on B-trees, but I can't find any insight, in the doc or the mailing list regarding the magic behind Mango. Mango essentially is white magic for me at the minute. Yet I can tell that having an in-depth knowledge of how the javascript views are indexed behind the scenes was massively helpful to avoid pitfalls, naive implementations as well as to optimize performances. Does anyone have any insight on how Mango works ? Are the indexes B-trees too ? When are the indexes updated since there is no longer design documents ? Where do the performance gains come from ? (these gains are counter-intuitive to me, since in my understanding, the performance of javascript queries came from the precomputed nature of Map functions)


What I'm essentially after is on the one hand some insight regarding Mango and on the other hand, an overview of how Mango and Map/Reduce are supposed to live together in the 2.x era.


","Answer from a core developer :



> 
> Some good questions. I don't think Mango will ever replace Map/Reduce
>  completely. It is an alternative querying tool. What is great about
>  the Mango query syntax is that it is a lot easier to understand and
>  get started. And we can use it in a lot of places outside of just
>  querying for documents. It can be used for replication filtering and
>  the changes feed. We hope to soon have support for validation doc
>  updates as well.
> 
> 
> Underneath Mango is using erlang map/reduce. Which means it is
>  creating a B-tree index just like map/reduce. What makes it faster is
>  that it is using erlang/native functions to create the B-Tree instead
>  of javascript. I wrote a blog post a long time ago about the internals
>  of PouchDB-find [1] which is the mango syntax for PouchDB. It might
>  help you understand a little more how the internals work. The key
>  thing to understand is that there is a Map query part which uses the
>  B-Tree and an in-memory filter. Ideally the less memory filtering you
>  do the faster your query will be.
> 
> 
> I would say that Mango is very much a work in process but the basic
>  ground work is done. There are definitely things we can improve on.
>  I've seen it used quite a bit when developers start a new project
>  because its quick and simple to do basic querying, like find by email
>  address or find all users with the name ""John Rambo"".
> 
> 
> Hope that helps.
> 
> 
> [1] <http://www.redcometlabs.com/blog/2015/12/1/a-look-under-the-covers-of-pouchdb-find>
> 
> 
> 


"
"Usage of SIM card GlobalPlatform keys in field
**Question 1**. When a SIM manufacture personalizes a SIM card and then the mobile operator hand it over to an end user, are there any usage for GlobalPlatform (GP) keys (ENC,MAC,KEK,...)?


As long as I know the SIM file structure or its applets are accessible by (KID/KIC/KIK) over OTA RFM and RAM. So there is no need to have GP keys when SIM is in hands of customers.


Are there anywhere special (e.g. during OTA interactions that we require GP keys or not)? As I understood correctly, GP keys are usable when we physically can access a SIM card and we can directly send APDUs to the SIM. Am I right?


**Question 2**. Is it possible to access SIM/USIM file structure using global platform commands (e.g. is it possible to read record, read binary) using GP keys and access which ISD has? Typically file structure is accessible based on access conditions defined in 3GPP TS 51.011/ETSI TS 151 011 using PIN and ADM keys. But is it possible using GP keys and access which ISD has?


","Q1:


I'm doing consulting for a MNO. All our plastic SIM cards are only using SCP80 (the OTA mechanism supporting [SMS, CAT\_TP, TCP, BIP](https://www.etsi.org/deliver/etsi_ts/102200_102299/102225/09.00.00_60/ts_102225v090000p.pdf)) and more recent versions are also supporting SCP81 (OTA using HTTPs using a pre-shared TLS keys [GlobalPlatform RAM](https://globalplatform.org/wp-content/uploads/2015/05/GPC_2.2_B_RAM_over_HTTP_v1.1.1.pdf)). The SCP02 and SCP03 keys are rotated and not used. When using [eSIM](https://trustedconnectivityalliance.org/technology_overview/esim/) SCP03 is not even enabled in our case.


Q2:


No. But you can use OTA messages with e.g. SCP80 and wrap the read APDUs into [RFM](https://www.etsi.org/deliver/etsi_ts/102200_102299/102226/16.00.00_60/ts_102226v160000p.pdf). With SCP02 /SCP03 this is not directly possible. But you could install an applet using SCP02/SCP03 with access to the [file system](https://www.etsi.org/deliver/etsi_ts/131100_131199/131130/08.01.00_60/ts_131130v080100p.pdf) and use this as tunnel if you do not have the ADMs.


"
"I get an opcode error when trying to connect to a vanilla mongodb docker container from windows robo3t
I'm using Windows 10 to develop Linux applications using Docker for Windows. I'm also using docker-compose. When I try to connect to a vanilla MongoDB 4.1 image from my Windows Robo 3T client, I get this peculiar error.


docker-compose.yml



```
version: ""3""
services:
  myapp-db:
    image: mongo:4.1
    ports:
      - 27017:27017
    networks:
      - myapp-net
    volumes:
      - vol-myapp-db:/data/db
    restart: always

networks:
  myapp-net:

volumes:
  vol-myapp-db:

```

When I run this configuration I get the following error only when I attempt to connect with Robo3T.


stdout:



```
$ docker-compose up --build

myapp-db_1          | 2019-02-16T02:59:05.825+0000 I NETWORK  [listener] connection accepted from 172.18.0.1:42622 #1 (1 connection now open)
myapp-db_1          | 2019-02-16T02:59:05.826+0000 I NETWORK  [conn1] received client metadata from 172.18.0.1:42622 conn1: { application: { name: ""robo3t"" }, driver: { name: ""MongoDB Internal Client"", version: ""3.4.3-10-g865d2fb"" }, os: { type: ""Windows"", name: ""Microsoft Windows 8"", architecture: ""x86_64"", version: ""6.2 (build 9200)"" } }
myapp-db_1          | 2019-02-16T02:59:05.828+0000 E -        [conn1] Assertion: Location34348: cannot translate opcode 2010 src/mongo/rpc/message.h 121
myapp-db_1          | 2019-02-16T02:59:05.828+0000 I NETWORK  [conn1] DBException handling request, closing client connection: Location34348: cannot translate opcode 2010
myapp-db_1          | 2019-02-16T02:59:05.828+0000 I NETWORK  [conn1] end connection 172.18.0.1:42622 (0 connections now open)

```

Anyone have an idea what's going on here?


","I used to have the same error with robo 3T 1.2.1


I updated the robo 3T version to 1.3 and it worked again,  

which uses the MongoDB Internal Client 4.0:



```
{ 
  application: { name: ""MongoDB Shell"" },   
  driver: { name: ""MongoDB Internal Client"", version: ""4.0.5-17-gd808df2233"" }
} 

```

But as stated on their website they only recommend up to Mongo 4.0.



> 
> Robo 3T 1.3 brings you support for MongoDB 4.0 and SCRAM-SHA-256, an
>  upgraded mongo shell, support for importing from MongoDB SRV
>  connection strings, among many other fixes and improvements.
> 
> 
> 


So i recommend using **Studio 3T.**


"
"Generate random number with my array values without repetition
I'm new to C# and I'm making an application with arrays. I have an array with the numbers shown below:



```
int[] array2 = new int[] { 1, 3, 5, 7, 9 };

```

What I need to do is change the order of these numbers in the array without repetitions, because when I use Random Function, this shows me repeated numbers.


I saw this method, but do not know how to apply it with numbers: <http://www.dotnetperls.com/shuffle>


","You can use the following LINQ chain:



```
int[] array2 = new int[] { 1, 3, 5, 7, 9 };
var random = new Random();
var total = (int)array2.
    OrderBy(digit => random.Next()).
    Select((digit, index) => digit*Math.Pow(10, index)).
    Sum();

```

First, it orders the elements randomly, then it selects each element multiplied by 10 raised to the power of its index, then sums them together and casts the result to an integer. Also, please note that I didn't provide an useful seed for your `Random` instance. You might want to do that, to produce pseudo-random results.


You might also want to use a method for exponentiation described [here](https://stackoverflow.com/questions/383587/how-do-you-do-integer-exponentiation-in-c), to avoid having to cast to an integer.


EDIT: As Rhumborl pointed out, you may just need the shuffled array. In that case:



```
var shuffledArray = array2.OrderBy(n => random.Next()).
   ToArray();

```

Should work for you.


"
"Can't install ATI proprietary drivers in 12.10
I have a laptop with ATI Radeon 6770M HD Hybrid graphics card. In Ubuntu 12.04, I installed the fglrx driver through ""additional drivers"" and it worked. (I can even switch GPUs). But in the new Ubuntu 12.10, after installing, Unity won't load. 


Only the mouse and the wallpaper. If I initialize the settings `sudo aticonfig --initial` then after rebooting it gives a warning saying I'm in low graphics mode! How to fix this?


Earlier I used software source to install the drivers. But when using the terminal, I got this warning : 



```
update-alternatives: warning: forcing reinstallation of alternative /usr/lib/fglrx/ld.so.conf because link group x86_64-linux-gnu_gl_conf is broken

```

Update : Filed a bug report in launchpad : <https://bugs.launchpad.net/fglrx/+bug/1068661>


","I started with a clean 12.10 install (not an upgrade)
My card is a Radeon HD 5450


First thing I done was to install Linux-source and Linux headers



```
sudo apt-get install linux-source
sudo apt-get install linux-headers-generic

```

I noticed that many with Nvidia cards were having to do this and am sure I saw mentioned some-where that it also helped ATI users.


Like everyone else if I installed either of the proprietary drivers after reboot I was able to login but lost unity.


One thing I did notice is that installing the driver did not install Catalyst control centre (not sure if it is the same for everyone else)


So I installed Synaptic package manager and installed fglrx-updates from there


![enter image description here](https://i.stack.imgur.com/yYMJb.png)


Now the Catalyst control centre was installed.


Upon rebooting however instead of being able to login with no Unity the system was hanging at the splash screen.


So I applied `nomodeset` and this time it booted with Unity present


Here is the result of fglrxinfo


![enter image description here](https://i.stack.imgur.com/fcaq4.png)


And here is what software sources says


![enter image description here](https://i.stack.imgur.com/wxTbt.png)


With `nomodeset` applied I might not be getting the splash screen on boot up but fglrx-updates is working well. I would go as far as to say that it is performing better than it was in 12.04


**More info**


Grub offers me 2 kernels


3.5.0-18-generic
and
3.5.0-17-generic


The later one '18' is good, the older one '17' still loads without Unity


"
"Multiply vector elements by a scalar value using STL and templates
I wrote a small scientific simulation where I wanted to manipulate vectors with the same kind of functionality that Python's NumPy vectors use. Namely, I wanted to be able to add vectors together, multiply by constants, etc. I have need for both vectors of type int and floating point, so I wanted to template the overloading. As an example (and for something to review), I overloaded scalar multiplication as:



```
template <class T, class Q>
vector <T> operator*(const Q c, const vector<T> &A) {
  vector <T> R(A.size());
  for(int i=0;i<A.size();++i) 
    R[i] = c*A[i];
  return R;
}

```

The STL algorithm way I hinted at looks like this:



```
template <class T, class Q>
vector <T> operator*(const Q c, const vector<T> &A) {

  vector <T> R(A.size());
  std::transform(A.begin(), A.end(), R.begin(),
                 std::bind1st(std::multiplies<T>(),c));
  return R;
}

```

I would appreciate all criticism relevant to code, style, flow. My project is in C++11, so in theory I could use the newer range-based for loops for readability (I do in other parts of my code). Additionally, I think I could rewrite this using [`std::transform`](https://stackoverflow.com/questions/3885095/multiply-vector-elements-by-a-scalar-value-using-stl), but I'm not sure if it would help improve the efficiency or the readability of my code.


","I would use the second version that you posted. Here are some problems with the first version:


**1. Prefer passing by value rather than making an explicit copy.**


This is a problem in both versions of your code. You pass `A` via constant reference. Normally this is a good thing, but then you then make a copy of `A` explicitly. In this case, it is possible to miss out on a compiler optimization called [copy elision](https://stackoverflow.com/questions/12953127/what-are-copy-elision-and-return-value-optimization) ([second reference here](http://en.cppreference.com/w/cpp/language/copy_elision)). Passing by value would not cause any variable aliasing problems.


*Update*: Imagine somebody (maybe you) using your function.



```
std::vector <int> v1 = // ... insert elements
auto v2 = 3 * v1 ;

```

For simplicity, let's say variable `v2` is located at address 0x10.
When you do `3 * v1`, you are calling your function. 


`R` would then be created at let's say address 0x20. When your function returns, the contents of `R` would have to be copied from 0x20 to `v2` at 0x10 and then the destructor for `R` would have to be called. Without copy elision, this would require copying every element from `R` into `v2`. 


Now in C++11, there is something called [move semantics](https://stackoverflow.com/questions/3106110/what-is-move-semantics). This means, that without copy elision, `R` would not have to copy every element into `v2`, it could swap pointers. This is much faster, but it is still slower than copy elision. 


With copy elision, instead of creating `R` at 0x20, it uses address 0x10 and builds `v2` in place. With copy elision, nothing has to be copied and no pointers have to be swapped because `v2` is built in place.


Here's how I would re-write the function:



```
template <class T, class Q>
std::vector <T> operator* (const Q c, std::vector <T> A)
{
    std::transform (A.begin (), A.end (), A.begin (),
                 std::bind1st (std::multiplies <T> () , c)) ;
    return A ;
}

```

**2. If you have to use an index-loop for some reason, rather than a range-loop or iterators, make sure you are comparing compatible types.**


You could do something like this:



```
typedef std::vector <T>::size_type size_type ;
for (size_type i = 0; i < A.size (); ++i)

```

C++11 has expanded the functionality of the `using` keyword. So instead of a `typedef`, you could also do:



```
using size_type = decltype (A.size ()) ;

```

*Update*: The reason you want to use compatible types is because overflowing a signed number is undefined behavior in C++. This means anything can happen. Consider this example:



```
const unsigned size = std::numeric_limits <int>::max () + 5 ;
std::vector <char> v1 (size) ;

for (int n = std::numeric_limits <int>::max (); n < v1.size (); ++n) {
    std::cout << n << ',' << v1.size () << '\n' ;
}

```

Different compilers on different platforms will do different things with the code above because there is undefined behavior (an `int` can never reach `std::numeric_limits <int>::max () + 5`).


**3. Consider wrapping the `std::vector` into a class.**


Programs typically use `std::vector` for many things. Overloading a global operator like you do would make me uncomfortable using your code.


**4. Obligatory [avoid `using namespace std;`](https://stackoverflow.com/questions/1452721/why-is-using-namespace-std-considered-bad-practice).**


**5. Use more meaningful variable names and typenames.**


Unless `A`, `R`, `Q`, etc are commonly understood names for special variables in your scientific community, use more meaningful names.


**6. Do not omit braces for if-statements or loops.**


This is more a subjective style guideline than anything.


"
"Search and replace contents of yanked text in vim
In my vimrc, I have a shortcut to copy the filename with its path to the clipboard.



```
:nmap cp :let @* = expand(""%"")

```

This works fine. Now, I want to replace the contents of this yanked text to



```
1) replace \ with /  
2) remove certain words from the yanked text.

```

I am familiar with search and replace on regular text but I don't know how to change my vimrc entry to do this every time on the yanked text when I use my shortcut. 


So, something like this?



```
:nmap cp :let @* = expand(""%"") || %s/\\/\/ || %s/<word>//

```

I am using gvim on windows.


","The `:substitute` command works on the buffer contents itself; that's not so useful here. (You could temporarily `:put` the register / file name, transform it, and then `:delete` it (back) into a register.) Fortunately, there's an equivalent low-level [`:help substitute()`](https://vimhelp.appspot.com/eval.txt.html#substitute%28%29) function that you can apply on a String:



```
:nnoremap cp :let @* = substitute(expand(""%""), '\\', '/', 'g')<CR>

```

In fact, `expand()` directly supports a special substitution [`:help filename-modifiers`](https://vimhelp.appspot.com/cmdline.txt.html#filename-modifiers), so this would be a (rather obscure) solution, too:



```
:nnoremap cp :let @* = expand(""%:gs?\\?/?"")<CR>

```

For the additional removal of `<word>`, you can use another layer of `substitute()` / append another `:gs???`. I'll leave that to you.


### Additional critique


- Your mapping is missing the trailing `<CR>` to conclude command-line mode.
- [You should use `:noremap`](http://learnvimscriptthehardway.stevelosh.com/chapters/05.html); it makes the mapping immune to remapping and recursion.
- `%` is relative to the current working dir. If you need the full absolute path, you can get that via `%:p`.
- Starting your mapping with `c` is unconventional. It's fine here, because the `c`hange command wants a `{motion}`, and `p` is not a valid one. In general, I'd rather avoid such clever overloading (as long as you have other unused mapping prefixes still available).


"
"How to get current date and time from DB using SQLAlchemy
I need to retrieve what's the **current date and time for the database I'm connected** with SQLAlchemy (not date and time of the machine where I'm running Python code). I've seen this functions, but they don't seem to do what they say: 



```
>>> from sqlalchemy import *
>>> print func.current_date()
CURRENT_DATE
>>> print func.current_timestamp()
CURRENT_TIMESTAMP

```

Moreover it seems they don't need to be binded to any SQLAlchemy session or engine. It makes no sense...


Thanks!


","I foud the solution: these functions cannot be used in the way I used (`print...`), but need to be called inside of the code that interacts with the database. For instance:



```
print select([my_table, func.current_date()]).execute()

```

or assigned to a field in an insert operation.
Accidentally I discovered that exists at least a couple of parameters for these functions:


- `type_` that indicates the type of the value to return, I guess
- `bind` that indicates a binding to an SQLAlchemy engine


Two examples of use:



```
func.current_date(type_=types.Date, bind=engine1)
func.current_timestamp(type_=types.Time, bind=engine2)

```

Anyway my tests seems to say these parameters are not so important.


"
"Problems after updating SVN to 1.8.3 on Ubuntu 12.10 (using WANdisco package)

```
apache2: Syntax error on line 203 of /etc/apache2/apache2.conf:
Syntax error on line 2 of /etc/apache2/mods-enabled/authz_svn.load:
Cannot load /usr/lib/apache2/modules/mod_authz_svn.so into server:
/usr/lib/apache2/modules/mod_authz_svn.so: undefined symbol: dav_svn_get_repos_path
Action 'configtest' failed.

```

I get this error after updating SVN on Ubunutu (12.10 Quantal) when I try to restart the server. All the modules dav, dav\_svn and authz\_svn are in place an running. Unfortunately., searching for the error in a search engine of my choice doesn't get me any further.
When accessing svn via a browser it still tells the version to be 1.7.5.x when do a 'svn help' it shows 1.8.3.x


After downgrading to 1.7.5 again, I get this:



```
Syntax error on line 203 of /etc/apache2/apache2.conf:
Syntax error on line 2 of /etc/apache2/mods-enabled/authz_svn.load: Cannot load
/usr/lib/apache2/modules/mod_authz_svn.so into server: libsvn_repos-1.so.0:
cannot open shared object file: No such file or directory

```

","I got similar errors after installing custom Subversion 1.7 Apache modules in Ubuntu 12.04. The modules must be loaded in this order: dav -> dav\_svn\_module -> authz\_svn\_module, for the libraries to load properly. The libapache2-svn Ubuntu package seems to handle this with 'Depends:' comments at the top of the .load files:


/etc/apache2/mods-available/dav\_svn.load:



```
# Depends: dav
LoadModule dav_svn_module /usr/lib/apache2/modules/mod_dav_svn.so

```

/etc/apache2/mods-available/authz\_svn.load



```
# Depends: dav_svn
LoadModule authz_svn_module /usr/lib/apache2/modules/mod_authz_svn.so

```

In my experience, my custom modules are enabled/disabled based on those dependencies, but Apache still tries to load /etc/apache2/mods-enabled/\* in alphabetical order, therefore loading authz\_svn.load first and getting the above error. I don't know how the libapache2-svn package avoids this problem while custom modules do not.


I've worked around the problem by removing the mods-available/mods-enabled dav\_svn.load and authz\_svn.load files entirely. Then I put this into /etc/apache2/httpd.conf:



```
LoadModule dav_svn_module /usr/lib/apache2/modules/mod_dav_svn.so
LoadModule authz_svn_module /usr/lib/apache2/modules/mod_authz_svn.so

```

Order matters, obviously. Apache loads httpd.conf after mods-enabled, so now Apache starts for me.


"
"Generic method without parameters
I was confused with my code that includes a generic method that takes no parameters, so what will be the return generic type of such a method, eg:



```
static <T> example<T> getObj() {
    return new example<T>() {

        public T getObject() {
            return null;
        }

    };
}

```

and this was called via:



```
example<String> exm = getObj(); // it accepts anything String like in this case or Object and everything

```

the interface `example's` defination is:



```
public interface example<T> {

    T getObject();
}

```

My question:`example<String> exm` is accepting String, Object and everything. So at what time generic return type is specified as String and how??


","The compiler infers the type of `T` from the concrete type used on the *LHS* of the assignment.


From [this link](http://www.angelikalanger.com/GenericsFAQ/FAQSections/TechnicalDetails.html#What%20happens%20if%20a%20type%20parameter%20does%20appear%20in%20the%20method%20parameter%20list?):



> 
> If the type parameter does not appear in the types of the method
>  arguments, then the compiler cannot infer the type arguments by
>  examining the types of the actual method arguments. If the type
>  parameter appears in the method's return type, then the compiler takes
>  a look at the context in which the return value is used. If the
>  method call appears as the righthand side operand of an assignment,
>  then the compiler tries to infer the method's type arguments from the
>  static type of the lefthand side operand of the assignment.
> 
> 
> 


The example code in the link is similar to the one in your question:



```
public final class Utilities { 
  ... 
  public static <T> HashSet<T> create(int size) {  
    return new HashSet<T>(size);  
  } 
} 
public final class Test 
  public static void main(String[] args) { 
    HashSet<Integer> hi = Utilities.create(10); // T is inferred from LHS to be `Integer`
  } 
}

```

"
"Type inference with piping or composition fails, where normal function call succeeds
I rarely have this struggle nowadays with F#, but then again type inheritance is much less common with F#, so perhaps I was just lucky. Or I am missing the obvious. Normally when the compiler complains about not knowing a certain type I reverse the order of pipes or composition operands and I'm done.


Basically, given a function call that works as `g(f x)`, it also works as `x |> f |> g` or `(f >> g) x`. But today it doesn't...


Here's a messy proof-of-concept of what I mean:



```
module Exc =
    open System

    type MyExc(t) = inherit Exception(t)

    let createExc t = new MyExc(t)
    type Ex = Ex of exn
    type Res = Success of string | Fail of Ex with
        static member createRes1 t = Ex(createExc(t)) |> Fail   // compiled
        static member createRes2 t =  t |> createExc |> Ex |> Fail  // FS0001
        static member createRes3 = createExc >> Ex >> Fail   // FS0001

```

Normally, this works (at least in my experience). The lines with ""fail"" throw:



> 
> error FS0001: Type mismatch. Expecting a MyExc -> 'a but given a exn -> Ex. The type 'MyExc' does not match the type 'exn'
> 
> 
> 


Not a big deal, not hard to workaround, but I happen to have to write a lot of code where composition is the easier/cleaner approach and I don't wish to write a bunch of utility functions that I have to put in everywhere.


I looked at *flexible types*, as I guess this is a contravariance problem, but I don't see how I can apply it here. Any ideas to keep this idiomatic?


Note, if I rearrange, i.e. as `Ex << createExc >> Fail` or using the pipe-backward operator I end up with the same error on a different part.


","The F# compiler behaves a bit irregularly in this case. In your example, you want to pass a value of type `MyExc` to a constructor that expects `exn`. Treating an object as a value of its base class is a valid coersion, but the F# compiler inserts such coersions in only very limited places.


In particular, it inserts coersion when you pass arguments to a function, but it does not insert them (for example) when creating a list or returning results from a function.


In your example, you need a coersion when passing value to a discriminated union constructor. It seems that this happens only when directly creating the union case, but it does *not* happen when treating the union case as a function:



```
// foo is a function that takes `obj` and Foo is a DU case that takes `obj`
let foo (o:obj) = o
type Foo = Foo of obj

foo(System.Random()) // Coersion inserted automatically
Foo(System.Random()) // Coersion inserted automatically

System.Random() |> foo // Coersion inserted automatically
System.Random() |> Foo // ..but not here!

```

So, the limited set of places where F# compiler applies coersions automatically includes various ways of calling functions, but only direct way of creating DU cases.


This is a bit funny behaviour - and I think that it would make sense to treat DU cases as ordinary functions including the automatic insertion of coersions when you use `|>`, but I'm not sure if there are any technical reasons that make that hard.


"
"Android ndk; How to create a 2d boolean array in c/c++ and return it to the Java?
So I've worked with the ndk/c/c++ a little bit, and I know about the jbooleanArray var type, but my question is how would I go about creating a 2d boolean array and returning it to the Java to be used? Here's an example:



```
#include ""retrieveboolarray.h""
#define  LOG_TAG    ""!""
//macros to call the Log functions to allow for debugging. the tag is declared on the line above
#define  LOGD(...)  __android_log_print(ANDROID_LOG_DEBUG,LOG_TAG,__VA_ARGS__)
#define  LOGE(...)  __android_log_print(ANDROID_LOG_ERROR,LOG_TAG,__VA_ARGS__)

JNIEXPORT jint JNICALL Java_com_example_ndktesting_generateBoolArrays_retrieveBoolArray(
        JNIEnv * env, jobject obj, jint refNum) {
    //refNum will eventually be used to access certain arrays
    return refNum;
}

```

so right now it's returning an integer, how would I generate a 2d bool array and return it to the Java?


update: here is some code I've tried previously, still with no luck unfortunately



```
JNIEXPORT jobjectArray JNICALL Java_com_example_ndktesting_generateBoolArrays_retrieveBoolArray(
        JNIEnv * env, jobject obj, jint refNum) {

    jboolean flag = JNI_TRUE;
    jclass stringClass = (*env)->FindClass(env, ""java/lang/Boolean"");
    jobjectArray row;
    jobjectArray rows;
    jboolean hello = 1;
    int MAXRESPONSE = 5;
    int MAXTEST = 5;

    jsize i, j;
    for (i = 0; i < 5; i++) {
        row = (*env)->NewObjectArray(env, MAXRESPONSE, stringClass, 0);
        for (j = 0; j < 3; j++) {
            (*env)->SetObjectArrayElement(env, row, j, hello);
        }

        if (flag == JNI_TRUE) {
            flag = JNI_FALSE;
            rows = (*env)->NewObjectArray(env, MAXTEST,
                    (*env)->GetObjectClass(env, row), 0);
        }

        (*env)->SetObjectArrayElement(env, rows, i, row);
    }

    return rows;
}

```

hopefully someone has an idea on how to fix this so I can make/return a bool array.


","Essentially, you need to use a jobjectArray that contains jbooleanArray instances. See [here](https://stackoverflow.com/questions/15972897/jni-pass-char-2d-array-to-java-code) for a similar example.


Update:


I wrote the code using C++ and converted to C here, but other than that it is tested and works.



```
JNIEXPORT jobjectArray JNICALL Java_com_example_ndktesting_generateBoolArrays_retrieveBoolArray(
    JNIEnv* env, jobject obj, jint refNum) {
    jclass arrayClass = (*env)->FindClass(env, ""[Z"");
    int numColumns = 5, numRows = 5;
    jobjectArray twoDArray = (*env)->NewObjectArray(env, numColumns, arrayClass, 0);
    jbooleanArray column;

    // Initialization of the source array
    jboolean source[numColumns][numRows];
    for (int i = 0; i < numColumns; ++i) {
        for (int j = 0; j < numRows; ++j)
            source[i][j] = true;
    }

    for (int i = 0; i < numColumns; ++i) {
        column = (*env)->NewBooleanArray(env, numRows);
        (*env)->SetBooleanArrayRegion(env, column, 0, numRows, source[i]);
        (*env)->SetObjectArrayElement(env, twoDArray, i, column);
    }
    return twoDArray;
}

```

In your Java code, you will get an `Object[]`. You can cast each of its elements to a `boolean[]`.


"
"Get the order fee item details in Woocommerce 3
I try to get the name of the fee attached to my order in Woocommerce, I get an array but I don't know how to get the name. 


I tried with function `get_name ()` but it does not work.



```
$the_order->get_items( array( 'line_item', 'fee', 'shipping' ) );

```

Raw data output:



```
[137] => WC_Order_Item_Fee Object
        (
            [extra_data:protected] => Array
                (
                    [tax_class] => 
                    [tax_status] => taxable
                    [amount] => 
                    [total] => 
                    [total_tax] => 
                    [taxes] => Array
                        (
                            [total] => Array
                                (
                                )

                        )

                )

            [data:protected] => Array
                (
                    [order_id] => 7795
                    [name] => Frais de réservation
                    [tax_class] => 0
                    [tax_status] => taxable
                    [amount] => 
                    [total] => 35
                    [total_tax] => 0
                    [taxes] => Array
                        (
                            [total] => Array
                                (
                                )

                        )

                )

```

","To access and to use properties on Order Fee items you need to use the [WC\_Order\_Item\_Fee](https://docs.woocommerce.com/wc-apidocs/class-WC_Order_Item_Fee.html) methods **using first a foreach loop** this way:



```
// (optional if not defined) An instance of the WC_Order object
$the_order = wc_get_order( $order_id );

// Iterating through order fee items ONLY
foreach( $the_order->get_items('fee') as $item_id => $item_fee ){

    // The fee name
    $fee_name = $item_fee->get_name();

    // The fee total amount
    $fee_total = $item_fee->get_total();

    // The fee total tax amount
    $fee_total_tax = $item_fee->get_total_tax();
}

```

Tested and works


"
"Translate Rails model association - not working
Anyone has some tips as how to translate model associations in Rails?


For example: I have a Person model, which can have many Phone. However, a Person needs to have at least one Phone. I'm not able to translate that validation. The best I could do was this:



```
validates_presence_of :phones, :message => ""At least one phone is required.""

```

And on my YAML, I replaced this line to omit the `%{attribute}`:



```
format: ! '%{message}'

```

This way only my message is displayed, and I avoid the un-translated field name to be displayed.


This is causing me a lot of headache, because some gems simply don't allow me to pass `:message => ""something describing the error""`, so I wanted to configure all the error messages through my YAML.


Also, with some models I'm able to translate their attributes, while with others I'm not. For example:



```
activerecord:  
  attributes:
    additional_info:
      account_manager: ""Manager""

```

This works. I can see on my form ""Manager"". However, when this field has an error, Rails will display it as `""Additional info account manager can't be blank""`.


I tried this:



```
activerecord:          
  errors:
    models:
      additional_info:
        attributes:
          account_manager: ""Manager""

```

But no luck. 


I did read the docs, but no clue on why it's happening.


","Rails 3.2 has changed this behavior. The way I've posted before is deprecated. 


Now, in order to translate associations, there is the need to add a slash (instead of nesting everything). So instead of this:



```
    activerecord:
      attributes:
        person:
          additional_info:
            account_manager: ""Manager""

```

The correct now is:



```
    activerecord:
      attributes:
        person:
          additional_info/account_manager: ""Manager""

```

Also, I figured out that `has_many` associations are being translated differently from that. If you want to translate those, the following example may help:



```
    activerecord:
      attributes:
         emails:
           address: ""E-mail field""

```

Instead of the model name, like I did above, you need to pass the association name, in this case `emails`.


Check this comment and pull request for more info:


<https://github.com/rails/rails/commit/c19bd4f88ea5cf56b2bc8ac0b97f59c5c89dbff7#commitcomment-619858>


<https://github.com/rails/rails/pull/3859>


"
"GWT Atmosphere + RequestFactory
I'd like to produce following behaviour:


User in the browser saves some entity. After entity is saved on the server side, it generates notification and it must notify all other clients by means of Atmosphere with updated entity instance. 


So the problem is **how I can push POJO entity through Atmosphere interface**? 
Last one supports Serializable types to transfer over wire whereas RequestFactory has another serialization mechanism.


So I need transform POJO entity to autobean as this is done by RF and send to client. Anybody knows how it can be done?


","You can just use create AutoBeans and use them instead of POJOs in your client code.
AutoBeans can be easily serialized to/from JSON in GWT and JSON can be transferred.
That's what I do in my web-socket based GWT app.


On the server you can create the same AutoBean factory that you use on the client. Instead of GWT.create we use AutoBeanFactorySource:



```
MyDataFactory factory = AutoBeanFactorySource.create(MyDataFactory.class)
AutoBean<MyData> myDataAutoBean = factory.myData();
MyData data = myDataAutoBean.as();
data.setValue(1); // call all kinds of setters

// to converto AutoBean to JSON
String json = AutoBeanCodex.encode(myDataAutoBean).getPayload();

```

On the client side we use the [AutoBeanCodex](http://code.google.com/p/google-web-toolkit/wiki/AutoBean#AutoBeanCodex) to decode json into an AutoBean


"
"fread Function in C Programming
I have two questions about C's `fread` function:


1. I have read that `fread` is used to read a binary file. However, when I read a binary file with `fgets` using read mode `""r""` and a text file with `fread` using `""rb""` mode, the results are the same as reading a text file with `fgets` and a binary file with `fread`. So, why are there different functions for reading binary and text files?
2. I am using `fread` to read 10 bytes of a file in one call. How should I stop reading at the end of file – i.e. how is `EOF` specified in `fread`?


","answer of 1 question >


1>**fread**



```
size_t fread ( void * ptr, size_t size, size_t count, FILE * stream );

```

**Read block of data from stream** (try to understand this)


Reads an array of count elements, each one with a size of size bytes, from the stream and stores them in the block of memory specified by ptr.
The postion indicator of the stream is advanced by the total amount of bytes read.
The total amount of bytes read if successful is (size \* count).


**2>fgets**



```
char * fgets ( char * str, int num, FILE * stream );

```

**Get string from stream** (try to understand this)


Reads characters from stream and stores them as a C string into str until (num-1) characters have been read or either a newline or a the End-of-File is reached, whichever comes first.
A newline character makes fgets stop reading, but it is considered a valid character and therefore it is included in the string copied to str.
A null character is automatically appended in str after the characters read to signal the end of the C string.




---


**answer of 2nd question** 
in fread return value is


The total number of elements successfully read is returned as a size\_t object, which is an integral data type.
**If this number differs from the count parameter, either an error occured or the End Of File was reached.**


**You can use either ferror or feof to check whether an error happened or the End-of-File was reached.**


"
"Save WooCommerce custom Session Variables as order meta data
I'm trying to store custom session variables in database. Then, show them inside new order email and order details in WooCommerce Admin.


I have custom variables inside session:



```
add_action( 'init', 'oturum_degiskeni_olustur' );
function oturum_degiskeni_olustur () {
  // Early initialize customer session
  if ( isset(WC()->session) && ! WC()->session->has_session() ) {
      WC()->session->set_customer_session_cookie( true );
  }

  if ( isset( $_GET['konumu'] ) || isset( $_GET['masa_no'] ) ) {
      $konum = isset( $_GET['konumu'] )  ? esc_attr( $_GET['konumu'] )  : '';
      $masa  = isset( $_GET['masa_no'] ) ? esc_attr( $_GET['masa_no'] ) : '';

      // Set the session data
      WC()->session->set( 'custom_data', array( 'konum' => $konum, 'masa' => $masa ) );
  }
}

```

Firstly, I added custom variables to database with this code;



```
// Storing session variables for using them in order notifications
add_action( 'woocommerce_checkout_create_order', 'oturum_degiskeni_kaydet' );

function oturum_degiskeni_kaydet( $order, $data ) { 
  if ( $_POST['konumu'] ) update_meta_data( $order_id, '_konum', esc_attr( $_POST['konumu'] ) );
  if ( $_POST['masa_no'] ) update_meta_data( $order_id, '_masa', esc_attr( $_POST['masa_no'] ) );
}

```

Second, I added this variable's data to a new order email for admin.



```
// Show this session variables in new order email for admin 
add_action( 'woocommerce_email_after_order_table', 'konumu_emaile_ekle', 20, 4 );

function konumu_emaile_ekle( $order, $sent_to_admin, $plain_text, $email ) {
  if ( get_post_meta( $order->get_id(), '_konum', true ) ) echo '<p><strong>Konum :</strong> ' . get_post_meta( $order->get_id(), '_konum', true ) . '</p>';
  if ( get_post_meta( $order->get_id(), '_masa', true ) ) echo '<p><strong>Masa Numarası :</strong> ' . get_post_meta( $order->get_id(), '_masa', true ) . '</p>';
}

```

Last part of the code is shown session variables data in WooCommerce order page;



```
// Show session variable in woocommerce order page
add_action( 'woocommerce_admin_order_data_after_billing_address', 'konumu_admine_ekle', 10, 1 );

function konumu_admine_ekle( $order ) {

 $order_id = $order->get_id();
 if ( get_post_meta( $order_id, '_konum', true ) ) echo '<p><strong>Konum :</strong> ' . get_post_meta( $order_id, '_konum', true ) . '</p>';
 if ( get_post_meta( $order_id, '_masa', true ) ) echo '<p><strong>Masa Numarası :</strong> ' . get_post_meta( $order_id, '_masa', true ) . '</p>';
}

```

But, it does not work. When customer made an order, it gave an error ""We were unable to process your order, please try again.""


","**Updated:** There are some mistakes in your code when you are trying to save your custom data from session as order meta data and display it on emails and admin order pages…


Your first function is correct (oturum\_degiskeni\_olustur)…


Assuming that data is passed through URL like: `website.com/?konumu=newyork&masa_no=12`


Here is the revisited code:



```
// Unchanged
add_action( 'init', 'oturum_degiskeni_olustur' );
function oturum_degiskeni_olustur () {
    // Early initialize customer session
    if ( isset(WC()->session) && ! WC()->session->has_session() ) {
        WC()->session->set_customer_session_cookie( true );
    }

    if ( isset( $_GET['konumu'] ) || isset( $_GET['masa_no'] ) ) {
        $konum = isset( $_GET['konumu'] )  ? esc_attr( $_GET['konumu'] )  : '';
        $masa  = isset( $_GET['masa_no'] ) ? esc_attr( $_GET['masa_no'] ) : '';

        // Set the session data
        WC()->session->set( 'custom_data', array( 'konum' => $konum, 'masa' => $masa ) );
    }
}

// Save custom session data as order meta data
add_action( 'woocommerce_checkout_create_order', 'oturum_degiskeni_kaydet' );
function oturum_degiskeni_kaydet( $order ) {
    $data = WC()->session->get( 'custom_data' ); // Get custom data from session

    if ( isset($data['konum']) ) {
        $order->update_meta_data( '_konum', $data['konum'] );
    }
    
    if ( isset($data['masa']) ) {
        $order->update_meta_data( '_masa', $data['masa'] );
    }
    
    WC()->session->__unset( 'custom_data' ); // Remove session variable
}

// Show this session variables in new order email for admin and in woocommerce order page
add_action( 'woocommerce_email_after_order_table', 'konumu_emaile_admine_ekle', 20 );
add_action( 'woocommerce_admin_order_data_after_billing_address', 'konumu_emaile_admine_ekle' );
function konumu_emaile_admine_ekle( $order ) {

    if ( $konum = $order->get_meta( '_konum' ) )
        echo '<p><strong>Masa Numarası :</strong> ' . $konum . '</p>';

    if ( $masa = $order->get_meta( '_masa' ) )
        echo '<p><strong>Konum :</strong> ' . $masa . '</p>';
}

```

Code goes in functions.php file of your active child theme (or active theme). Tested and Works.


**On the email notification before customer details:**


[![enter image description here](https://i.stack.imgur.com/Rrt9v.png)](https://i.stack.imgur.com/Rrt9v.png)


**On admin order edit page *(under billing phone)*:**


[![enter image description here](https://i.stack.imgur.com/XokIz.png)](https://i.stack.imgur.com/XokIz.png)


*Tested on WooCommerce 4.2+ under storefront theme*


"
"How to fix ""Error while parsing the 'sandbox' attribute"" flags error with HtmlService?
I have just tried many of the sample [Google Apps Script scripts](https://developers.google.com/apps-script/guides/html/)


But I always get the error:



> 
> ""Error while parsing the 'sandbox' attribute: 'allow-modals', 'allow-popups-to-escape-sandbox' are invalid sandbox flags.""
> 
> 
> 


I have tried with Chrome and Safari on Mac OS and Chrome on Win 8.


","This error can be safely ignored, the script will continue to run fine. These are two new flags which have been implemented in Chrome 46 (the current stable release). If you're testing in a browser that doesn't support them, you'll see that error (which is really just a warning).


The `allow-modals` flag is required because Chrome will, by default, [block modal dialogs within a sandboxed iframe](https://www.chromestatus.com/features/4747009953103872). Adding this flag allows modals to work in your apps scripts.


The `allow-popups-to-escape-sandbox` flag is somewhat similar; its purpose is to [permit new windows to be created without them being subject to the same sandboxing restrictions](https://www.chromestatus.com/features/5708368589094912).


"
"In what sense const allows only atomic changes to mutable member variables?
I'm reading Functional Programming in C++ from [Ivan Čukić](https://cukic.co/), and I am having a hard time interpreting a point in the summary of Chapter 5:



> 
> - When you make a member function `const`, you promise that the function won't change any data in the class (not a bit of the object will change), or that any changes to the object *(to members declared as `mutable`) will be atomic as far as the users of the object are concerned*.
> 
> 
> 


If the part in *italic* was simply *are limited to members declared as `mutable`* I would have been happy with it. However, this rewording of mine seems to correspond to what the author put in parenthesis. What is out of parenthesis is what is puzzling me: what is the meaning of *atomic* in that sentence?


","The author is making a claim about best practices, not about the rules of the language.


You *can* write a class in which `const` methods alter `mutable` members in ways that are visible to the user, like this:



```
struct S {
    mutable int value = 0;
    int get() const {
        return value++;
    }
};
const S s;
std::cout << s.get();  // prints 0
std::cout << s.get();  // prints 1
// etc

```

You can do that, and it wouldn't break any of the rules of the language. However, you *shouldn't*. It violates the user's expectation that the `const` method should not change the internal state in an observable way.


There are legitimate uses for `mutable` members, such as memoization that can speed up subsequent executions of a `const` member function.


The author suggests that, as a matter of best practices, such uses of `mutable` members by `const` member functions *should* be atomic, since users are likely to *expect* that two different threads can call `const` member functions on an object concurrently.


If you violate this guideline, then you're not directly breaking any rules of the language. However, it makes it likely that users will use your class in a way that will cause data races (which are undefined behaviour). It takes away the user's ability to use the `const` qualifier to reason about the thread-safety of your class.


"
"Apply function to all elements of parameter pack from a variadic function
Consider the following (not working!) example:



```
#include <iostream>

template <typename type> void print(const type & item)
{
    std :: cout << item << std :: endl;
}

template <typename... types> void printall(const types & ... items)
{
    print(items)...;
}

int main()
{
    printall(1, 2, ""hello"");
}

```

Here I have a function `print` that simply prints out its argument, and a variadic function `printall` that accepts a pack of arguments. Now, what I would like to do is to simply have `printall` apply `print` to each element of the pack `items`. How can I get that done?


**Note**: I am *not* asking how to print a pack of values. I am aware of the existence of fold expressions and I know I could just throw the whole `items` in `std::cout` using them. Here `print` is just an example, could be any function.


How can I get that done? This sounds like something that should remarkably simple to do and yet I couldn't find any (reasonable) syntax to do it.


","
> 
> What I would like to do is to simply have `printall` apply `print` to
>  each element of the pack items. How can I get that done?
> 
> 
> 


## Option - 1


As user [@liliscent](https://stackoverflow.com/users/8605791/liliscent) and user [@max66](https://stackoverflow.com/users/6022656/max66) suggested in the comments, 
in C++11/ C++14 you can use the following hacky-way which act act like *fold expressions* in C++17.


`[SEE HERE](https://wandbox.org/permlink/MWsC2zQsSkh1G9xe)`



```
#include <iostream>

template <typename type> void print(const type& item)
{
    std::cout << item << '\n';
}

template <typename... types> 
void printall (const types&... items)
{
    using dummy = int[];
    (void)dummy { 0, (print(items), 0)... };
}

```



---


## Option - 2


If the above does not look like good enough, provide a classical variadic templates overload as *a wrapper/ helper* between your `printall()` and `print()` functions, so that each template function arguments can be accessed in `print()`.


`[SEE HERE](https://wandbox.org/permlink/nGtbaHaWCuvZbkdG)`



```
#include <iostream>

template <typename Type> void print(const Type& item)
{
    std::cout << item << '\n';  // print each argument
}

namespace helper 
{
    void printall() {}          // nontemplate overload for last call(i.e, no arguments call)

    template<typename FirstArg, typename... Types>
    void printall(const FirstArg& firstItem, Types&&... items)  
    {
        ::print(firstItem);                             // call print() for each argument
        helper::printall(std::forward<Types>(items)...);// calls the same wrapper::printalll()
    }
}

template <typename... Types> void printall(const Types& ... items)
{
    helper::printall(items...); // calls the wrapper::printall()
}

```



---


## Option - 3


However, if you have access to the C++17, just use [fold expressions](https://en.cppreference.com/w/cpp/language/fold). That provides a clean(non-hacky) and less amount of code.


`[SEE HERE](https://wandbox.org/permlink/bsG53MlVGlvuNfGq)`



```
template <typename type> void print(const type& item)
{
    std::cout << item << '\n';
}

template <typename... types> void printall(const types&... items)
{
    (print(items),...);
}

```

"
"Filter data to get only first day of the month rows
I have a dataset of daily data. I need to get only the data of the first day of each month in the data set (The data is from 1972 to 2013). So for example I would need Index `20`, Date `2013-12-02` value of `0.1555` to be extracted.
The problem I have is that the first day for each month is different, so I cannot use a step such as `relativedelta(months=1)`, how would I go about of extracting these values from my dataset?


Is there a similar command as I have found in another post for R?


[R - XTS: Get the first dates and values for each month from a daily time series with missing rows](https://stackoverflow.com/questions/22592193/r-xts-get-the-first-dates-and-values-for-each-month-from-a-daily-time-series)



```
17 2013-12-05 0.1621
18 2013-12-04 0.1698
19 2013-12-03 0.1516
20 2013-12-02 0.1555
21 2013-11-29 0.1480
22 2013-11-27 0.1487
23 2013-11-26 0.1648

```

","I would groupby the month and then get the zeroth (nth) row of each group.


First set as index (I think this is necessary):



```
In [11]: df1 = df.set_index('date')

In [12]: df1
Out[12]:
             n     val
date
2013-12-05  17  0.1621
2013-12-04  18  0.1698
2013-12-03  19  0.1516
2013-12-02  20  0.1555
2013-11-29  21  0.1480
2013-11-27  22  0.1487
2013-11-26  23  0.1648

```

Next sort, so that the first element is the first date of that month (Note: this doesn't *appear* to be necessary for nth, but I think that's actually a bug!):



```
In [13]: df1.sort_index(inplace=True)

In [14]: df1.groupby(pd.TimeGrouper('M')).nth(0)
Out[14]:
             n     val
date
2013-11-26  23  0.1648
2013-12-02  20  0.1555

```

another option is to resample and take the first entry:



```
In [15]: df1.resample('M', 'first')
Out[15]:
             n     val
date
2013-11-30  23  0.1648
2013-12-31  20  0.1555

```



---


Thinking about this, you can do this much simpler by extracting the month and then grouping by that:



```
In [21]: pd.DatetimeIndex(df.date).to_period('M')
Out[21]:
<class 'pandas.tseries.period.PeriodIndex'>
[2013-12, ..., 2013-11]
Length: 7, Freq: M

In [22]: df.groupby(pd.DatetimeIndex(df.date).to_period('M')).nth(0)
Out[22]:
    n       date     val
0  17 2013-12-05  0.1621
4  21 2013-11-29  0.1480

```

This time the sortedness of `df.date` *is* (correctly) relevant, if you know it's in descending date order you can use `nth(-1)`:



```
In [23]: df.groupby(pd.DatetimeIndex(df.date).to_period('M')).nth(-1)
Out[23]:
    n       date     val
3  20 2013-12-02  0.1555
6  23 2013-11-26  0.1648

```

*If this isn't guaranteed then sort by the date column first: `df.sort('date')`.*


"
"Disable form rising to the top
I want a form that when I set at the bottom of the z-order it stays there. I tried:



```
SetWindowPos(Handle,HWND_BOTTOM,Left,Top,Width,Height,SWP_NOZORDER);

```

and when I overlap it with some other apps it stays at the bottom as I need. However when I click on it, it rises to the top. I then tried:



```
SetWindowPos(Handle, HWND_BOTTOM, Left, Top, Width, Height,
             SWP_NOACTIVATE or SWP_NOZORDER);

```

and various other switches from this website...
<http://msdn.microsoft.com/en-us/library/ms633545.aspx>


But it still rises to the top.


","`SetWindowPos` sets the position of a window only when it is called, it does not establish a state. Handling [`WM_WINDOWPOSCHANGING`](http://msdn.microsoft.com/en-us/library/windows/desktop/ms632653%28v=vs.85%29.aspx) is the correct way to do this:



> 
> While this message is being processed, modifying any of the values in
>  WINDOWPOS affects the window's new size, position, or place in the Z
>  order. An application can prevent changes to the window by setting or
>  clearing the appropriate bits in the flags member of WINDOWPOS.
> 
> 
> 


  




```
type
  TForm1 = class(TForm)
    ..
  private
    procedure WindowPosChanging(var Msg: TWMWindowPosMsg);
        message WM_WINDOWPOSCHANGING;
  end;

..

procedure TForm1.WindowPosChanging(var Msg: TWMWindowPosMsg);
begin
  if Msg.WindowPos.flags and SWP_NOZORDER = 0 then
    Msg.WindowPos.hwndInsertAfter := HWND_BOTTOM;
  inherited;
end;

```

"
"why there is a deadlock in multithreaded program given below
I am new to multi threaded programming  

What is the reason for dead lock in this approach if one Thread has to print odd numbers from 0 to 1000 and other has to print even numbers?



```
using System;
using System.Collections.Generic;
using System.Linq;
using System.Text;
using System.Threading;

namespace ConsoleApplication9
{
    class Program
    {
        static int count1;
        static int count2;
        static Thread t1, t2;
        static void MulOf2()
        {
            while (count1 < 1000)
            {
                Console.Write(""Th1"" + (2 * count1) + ""\n"");
                count1++;
                if (t2.IsBackground)
                {
                    if (!t2.IsAlive)
                    {
                        t2.Resume();
                    }
                }
                t1.Suspend();
            }
        }
        static void Main(string[] args)
        {

            t1 = new Thread(MulOf2);
            t2 = new Thread(MulOf2Plus1);
            t1.Start();
            t2.Start();
        } 

         static void MulOf2Plus1()
        {
            while (count2 < 1000)
            {
                Console.Write(""Th2"" + ((2 * count2) + 1) + ""\n"");
                count2++;
                if (t1.IsBackground)
                {
                    if (!t1.IsAlive)
                    {
                        t1.Resume();
                    }
                }
                t2.Suspend();
            }
        }
    }
}

```

I modified the code to prevent crashes


","Assuming you swallow enough exceptions to even get the code running, it can deadlock when the operations execute in the following order:



```
t2.Resume() //on t1
t1.Resume() //on t2
t2.Suspend() //on t2
t1.Suspend() //on t1

```

As a result both threads remain suspended.


This is generally not the way to handle thread synchronization. Personally, I have never had to use `Resume` or `Suspend` on threads.


You should read about synchronization mechanisms in .NET, starting with the `lock` statement. I recommend the [free chapters](http://www.albahari.com/threading/) from Albahari's [c# in a Nutshell](http://www.albahari.com/nutshell/): 


"
"C3 chart not rendering in react project
I want to create a char using c3 in React to, later on, be updated each second.


I am trying to follow the example provided by 
[Updating C3 charts on props change with React](https://stackoverflow.com/questions/30018487/updating-c3-charts-on-props-change-with-react), but the first step, which is to create the chart, is not happening.


This is my fiddle:


<https://jsfiddle.net/69z2wepo/227446/>



```
import c3 from 'c3';
import React from ""react"";
import ReactDOM from ""react-dom"";

class Hello extends React.Component {

  renderChart() {
        this.chart = c3.generate({
            bindto:""#chart1"",
            data: {
                columns: [
                    ['data1', 30, 200, 100, 400, 150, 250],
                    ['data2', 50, 20, 10, 40, 15, 25]
                ]
            }
        });
    }



  render() {

    this.renderChart()

    return <div id=""chart1""></div>;
  }
}

ReactDOM.render(
  <Hello />,
  document.getElementById('container')
);

```

I installed c3 with npm and am importing it in the component.


Thanks for the help.


","In your example, it looks like the chart is generated before the selector div is even rendered, so the chart has no where to go. Instead of calling `this.renderChart()` in `render()`, you can call it `componentDidMount`. In that case, render will be called on the initial load, your `<div id=""chart1""></div>` will be rendered and then `renderChart` will run, adding the SVG to the div.


As for updating the data, you can move the column data itself to state, then call `setState` with some new data and use `componentDidUpdate` to rerender the chart. That might look something like this:



```
class Chart extends React.Component {
  constructor(props) {
    super(props);

    this.state = {
      column1: ['data1', 30, 200, 100, 400, 150, 250],
      column2: ['data2', 50, 20, 10, 40, 15, 25],
    };

    this.changeData = this.changeData.bind(this);
  }

  renderChart() {
    c3.generate({
      bindto: ""#chart1"",
      data: {
        columns: [this.state.column1, this.state.column2],
      },
    });
  }

  componentDidMount() {
    this.renderChart();
  }

  componentDidUpdate() {
    this.renderChart();
  }

  // Changes data to something arbitrary on button click
  changeData() {
    this.setState({
      column1: ['data1', 70, 120, 30, 300, 230, 300],
      column2: ['data2', 100, 120, 50, 140, 150, 80],
    });
  }

  render() {
    return (
      <div>
        <div id=""chart1""></div>
        <button onClick={this.changeData}>Change</button>
      </div>
    );
  }
}

```

React Lifecycle methods are key here. Here's a handy chart linked from the docs: <http://projects.wojtekmaj.pl/react-lifecycle-methods-diagram/>


"
"Better understanding of the SimpleAdapter's ViewBinder
I have 4 `TextViews`, 2 `ImageViews`, 2 `Buttons` and 2 widgets that are part of a row definition in a `ListView`. The data comes from XML and a `SimpleAdapter`. To access these `TextViews` I implement the `ViewBinde`r in a custom class and override the `setViewValue`. This works and the two `TextViews` I want to dynamically change are handled in the `setViewValue`. What is confusing to me is why my other two TextViews don't get passed through `setViewValue`. I say this based upon setting a breakpoint where the execution thread only enters twice. I was expecting to see it 4 or more times?


Here is the `setViewValue` where I have a breakpoint set.



```
@Override     
public boolean setViewValue(View view, Object data, String text)
{
    if(view.getId() == R.id.txtvw1)
    {             
//blah do some stuff
    }
    else if (view.getId() == R.id.txtvw2)
    {
//Blah do some stuff
    }

    return true;
} 

```

xml declaration of the `TextViews`(1 shows and 4 doesn't):



```
<TextView
        android:id=""@+id/txtvw1""
        android:layout_centerHorizontal=""true""
        android:layout_width=""185dp""
        android:layout_height=""25dp""
        android:textSize=""20sp""
        android:layout_marginTop=""60dp""
        android:gravity=""center""
        android:inputType=""none""
        android:text=""@string/str_StaticA""
        android:textColor=""#C0F700"" />

    <TextView
        android:id=""@+id/txtvw4""
        android:layout_alignParentLeft=""true""
        android:layout_marginLeft=""35dp""
        android:layout_width=""95dp""
        android:layout_height=""50dp""
        android:textSize=""18dp""
        android:layout_marginTop=""110dp""
        android:gravity=""center""
        android:inputType=""none""
        android:text=""IMHO:""
        android:textColor=""#FFBA19"" />

```

So in summary why does the execution enter the override only twice AND it just so happens to be the exact two I want to update?


Well through additional reading and deductive reasoning a few more pieces of information have come to light. In my mind ever View in the `ListView` row XML would pass through the `ViewBinder` but that was WRONG! From what I can figure you attach the `ViewBinder` to the DataAdapter via `setViewBinder`. Well since the DataAdapter is told what View(s) to use to populate the data in it it makes sense that only the two were showing. I was only revealing two to it.


Here is a snippet where I specify the Views involved in the populating process and pass them on the constructor of the DataAdapter:



```
String[] from = new String[] {""txtvw_PrevLift"", ""txtvw_PrevReps"", ""ActuLiftPikr"", ""ActulRepsPikr"" };
int[] to = {R.id.txtvw_PrevLift, R.id.txtvw_PrevReps, R.id.ActuLiftPikr, R.id.ActulRepsPikr };

LiftDataAdapter LiftDataAdapter = new LiftDataAdapter(this, LiftDataMaps, R.layout.liftdatalayout, from, to);

```

While this seems to make sense I don't know that my deductive reasoning is accurate. Will update as I find more information.


","If you set a `ViewBinder` on `SimpleAdapter` **all** the `Views` you declared(in the `to` array in your case) will be passed to the `ViewBinder`'s `setViewValue()` method no matter what. If the `setViewValue` method doesn't return `true` meaning the data binding for that `View` has failed(for whatever reason) then you get the default action(for a `TextView`), setting the text from the data `Hashmap`. Right now you declared the `setViewValue` to do something for only two `TextViews`(checking by their ids) and also to return `true` **no matter what**. When you'll enter the `setViewValue` for the other `Views` in your row layout you'll again enter the `setViewValue`, don't match the ids(so no updates for those `Views`) and return `true`(which will tell the adapter that the binding was **successful** and this `View` doesn't require any more work). [A look at the source code](http://grepcode.com/file/repository.grepcode.com/java/ext/com.google.android/android/1.5_r4/android/widget/SimpleAdapter.java#SimpleAdapter.bindView%28int,android.view.View%29) for the `SimpleAdapter.bindView` method will show you how the `ViewBinder` is used.


Also, I see that you use a custom adapter, are you sure you don't interfere with the normal `SimpleAdapter` logic?(also don't use the same name for the class and variable name)


"
"Is it guaranteed that array elements in C will be stored consecutively, with no padding?
In other words: is it guaranteed that if I have an array allocated this way:



```
void *arr = calloc(nmemb, sizeof(some_type))

```

Then `elta`, `eltb`, `eltc` will all point to the same location in memory, which will be the second element of type `some_type` of this array?



```
some_type *elta = &((some_type*)arr)[1];
some_type *eltb = ((some_type*)arr)+1;
some_type *eltc = (char*)arr+sizeof(some_type);

```

The reason I’m asking this is because I’m trying to do a “container” in C, and if this doesn’t hold then I’m out of ideas how to return a pointer to any other element than the first one.


","Yes, it is guaranteed. *If* padding bytes are added, they are added *within* `struct some_type`, but not in between two array elements.


E. g.:



```
struct S
{
    int n;
    short s;

// this is just for illustration WHERE byte padding (typically) would occur!!!
#if BYTE_ALIGNMENT >= 4
    unsigned char : 0;
    unsigned char : 0;
#endif
};
struct S s[2];
size_t d = (char*)(s + 1) - (char*)s;

```

With byte alignment adjusted to 4 or 8 (or even larger powers of 2), this struct will have size of 8 and d will be equally 8, with byte alignment set to 1 or 2, the struct will have size of 6 just as will be d...


Note: This is not the only place where padding bytes can occur: If you switched members `n` and `s`, padding bytes would be needed in between `s` and `n` to get `n` correctly aligned. On the other hand, no padding bytes would be necessary after n any more as the structure size would assure correct alignment already.


Referring to the standard: C11, 6.2.5.20:



> 
> An array type describes a **contiguously allocated** nonempty set of objects with a particular member object type, called the element type. 36) Array types are characterized by their element type and by the number of elements in the array. [...]
> 
> 
> 


(Highlighting by me!).


"
"Screen tearing in 11.10 with intel graphics
I have a slight screen tearing when watching videos in Ubuntu 11.10 x64, really annoying. Both in Unity and GNOME-Shell. Please focus the answer to GS since it's my primary DE and it's where I want it to work.


This is what I use:


- OpenGL vendor string: Tungsten Graphics, Inc
- OpenGL renderer string: Mesa DRI Intel(R) Sandybridge Mobile
- OpenGL version string: 2.1 Mesa 7.11
- OpenGL shading language version string: 1.20
- OpenGL extensions:


I guess it's because I can't get the screen to sync vertically but I don't know how to fix it. I found a thread talking about driconf and I've set the settings there to ""always sync vertically"". What can I do more to get this fixed?


Please help, any answer is welcome :) it would suck if I can't watch videos comfortably on my new laptop.


**Update:**


I thought it might have something to do with the screen refresh rate so I tried to change that manually. I found an online documentation for my laptop (DELL LATITUDE E5420):


<http://support.dell.com/support/edocs/systems/late5520/en/SM_EN/GUID-89CF3778-9361-4F01-A0A4-1D257AE39CAD.html>


There I found that the screen refresh rate was 60 Hz but I don't know how to change that manually in GNOME-Shell, so I tried changing the compiz settings in Unity to see if there was any difference. I still had the issue but I think it was not as severe, it's hard to tell.


**Update 2:**


The problem exists also in Unity 2D and Gnome-classic. So this problem is perhaps restricted to a display setting or driver problem?


**Update 3:**


The solution for GNOME-Shell is in the comments to the accepted answer.


","This is a bug with Intel Sandy Bridge graphics.


- [Launchpad bug report](https://bugs.launchpad.net/xserver-xorg-video-intel/+bug/755841)
- [Upstream bug report](https://bugs.freedesktop.org/show_bug.cgi?id=37686)


# Unity


A workaround for Unity (or Unity 2D with Compiz) is to enable the options *Force full screen redraws (buffer swap) on repaint* and maybe also *Don't wait for video sync* in the *Workarounds* section of CompizConfig Settings Manager:


![enter image description here](https://i.stack.imgur.com/kWaaa.png)


# GNOME Shell


In GNOME Shell, you can work around the bug by setting the `CLUTTER_PAINT` environment variable in */etc/environment*:



```
CLUTTER_PAINT=disable-clipped-redraws:disable-culling
CLUTTER_VBLANK=True

```

"
"Using 'switch' with strings in resource file
I have a bunch of strings in my resource(.resx) file. I am trying to directly use them as part of switch statement (see the sample code below).



```
class Test
{
    static void main(string[] args)
    {
        string case = args[1];
        switch(case)
        {
            case StringResources.CFG_PARAM1: // Do Something1 
                break;
            case StringResources.CFG_PARAM2: // Do Something2
                break;
            case StringResources.CFG_PARAM3: // Do Something3
                break;              
            default:
                break;
        }
    }
}

```

I looked at some of the solutions, most of them seem to suggest that I need to declare them as `const string` which I personally dislike. 
I liked the top voted solution for this question: [using collection of strings in a switch statement](https://stackoverflow.com/questions/3839978/is-there-a-way-in-c-to-use-strings-in-a-switch-statement-that-are-in-an-array-or/3840046#3840046). But then I need to make sure that my `enum` and `strings` in resource file are tied together. I would like to know a neat way of doing that.


**Edit:**
Also found [this great answer](https://stackoverflow.com/questions/371054/uses-of-action-delegate-in-c/371080#371080) while researching how to use `Action`:


","You could use a `Dictionary<string, Action>`. You put an `Action` (a delegate to a method) for each string in the Dictionary and search it. 



```
var actions = new Dictionary<string, Action> {
    { ""String1"", () => Method1() },
    { ""String2"", () => Method2() },
    { ""String3"", () => Method3() },
};

Action action;

if (actions.TryGetValue(myString, out action))
{
    action();
}
else
{
    // no action found
}

```

As a sidenote, if `Method1` is already an `Action` or a `void Method1()` method (with no parameters and no return value), you could do 



```
    { ""String1"", (Action)Method1 },

```

"
"what is the difference between installing prettier as a NPM package and installing prettier extension in VS Code
I'm new to web programming and I learned that we can type `npm i -D prettier` to install prettier as one of the `devDependency` and use it to format our code. Then I found that there is a VS Code extension Pretty that does exactly the same thing. 


I don't know what the differences are between the two.


If I only install the extension, can I format the codebase where the prettier npm package is not installed? 


Also does the configuration process differ for these two? Which one is preferred?


","Functionally there is no difference, they will both work.


The VS Code extension [Prettier](https://marketplace.visualstudio.com/items?itemName=esbenp.prettier-vscode#prettier-resolution) (not Pretty Formatter, that's different) includes a recent copy of the `prettier` npm package inside it, which it will use by default if you don't have the package installed via npm in your repo. See the extension page's section on [Prettier Resolution](https://marketplace.visualstudio.com/items?itemName=esbenp.prettier-vscode).


I want to use an example to illustrate why you might use npm over the extension.


Lets say you worked on a team and you wanted everyone to have the same version of prettier. It would be pretty hard to get everyone on the same version by VS Code extension. Developers would have to manage the versioning on their own, and would be difficult to all be on the same version (if needed). But If you used npm and had the package.json/lock file in a repository you could all easily be on the same version. If you needed to update to the same version you can just `npm install` and that will bring everyone on the same page. It might not be so important for prettier, but think about packages with breaking changes to the latest version.


so to answer your questions:



> 
> If I only install the extension, can I format the codebase where the prettier npm package is not installed?
> 
> 
> 


When using the VS Code extension, you can use it to format any file. If you don't have the npm package installed in a given project, the extension will use its built-in default package.



> 
> Also does the configuration process differ for these two? Which one is preferred?
> 
> 
> 


Both use `.prettierrc` for the (small) amount of things you can configure in Prettier. The VS Code extension comes with some additional configuration for how it should interact with your editor.


If you want your editor to do things like run prettier on save, or to be able to run it from the VS Code command palette, you should use the VS Code extension. Additionally, if you need a specific version of Prettier, or you need to run it from the command line, you can install it as an npm package in your repo. These two options are not mutually exclusive.


"
"Java SWT : Badge Notifications
I have a desktop based UI application written in Java SWT running on windows.


I want to add a button on the UI screen whose behaviour should be similar to the badges on an iphone or facebook notifications as shown in the images below.


The number on the badge will be dynamic and will increase or decrease based on the number of pending notifications.


How can I implement something similar in SWT/AWT?


### IOS Badge:


![enter image description here](https://i.stack.imgur.com/ZxPB5.jpg)
  
  



### Facebook Notification:


![enter image description here](https://www.techlicious.com/images/computers/facebook-notification-icon-100px.jpg)


","I've implemented something like that recently. You can simply paint a custom image with `GC`, and the overlay on your desired icon.


I'm including my helper class here. It's not the cleanest code (a lot of stuff is hardcoded), but you'll get the point. The notification bubble resizes itself depending on the number of notifications (max 999).


**How to use (Remember to cache and/or dispose your images!):**



```
Image decoratedIcon = new ImageOverlayer()
        .baseImage(baseImage) // You icon/badget
        .overlayImage(ImageOverlayer.createNotifImage(5)) // 5 notifications 
        .overlayImagePosition(OverlayedImagePosition.TOP_RIGHT)
        .createImage();

```

  




```
/**
 * <pre>
 * The difference between this and the ImageBuilder is 
 * that ImageOverlayer does not chain the images, rather
 * just overlays them one onto another.
 * 
 * 
 * Rules:
 * 
 * 1.) Images are not disposed. Resource handing must be done externally.
 * 2.) Only two images allowed, for now.
 * 3.) The size of the composite image should normally be the size of the
 *     base image, BUT: if the overlaying image is larger, then larger
 *     parameters are grabbed, and the base image is still underneath.
 * 4.) Use the builder APIs to set the base and overlaying images. The 
 *     position of the overlaying image is optional, and CENTER by default.
 *     When you've set these, simply call createImage()
 * 
 * Further improvements:
 * 
 * - Combine this with ImageBuilder. These two composers should be welded.
 * 
 * </pre>
 * 
 * @author grec.georgian@gmail.com
 *
 */
public class ImageOverlayer extends CompositeImageDescriptor
{

    // ==================== 1. Static Fields ========================

    public enum OverlayedImagePosition
    {
        TOP_LEFT, TOP_RIGHT, BOTTOM_LEFT, BOTTOM_RIGHT, CENTER;
    }


    // ====================== 2. Instance Fields =============================

    private ImageData baseImageData;

    private ImageData overlayedImageData;

    private OverlayedImagePosition overlayedImagePosition = OverlayedImagePosition.CENTER;



    // ==================== 3. Static Methods ====================

    /**
     * Creates a red circle with a white bold number inside it.
     * Does not cache the final image.
     */
    public static final Image createNotifImage(final int numberOfNotifications)
    {
        // Initial width and height - hardcoded for now
        final int width = 14;
        int height = 14;

        // Initial font size
        int fontSize = 100;

        int decorationWidth = width;

        String textToDraw = String.valueOf(numberOfNotifications);

        final int numberLength = Integer.toString(numberOfNotifications).length();

        if(numberLength > 3)
        {
            // spetrila, 2014.12.17: - set a width that fits the text
            //                       - smaller height since we will have a rounded rectangle and not a circle
            //                       - smaller font size so the new text will fit(set to 999+) if we have
            //                         a number of notifications with more than 3 digits
            decorationWidth += numberLength * 2;
            height -= 4;

            fontSize = 80;
            textToDraw = ""999+""; //$NON-NLS-1$
        }
        else if (numberLength > 2)
        {
            // spetrila, 2014.12.17: - set a width that fits the text
            //                       - smaller height since we will have a rounded rectangle and not a circle
            decorationWidth += numberLength * 1.5;
            height -= 4;
        }

        final Font font = new Font(Display.getDefault(), ""Arial"", width / 2, SWT.BOLD); //$NON-NLS-1$

        final Image canvas = new Image(null, decorationWidth, height);

        final GC gc = new GC(canvas);

        gc.setAntialias(SWT.ON);
        gc.setAlpha(0);
        gc.fillRectangle(0, 0, decorationWidth, height);

        gc.setAlpha(255);
        gc.setBackground(Display.getDefault().getSystemColor(SWT.COLOR_RED));

        // spetrila, 2014.12.17: In case we have more than two digits in the number of notifications,
        //                       we will change the decoration to a rounded rectangle so it can contain
        //                       all of the digits in the notification number
        if(decorationWidth == width)
            gc.fillOval(0, 0, decorationWidth - 1, height - 1);
        else
            gc.fillRoundRectangle(0, 0, decorationWidth, height, 10, 10);

        final FontData fontData = font.getFontData()[0];
        fontData.setHeight((int) (fontData.getHeight() * fontSize / 100.0 + 0.5));
        fontData.setStyle(SWT.BOLD);

        final Font newFont = new Font(Display.getCurrent(), fontData);

//      gc.setFont(AEFUIActivator.getDefault().getCustomizedFont(font, fontSize, SWT.BOLD));
        gc.setFont(newFont);
        gc.setForeground(Display.getDefault().getSystemColor(SWT.COLOR_WHITE));

        final Point textSize = gc.stringExtent(textToDraw);
        final int xPos = (decorationWidth - textSize.x) / 2;
        final int yPos = (height - textSize.y) / 2;
        gc.drawText(textToDraw, xPos + 1, yPos, true);

        gc.dispose();

        final ImageData imgData = canvas.getImageData();

        // Remove white transparent pixels
        final int whitePixel = imgData.palette.getPixel(new RGB(255,255,255));
        imgData.transparentPixel = whitePixel;
        final Image finalImage = new Image(null, imgData);

        canvas.dispose();
        font.dispose();
        newFont.dispose();

        return finalImage;
    }


    // ==================== 5. Creators ====================

    @Override
    public Image createImage()
    {
        if (baseImageData == null || overlayedImageData == null)
            throw new IllegalArgumentException(""Please check the ImageOverlayer. One of the overlaying images is NULL.""); //$NON-NLS-1$

        return super.createImage();
    }


    // ==================== 6. Action Methods ====================

    @Override
    protected void drawCompositeImage(final int width, final int height)
    {
        /*
         * These two determine where the overlayed image top left
         * corner should go, relative to the base image behind it.
         */
        int xPos = 0;
        int yPos = 0;

        switch (overlayedImagePosition)
        {
            case TOP_LEFT:
                break;

            case TOP_RIGHT:
                xPos = baseImageData.width - overlayedImageData.width;
                break;

            case BOTTOM_LEFT:
                yPos = baseImageData.height - overlayedImageData.height;
                break;

            case BOTTOM_RIGHT:
                xPos = baseImageData.width - overlayedImageData.width;
                yPos = baseImageData.height - overlayedImageData.height;
                break;

            case CENTER:
                xPos = (baseImageData.width - overlayedImageData.width) / 2;
                yPos = (baseImageData.height - overlayedImageData.height) / 2;
                break;

            default:
                break;
        }

        drawImage(baseImageData, 0, 0);
        drawImage(overlayedImageData, xPos, yPos);
    }


    // ==================== 7. Getters & Setters ====================

    final public ImageOverlayer overlayImagePosition(final OverlayedImagePosition overlayImagePosition)
    {
        this.overlayedImagePosition = overlayImagePosition;
        return this;
    }


    final public ImageOverlayer baseImage(final ImageData baseImageData)
    {
        this.baseImageData = baseImageData;
        return this;
    }


    final public ImageOverlayer baseImage(final Image baseImage)
    {
        this.baseImageData = baseImage.getImageData();
        return this;
    }


    final public ImageOverlayer baseImage(final ImageDescriptor baseImageDescriptor)
    {
        this.baseImageData = baseImageDescriptor.getImageData();
        return this;
    }


    final public ImageOverlayer overlayImage(final ImageData overlayImageData)
    {
        this.overlayedImageData = overlayImageData;
        return this;
    }


    final public ImageOverlayer overlayImage(final Image overlayImage)
    {
        this.overlayedImageData = overlayImage.getImageData();
        return this;
    }


    final public ImageOverlayer overlayImage(final ImageDescriptor overlayImageDescriptor)
    {
        this.overlayedImageData = overlayImageDescriptor.getImageData();
        return this;
    }


    @Override
    protected Point getSize()
    {
        // The size of the composite image is determined by the maximum size between the two building images,
        // although keep in mind that the base image always comes underneath the overlaying one.
        return new Point( max(baseImageData.width, overlayedImageData.width), max(baseImageData.height, overlayedImageData.height) );
    }

}

```

"
"Import .css file in Less using nodejs gulp
According to [this question](https://stackoverflow.com/questions/11196915/import-css-file-into-less-file) it's possible to import css file from less v.1.4.0.  

I tried to use `gulp-less` task, and it's not working.  

There's any way to import css file like this: `@import ""../spritesheet.css"";` with nodeJS gulp?


","There are several ways to include files in LESS, depending on what you need, using `@import ([keyword]) ""filename""`.


Allowed keywords are:



> 
> - reference: use a Less file but do not output it
> - inline: include the source file in the output but do not process it
> - less: treat the file as a Less file, no matter what the file extension
> - css: treat the file as a CSS file, no matter what the file extension
> - once: only include the file once (this is default behavior)
> - multiple: include the file multiple times
> - optional: continue compiling when file is not found
> 
> 
> 


[Lesscss.org - Import Options](http://lesscss.org/features/#import-options)


It's not clear to me what your use case is, but if you want to use the contents of `spritesheet.css` further on in your LESS file (for example extending classes defined in your spritesheet) you need the `(less)` keyword.


If you just want to include the contents of `spritesheet.css` into your output file without accessing them from within the LESS file at all, stick with `(inline)`.


"
"Scrapy hxs.select() not selecting all results
I am trying to scrapy to scrape odds from [here](http://www.paddypower.com/football/football-matches/premier-league).


Currently just trying to log the results with the following spider : 



```
def parse(self, response):         
   log.start(""LogFile.txt"", log.DEBUG);

   hxs = HtmlXPathSelector(response)
   sites = hxs.select('//div[@class=""fb_day_type_wrapper""]')

   items = []
   for site in sites:
       siteAddress = urlparse.urljoin(response.url, site.extract())
       self.log('Found category url: %s' % siteAddress)

```

This only logs the entry: This market is currently unavailable....
Not the other elements which contain the odds. 


I have tried a few different selectors with no luck. It looks like once I try and get inside of the element `div[@class=""fb_day_type_wrapper""]` I get nothing returned. I have the same results using the scrapy shell.


","The site uses javascript to generate the data table. There are some alternatives like [scrapyjs](https://github.com/scrapinghub/scrapyjs) or [splash](https://github.com/scrapinghub/splash) that allows to get the js-rendered html page. If you only need to scrape one page, you might be better off using Selenium.


Otherwise, you might need to go into hardcore mode and reverse engineer what is going on in the site with the data. I will show you how to do that.


First, start the `scrapy shell` so we can explore the web page:



```
scrapy shell http://www.paddypower.com/football/football-matches/premier-league

```

**Note:** I'm using python 2.7.4, ipython 0.13.2 and scrapy 0.18.0.


If you lookup in the source for ""Crystal Palace v Fulham"" in your browser, you will see there is a javascript code that have that reference. The `<script>` block looks like:



```
document.bodyOnLoad.push(function() {
    lb_fb_cpn_init(
        """",
        ""html"",
        ""MR_224"",
        {category: 'SOCCER',

```

We lookup in the shell for this element:



```
In [1]: hxs.select('//script[contains(., ""lb_fb_cpn_init"")]')
Out[1]: [<HtmlXPathSelector xpath='//script[contains(., ""lb_fb_cpn_init"")]' data=u'<script type=""text/javascript"">\n/* $Id: '>]

```

If you lookup into the `lb_fb_cpn_init` arguments, you will see the data we are looking for is passed as an argument in this form:



```
[{names: {en: 'Newcastle v Liverpool'}, ...

```

In fact there are three arguments like that:



```
In [2]: hxs.select('//script[contains(., ""lb_fb_cpn_init"")]').re('\[{names:')
Out[2]: [u'[{names:', u'[{names:', u'[{names:']

```

So we extract all of them, notice that we use a lot of regular expressions:



```
In [3]: js_args = hxs.select('//script[contains(., ""lb_fb_cpn_init"")]').re(r'(\[{names:(?:.+?)\]),')

In [4]: len(js_args)
Out[4]: 3

```

The idea here is that we want to parse the javascript code (which is a literal object) into python code (a dict). We could use `json.loads` but to do so the js code must be a valid json object, that is, have field names and strings enclosed in `""""`.


We proceed to do so. First I join the arguments in a single string as a javascript list:



```
In [5]: args_raw = '[{}]'.format(', '.join(js_args))

```

Then we enclose the field names into `""""` **and** replace with single quotes with double quotes:



```
In [6]: import re

In [7]: args_json = re.sub(r'(,\s?|{)(\w+):', r'\1""\2"":', args_raw).replace(""'"", '""')

```

This might not always work in all cases as the javascript code might have patterns that are not so easy to replace with a single `re.sub` and/or `.replace`.


We are ready to parse the javascript code as a json object:



```
In [8]: import json

In [9]: data = json.loads(args_json)

In [10]: len(data)
Out[10]: 3

```

Here, I'm just looking for the event name and odds. You can take a look to the `data` content to see what it looks like.


Luckily, the data seems to have a correlation:



```
In [11]: map(len, data)
Out[11]: [20, 20, 60]

```

You could as well build a single `dict` from the three of them by using the `ev_id` field. I will just assume that `data[0]` and `data[1]` hava a direct correlation and that `data[2]` contains 3 items per event. This can be easily verified with:



```
In [12]: map(lambda v: v['ev_id'], data[2])
Out [12]:
[5889932,
 5889932,
 5889932,
 5889933,
 5889933,
 5889933,
 ...

```

With some python-fu, we can merge the records:



```
In [13]: odds = iter(data[2])

In [14]: odds_merged = zip(odds, odds, odds)

In [15]: data_merged = zip(data[0], data[1], odds_merged)

In [16]: len(data_merged)
Out[16]: 20

```

Finally, we collect the data:



```
In [17]: get_odd = lambda obj: (obj['names']['en'], '/'.join([obj['lp_num'], obj['lp_den']]))

In [18]: event_odds = []

In [19]: for event, _, odds in data_merged:
   ....:     event_odds.append({'name': event['names']['en'], 'odds': dict(map(get_odd, odds)), 'url': event['url']})
   ....:     

In [20]: event_odds
Out[20]: 
[{'name': u'Newcastle v Liverpool',
  'odds': {u'Draw': u'14/5', u'Liverpool': u'17/20', u'Newcastle': u'3/1'},
  'url': u'http://www.paddypower.com/football/football-matches/premier-league-matches/Newcastle%2dv%2dLiverpool-5889932.html'},
 {'name': u'Arsenal v Norwich',
  'odds': {u'Arsenal': u'3/10', u'Draw': u'9/2', u'Norwich': u'9/1'},
  'url': u'http://www.paddypower.com/football/football-matches/premier-league-matches/Arsenal%2dv%2dNorwich-5889933.html'},
 {'name': u'Chelsea v Cardiff',
  'odds': {u'Cardiff': u'10/1', u'Chelsea': u'1/4', u'Draw': u'5/1'},
  'url': u'http://www.paddypower.com/football/football-matches/premier-league-matches/Chelsea%2dv%2dCardiff-5889934.html'},
 {'name': u'Everton v Hull',
  'odds': {u'Draw': u'10/3', u'Everton': u'4/9', u'Hull': u'13/2'},
  'url': u'http://www.paddypower.com/football/football-matches/premier-league-matches/Everton%2dv%2dHull-5889935.html'},
 {'name': u'Man Utd v Southampton',
  'odds': {u'Draw': u'3/1', u'Man Utd': u'8/15', u'Southampton': u'11/2'},
  'url': u'http://www.paddypower.com/football/football-matches/premier-league-matches/Man%2dUtd%2dv%2dSouthampton-5889939.html'},
 ...

```

As you can see, web scraping can be very challenging (and fun!). All it depends how the website displays the data. Here you could save time by just using Selenium, but if you are looking to scrape a large website, Selenium will be very slow compared to Scrapy.


Also you have to consider whether the site will get code updates very often, in that case you will spend more time reverse engineering the js code. In that case a solution like [scrapyjs](https://github.com/scrapinghub/scrapyjs) or [splash](https://github.com/scrapinghub/splash) can be a better option.


Final remarks:


- Now you have all the code required to extract the data. You need to integrate this into your spider callback and build your item.
- Don't use `log.start`. Use the setting `LOG_FILE` (command line argument: `--set LOG_FILE=mylog.txt`).
- Remeber that `.extract()` always returns a list.


"
"Computing the sum of squares from a file in Python
I'm trying to implement a program that computes the sum of the squares of numbers read from a file.


The numbers in the .txt file are:



```
37
42
59
14
25
95
6

```

So far I have:



```
def squareEach(nums):
    nums = nums * nums

def sumList(nums):
    nums = nums + nums

def toNumbers(strList):
    while line != """":
            line = int(line)
            line = infile.readline()

def main():
    print ""** Program to find sum of squares from numbers in a file **""
    fileName = raw_input(""What file are the numbers in?"")
    infile = open(fileName, 'r')
    line = infile.readline()
    toNumbers(line)
    squareEach(line)
    sumList(line)
    print sum

```

When I run the program through I get:



```
main()
** Program to find sum of squares from numbers in a file **
What file are the numbers in?C:\Users\lablogin\Desktop\mytextfile.txt

Traceback (most recent call last):
File ""<pyshell#17>"", line 1, in <module>
main()
File ""<pyshell#16>"", line 6, in main
toNumbers(strList)
NameError: global name 'strList' is not defined

```

","Normally, I would include a much more detailed explanation, but since I have to run, I'll leave you with an amended version of your code that will work



```
def squareEach(nums):
    answer = []
    for num in nums:
        answer.append(num*num)
    return answer

def sumList(nums):
    answer = 0
    for num in nums:
        answer += num
    return answer

def toNumbers(strList):
    answer = []
    for numStr in strList.split():
        try:
            num = int(numStr)
            answer.append(num)
        except:
            pass
    return answer

def main():
    print ""** Program to find sum of squares from numbers in a file **""
    fileName = raw_input(""What file are the numbers in? "")
    sum = 0
    with open(fileName, 'r') as infile:
        for line in infile:
            nums = toNumbers(line)
            squares = squareEach(nums)
            sum += sumList(squares)
        print sum

```

Of course, you could do all of it in one line:



```
print ""the sum is"", sum(int(num)**2 for line in open(raw_input(""What file are the numbers in? "")) for num in line.strip()split())

```

Hope this helps


"
"numpy array concatenate: ""ValueError: all the input arrays must have same number of dimensions""
How to concatenate these `numpy` arrays?


first `np.array` with a shape `(5,4)`



```
[[  6487    400 489580      0]
 [  6488    401 492994      0]
 [  6491    408 489247      0]
 [  6491    408 489247      0]
 [  6492    402 499013      0]]

```

second `np.array` with a shape `(5,)`



```
[  16.   15.   12.  12.  17. ]

```

final result should be



```
[[  6487    400    489580    0   16]
 [  6488    401    492994    0   15]
 [  6491    408    489247    0   12]
 [  6491    408    489247    0   12]
 [  6492    402    499013    0   17]]

```

I tried `np.concatenate([array1, array2])`
but i get this error


`ValueError: all the input arrays must have same number of dimensions`


What am I doing wrong?


","To use `np.concatenate`, we need to extend the second array to `2D` and then concatenate along `axis=1` -



```
np.concatenate((a,b[:,None]),axis=1)

```

Alternatively, we can use `np.column_stack` that takes care of it -



```
np.column_stack((a,b))

```

Sample run -



```
In [84]: a
Out[84]: 
array([[54, 30, 55, 12],
       [64, 94, 50, 72],
       [67, 31, 56, 43],
       [26, 58, 35, 14],
       [97, 76, 84, 52]])

In [85]: b
Out[85]: array([56, 70, 43, 19, 16])

In [86]: np.concatenate((a,b[:,None]),axis=1)
Out[86]: 
array([[54, 30, 55, 12, 56],
       [64, 94, 50, 72, 70],
       [67, 31, 56, 43, 43],
       [26, 58, 35, 14, 19],
       [97, 76, 84, 52, 16]])

```

If `b` is such that its a `1D` array of `dtype=object` with a shape of `(1,)`, most probably all of the data is contained in the only element in it, we need to *flatten* it out before concatenating. For that purpose, we can use `np.concatenate` on it too. Here's a sample run to make the point clear -



```
In [118]: a
Out[118]: 
array([[54, 30, 55, 12],
       [64, 94, 50, 72],
       [67, 31, 56, 43],
       [26, 58, 35, 14],
       [97, 76, 84, 52]])

In [119]: b
Out[119]: array([array([30, 41, 76, 13, 69])], dtype=object)

In [120]: b.shape
Out[120]: (1,)

In [121]: np.concatenate((a,np.concatenate(b)[:,None]),axis=1)
Out[121]: 
array([[54, 30, 55, 12, 30],
       [64, 94, 50, 72, 41],
       [67, 31, 56, 43, 76],
       [26, 58, 35, 14, 13],
       [97, 76, 84, 52, 69]])

```

"
"Matplotlib histogram with multiple legend entries
I have this code that produces a histogram, identifying three types of fields; ""Low"", ""medium"" , and ""high"":



```
import pylab as plt
import pandas as pd


df = pd.read_csv('April2017NEW.csv', index_col =1)
df1 = df.loc['Output Energy, (Wh/h)']  # choose index value and Average
df1['Average'] = df1.mean(axis=1)

N, bins, patches = plt.hist(df1['Average'], 30)

cmap = plt.get_cmap('jet')
low = cmap(0.5)
medium =cmap(0.25)
high = cmap(0.8)


for i in range(0,4):
    patches[i].set_facecolor(low)
for i in range(4,11):
    patches[i].set_facecolor(medium)
for i in range(11,30):
    patches[i].set_facecolor(high)

plt.xlabel(""Watt Hours"", fontsize=16)  
plt.ylabel(""Households"", fontsize=16)
plt.xticks(fontsize=14)  
plt.yticks(fontsize=14)
ax = plt.subplot(111)  
ax.spines[""top""].set_visible(False)  
ax.spines[""right""].set_visible(False)

plt.show()

```

that produces this:


[![enter image description here](https://i.stack.imgur.com/MXdGq.png)](https://i.stack.imgur.com/MXdGq.png)


How do I get a legend in there for the three different colors?


","You would need to create the legend yourself. To this end, create some rectangles, which are not shown in the figure (so called proxy artists). 



```
#create legend
handles = [Rectangle((0,0),1,1,color=c,ec=""k"") for c in [low,medium, high]]
labels= [""low"",""medium"", ""high""]
plt.legend(handles, labels)

```

Complete example:



```
import matplotlib.pyplot as plt
import numpy as np
from matplotlib.patches import Rectangle

data = np.random.rayleigh(size=1000)*35

N, bins, patches = plt.hist(data, 30, ec=""k"")

cmap = plt.get_cmap('jet')
low = cmap(0.5)
medium =cmap(0.25)
high = cmap(0.8)


for i in range(0,4):
    patches[i].set_facecolor(low)
for i in range(4,11):
    patches[i].set_facecolor(medium)
for i in range(11,30):
    patches[i].set_facecolor(high)

#create legend
handles = [Rectangle((0,0),1,1,color=c,ec=""k"") for c in [low,medium, high]]
labels= [""low"",""medium"", ""high""]
plt.legend(handles, labels)

plt.xlabel(""Watt Hours"", fontsize=16)  
plt.ylabel(""Households"", fontsize=16)
plt.xticks(fontsize=14)  
plt.yticks(fontsize=14)

plt.gca().spines[""top""].set_visible(False)  
plt.gca().spines[""right""].set_visible(False)

plt.show()

```

[![enter image description here](https://i.stack.imgur.com/FpzNm.png)](https://i.stack.imgur.com/FpzNm.png)


"
"ParallelEnumerable.Range vs Enumerable.Range.AsParallel?
What is the differences between `ParallelEnumerable.Range` and `Enumerable.Range(...).AsParallel()`?


`ParallelEnumerable.Range` creates range partition( best for operations where cpu time is equal foreach item)


where as `Enumerable.Range(...).AsParallel()` might be executes as `range` or `chunk`


Is there any performance difference ? When should I use which ?


","If you're performing an operation where it's CPU pressure grows as it iterates then you're going to want to use `AsParallel` because it can make adjustments as it iterates. However, if it's something that you just need to run in parallel and the operation doesn't create more pressure as it iterates, then use `ParallelEnumerable.Range` because it can partition the work immediately.


Reference [this article](http://blogs.msdn.com/pfxteam/archive/2007/12/02/6558579.aspx) for a more detailed explanation.


So, let's assume you're performing some complex math in each iteration, and as the input values grow the math takes longer, you're going to want to use `AsParallel` so that it can make adjustments. But if you use that same option with something like setting a registry setting, you'll see a performance decrease because there is overhead associated with `AsParallel` that is not necessary.


"
"Why is upper casing not enough for case-insensitive comparison?
To compare two strings case insensitively, one correct way is to case
fold them first. How is this better than upper casing or lower casing?


I find examples where lower casing doesn't work right online. For
example ""σ"" and ""ς"" (two forms of ""Σ"") don't become the same when
converted to lower case. But I've failed to find why case folding is
better than mapping to upper case. Is there a case where two strings
that should match case insensitively don't upper case to the same
strings?


Another scenario is when I want to store a case insensitive index. The
recommended way seems to be case folding and then normalizing. What are
its advantages over storing the string mapped to upper case and
normalized? The specs say mapping to upper case is not guaranteed to be
stable across versions of Unicode while case folding is. But are there
any cases where mapping to upper case gives a different string in an
earlier version of Unicode?


","As per [Unicode stability policy](https://www.unicode.org/policies/stability_policy.html#Case_Pair), case mappings are only stable for case pairs, i.e. pairs of characters *X* and *Y* where *X* is the full uppercase mapping of *Y*, and *Y* is the full lowercase mapping of *X*. Only when both these characters exist with these properties is the casing relation between them set in stone.


However, Unicode contains many “incomplete” case pairs where only the lowercase form has been encoded and the uppercase form is missing completely. This is usually the case for letters used in transcription systems that are traditionally lowercase-only. Should capital forms be discovered and subsequently added to Unicode, these letters would then receive a new uppercase mapping.


The most recent characters this has happened to are “ʂ” (from Unicode 1.1), “ᶎ” (from Unicode 4.1), and “ꞔ” (from Unicode 7.0), which all got brand new uppercase forms (Ꞔ, Ʂ, Ᶎ) in Unicode 12.0 two years ago.


Because case mappings do not have to be unique, this makes uppercasing a poor substitute for proper case-folding. For example, both U+0434 (д) and U+1C81 (ᲁ) uppercase to U+0414 (Д), but only the former is locked into a case pair by virtue of being U+0414’s full lowercase mapping. If someone were to find a dedicated capital letter version of U+1C81 in some old manuscript, it would be given a new uppercase mapping, resulting in U+0434 and U+1C81 suddenly no longer comparing equal under that operation.


**EDIT:** I have just remembered a current example of uppercasing not being sufficient for case-insensitive matching: U+1E9E (ẞ) is already a capital letter and thus uppercases to itself. Its lowercase counterpart is U+00DF (ß), but the uppercase mapping of U+00DF is the sequence <U+0053, U+0053> (SS).



```
uppercase(""ẞ"") ≠ uppercase(lowercase(""ẞ""))

```

"
"How to create a click event for polymer-based webpage
Apologies in advanced, but I'm a complete noob at Polymer.


So far I've done an HTML file (treated as external dialog box), with some stuff in it.


First off, the user clicks a button from the main page, which then displays a dialog box. From the dialog box, the user will type in text (at this point, anything), and then click on the submit button. At which point, either it succeeds, or if the textbox is empty, an error is thrown, the textbox is selected, and (if possible) turns red / displays an error message.


Unfortunately, even getting the polymer event down escapes me.


Here's my code so far:


The main page (works fine):



```
<!DOCTYPE HTML>
<html>
<head>
    <link rel=""import"" href=""elements/accountability-ticket.html""/>
</head>
<body>
    <section>
        <paper-button raised label=""Buy Now"" id=""ticket"" onclick=""clickHandler('diagTicket')"">Buy Now</paper-button>
    </section>
    <section><accountability-ticket></accountability-ticket></section>
    <script>
    window.addEventListener('WebComponentsReady', function(e) {
      // imports are loaded and elements have been registered
      console.log('Components are ready');
    });


    function clickHandler(e) {
        var dialog = document.getElementById(e);
        if (dialog) {
            dialog.open();
        }
    }
    </script>
</body>
</html>

```

There's a button there, ""Buy Now"", which calls the below code as a pop-up dialog form:



```
<dom-module id=""accountability-ticket"">
    <template>
        <paper-dialog with-backdrop entry-animation=""scale-up-animation""
                      exit-animation=""fade-out-animation"" id=""diagTicket"">
            <h2>I Own It Ticket </h2>
            <div class=""ctrlButtons flex"">  
                <paper-button dialog-dismiss>Cancel</paper-button>
                <paper-button dialog-confirm>Submit</paper-button>
            </div>
        </paper-dialog>
    </template>
</dom-module>
<script>
Polymer({
    is: ""accountability-ticket""
});

</script>

```

According to my searches online, to get the button to check if the textbox is empty on click, I'll need an eventHandler first and foremost. I understand I'm supposed to put something like this:



```
on-click=""confirmClick""

```

in the `<paper-button dialog-confirm>Submit</paper-button>`.


And presumably in the `<script>` part, this:



```
Polymer('my-element', {
    confirmClick: function() {
    alert('Click event triggered');
});

```

But given that I already have this:



```
Polymer({
    is: ""accountability-ticket""
});

```

above, and so far my attempts at putting `confimClick` in it results in the entire dialog box not appearing when the user clicks the ""Buy Now"", I'm not sure what to do.


For now, how do I setup the click event? Thanks.


","To register event handlers, remove the double braces. 



```
<paper-button on-click=""confirmClick"">Confirm</paper-button>

```

**Update**


Thanks for the update. I now realise that you are trying to set up the click handler on your main page outside of a Polymer element. The syntax above only works inside `<dom-module>`. You can however use the ""traditional"" approach for setting up event handlers.



```
window.addEventListener('WebComponentsReady', function(e) {
  // imports are loaded and elements have been registered
  document.querySelector(""#ticket"").addEventListener(""click"", clickHandler);
});

function clickHandler(e) {
    // your code here
}

```

**Update**


If on the other hand you are inside a Polymer element, you can add the event handler to the prototype. The `id` in the `dom-module` where the handler is registered must match the `is` property in the prototype (in this case accountability-ticket). So assuming that your module looks like this: 



```
<dom-module id=""accountability-ticket"">
    <template>
        <paper-button on-click=""confirmClick"">Confirm</paper-button>
    </template>
</dom-module>

```

You add the handler to the prototype in the following way:



```
<script>
    Polymer({
        is: ""accountability-ticket"",
        confirmClick: function(event){
            console.log(event);
        }
    });

</script>>

```

"
"How to add JWT to authorization header?
As described in the following slide, it is necessary that the client sends the `jwt` back to the server by an `Authorization Header` at the next request. 


[![enter image description here](https://i.stack.imgur.com/cXxYn.png)](https://i.stack.imgur.com/cXxYn.png)


But how can I define the `Authorization Header` and add the JWT to the server?


My current status is:


1. User sends `username` and `password` to the server by a `POST` request.
2. The server creates the the `JWT`.
3. The server sends the signed `JWT` back to the client and saves it in a cookie.


Now my questions:


- **In case of a Login:**


As I understand it, now its necessary to send the JWT back to the server. The server verifies the token and sends it back to finish the login process.


How can I add the `JWT` to the `Authorization Header`?
- **In case of running a process and receiving data from a calculation:**


Do I understand right, that the client has to send the `JWT` from the login to the server and a second `JWT` with the data; or can I send the data by `POST` request?


","So, You are pretty much correct with JWT. All you need to do when sending data from client to server (after JWT creation), is to add it to the request header. Many folks will try to keep along the same path as OAuth and add a Bearer token similar to the node snippet below:



```
var rp = require('request-promise');
options = {
  method: GET,
  uri: 'https://www.example.com/api/sample',
  headers: {
    Authorization: ""Bearer <insert_your_JWT_here>""
  }
}
rp(options).then(function(res){
  <handle_response>
}

```

Granted I know you mentioned PHP, but the workflows are the same, its just the syntax is different.


Now, to verify that this token is present, the server would need to *verify()* that the token is valid with the secret that was defined. In every request made by the client, for an authorized endpoint, you would need to send this token everytime.


"
"Oracle SQL can't find the cause of the error
I keep getting an invalid identifier error on part of my code for creating tables. Every other part of my code executes fine but this one table does not.



```
CREATE TABLE saleinv
(saleinv VARCHAR(6)
,cname VARCHAR (20) NOT NULL
,salesman VARCHAR(20) NOT NULL
,saledate DATE
,serial VARCHAR(8) NOT NULL
,totalprice NUMERIC(9,2)
,discount NUMERIC(8,2)
,net NUMERIC(9,2)
,tax NUMERIC(8,2)
,licfee NUMERIC(6,2)
,comission NUMERIC(8,2)
,tradeserial VARCHAR(8) NOT NULL
,tradeallow NUMERIC(9,2)
,fire VARCHAR(1)
      CHECK (fire IN ('Y','N'))
,collision VARCHAR(1)
      CHECK (collision IN ('Y','N'))
,liability VARCHAR(1)
      CHECK (liability IN ('Y','N'))
,property VARCHAR(1)
      CHECK (property IN ('Y','N'))
,CONSTRAINT salesinv_saledate_nn NOT NULL
      CHECK (saledate > TO_DATE('01-JAN-1990','DD-MON-YYYY'))
,CONSTRAINT saleinv_pk PRIMARY KEY (saleinv)
,CONSTRAINT saleinv_cname_fk FOREIGN KEY (cname) REFERENCES customer (cname)
,CONSTRAINT saleinv_salesman_fk FOREIGN KEY (salesman) REFERENCES employee (empname)
,CONSTRAINT saleinv_serial_fk FOREIGN KEY (serial) REFERENCES car (serial)
,CONSTRAINT saleinv_tradeserial_fk FOREIGN KEY (tradeserial) REFERENCES car (serial)
)

```

This is the output:



```
Error report -
ORA-00904: : invalid identifier
00904. 00000 -  ""%s: invalid identifier""
*Cause:    
*Action:

```

If anyone could help me figure out why this error is happening would be greatly appreciated.


","SQL\*Plus is such a nice tool. Those who live by GUI, die by GUI. 


Scroll down; then remove ""NOT NULL"" from line 23 and put it into line 5.



```
SQL> CREATE TABLE saleinv
  2  (saleinv VARCHAR(6)
  3  ,cname VARCHAR (20) NOT NULL
  4  ,salesman VARCHAR(20) NOT NULL
  5  ,saledate DATE
  6  ,serial VARCHAR(8) NOT NULL
  7  ,totalprice NUMERIC(9,2)
  8  ,discount NUMERIC(8,2)
  9  ,net NUMERIC(9,2)
 10  ,tax NUMERIC(8,2)
 11  ,licfee NUMERIC(6,2)
 12  ,comission NUMERIC(8,2)
 13  ,tradeserial VARCHAR(8) NOT NULL
 14  ,tradeallow NUMERIC(9,2)
 15  ,fire VARCHAR(1)
 16        CHECK (fire IN ('Y','N'))
 17  ,collision VARCHAR(1)
 18        CHECK (collision IN ('Y','N'))
 19  ,liability VARCHAR(1)
 20        CHECK (liability IN ('Y','N'))
 21  ,property VARCHAR(1)
 22        CHECK (property IN ('Y','N'))
 23  ,CONSTRAINT salesinv_saledate_nn NOT NULL
 24        CHECK (saledate > TO_DATE('01-JAN-1990','DD-MON-YYYY'))
 25  ,CONSTRAINT saleinv_pk PRIMARY KEY (saleinv)
 26  ,CONSTRAINT saleinv_cname_fk FOREIGN KEY (cname) REFERENCES customer (cname)
 27  ,CONSTRAINT saleinv_salesman_fk FOREIGN KEY (salesman) REFERENCES employee (empname)
 28  ,CONSTRAINT saleinv_serial_fk FOREIGN KEY (serial) REFERENCES car (serial)
 29  ,CONSTRAINT saleinv_tradeserial_fk FOREIGN KEY (tradeserial) REFERENCES car (serial)
 30  );
,CONSTRAINT salesinv_saledate_nn NOT NULL
                                 *
ERROR at line 23:
ORA-00904: : invalid identifier


SQL>

```

"
"C# - Form2 value to Form1
I'm having trouble passing values ​​entered in form2(citacao) to form1(principal).


Principal.cs (form1)



```
richEditControl1.Document.AppendText(citacao.valor_edit[0]);

```

Citacao.cs (form2)



```
public string[] valor_edit = new string[3];
private void simpleButton2_Click(object sender, EventArgs e)
{
    valor_edit[0] = memoEdit1.Text;
    valor_edit[1] = comboBox1.SelectedItem.ToString();
    valor_edit[2] = textEdit1.Text;

} 

```

But when I click the button nothing happens , the values ​​are not inserted into the richedit I like it.


I already have this on form (Pass DataGrid to ComboBox)


Form1 (principal)



```
private void barButtonItem1_ItemClick(object sender, DevExpress.XtraBars.ItemClickEventArgs e)
{
    citacao cita = new citacao(this);
    cita.Show();
}

```

form2(citação)



```
public citacao(principal gridForm)
{
    InitializeComponent();
    frm1 = gridForm;
}

// LOAD ALL FONTS (Referencias);
private void citacao_Load(object sender, EventArgs e)
{
    comboBox1.Items.Clear();
    foreach (DataGridViewRow row in frm1.DataGridView1.Rows)
    {
        comboBox1.Items.Add(row.Cells[0].Value.ToString());
    }
    comboBox1.SelectedIndex = 0;
}

```

","let's see whether I understood your situation :)


declare your variable in Form 1 as a class variable 



```
private citacao cita;

```

then initialize it in the button press event



```
private void barButtonItem1_ItemClick(object sender, DevExpress.XtraBars.ItemClickEventArgs e)
{
    cita = new citacao(this);
    // subscribe to the closing event
    cita.FormClosing += form_FormClosing;
    cita.Show();
}

// when Form 2 will be closed you can execute your important line in the event
void form_FormClosing(object sender, FormClosingEventArgs e)
{
    // BUT! you have to use the variable name!
    richEditControl1.Document.AppendText(cita.valor_edit[0]);    
}

```

**EDIT:**


Ok after looking at the entire code:
please remove the button3! and this entire code:



```
private void button3_Click(object sender, EventArgs e)
{
    cita = new citacao(this);
    richEditControl1.Document.AppendText(citacao.valor_edit); // this line is the problem!
}

```

The function `AppendText` probably needs a `string` as parameter and you give the entire array! 
If you subscribe to the closing event in Form1 / principal and also implement
the event, your data will be transmitted automatically as soon as the Form 2 disappears from the screen :)


"
"possible to index a set by a variable?
I am trying to do something that logically should be possible to do. However, I am not sure how to do this within the realm of linear programming. I am using ZMPL/SCIP, but this should be readable to most.



```
set I := {1,2,3,4,5};
param u[I] := <1> 10, <2> 20, <3> 30, <4> 40, <5> 50;

var a;
var b;

subto bval:
  b == 2;

subto works:
  a == u[2];

#subto does_not_work:
#  a == u[b];

```

I am trying to make sure that the variable `a` is equal to the value at the index `b` in `u`. So for example, I ensure that `b == 2` and then I try to set the constraint that `a == u[b]`, but that does not work. It complains that I am trying to index with a variable. I am able to just do `a == u[2]` however, which makes `a` equal to `20`. 


Is there a way to easily access `u` at an index specified by a variable? Thanks for any help/guidance.




---


**EDIT**: I think the consensus is that this is not possible because it no longer becomes an LP. In that case, can anyone think of another way to write this so that, depending on the value of `b`, I can get an associated value from the set `u`? This would have to avoid directly indexing it.




---


**SOLUTION**: Based on the response from Ram, I was able to try it out and found that it was definitely a viable and linear solution. Thanks, Ram! Here is sample solution code in ZMPL:



```
set I := {1,2,3,4,5};
param u[I] := <1> 10, <2> 20, <3> 30, <4> 40, <5> 50;

var a;
var b;
var y[I] binary;

subto bval:
  b == 4;

subto only_one:
  sum <i> in I : y[i] == 1;

subto trick:
  b == (sum <i> in I : y[i] * i);

subto aval:
  (sum <i> in I : u[i]*y[i]) == a;

```

","Yes, you can rewrite and *linearize* your constraints, by introducing a few extra 0/1 variables (indicator variables). These kinds of tricks are not uncommon in Integer Programming.


**Constraints In English**


`b` can take on values from 1 through 5. b = {1..5}


and depending on b's value, the variable `a` should become `u[b]`


**Indicator Variables**


Let's introduce 5 `Y` variables - Y1..Y5 (one for each possible value of b)


Only one of them can be true at any given time.



```
 Y1 + Y2 + Y3 + Y4 + Y5 = 1
 All Y's are binary {0,1}

```

Here's the trick. We introduce one linear constraint to ensure that the corresponding Y variable will take on value 1, only when b is that value. 



```
b - 1xY1 - 2xY2 - 3xY3 - 4xY4 - 5xY5 = 0

```

(For example, if b is 3, the constraint above will force Y3 to be 1.)


Now, we want `a` to take on the value u[b].



```
 a = u[1]xY1 + u[2]xY2 + u[3]xY3 + u[4]xY4 + u[5]xY5 

```

Since u[ 1] ...u[5] are constants known beforehand, the constraint above is also linear.


Here is [one reference](http://agecon2.tamu.edu/people/faculty/mccarl-bruce/mccspr/new15.pdf) on these kinds of IF-THEN conditions in Integer Programming. Many of these tricks involve the Big-M, though we didn't need it in this case.


Hope that helps you move forward.


"
"mode returns Exception: Must produce aggregated value
for this dataframe



```
   values  ii
0     3.0   4
1     0.0   1
2     3.0   8
3     2.0   5
4     2.0   1
5     3.0   5
6     2.0   4
7     1.0   8
8     0.0   5
9     1.0   1

```

This line returns ""Must ptoduce aggregated values



```
bii2=df.groupby(['ii'])['values'].agg(pd.Series.mode)

```

While this line works



```
bii3=df.groupby('ii')['values'].agg(lambda x: pd.Series.mode(x)[0])

```

Could you explain why is that?


","Problem is `mode` return sometimes 2 or more values, check solution with [`GroupBy.apply`](http://pandas.pydata.org/pandas-docs/stable/reference/api/pandas.core.groupby.GroupBy.apply.html):



```
bii2=df.groupby(['ii'])['values'].apply(pd.Series.mode)
print (bii2)
ii   
1   0    0.0
    1    1.0
    2    2.0
4   0    2.0
    1    3.0
5   0    0.0
    1    2.0
    2    3.0
8   0    1.0
    1    3.0
Name: values, dtype: float64

```

And pandas `agg` need scalar in output, so return error. So if select first value it working nice



```
bii3=df.groupby('ii')['values'].agg(lambda x: pd.Series.mode(x).iat[0])
print (bii3)
ii
1    0.0
4    2.0
5    0.0
8    1.0
Name: values, dtype: float64

```

"
"Parse c# class file to get properties and methods

> 
> **Possible Duplicate:**  
> 
> [Parser for C#](https://stackoverflow.com/questions/81406/parser-for-c-sharp) 
> 
> 
> 


Say if I had a simple class such as inside a textbox control in a winforms application:



```
public class Person
{
    public string Name { get; set; }
    public int Age { get; set; }

    public void DoSomething(string x)
    {
        return ""Hello "" + x;
    }
}

```

Is there an easy way I can parse the following items from it:


- Class Name
- Properties
- Any public methods


Any ideas/suggestions appreciated.


","You can use [Reflection](http://msdn.microsoft.com/en-us/library/ms173183%28v=vs.80%29.aspx) for that:



```
Type type = typeof(Person);
var properties = type.GetProperties(); // public properties
var methods = type.GetMethods(); // public methods
var name = type.Name;

```

**UPDATE** First step for you is compiling your class 



```
sring source = textbox.Text;

CompilerParameters parameters = new CompilerParameters() {
   GenerateExecutable = false, 
   GenerateInMemory = true 
};

var provider = new CSharpCodeProvider();       
CompilerResults results = provider.CompileAssemblyFromSource(parameters, source);

```

Next you should verify if your text is a valid c# code. Actually your code is not valid - method DoSomething marked as void, but it returns a string.



```
if (results.Errors.HasErrors)
{
    foreach(var error in results.Errors)
        MessageBox.Show(error.ToString());
    return;
}

```

So far so good. Now get types from your compiled in-memory assembly:



```
var assembly = results.CompiledAssembly;
var types = assembly.GetTypes();

```

In your case there will be only `Person` type. But anyway - now you can use [Reflection](http://msdn.microsoft.com/en-us/library/ms173183%28v=vs.80%29.aspx) to get properties, methods, etc from these types:



```
foreach(Type type in types)
{
    var name = type.Name;  
    var properties = type.GetProperties();    
}

```

"
"Sequelize doesn't create foreign keys as constraints
We're trying to create tables automatically inside PostgreSQL using Sequelize. Unfortunately, it doesn't create the foreign keys as constraints. Here is an example of one of my models:



```
module.exports = function(schema, DataTypes) {

    var definition = {
        id: {
            type: DataTypes.INTEGER,
            primaryKey: true,
            autoIncrement: true
        },
        deal_id: {
            type: DataTypes.INTEGER
        },
        image: DataTypes.STRING,
    };


    var image = schema.define('Image', definition, {
        tableName: 'images', // this will define the table's name
        timestamps: false, // this will deactivate the timestamp columns
        syncOnAssociation: true,

        classMethods: {
            getDefinition: function() {
                return definition;
            },
            associate: function(_models) {
                image.belongsTo(_models.Product, {
                    foreignKey: 'deal_id'
                });
            }
        }
    });

    return image;
};

```

Am I doing something wrong?


","First, it's not uncommon for ORMs to handle this kind of thing internally rather than by using foreign key constraints in the database.


Second, it's not uncommon for ORMs to require a *pair* of association statements to trigger all the internal handling you might expect.



```
var Task = this.sequelize.define('Task', { title: Sequelize.STRING })
  , User = this.sequelize.define('User', { username: Sequelize.STRING })

User.hasMany(Task)
Task.belongsTo(User)

```

Finally, Sequelize will actually write foreign key declarations into the database, but only if you also declare some kind of action (or inaction) with `onUpdate` or `onDelete`.



> 
> To add references constraints to the columns, you can pass the options
>  onUpdate and onDelete to the association calls. . . .
> 
> 
> 



```
User.hasMany(Task, { onDelete: 'SET NULL', onUpdate: 'CASCADE' })

CREATE TABLE IF NOT EXISTS `Task` (
  `id` INTEGER PRIMARY KEY, 
  `title` VARCHAR(255), 
  `user_id` INTEGER REFERENCES `User` (`id`) ON DELETE SET NULL ON UPDATE CASCADE
);

```

[Code example source](http://sequelizejs.com/docs/1.7.8/associations#foreign-keys)


"
"Webstorm external libraries explanation
I read this page that explains what external libraries in Webstorm are, but I still have more questions: <https://www.jetbrains.com/webstorm/help/configuring-javascript-libraries.html>


Are external libraries only for code completion and help in production?
Can they be used to link libraries for example in index.html? (Probably not, because they are not found in the project folder).


I guess my question is what are the external libraries for, other than what I said in the beginning?


","Javascript libraries configured in Settings/Languages&Frameworks/javaScript/Libraries (and shown as external libraries in the Project window) have absolutely nothing to do with references in your `<script>` tag. The former are used by IDE for code completion/navigation/error highlighting, the latter are used by browser in runtime. The browser knows nothing about javascript libraries configured in IDE, the IDE doesn't use `<script>` references in your HTML files.


Let me try to clear the things up:


What libraries are supposed to be used for: by default, completion works for all JavaScript files located under your project root. So, if you do already have the library .js files in your project structure, it's enough to get the completion. If they are missing, and you don't like to clatter your project with all these files, you can store them externally (outside of your project) and configure them as libraries to make them available to WebStorm. Note also that libraries are 'light-weight' as compared to .js files in your project - they are treated as read-only and have inspections turned off. Plus, you can assign documentation URLs to them, enabling external documentation for library code. So, even if you have your library files in your project, it might make sense to add them as libraries


So, to summarize: 


- library files placed next to your sources files in the project
structure are available to both WebStorm and browser, whether or not
they are added to javascript libraries in Settings
- online library referenced via CDN link in your HTML is available to
the browser at runtime, but can't be used for completion in the  

IDE
- a library files placed outside of the project and configured as
javascript libraries will be available to WebStorm for completion,  

but won't be loaded by a browser


"
"How to Send Json String to Controller in mvc4 and Deserialize json
I have the json object like below



```
Extension = {
""BookMarks"":
[{""Name"":""User1"",""Number"":""101""},
{""Name"":""User2"",""Number"":""102""},
{""Name"":""User3"",""Number"":""103""}]}

```

I want to send this json string to my controller Action method and Deserialize the data 


I want to pass the data to the partialview 



```
 public ActionResult ExtensionsDialog(var data)
        {
            return PartialView(data); 
        }

```

Any help
Thanks in advance..


","In your **View**: 



```
function SendData(){
        var dataToSend = JSON.stringify(data);

        $.ajax({
            type: ""POST"",
            url: '@Url.Action(""YourAction"", ""YourController"")',
            dataType: ""json"",
            data: dataToSend,
            contentType: ""application/json; charset=utf-8"",

        });
}

$(""#Updatebtn"").click(function () {

             sendData(); 
});

```

In you **Model**:



```
public class YourModel
{
  public String Name { get; set; }
  public int Number { get; set; }
}

```

In your **Controller**:



```
[HttpPost]
public ActionResult YourAction()
        {
            var resolveRequest = HttpContext.Request;
            List<YourModel> model = new List<YourModel>();
            resolveRequest.InputStream.Seek(0, SeekOrigin.Begin);
            string jsonString = new StreamReader(resolveRequest.InputStream).ReadToEnd();
            if (jsonString != null)
            {
                JavaScriptSerializer serializer = new JavaScriptSerializer();
                model = (List<YourModel>)serializer.Deserialize(jsonString, typeof(List<YourModel>);
            }
          //Your operations..
         }

```

Hope this helps.


"
"Cannot change system file
Need to be able to edit this file: /sys/class/net/eth0/queues/rx-0/rps\_cpus


The file permissions:



```
cd:/sys/class/net/eth0/queues/rx-0$ ls -l
total 0
-rw-r--r-- 1 root root 4096 2012-02-13 17:23 rps_cpus
-rw-r--r-- 1 root root 4096 2012-02-13 12:54 rps_flow_cnt
cd :/sys/class/net/eth0/queues/rx-0$**

```

Tried this:



```
$ sudo rm /sys/class/net/eth0/queues/rx-0/rps_cpus; sudo sh -c ""echo 1 > /sys/class/net/eth0/queues/rx-0/rps_cpus""
rm: cannot remove `/sys/class/net/eth0/queues/rx-0/rps_cpus': Operation not permitted

```

","The files under /sys/ aren't real files, they're representations of the kernel's internal state. You can tweak kernel parameters by writing stuff to them, but as they are not actual files, you cannot delete them.


If what you want is to put a ""1"" there, try this:



```
echo 1 | sudo tee /sys/class/net/eth0/queues/rx-0/rps_cpus

```

See the answer to [this question](https://askubuntu.com/questions/89112/cant-create-file-proc-sys-net-ipv3-ip-forward-to-ip-forward-before-arpspoof) for an explanation of the ""sudo tee"" trick and why it's useful in this case.


Read more about /sys/ here: <http://en.wikipedia.org/wiki/Sysfs>


"
"How to split a string by the third .(dot) delimiter
I did a few research, I can split the string by '.' but the only thing I want to replace is the third '.', and the version number is actually variable so how can I do the replace/split?



```
version: 1.8.0.110

```

but What I want is the output like this:



```
version: 1.8.0-110

```

","Used `sed` for example:



```
$ echo 'version: 1.8.0.110' | sed 's/\./-/3'
version: 1.8.0-110

```

Explanation:


`sed s/search/replace/x` searches for a string and replaces it with another string. `x` determines which occurence to replace - here the 3rd. Often `g` is used for `x` to mean all occurances.


Here we wish to replace the dot `.` but this is a special character in the regular expression `sed` expects in the `search` term. Therefore we backslashify the `.` to `\.` to specify a literal `.`.


Since we use special characters in the argument to `sed` (here, the backslash `\`) we need to put the whole argument in single quotes `''`. Many people always use quotes here so as not to run into problems when using characters that might be special to the shell (like space ).


"
"How to generate random failure time conditional to a lower bound
For my Monte Carlo simulation (job related) I needed to generate random numbers with conditional probability of related distribution. (***example***: given that an item has been working for 1200 hours, generating a proper random failure time related to its time-to-failure distribution.)


I searched the web and couldn't find a formula or process to generate this kind of dataset.


In my case I am only interested in 2 distributions:


1. exponential
2. weibull


**Question 1**: My opinion **for exponential** is ""*Since exponential distribution has memoryless property then generating random number with conditional probability is still same as generating random number without any condition.""*


Am I right?


**Question 2**: I couldn't find any source for weibull distribution. Can you please help me on how to generate weibull random number with conditional probability.


","If I understand correctly, you are correct concerning exponential distribution: the failure time $T$, conditional to $T > t\_1$, is $t\_1 + T'$ with $T' \sim \mathcal E(\lambda)$.


Weibull distribution has the property (which may be used as a definition) that
$$ Pr( T > t ) = e^{-\left( {t\over \lambda} \right)^k}. $$
Thus, the distribution of the failure time conditional to $T > t\_1$ verifies
$$ Pr( T > t\_1 + t | T > t\_1) = e^{ -{1\over \lambda^k}\left( (t\_1+t)^k - t\_1^k\right) }. $$
You can generate $T$ as $t\_1 + T'$ with $T'$ having the above distribution. This can be done by inverse transform sampling.


Concretely, if $U$ is uniform on $(0,1)$, let $T' = F^{-1}(U)$ with $F(t) = 1-e^{ -{1\over \lambda^k}\left( (t\_1+t)^k - t\_1^k\right) }$, that is
$$ T' = \left(t\_1^k - \lambda^k \log(1-U) \right)^{1\over k} - t\_1, $$
where $\log$ is the natural logarithm. You can simplify this by noting that $U$ and $1-U$ have the same distribution so $\log U$ will do it.


"
"Combine multiple JSON rows into one JSON object in PostgreSQL
I want to combine the following JSON from multiple rows into one single JSON object as a row.



```
{""Salary"": """"}
{""what is your name?"": """"}
{""what is your lastname"": """"}

```

**Expected output**



```
{
  ""Salary"": """",
  ""what is your name?"": """",
  ""what is your lastname"": """"
}

```

","With only built-in functions, you need to expand the rows into key/value pairs and aggregate that back into a single JSON value:



```
select jsonb_object_agg(t.k, t.v)
from the_table, jsonb_each(ob) as t(k,v);

```

If your column is of type `json` rather than `jsonb` you need to cast it:



```
select jsonb_object_agg(t.k, t.v)
from the_table, jsonb_each(ob::jsonb) as t(k,v);

```

A slightly more elegant solution is to define a new aggregate that does that:



```
CREATE AGGREGATE jsonb_combine(jsonb) 
(
    SFUNC = jsonb_concat(jsonb, jsonb),
    STYPE = jsonb
);

```

Then you can aggregate the values directly:



```
select jsonb_combine(ob)
from the_table;

```

(Again you need to cast your column if it's `json` rather than `jsonb`)


[Online example](https://dbfiddle.uk/?rdbms=postgres_9.5&fiddle=aad32d85449c7b48d61c2faee2e4adc0)


"
"Using Moq, how do I set up a method call with an input parameter as an object with expected property values?

```
 var storageManager = new Mock<IStorageManager>(); 
 storageManager.Setup(e => e.Add(It.IsAny<UserMetaData>()));

```

The Add() method expects a UserMetaData object which has a FirstName property.


I'd like to make sure that an object of type UserMetaData with the FirstName of ""FirstName1"" has been passed.


","You can use `Verify`.


**Examples:**


Verify that `Add` was never called with an `UserMetaData` with `FirstName`!= `""FirstName1""`:



```
storageManager.Verify(e => e.Add(It.Is<UserMetaData>(d => d.FirstName!=""FirstName1"")), Times.Never());

```

Verify that `Add` was called at least once with an `UserMetaData` with `FirstName`== `""FirstName1""`:



```
storageManager.Verify(e => e.Add(It.Is<UserMetaData>(d => d.FirstName==""FirstName1"")), Times.AtLeastOnce());

```

Verify that `Add` was called exactly once with `FirstName` == `""Firstname1""` and `LastName` == `""LastName2""`:



```
storageManager.Setup(e => e.Add(It.Is<UserMetaData>(data => data.FirstName == ""FirstName1""
                                                         && data.LastName  == ""LastName2"")));

...

storageManager.VerifyAll();

```

"
"Value-based color scheme for year-by-year histograms in R
I'm utilizing ggplot2 in R in order to build histograms of some test results. Right now I have a year-by-year anaylsis of the table Glucose\_Results2, which has one Value column and one Year column (and a LOT of rows).



```
Gval2 <- read.csv('Glucose_Results2.csv', sep = ',', header=T)
Gdist = qplot(Value, binwidth=10, data=Gval2, geom=""histogram"", 
  main=""Distribution of Glucose Levels"",
  xlab=""Glucose Levels in mg/dl"",
  ylab=""Frequency"")
Gval2$Year = factor(Gval2$Year)
Gdist + facet_wrap( ~ Year)

```

This gives me plots for 2005 through 2012. I'm interested in altering the color scheme of these plots to indicate values that fall within the reference range for each year (see below). So for 2005 I want the bins for values between 70 and 110 to be, say, red, whereas the bins outside that range can stay black.



```
>  Gref2
  REFERENCE_LOW REFERENCE_HIGH REFERENCE_YEAR
1            70            110           2005
2            70            110           2006
3            70            110           2007
4            70            110           2008
5            70            110           2009
6            70            100           2010
7            70            100           2011
8            70            100           2012

```

Having only a few weeks' experience in R, I really have no idea how to do this. 


Ideally this would include labels for reference high and reference low, or something to the effect of red = within reference range, but I'm not overly concerned about that. What I do want is a solution that reads the above table and selects the appropriate reference ranges based on the year of the histogram (though if it comes down to it, I can enter the numbers manually).


Thank you for reading this and for any responses!


","Using ggplot, to control the color, the easiest way is to get a variable into the data frame which indicates which group it is a part of (the within or outside reference range group). This can be accomplished by a series of steps. First, for each case, match up the appropriate limits based on the year.



```
Gboth <- merge(Gval2, Gref2, by.x=""Year"", by.y=""REFERENCE_YEAR"")

```

This matches, for each row of `Gval2`, the appropriate row of `Gref2` where appropriate is defined by matching the `Year` column in `Gval2` with the `REFERENCE_YEAR` column in `Gref2`. Then an indicator column can be made using Value and the reference limits. I've made it a labelled factor to make things easier later, but that is not strictly necessary.



```
Gboth$within <- factor(Gboth$REFERENCE_LOW < Gboth$Value & 
                       Gboth$Value <= Gboth$REFERENCE_HIGH,
                       levels = c(FALSE, TRUE),
                       labels = c(""Outside reference range"",
                                  ""Within reference range""))

```

Now I can use this variable to specify the color (really, the fill, since for a histogram colour is the color of the lines around the bars and fill is what fills in the bars). Also, I've translated from `qplot` to full `ggplot` syntax. `qplot` is helpful for transitioning from `plot` commands until you get more familiar with the syntax, but it is not as expressive. If you are learning from scratch, you might as well skip it.



```
ggplot(Gboth, aes(x=Value)) +
  geom_histogram(aes(fill=within), binwidth=10) +
  facet_wrap( ~ Year) +
  scale_x_continuous(""Glucose Levels in mg/dl"") +
  scale_y_continuous(""Frequency"") +
  scale_fill_manual("""", breaks=c(""Outside reference range"", 
                                 ""Within reference range""),
                    values = c(""black"", ""red"")) +
  opts(title = ""Distribution of Glucose Levels"")

```

For some made up values for `Gval2`, I get a graph that looks like


![enter image description here](https://i.stack.imgur.com/ymhO9.png)


"
"MipMapping problems in OpenGL
I'm loading 3D objects (obj or 3ds or collada files) into my openGL application. The 3 environment is quite large (a few hundred metres in all axis'). 
My problem is that smaller 3D objects (i.e. in the order of ~< 1-2m ) don't appear to be depth-tested properly. Depending on the zoom of the camera, I can sometimes see the back face of the object (I have been using a simple cube for testing) or other faces becoming visible/invisible/torn. Please see the attached images for a better explanation.
I am led to believe the problem is due to mipmapping being enabled. I would either like to disable mipmapping (can someone suggest a simple, fast way to do this) or set the resolution to be greater for the mipmapped objects. Or am I barking up the wrong tree completely?


![alt text](https://i.stack.imgur.com/Hu6dJ.png)
![alt text](https://i.stack.imgur.com/xHy44.png)
![alt text](https://i.stack.imgur.com/T9ccw.png) 


Thanks


Chris


","That's the result of insufficient z-buffer precission, which is an issue in games that have huge worlds but (relatively) small objects. The immediate solution would be to try using a 24 bit z-buffer instead of a 16 bit one. Another way to tackle this would be to render the game world it two steps, first the big distant objects, then clearing the zbuffer and then drawing the closer objects.


This specific problem is called z-fighting by the way, here's a great resource on this issue: <http://www.codermind.com/articles/Depth-buffer-tutorial.html>


The take-away is the last paragraph of the article above:



> 
> the true issue is that you can't draw
>  both objects that are very far and
>  objects that are very near with the
>  same depth buffer equations. If you
>  want to draw very far objects then you
>  need to sacrifice your near view by
>  pushing it further. To avoid clipping
>  artifacts you can make your collision
>  envelope large enough so that your
>  clip plane will never intercept an
>  existing object within your frustum.
>  Or you can make object gradually
>  disappear with transparency as they
>  come near your clip plane.
> 
> 
> If you want to keep near objects and
>  at the same time draw mountains (or
>  planets) in the far distance, then you
>  can cut your rendering in parts. First
>  drawing your far objects, then
>  clearing the depth buffer and
>  rendering the near objects with a
>  different z buffer.
> 
> 
> 


"
"Body attributes (onkeydown, onkeyup...) not rendered after upgrading to Primefaces 5.1, from 4.0
I have set up a simple maven project with the following dependencies:



```
<dependencies>
   <dependency>
      <groupId>org.glassfish</groupId>
      <artifactId>javax.faces</artifactId>
      <version>2.2.2</version>
   </dependency>
   <dependency>
      <groupId>javax.servlet</groupId>
      <artifactId>javax.servlet-api</artifactId>
      <version>3.0.1</version>
      <scope>provided</scope>
   </dependency>
   <dependency>
      <groupId>javax.el</groupId>
      <artifactId>javax.el-api</artifactId>
      <version>3.0.1-b04</version>
      <scope>provided</scope>
   </dependency>
   <dependency>
      <groupId>javax.servlet.jsp.jstl</groupId>
      <artifactId>jstl-api</artifactId>
      <version>1.2</version>
   </dependency>

   <dependency>
      <groupId>org.hibernate</groupId>
      <artifactId>hibernate-entitymanager</artifactId>
      <version>4.3.6.Final</version>
   </dependency>

   <dependency>
      <groupId>org.hibernate</groupId>
      <artifactId>hibernate-c3p0</artifactId>
      <version>4.3.6.Final</version>
   </dependency>

   <dependency>
      <groupId>org.hibernate</groupId>
      <artifactId>hibernate-validator</artifactId>
      <version>5.1.2.Final</version>
   </dependency>

   <dependency>
      <groupId>org.primefaces</groupId>
      <artifactId>primefaces</artifactId>
      <version>5.1</version>
   </dependency>

</dependencies>

```

And this page:



```
<?xml version=""1.0"" encoding=""UTF-8""?>
<!DOCTYPE html PUBLIC ""-//W3C//DTD XHTML 1.0 Transitional//EN""
        ""http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"">
<html xmlns=""http://www.w3.org/1999/xhtml""
      xmlns:h=""http://xmlns.jcp.org/jsf/html""
      xmlns:ui=""http://xmlns.jcp.org/jsf/facelets""
      xmlns:f=""http://xmlns.jcp.org/jsf/core"">
<h:body onkeydown=""alert('You pressed some key!')"">
    <h:outputLabel value=""Hello, young fellows!""/>
</h:body>
</html>

```

This is the generated html, notice the body tag without methods:



```
<?xml version=""1.0"" encoding=""UTF-8""?>
<!DOCTYPE html>
<html xmlns=""http://www.w3.org/1999/xhtml""><body><label>Hello, young fellows!</label></body>
</html>

```

I tried to rollback the primefaces version to 4.0 by changing the dependency version in pom.xml, like this:



```
<dependency>
   <groupId>org.primefaces</groupId>
   <artifactId>primefaces</artifactId>
   <version>4.0</version>
</dependency>

```

And then it works as expected:



```
<?xml version=""1.0"" encoding=""UTF-8""?>
<!DOCTYPE html>
<html xmlns=""http://www.w3.org/1999/xhtml""><body onkeydown=""alert('You pressed some key!')""><label>Hello, young fellows!</label></body>
</html>

```

What is the problem? Is it a bug on PF 5.1 or something else?


","Starting from 5.0 the `h:body` has a [renderer](https://github.com/primefaces/primefaces/blob/5.0.x/src/main/resources-maven-jsf/standard-renderers.xml#L10) in PrimeFaces, [`BodyRenderer`](https://github.com/primefaces/primefaces/blob/5.1.x/src/main/java/org/primefaces/renderkit/BodyRenderer.java).


In the `encodeBegin` of that renderer an array of [attributes](https://github.com/primefaces/primefaces/blob/5.1.x/src/main/java/org/primefaces/renderkit/BodyRenderer.java#L44) is being passed, [`HTML.BODY_ATTRS`](https://github.com/primefaces/primefaces/blob/5.1.x/src/main/java/org/primefaces/util/HTML.java#L144).



```
public static final String[] BODY_ATTRS = {
    ""dir"",
    ""lang"",
    ""style"",
    ""onload"",
    ""onunload""
};

```

And as you can see that this array for some reason doesn't cover all the actual attributes of the `h:body` tag, thus some of the attributes are being ignored.


In order to avoid this problem you can simply extend that renderer into a one which accepts all the actual attributes.



```
public class CustomBodyRenderer extends BodyRenderer{

   //our array with all the attributes of h:body tag
   public static final String[] BODY_ATTRS = {
       ""dir"",
       ""lang"",
       ""onclick"",
       ""ondblclick"",
       ""onkeydown"",
       ""onkeypress"",
       ""onkeyup"",
       ""onmousedown"",
       ""onmousemove"",
       ""onmouseout"",
       ""onmouseover"",
       ""onmouseup"",
       ""style"",
       ""title"",
       ""onload"",
       ""onunload""
   };

   @Override
   public void encodeBegin(FacesContext context, UIComponent component) throws IOException {
       ResponseWriter writer = context.getResponseWriter();
       String clientId = component.getClientId(context);
       writer.startElement(""body"", component);

       if (shouldWriteId(component)) {
           writer.writeAttribute(""id"", clientId, ""id"");
       }

       String styleClass = (String) component.getAttributes().get(""styleClass"");
       if (styleClass != null && styleClass.length() != 0) {
           writer.writeAttribute(""class"", styleClass, ""styleClass"");
       }
       //the only changed line from the original renderer
       renderPassThruAttributes(context, component, BODY_ATTRS);
   }

}

```

Then register it in the **faces-config**



```
<render-kit>
    <renderer>
        <component-family>javax.faces.Output</component-family>
        <renderer-type>javax.faces.Body</renderer-type>
        <renderer-class>com.hatemalimam.CustomBodyRenderer</renderer-class>
    </renderer>
</render-kit>

```

This way you would get the expected outcome in a ""minimal"" changes.


I have submitted an [issue](https://code.google.com/p/primefaces/issues/detail?id=7621) on the tracker.


**Update:**
*This has been fixed in PF 5.2.*


"
"Simple Injector conditional injection
Lets say I have two Controllers: ControllerA and ControllerB. Both of those controllers accept as parameter IFooInterface. Now I have 2 implementation of IFooInterface, FooA and FooB. I want to inject FooA in ControllerA and FooB in ControllerB. This was easily achieved in Ninject, but I'm moving to Simple Injector due to better performance. So how can I do this in Simple Injector? Please note that ControllerA and ControllerB resides in different assemblies and are loaded dynamically.


Thanks


","The SimpleInjector documentation calls this [context-based injection](https://simpleinjector.readthedocs.org/en/latest/advanced.html#context-based-injection). As of version 3, you would use `RegisterConditional`. As of version 2.8, this feature isn't implemented in SimpleInjector, however the documentation contains [a working code sample implementing this feature](http://simpleinjector.readthedocs.org/en/2.8/ContextDependentExtensions.html) as extensions to the `Container` class.


Using those extensions methods you would do something like this:



```
Type fooAType = Assembly.LoadFrom(@""path\to\fooA.dll"").GetType(""FooA"");
Type fooBType = Assembly.LoadFrom(@""path\to\fooB.dll"").GetType(""FooB"");
container.RegisterWithContext<IFooInterface>(context => {
    if (context.ImplementationType.Name == ""ControllerA"") 
    {
        return container.GetInstance(fooAType);
    } 
    else if (context.ImplementationType.Name == ""ControllerB"") 
    {
        return container.GetInstance(fooBType)
    } 
    else 
    {
        return null;
    }
});

```

"
"Different approach of using IN clause in MySql
Today I have posted an answer with a query like this



```
SELECT * FROM table_name where column_name IN (val1,val2,...)

```

Some another user has posted the answer a query like this



```
SELECT * FROM table_name where val1 IN (column_name)

```

As you can see here the position of the column\_name and values are interchanged.


## From Mysql Docs



> 
> expr IN (value,...)
> 
> 
> Returns 1 if expr is equal to any of the values in the IN list, else returns 0. If all values are constants, they are evaluated according to the type of expr and sorted. The search for the item then is done using a binary search. This means IN is very quick if the IN value list consists entirely of constants. 
> 
> 
> 



```
mysql> SELECT 2 IN (0,3,5,7);
-> 0

mysql> SELECT 'wefwf' IN ('wee','wefwf','weg');
-> 1

```

As it clearly says that the above one(my query) is correct. but both the above queries produce the same output.


Also why not the other approach in listed in **[Mysql Documentation](http://dev.mysql.com/doc/refman/5.0/en/comparison-operators.html#function_in)**?



> 
> This question serves as a canonical information source regarding the use of IN. Its purpose is to have detailed, high quality answers detailing the proper use on IN in queries.
> 
> 
> 


","you raised a question that is connected with my answer [**here**](https://stackoverflow.com/a/16117259/491243).


In a simple explanation using this statements below,



```
SELECT * FROM TableName WHERE column1 IN (1, 2, 3, 4)
-- versus
SELECT * FROM TableName WHERE 1 IN (column1, column2, column3, column4)

```

The first statement involves only **ONE COLUMN** that is being *compared to multiple values*.



```
SELECT  *
FROM   TableName
WHERE  column1 = 1 OR
       column1 = 2 OR
       column1 = 3 OR
       column1 = 4

```

while the second statement is **A VALUE** that is *compared to multiple columns*.



```
SELECT  *
FROM   TableName
WHERE  column1 = 1 OR
       column2 = 1 OR
       column3 = 1 OR
       column4 = 1

```

*which is a bit different from one another.*




---


**UPDATE 1**


Here's the third form of `IN` clause:


- [Multi-Column IN clause – Unexpected MySQL Issue](http://www.mysqlperformanceblog.com/2008/04/04/multi-column-in-clause-unexpected-mysql-issue/)


"
"Chrome Extension ContextMenu for PDF
I'm developing a extension for the chrome browser and i want to add a specified contextmenu for pdf documents. I also add to specified contextmenus for the type ""page"" and ""image"".
If i set the type to ""all"" then there is a contextmenu, but not specified for pdf documents.


Is it possible to add a specified contextmenu for pdf documents or should i use a the type ""all"" an it make switch case in the clickEventHandler?!


See more at:
<http://developer.chrome.com/extensions/contextMenus.html>


These are the ""file"" types:



```
contexts ( optional array of enum of ""all"", ""page"", ""frame"", ""selection"", ""link"", ""editable"", ""image"", ""video"", ""audio"", or ""launcher"" )

```

","I'm guessing that you want to add a context menu only when a PDF is shown in a tab, right? Just asking because I thought at first that you wanted to add the context menu on links to PDF files, which is indeed possible\*. (as you probably know)


I couldn't find a way to do this directly, however one alternative could be to listen to [`chrome.tabs.onActivated`](http://developer.chrome.com/extensions/tabs.html#event-onActivated) and add or remove your context menu based on if the current URL matches a PDF file. One drawback is that it means asking for the `tabs` permission which might looks scary to users. (""This extension can access your tabs and browsing activity"" or something like that)


\*for the curious, you do it like this:



```
chrome.contextMenus.create({
    title: ""Hello world"",
    contexts: [""link""],
    targetUrlPatterns: [""*://*/*.pdf""]
});

```

(you would add the other options that interest you of course)


"
"Case-sensitive entity recognition
I have keywords that are all stored in lower case, e.g. ""discount nike shoes"", that I am trying to perform entity extraction on. The issue I've run into is that spaCy seems to be case sensitive when it comes to NER. Mind you , I don't think that this is spaCy specific. 


When I run...



```
doc = nlp(u""i love nike shoes from the uk"")

for ent in doc.ents:
    print(ent.text, ent.start_char, ent.end_char, ent.label_)

```

... nothing is returned.


When I run...



```
doc = nlp(u""i love Nike shoes from the Uk"")

for ent in doc.ents:
    print(ent.text, ent.start_char, ent.end_char, ent.label_)

```

I get the following results...



```
Nike 7 11 ORG
Uk 25 27 GPE

```

Should I just title case everything? Is there another workaround that I could use?


","spaCy's [pre-trained statistical models](https://spacy.io/models/en) were trained on a large corpus of general news and web text. This means that the entity recognizer has likely only seen very few all-lowercase examples, because that's much less common in those types of texts. In English, capitalisation is also a strong indicator for a named entitiy (unlike German, where all nouns are typically capitalised), so the model probably tends to pay more attention to that.


If you're working with text that doesn't have proper capitalisation, you probably want to fine-tune the model to be less sensitive here. See the docs on [updating the named entity recognizer](https://spacy.io/usage/training#ner) for more details and code examples.


Producing the training examples will hopefully not be very difficult, because you can use existing annotations and datasets, or create one using the pre-trained model, and then lowercase everything. For example, you could take text with proper capitalisation, run the model over it and extract all entitiy spans in the text. Next, you lowercase all the texts, and update the model with the new data. Make sure to also mix in text with proper capitalisation, because you don't want the model to learn something like ""Everything is lowercase now! Capitalisation doesn't exist anymore!"".


Btw, if you have entities that can be defined using a list or set of rules, you might also want to check out the [`EntityRuler` component](https://spacy.io/usage/rule-based-matching#entityruler). It can be combined with the statistical entity recognizer and will let you pass in a dictionary of exact matches or abstract token patterns that can be case-insensitive. For instance, `[{""lower"": ""nike""}]` would match one token whose lowercase form is ""nike"" – so ""NIKE"", ""Nike"", ""nike"", ""NiKe"" etc.


"
"Are there any Ruby ORMs which use cursors or smart fetch?
I'm looking for a Ruby ORM to replace ActiveRecord. I've been looking at Sequel and DataMapper. They look pretty good however none of them seems to do the basic: not loading everything in memory when you don't need it.


I mean I've tried the following (or equivalent) on ActiveRecord and Sequel on table with lots of rows:



```
 posts.each { |p| puts p }

```

Both of them go crazy on memory. They seem to load everything in memory rather than fetching stuff when needed. I used the `find_in_batches` in ActiveRecord, but it's not an acceptable solution:


1. ActiveRecord is not an acceptable solution because we had too many problems with it.
2. Why should my code be aware of a paging mechanism? I'm happy to configure somewhere the size of the page but that's it. With `find_in_batches` you need to do something like:


post.find\_in\_batches { |batch| batch.each { |p| puts p } }


But that should be transparent.


So is there somewhere a reliable Ruby ORM which does the fetch properly?




---


Update:


As Sergio mentioned, in Rails 3 you can use `find_each` which exactly what I want. However as ActiveRecord is not an option, except if someone can really convince me to use it, the questions are:


1. Which ORMs support the equivalent of find\_each?
2. How to do it?
3. Why do we need a `find_each`, while `find` should do it, shouldn't it?


","Sequel's `Dataset#each` does yield individual rows at a time, but most database drivers will load the entire result in memory first.


If you are using Sequel's Postgres adapter, you can choose to use real cursors:



```
posts.use_cursor.each{|p| puts p}

```

This fetches 1000 rows at a time by default, but you can use an option to specify the amount of rows to grab per cursor fetch:



```
posts.use_cursor(:rows_per_fetch=>100).each{|p| puts p}

```

If you aren't using Sequel's Postgres adapter, you can use Sequel's pagination extension:



```
Sequel.extension :pagination
posts.order(:id).each_page(1000){|ds| ds.each{|p| puts p}}

```

However, like ActiveRecord's `find_in_batches`/`find_each`, this does separate queries, so you need to be careful if there are concurrent modifications to the dataset you are retrieving.


The reason this isn't the default in Sequel is probably the same reason it isn't the default in ActiveRecord, which is that it isn't a good default in the general case. Only queries with large result sets really need to worry about it, and most queries don't return large result sets.


At least with the Postgres adapter cursor support, it's fairly easy to make it the default for your model:



```
Post.dataset = Post.dataset.use_cursor

```

For the pagination extension, you can't really do that, but you can wrap it in a method that makes it mostly transparent.


"
"Libsass with Susy not working + Gulp
I've no doubt others have got this working... I just haven't quite worked out how right now.


**Gulp**



```
var gulp = require('gulp');
var sass = require('gulp-sass');
var handleErrors = require('../util/handleErrors');
var sourcemaps = require('gulp-sourcemaps');
var minifyCSS = require('gulp-minify-css');

gulp.task('sass', function () {
    return gulp.src('./frontend/sass/style.scss')
        .pipe(sourcemaps.init())
        .pipe(sass({
            'require':'susy'
        }))
        //.pipe(minifyCSS())
        .pipe(sourcemaps.write('./app/www/css'))
        .on('error', handleErrors)
        .pipe(gulp.dest('./app/www/css'))
});

```

**Sass**



```
@import ""susy"";

```

**Gulp Running**



```
[23:58:08] Starting 'sass'...

stream.js:94
      throw er; // Unhandled stream error in pipe.
            ^
Error: file to import not found or unreadable: susy
Current dir: C:/var/www/rnli.hutber.com/frontend/sass/

```

I have installed susy@2.2.2 in the root along side the `gulp` file with the following: `gem install susy`


**Current working setup with gulp-ruby-sass**


I do however have it working, which confirms that susy is working via the gem installed with the following code:



```
var gulp = require('gulp');
var sass = require('gulp-ruby-sass');
var handleErrors = require('../util/handleErrors');
var minifyCSS = require('gulp-minify-css');

gulp.task('sass', function () {
    return gulp.src('./frontend/sass/style.scss')
        .pipe(sass({
            'sourcemapPath':'./app/www/css',
            'require':'susy'
        }))
        //.pipe(minifyCSS())
        .on('error', handleErrors)
        .pipe(gulp.dest('./'))
});

```

**NOTE** The above code will not working using `gulp-ruby-sass@1.0.0alpha' It will only work as I can tell with`v0.7.1`


**package.json**



```
  ""devDependencies"": {
    ""gulp"": ""~3.8.10"",
    ""gulp-sass"": ""~1.3.2"",
    ""gulp-ruby-sass"": ""~0.7.1"",
    ""gulp-sourcemaps"": ""~1.3.0"",
    ""susy"":""~2.2.1""
  }

```

How do I get `susy` working correctly and able to compile into `css`?


","I've managed to get **susy** working with **libsass** (gulp-sass)


**gulp-sass** needs to know where to find files specified with an `@import` directive. You can try set the [includesPath](https://github.com/sass/node-sass#includepaths) option in gulp-sass to point to your local copy of susy. 


As I've installed susy with bower, I have something like this:



```
var sass = require('gulp-sass');

gulp.task('sass', function () {
  return gulp.src(config.src)
    .pipe(sass({
      includePaths: [
        'bower_components/susy/sass'
      ]
    }))
    ...
});

```

"
"Syntax question from `gin-gonic` documentation
I have been developing in Node/Python for years, and over the holiday I have been trying to expand my knowledge of Go. I have a pet project that I've been working on to learn it. 


As I have been reading the [gin-gonic documentation](https://github.com/gin-gonic/gin#grouping-routes), I found syntax that I can't wrap my head around.



```
func main() {
    router := gin.Default()

    // Simple group: v1
    v1 := router.Group(""/v1"")
    {
        v1.POST(""/login"", loginEndpoint)
        v1.POST(""/submit"", submitEndpoint)
        v1.POST(""/read"", readEndpoint)
    }

    // Simple group: v2
    v2 := router.Group(""/v2"")
    {
        v2.POST(""/login"", loginEndpoint)
        v2.POST(""/submit"", submitEndpoint)
        v2.POST(""/read"", readEndpoint)
    }

    router.Run("":8080"")
}

```

Basically, it looks like the `router.Group()` method takes in a struct, but I am not quite sure what the mechanics are here... It's not wrapped in parentheses. How is this object being passed in/handled by the `Group` method?


","
> 
> [The Go Programming Language Specification](https://golang.org/ref/spec)
> 
> 
> [Blocks](https://golang.org/ref/spec#Blocks)
> 
> 
> A block is a possibly empty sequence of declarations and statements
>  within matching brace brackets.
> 
> 
> [Declarations and scope](https://golang.org/ref/spec#Declarations_and_scope)
> 
> 
> Go is lexically scoped using blocks.
> 
> 
> 




---


For example, the `v1.POST` method call statements form a block:



```
// Simple group: v1
v1 := router.Group(""/v1"")
{
    v1.POST(""/login"", loginEndpoint)
    v1.POST(""/submit"", submitEndpoint)
    v1.POST(""/read"", readEndpoint)
}

```

This is an unusual use of blocks.


"
"SProxy in purescript?
What's the use of ***Sproxy*** in purescript?


In Pursuit, it's written as



```
data SProxy (sym :: Symbol)
--| A value-level proxy for a type-level symbol.

```

and what is meant by Symbol in purescipt?


","*First, please note that PureScript now has polykinds since version 0.14 and most functions now use `Proxy` instead of `SProxy`. `Proxy` is basically a generalisation of `SProxy`.*


## About Symbols and Strings


PureScript knows value level strings (known as `String`) and type level strings (known as `Symbol`).


- A `String` can have any string value at runtime. The compiler does not track the value of the string.
- A `Symbol` is different, it can only have one value (but remember, it is on the type level). The compiler keeps track of this string. This allows the compiler to type check certain expressions.


## Symbols in Practice


The most prominent use of Symbols is in records. The difference between a Record and a String-Map is that the compiler knows about the keys at compile time and can typecheck lookups.


Now, sometimes we need to bridge the gap between these two worlds: The type level and the value level world. Maybe you know that PureScript records are implemented as JavaScript objects in the official compiler. This means we need to somehow receive a string value from our symbol. The magical function `reflectSymbol` allows us to turn a symbol into a string. But a symbol is on the type level. This means we can only write a symbol where we can write types (so for example in type definition after `::`). This is where the Proxy hack comes in. The `SProxy` is a simple value that ""stores"" the type by applying it.


For example the `get` function from `purescript-records` allows us to get a value at a property from a record.



```
get :: forall proxy r r' l a. IsSymbol l => Cons l a r' r => proxy l -> Record r -> a

```

If we apply the first paramerter we get:



```
get (Proxy :: Proxy ""x"") :: forall r a. { x :: a | r } -> a

```

Now you could argue that you can get the same function by simply writing:



```
_.x :: forall r a. { x :: a | r } -> a

```

It has exactly the same type. This leads to one last question:


## But why?


Well, there are certain meta programming szenarios, where you don't programm for a specific symbol, but rather for any symbol. Imagine you want to write a JSON serialiser for any record. You might want to ""iterate"" over every property of the record, `get` the value, turn the value itself into JSON and then concatinate the key value pair with all the other keys and values.


An example for such an implementation can be found [here](https://github.com/justinwoo/purescript-simple-json/blob/v8.0.0/src/Simple/JSON.purs#L322-L358)


This is maybe not the most technical explanation of it all, but this is how I understand it.


"
"Shall we use poll() or select()?
I'm fully aware of the major differences between `poll()` and `select()`:


- `select()` only supports a fixed amount of file descriptors
- `select()` is supposedly supported on more systems
- `poll()` allows slightly more fine-grained control of event types
- `poll()` implementations may differ in certain details


However, they both accomplish the same task in roughly the same way. So:


Shall we use `poll()` or `select()`?


  

**EDIT:** I might add that I'm not interested in `epoll()` since portability is of concern to me. Furthermore, `libev(ent)` is not an option either, since I'm asking this question because I'm writing my own replacement library for `libev(ent)`.


","All remotely modern systems have `poll`, and it's a greatly superior interface to `select`/`pselect` in almost all ways:


- `poll` allows more fine-grained detection of status than `select`.
- `poll` does not have limits on the max file descriptor you can use (and more importantly, does not have critical vulnerabilities when you fail to check for file descriptors past the `FD_SETSIZE` limit).


The only disadvantages I can think of to using `poll` are that:


- unlike `pselect`, `poll` cannot atomically unmask/mask signals, so you can't use it for waiting for a set of events that includes both file descriptor activity and signals unless you resort to the self-pipe trick.
- `poll` only has millisecond resolution for the wait timeout, rather than microsecond (`select`) or nanosecond (`pselect`).


Certainly portability of `poll` is not a consideration anymore. Any system old enough to lack `poll` is full of so many vulnerabilities it should not be connected to a network.


In summary, unless you have very special needs (tiny timeout intervals, nasty signal interactions, scaling to millions of persistent connections, etc.) I would simply use `poll` and be done with it. As others have mentioned, `libevent` is also an option, but it's not clean/safe code (its use of `select` actually invokes dangerous UB trying to workaround the limitations of `select`!) and I find code that uses `libevent` is generally a lot more unnecessarily complicated than code that simply uses `poll` directly.


"
"Non-WSDL php call
I'm new to php and looking to integrate a website with some rental software. They use non-WSDL mode services and have supplied this piece of code but I'm a bit confused about the ""this-is-the-action-uri"". My guess is that's a method I need to call.



```
    $client = new SoapClient(NULL, array(
    'location' => '21.ip2.ip3.ip4/r2ws_v5/servlet/messagerouter',
    'uri' => 'urn:this-is-the-action-uri',
    'exceptions' => 1,
    );

```

I can call this and get a response.



```
http://21.ip2.ip3.ip4:8080/r2ws_v5/jsp/UBS_GetAvailability.jsp

```

The reposnse starts like this.



```
<SOAP-ENV:Envelope xmlns:SOAP-ENV=""http://schemas.xmlsoap.org/soap/envelope/"" xmlns:xsi=""http://www.w3.org/2001/XMLSchema-instance"" xmlns:xsd=""http://www.w3.org/2001/XMLSchema"">
<SOAP-ENV:Body>
<ns1:getAvailabilityByItemResponse xmlns:ns1=""UBS/R2"" SOAP-ENV:encodingStyle=""http://schemas.xmlsoap.org/soap/encoding/"">
<response>
<product>
<productID xsi:type=""xsd:string"">TESTUB</productID>
<level xsi:type=""xsd:int"">1</level>
<description xsi:type=""xsd:string"">
<![CDATA[ testub ]]>
</description>
</level>

```

Can you give me any idea of what the 'uri' parameter should be in this case or how it should be formatted?


","Hi just create a PHP file and put below code into your PHP file.



```
$xmlRequest = ""<SOAP:Envelope xmlns:SOAP=\""http://schemas.xmlsoap.org/soap/envelope/\"" xmlns:xsi=\""http://www.w3.org/2001/XMLSchema-instance\"" xmlns:xsd=\""http://www.w3.org/2001/XMLSchema\"">"" .
    ""<SOAP:Body>"" .
    ""<getAvailabilityByItem xmlns=\""UBS/R2WebServices/AvailabilityService\"">"" .
        ""<request>"" .
            ""<detail xsi:type=\""xsd:string\"">0</detail>"" .
            ""<products>"".
                ""<product>"" .
                    ""<type xsi:type=\""xsd:int\"">0</type>"" .
                    ""<productID xsi:type=\""xsd:string\""><![CDATA[DE014SML]]></productID>"" .
                    ""<startDate xsi:type=\""xsd:date\"">12/07/2014</startDate>"" .
                    ""<startTime xsi:type=\""xsd:timeInstant\"">8:00 AM</startTime>"" .
                    ""<endDate xsi:type=\""xsd:date\"">12/08/2014</endDate>"" .
                    ""<endTime xsi:type=\""xsd:timeInstant\"">12:00 PM</endTime>"" .
                ""</product>"".
            ""</products>"" .
        ""</request>"" .
    ""</getAvailabilityByItem>"" .
    ""</SOAP:Body>"" .
    ""</SOAP:Envelope>"" ;

$location_URL = 'http://21.ip2.ip3.ip4/r2ws_v5/servlet/messagerouter';

$client = new SoapClient(null, array(
'location' => $location_URL,
'uri' => """",
'trace' => 1,
    ));
$order_return = $client->__doRequest($xmlRequest, $location_URL, '' , 1);

```

Hope you found the solution ;)


"
"UnsatisfiedLinkError at JNI invocation
I've already found a lot of things and questions about UnsatisfiedLinkError in combination with JNI but none of them could help me out though.
I'm trying to hide a window using Java, so I need 3 WinAPI functions.


I have the following Java source:



```
package hide.window;
public class HideWindow {

    public native static int hideWindow(char[] windowTitle);

    static {
        System.loadLibrary(""hideWindow"");
    }

    public static void main(String[] args) {
        System.out.println(""loadLibrary was successful"");
        System.out.println(hideWindow(""Session Window - Windows Internet Explorer"".toCharArray()));
    }
}

```

Then I have my hide\_window\_HideWindow.h generated by javah.
There's one thing I had to change, because it didn't find the jni.h file in the library path so I just put it into my project and included it the following way:



```
#include ""jni.h""

```

**May this cause my Error??**
The whole file:



```
/* DO NOT EDIT THIS FILE - it is machine generated */
#include ""jni.h""
/* Header for class hide_window_HideWindow */

#ifndef _Included_hide_window_HideWindow
#define _Included_hide_window_HideWindow
#ifdef __cplusplus
extern ""C"" {
#endif
/*
 * Class:     hide_window_HideWindow
 * Method:    hideWindow
 * Signature: ([C)I
 */
JNIEXPORT jint JNICALL Java_hide_window_HideWindow_hideWindow
  (JNIEnv *, jobject, jcharArray);

#ifdef __cplusplus
}
#endif
#endif

```

The code I compiled to my DLL looks as follows:



```
#include <windows.h>
#include ""hide_window_HideWindow.h""
using namespace std;

JNIEXPORT jint JNICALL Java_hide_window_HideWindow_hideWindow(char windowTitle[20])
{
    HWND windowHandle = FindWindowA(NULL, windowTitle);
    if (windowHandle){
        if (IsWindowVisible(windowHandle)){
            ShowWindow(windowHandle, 0);
            return 0;
        }
        else{
            ShowWindow(windowHandle, 1);
            return 1;
        }
    }
    return -1;
}

```

There's no difference between a call from eclipse and the following call via cmd (Path is correct):



```
java -Djava.library.path=.. hide.window.HideWindow

```

The resulting output looks like that:



> 
> 
> ```
> loadLibrary was successful
> Exception in thread ""main"" java.lang.UnsatisfiedLinkError: hide.window.HideWindow.hideWindow([C)I
>   at hide.window.HideWindow.hideWindow(Native Method)
>   at hide.window.HideWindow.main(HideWindow.java:12)
> 
> ```
> 
> 


That indicates to me that the Java program is able to load the DLL but it fails at the function call. 
I use Eclipse as Java and Codeblocks as C++ IDE, and I have already added the following to the mingw build options.



```
-Wl,--kill-at

```

I hope this question is not too stupid and thank you in advance,
Tom


","Look at the prototype:



```
JNIEXPORT jint JNICALL Java_hide_window_HideWindow_hideWindow
    (JNIEnv *, jobject, jcharArray);

```

and your implementation:



```
JNIEXPORT jint JNICALL Java_hide_window_HideWindow_hideWindow
    (char windowTitle[20])

```

Of course there's an `UnsatisfiedLinkError` because you didn't implement the correct function.


So you would need something like:



```
JNIEXPORT jint JNICALL Java_hide_window_HideWindow_hideWindow
    (JNIEnv* env, jobject thiz, jcharArray windowTitle)

```

And to get the chars out of the array have a look at `GetPrimitiveArrayCritial` or `GetCharArrayRegion`.


The generated header also look suspicious to me, because you defined the Java method to be `static` but the generated prototype wants to pass a `jobject` ... static methods need `jclass` (because there is no instance object for static method calls):



```
JNIEXPORT jint JNICALL Java_hide_window_HideWindow_hideWindow
    (JNIEnv* env, jclass clazz, jcharArray windowTitle)

```

Another thing: Java's `char` (`jchar`) is an `unsigned int16`, C/C++ `char` is signed or unsigned `int8`. So it might be better to pass a `String` (`jstring`) and get the contents as an UTF8 string using `GetStringUTFChars`.


"
"StartFlipping for ViewFlipper after 60seconds from last touch
My app contains a `ViewFlipper` with some images. when app starts, the `ViewFlipper` `startflipping()`. When user touch the screen `ViewFlipper stopflipping()`. I must do that after 60 seconds from last touch, ViewFlipper to start again flipping. My class implements `onTouchListener` and I have this method `onTouch`:



```
public boolean onTouch(View arg0, MotionEvent arg1) {


        switch (arg1.getAction()) {
        case MotionEvent.ACTION_DOWN: {

            downXValue = arg1.getX();
            break;
        }

        case MotionEvent.ACTION_UP: {

            currentX = arg1.getX();


            if (downXValue < currentX) {
                // Set the animation
                vf.stopFlipping();
                vf.setOutAnimation(AnimationUtils.loadAnimation(this,
                        R.anim.push_right_out));
                vf.setInAnimation(AnimationUtils.loadAnimation(this,
                        R.anim.push_right_in));
                // Flip!
                vf.showPrevious();
            }


            if (downXValue > currentX) {
                // Set the animation
                vf.stopFlipping();
                vf.setOutAnimation(AnimationUtils.loadAnimation(this,
                        R.anim.push_left_out));
                vf.setInAnimation(AnimationUtils.loadAnimation(this,
                        R.anim.push_left_in));
                // Flip!
                vf.showNext();
            }

            if (downXValue == currentX) {
                final int idImage = arg0.getId();

                vf.stopFlipping();
                System.out.println(""id"" + idImage);
                System.out.println(""last touch ""+getTimeOfLastEvent());

            }
            break;
        }
        }

        // if you return false, these actions will not be recorded
        return true;
    }

```

and I found this method, for finding time of the last touch :



```
static long timeLastEvent=0;
public long getTimeOfLastEvent() {

        long duration = System.currentTimeMillis() - timeLastEvent;
        timeLastEvent = System.currentTimeMillis();
        return duration;
    }

```

My question is : where should I call the `getTimeOfLastEvent()` ? If I put it on `onTouch()` I will never catch the moment when getTimeOfLastEvent==60000, right?


","What you should do is create a [`Handler`](http://developer.android.com/reference/android/os/Handler.html) (should be an instance variable of your `Activity` and should be initialized during `onCreate`):



```
Handler myHandler = new Handler();

```

Also you will need a `Runnable` that can start the flipping again (also needs to be declared within your `Activity`):



```
private Runnable flipController = new Runnable() {
  @Override
  public void run() {
    vf.startFlipping();
  }
};

```

Then in your `onClick` you just post the `Runnable` on the `Handler` but delayed by 60 seconds:



```
myHandler.postDelayed( flipController, 60000 );

```

Posting it delayed means: ""Run this code in 60 seconds"".


"
"Oracle xmltype column indexing based on xml attribute
I am using structured storage type for xmltype column. I am trying to create an index on the attribute but somehow Oracle is not taking the index and doing a full table scan. I have checked the documentation at <http://docs.oracle.com/cd/B28359_01/appdev.111/b28369/xdb_indexing.htm#CHDCJIJB> but it doesnot give an example on creating index on attribute.
Can someone please give me an example of creating the xml index on attribute?


Thanks
Kevin


","you'd need to show some code here. 


here's an example of creating an `xmlindex` on an attribute (and its descendant nodes):



```
SQL> create table xmltest ( a xmltype);

Table created.

SQL> insert into xmltest values ('<root>
  2   <a test=""foo"" test2=""asd"">
  3     <b>hi</b>
  4   </a>
  5   <a test=""foo2"" test2=""asd2"">hi</a>
  6  </root>');

1 row created.

SQL> commit;

Commit complete.

SQL> create index test_xmlindex on xmltest (a)
  2  indextype is XDB.XMLIndex
  3  parameters ('paths (include (/root/a/@test))');

Index created.

SQL> set pagesize 100 lines 200 heading off
SQL> explain plan for select * from xmltest where xmlexists('/root/a[@test=""foo""]' passing a);

Explained.

SQL> select * from table(dbms_xplan.display());

Plan hash value: 2925043098

-------------------------------------------------------------------------------------------------------------------
| Id  | Operation                        | Name                           | Rows  | Bytes | Cost (%CPU)| Time     |
-------------------------------------------------------------------------------------------------------------------
|   0 | SELECT STATEMENT                 |                                |     1 |  2026 |     6  (17)| 00:00:01 |
|   1 |  NESTED LOOPS                    |                                |     1 |  2026 |     6  (17)| 00:00:01 |
|   2 |   VIEW                           | VW_SQ_1                        |     1 |    12 |     4   (0)| 00:00:01 |
|   3 |    HASH UNIQUE                   |                                |     1 |  5046 |            |          |
|   4 |     NESTED LOOPS                 |                                |       |       |            |          |
|   5 |      NESTED LOOPS                |                                |     1 |  5046 |     4   (0)| 00:00:01 |
|*  6 |       TABLE ACCESS BY INDEX ROWID| SYS82641_TEST_XMLIN_PATH_TABLE |     1 |  3524 |     2   (0)| 00:00:01 |
|*  7 |        INDEX RANGE SCAN          | SYS82641_TEST_XMLIN_VALUE_IX   |     1 |       |     1   (0)| 00:00:01 |
|*  8 |       INDEX RANGE SCAN           | SYS82641_TEST_XMLIN_PIKEY_IX   |     1 |       |     1   (0)| 00:00:01 |
|*  9 |      TABLE ACCESS BY INDEX ROWID | SYS82641_TEST_XMLIN_PATH_TABLE |     1 |  1522 |     2   (0)| 00:00:01 |
|  10 |   TABLE ACCESS BY USER ROWID     | XMLTEST                        |     1 |  2014 |     1   (0)| 00:00:01 |
-------------------------------------------------------------------------------------------------------------------

```

by indexing `paths (include (/root/a/@test))` it would contain the child `<b>` node in the index but not the `test2` attribute. ommitting `/@test` would have the index pick up the `test2` attribute as well.


"
"I'm trying to build site navigation using django-mptt
I've used django-mptt to model the sections and sub-sections etc of the headings I want to store my articles under. Some sections and sections will have children others won't.
 eg:



```
[Home]
 [News]
    [Politics][Crime][Policing]  etc 
 then [News] > [Politics]
                 [UKIP][Labour] etc

```

I'm sure you get the idea.
I've created the Section models



```
class Section(MPTTModel):
    name = models.CharField(max_length=50)
    parent = TreeForeignKey('self',null=True, blank=True, related_name='sub_sections')
    tree = SectionTreeManager()

```

and by passing Section.objects.all() into the template :- 



```
<ul class=""headernav"" >

    {% recursetree sections %}

        <li>
            <a href=""/section/{{node.name}}"" >{{ node.name }}</a>

             {% if node.is_leaf_node %}
                <ul  class=""headernav""> 
                    {{ children }}
                </ul>
            {% endif %}
           </li>
    {% endrecursetree %}
</ul>

```

I can display all nodes. What I want to do is , based on the current url ie /section/News
 get all siblings on current level all ancestors on previous level and all related descendants on the level below. so if I selected News I'd get



```
[home][**News**][World][Sports][Culture] <- Level 0

[Politics][Crime][Education] <- Children on Level 1

```

How can I do this? Any help would be much appreciated. I'm learning Django so if it's obvious I apologise in advance.


","[django-page-cms](https://github.com/batiste/django-page-cms) has solved this type of problem using `django-mptt`, so I suggest having a look at the way it's done there.


Specifically, the [template to render the tree menu](https://github.com/batiste/django-page-cms/blob/master/pages/templates/pages/dynamic_tree_menu.html) is recursively included via a [custom template tag](https://docs.djangoproject.com/en/dev/howto/custom-template-tags/) called [`pages_dynamic_tree_menu`](https://github.com/batiste/django-page-cms/blob/master/pages/templatetags/pages_tags.py#L225).


This sounds complicated (custom ta-whaaa!), so I'll break it down a bit more.


We can consider a [`Page` in `django-page-cms`](https://github.com/batiste/django-page-cms/blob/master/pages/models.py#L27) to be roughly the same as your `Section` class.


In the base template which is shared between 99% of your pages, `django-page-cms` [gets you to insert the following snippet](http://packages.python.org/django-page-cms/navigation-template-tags.html#pages-dynamic-tree-menu):



```
<ul>
{% for page in pages_navigation %}
    {% pages_dynamic_tree_menu page %}
{% endfor %}
</ul>

```

`pages_navigation` [is set by](https://github.com/batiste/django-page-cms/blob/master/pages/context_processors.py#L16) a [context processor](https://docs.djangoproject.com/en/dev/ref/settings/#std:setting-TEMPLATE_CONTEXT_PROCESSORS) (a way to add items to a [template context](https://docs.djangoproject.com/en/dev/ref/templates/api/#playing-with-context-objects) for all views in a django project) and contains the top level pages (i.e. those without parents).


The `pages_dynamic_tree_menu` tag is defined and registered as follows:



```
def pages_dynamic_tree_menu(context, page, url='/'):
    """"""
    Render a ""dynamic"" tree menu, with all nodes expanded which are either
    ancestors or the current page itself.

    Override ``pages/dynamic_tree_menu.html`` if you want to change the
    design.

    :param page: the current page
    :param url: not used anymore
    """"""
    lang = context.get('lang', pages_settings.PAGE_DEFAULT_LANGUAGE)
    page = get_page_from_string_or_id(page, lang)
    children = None
    if page and 'current_page' in context:
        current_page = context['current_page']
        # if this node is expanded, we also have to render its children
        # a node is expanded if it is the current node or one of its ancestors
        if(page.tree_id == current_page.tree_id and
            page.lft <= current_page.lft and
            page.rght >= current_page.rght):
            children = page.get_children_for_frontend()
    context.update({'children': children, 'page': page})
    return context
    pages_dynamic_tree_menu = register.inclusion_tag(
        'pages/dynamic_tree_menu.html',
        takes_context=True
    )(pages_dynamic_tree_menu)

```

The whole point of it is to set the `page` and the `children` variables. 


The magic is in the `if` statement, which ensures the children are only displayed if the page is the current one (the `current_page` template variable is set by the `django-page-cms` view which renders a `Page`), or if the page is an ancestor. The `lft` & `rght` attributes are provided by `django-mptt`, and two of the properties of this MPTT are that *the left value of a parent tree node will be less than its child node's left value* and *the right value of a parent node will be more than the right value of its child node*, which you can verify visually by looking at the image below from the article [Trees in SQL](http://www.ibase.ru/devinfo/DBMSTrees/sqltrees.html):


![picture of MPTT](https://i.stack.imgur.com/P4q4w.gif)


The `get_children_for_frontend()` method of `Page` calls `MPTTModel.getChildren()` with some filtering to hide unpublished `Page` objects, so nothing special there.


The tag then does the [tag equivalent](https://docs.djangoproject.com/en/dev/howto/custom-template-tags/#inclusion-tags) of [`{% include %}`](https://docs.djangoproject.com/en/dev/ref/templates/builtins/#include) with the [`pages/dynamic_tree_menu.html`](https://github.com/batiste/django-page-cms/blob/master/pages/templates/pages/dynamic_tree_menu.html) template, which in turn calls the `pages_dynamic_tree_menu` tag again, causing the whole thing to recurse:



```
{% load pages_tags cache %}
{% if page.calculated_status %}
<li {% ifequal page current_page %}class=""selected""{% endifequal %}>
<a href=""{% show_absolute_url page %}"">{% show_content page ""title"" %}</a>
{% if children %}
    <ul>{% for child in children %}{% pages_dynamic_tree_menu child url %}{% endfor %}</ul>
{% endif %}
</li>
{% endif %}

```

So if you define a similar setup (provide top level sections and the current section to your template, define an ""inclusive"" template tag and matching template for the tag) and apply some styling, then you should be able to use this approach to accomplish your goal.


"
"Haskell: runInteractiveCommand: how to ignore the output given so far?
i run an external interactive program in haskell via ""runInteractiveCommand"" and want read the output of the program until it shows the programs prompt. So far i have



```
checkConsistency cal nets = do
    (sparqIn, sparqOut, sparqErr, sparqId) <- runInteractiveCommand ""sparq -i""
    mapM_ (flip hSetBinaryMode False) [sparqIn, sparqOut]
    hSetBuffering sparqIn LineBuffering
    hSetBuffering sparqOut NoBuffering
    hPutStrLn sparqIn (""load-calculus "" ++ cal)
-- Here i need to wait for output and ignore it until the prompt is shown!
    sparqAnswers <- mapM (checkConsistencyWithSparq sparqIn sparqOut) nets
    return sparqAnswers

```

I think i should use ""hReady"" and ""hGetChar"" in a while-loop but i am not sure how to do it exactly. Or maybe there is a better way?


Kind regards,
Annaluise


","What it seems you want to do is [Parse](http://en.wikipedia.org/wiki/Parsing) the output of your interactive program. While this can get very hairy (depending on the format of the output, your semantic requirements etc.), often you can get away with some very simple constructs.


Like the recursive `parseUntilPrompt` in the following example:



```
import Control.Applicative                                   
import System.IO                                             
import System.Process    

main = do                                                    
    (inn, out, err, idd) <- runInteractiveCommand ""mongo""    
    mapM_ (flip hSetBinaryMode False) [inn, out]             
    hSetBuffering inn LineBuffering                          
    hSetBuffering out NoBuffering                            
    hPutStrLn inn ""help""                                    
    parsedIntro <- parseUntilPrompt out                      
    mapM_ (putStrLn . \x -> ""PARSED:: "" ++  x) parsedIntro   

parseUntilPrompt :: Handle -> IO [String]                    
parseUntilPrompt out = do                                    
  latest <- hGetLine out                                     
  if latest == """"                                            
    then return []                                           
    else (:) <$> return latest <*> parseUntilPrompt out

```

which happily parses the output of my `help` command:



```
*Interactive> main
PARSED:: MongoDB shell version: 1.8.0                                     
PARSED:: connecting to: test                                              
PARSED:: > help                                                           
PARSED::        db.help()                    help on db methods           
PARSED::        db.mycoll.help()             help on collection methods   
PARSED::        rs.help()                    help on replica set methods  
PARSED::        help connect                 connecting to a db help      
PARSED::        help admin                   administrative help          
PARSED::        help misc                    misc things to know          
PARSED::        help mr                      mapreduce help               
*Interactive>                                            

```

"
"first\_or\_create vs find\_or\_create\_by
`first_or_create` seems much less documented, so I wondered if it was just because the two methods are synonymous. 


","Basically:



```
Foo.where(attributes).first_or_create

```

Is the same as:



```
Foo.find_or_create_by(attributes)

```



---


`#first_or_create` is sometimes misunderstood as people expect it to search by the attributes given, but that is not the case. Aka



```
Foo.first_or_create(attributes)

```

Will not search for a `foo` that satisfies the `attributes`. It will take the first `foo` if any are present. It's useful if the conditions for finding are a subset of those used for creation. Aka



```
Foo.where(something: value).first_or_create(attributes)

```

Will find the first `foo` where `something: value`. If none is present, it will use `attributes` to create it.


"
"How can I retrieve all MC in a Stage in AS3?
How can I retrieve all MC in a Stage?


I want to call addEventListener for all MovieClip present in my stage and if it's possible in a **selected frame number of my scenario**
Something like that pseudo code



```
for(i=0; nbOfChild ; i++)
   if(stage.childAt(i) is MC and isInTheFrameWithLabel('foo') )
     stage.childAt(i).addEventListerner(MyStuff)

```

","Pretty much exactly what you have..



```
for(var i:int = 0; i<numChildren; i++)
{
    var e:DisplayObject = getChildAt(i);
    if(e is MovieClip)
    {
        // do stuff with e
        trace(e);
    }
}

```

You could also be a tricky trickster and do something like this:



```
/**
 * Iterate through MovieClips within a container and parse them through handler
 * @param container The container to iterate through and find MovieClips in
 * @param handler A function that accepts MovieClip as its only parameter
 */
function each(container:DisplayObjectContainer, handler:Function):void
{
    for(var i:int = 0; i<container.numChildren; i++)
    {
        var e:DisplayObject = container.getChildAt(i);
        if(e is MovieClip) handler(e);
    }
}

// Example
each(this, function(mc:MovieClip):void
{
    trace(mc);
    mc.x += 10;
});

```

"
"jquery validate with html5 pattern
How do i make jQuery validate/support an HTML 5 pattern? It would be great if it did without specifying the pattern on each field.


I know I could do this (below), but is there better way?



```
$(""form"").validate({
  rules: {
    ""password"": {
      pattern: /[A-Za-z0-9\w]{4,20}/,
    }
  }
});

```

See Fiddle <http://jsfiddle.net/2ma8ec6j/>


","**Quote**:



> 
> *""How do i make jQuery validate/support an HTML 5 pattern?""*
> 
> 
> 


It already does. See below



> 
> *""It would be great if it did without specifying the pattern on each field.""*
> 
> 
> 


How will it know which pattern goes to which field if you don't specify this someplace?




---



> 
> *""...is there better way?""*
> 
> 
> 


You can declare *some* rules via inline HTML attributes. However, you can declare *all* rules via the `rules` option within `.validate()` or the `.rules()` method.


I don't know if this is ""better"" as that is purely a matter of opinion, but as long as you include the [`additional-methods.js`](http://ajax.aspnetcdn.com/ajax/jquery.validate/1.13.0/additional-methods.js) OR the [`pattern.js`](https://github.com/jzaefferer/jquery-validation/blob/master/src/additional/pattern.js) file, the jQuery Validate plugin will pickup and use the HTML5 `pattern` attribute.



```
<input type=""text"" name=""somename"" pattern=""[A-Za-z0-9\w]{4,20}"" />

```

**DEMO: <http://jsfiddle.net/fLxgz9dn/>**


See: <https://github.com/jzaefferer/jquery-validation/issues/785>


"
"Why does getLocalName() return null?
I'm loading some XML string like this:



```
Document doc = getDocumentBuilder().parse(new InputSource(new StringReader(xml)));

```

Later, I extract a node from this `Document`:



```
XPath xpath = getXPathFactory().newXPath();
XPathExpression expr = xpath.compile(expressionXPATH);
NodeList nodeList = (NodeList)expr.evaluate(doc, XPathConstants.NODESET);

Node node = nodeList.item(0);

```

Now I want to get the local name of this node but I get `null`.



```
node.getLocalName(); // return null

```

With the debugger, I saw that my node has the following type: [DOCUMENT\_POSITION\_DISCONNECTED](https://docs.oracle.com/javase/7/docs/api/org/w3c/dom/Node.html#DOCUMENT_POSITION_DISCONNECTED).


[The Javadoc](https://docs.oracle.com/javase/7/docs/api/org/w3c/dom/Node.html#getLocalName()) states that `getLocalName()` returns `null` for this type of node.


- Why node is of type DOCUMENT\_POSITION\_DISCONNECTED and not ELEMENT\_NODE?
- How to ""convert"" the type of the node?


","As the documentation <https://docs.oracle.com/javase/7/docs/api/org/w3c/dom/Node.html#getLocalName()> states:



> 
> for nodes created with a DOM Level 1 method, [...] this is always null
> 
> 
> 


so make sure you use a namespace aware `DocumentBuilderFactory` with `setNamespaceAware(true)`, that way the DOM is supporting the namespace aware DOM Level 2/3 and will have a non-null value for `getLocalName()`.


A simple test program



```
    String xml = ""<root/>"";

    DocumentBuilderFactory db = DocumentBuilderFactory.newInstance();

    Document dom1 = db.newDocumentBuilder().parse(new InputSource(new StringReader(xml)));

    System.out.println(dom1.getDocumentElement().getLocalName() == null);

    db.setNamespaceAware(true);

    Document dom2 = db.newDocumentBuilder().parse(new InputSource(new StringReader(xml)));

    System.out.println(dom2.getDocumentElement().getLocalName() == null);

```

outputs



```
true
false

```

so (at least) the local name problem you have is caused by using a DOM Level 1, not namespace aware document (builder factory).


"
"Is there a link format for physical street address?
Mobile devices have started to do link detection fairly well. They now detect links, email address, phone numbers and physical street addresses.


The problem is that link detection is not 100% correct. So to be more explicit, for most of these, we can use a link.



```
<a href=""tel:+1-408-555-5555"">1-408-555-5555</a>

```

Is there an equivalent link format for a physical street address?


","There doesn't seem to be one defined, [according to the IANA](http://www.iana.org/assignments/uri-schemes.html). Here is a [discussion](http://lists.w3.org/Archives/Public/uri/2001Mar/0012.html) about creating one, but it doesn't seem it ever went anywhere. The closest I can find is [geo](http://www.rfc-editor.org/rfc/rfc5870.txt), which is just latitude and longitude, not a postal address. I see from the update that you want to display a map, so what you want is to [geocode](http://en.wikipedia.org/wiki/Geocoding) the address to get a geo URI. You could use [Google's API](https://developers.google.com/maps/documentation/geocoding/) but I don't know if there is any need for the geo URI at that point. 


"
"ScrollArea in QT Issue?
I have one form where i am applying ScrollArea but it is not getting applied properly ?
I want to place the four buttons in Scrollable Area and when buttons gets increased a scroll bar should appear there to scorll down vertically.


Currently when i am trying to add more buttons the buttons overlap with the combobox and does not grow downward.


Here is my code :



```
 QPushButton *b1 = new QPushButton(strsearch);
           QPushButton *b2 = new QPushButton(strsearch);
           QPushButton *b3 = new QPushButton(strsearch);
           QPushButton *b4 = new QPushButton(strsearch);

 b1->setStyleSheet(

               ""background-image: url(:/user.gif);""
               ""background-repeat:no-repeat;""
               ""background-attachment:fixed;""
               ""background-position:left top;""
               ""border-style: outset;""
               ""background-color : black;""
               ""border-width: 2px;""
               ""border-radius: 10px;""
               ""border-color: black;""
               ""font: bold 16px;""
               ""color: black;""
               ""min-width: 10em;""
               ""min-height: 0.75em;""
               ""margin: 0 1px 0 1px;""
               ""color:rgb(255,255,255);""
               ""padding: 6px;""
               );



b2->setIcon(QIcon("":/user.gif""));
           b2->setIconSize(QSize(160, 26));
           b3->setIcon(QIcon("":/user.gif""));
           b3->setIconSize(QSize(160, 26));
           b4->setIcon(QIcon("":/user.gif""));
           b4->setIconSize(QSize(160, 26));

 QGridLayout *layout = new QGridLayout;
 layout->addWidget(b1, 1, 0);
           layout->addWidget(b2, 2, 0);
            layout->addWidget(b3, 3, 0);
            layout->addWidget(b4, 4, 0);
layout->addWidget(scrollArea);
  layout->setAlignment(Qt::AlignBottom);
  setLayout(layout);

```

","
```
//Create the buttons
QPushButton *b1 = new QPushButton(""Button 1"");
QPushButton *b2 = new QPushButton(""Button 2"");
QPushButton *b3 = new QPushButton(""Button 3"");
QPushButton *b4 = new QPushButton(""Button 4"");

//Add the buttons to a vertical layout (faster than grid layout)
QVBoxLayout *scrollLayout = new QVBoxLayout;
scrollLayout->addWidget(b1);
scrollLayout->addWidget(b2);
scrollLayout->addWidget(b3);
scrollLayout->addWidget(b4);

//Create a viewport widget that contains the layout with buttons
QWidget *viewport = new QWidget;
viewport->setLayout(scrollLayout);

//Add the viewport to the scroll area
QScrollArea *scrollArea = new QScrollArea;
scrollArea->setWidget(viewport);

//Add the scroll area to your main window's layout
mainLayout->addWidget(scrollArea);

```

"
"Cannot run migrations when postgres extension added in rails 5 application
I want to use pg\_trgm extension of Postgres for searching using query



```
SELECT * FROM USERS WHERE location like '%new%' ORDER BY location DESC;

```

Since Postgres already don't have pg\_trgm I need to execute a command to install it. So my migration to do so is



```
class Addtrigramindexlocationtousers < ActiveRecord::Migration[5.1]
  def change
    reversible do |direction|
        direction.up {
            execute %{
                CREATE EXTENSION IF NOT EXISTS pg_trgm;
                CREATE INDEX index_users_trigram_on_location ON users USING gin (location gin_trgm_ops);
            }
        }
        direction.down {
            execute %{
                DROP INDEX index_users_trigram_on_location;
            }           
        }
    end
  end
end

```

So when I run this migration it is giving me this error:



```
    ActiveRecord::StatementInvalid: PG::InsufficientPrivilege: ERROR:  permission denied to create extension ""pg_trgm""
HINT:  Must be superuser to create this extension.
: 
                DROP INDEX index_users_on_location;
                CREATE EXTENSION IF NOT EXISTS pg_trgm;
                CREATE INDEX index_users_trigram_on_location ON users USING gin (location gin_trgm_ops);

```

This works if I manually enter to database and execute the command but I need to run it from migration. 


","When enabling extensions its a good idea to create a separate migration as it makes it much easier to troubleshoot:



```
class EnableTrigramIndexLocationExtension < ActiveRecord::Migration[5.1]
  def change
    enable_extension ""pg_trgm""
  end
end

```


> 
> Loading an extension requires the same privileges that would be
>  required to create its component objects. For most extensions this
>  means superuser or database owner privileges are needed. The user who
>  runs CREATE EXTENSION becomes the owner of the extension for purposes
>  of later privilege checks, as well as the owner of any objects created
>  by the extension's script.
> 
> 
> 


The simplest way to solve this is by using `ALTER ROLE myapp SUPERUSER;` which is not a very secure solution but works for development. For a production server you should instead use [PostgreSQL Extension Whitelisting](https://github.com/dimitri/pgextwlist).


When generating migrations make sure to use either snakecase or camelcase to make the name readable `rails g migration AddTrigramIndexLocationToUsers` or `rails g migration add_trigram_index_location_to_users`.



```
class AddTrigramIndexLocationToUsers < ActiveRecord::Migration[5.1]
  def up
    execute %{ 
      CREATE INDEX index_users_trigram_on_location ON users USING gin (location gin_trgm_ops); 
    }
  end

  def down
    remove_index :users, :index_users_trigram_on_location
  end
end

```

"
"Mono on Mac: DllNotFoundException despite SQLite.Interop.dll being in dllmap
I have a C# application that uses SQLite and works fine on Windows.


The same Visual Studio project compiles fine in Xamarin Studio, but when running I get:



```
DllNotFoundException: SQLite.Interop.dll

```

Despite:


- `libsqlite3.0.dylib` is in `/usr/lib` and also in the same folder as the executable and other DLLs
- `.` is part of the `$DYLD_LIBRARY_PATH`
- The executable and all SQLite-using DLLs have a matching `<the_exe_or_dll_including_filename_extension>.config` file containing:


`<configuration>
 <dllmap dll=""sqlite"" target=""libsqlite.0.dylib"" os=""osx""/>
 <dllmap dll=""sqlite3"" target=""libsqlite3.0.dylib"" os=""osx""/>
</configuration>`


I also tried adding `<dllmap dll=""SQLite.Interop.dll"" target=""libsqlite3.0.dylib"" os=""osx""/>`, not better.


What is the problem?


","You can easily find where mono is looking for that native library by setting the MONO\_LOG\_LEVEL to debug and MONO\_LOG\_MASK filtering to only DLL related messages.



```
export MONO_LOG_LEVEL=debug
export MONO_LOG_MASK=dll
mono yourprogram.exe

```

or as a one liner so you do not have to unset env vars:



```
MONO_LOG_LEVEL=debug MONO_LOG_MASK=dll mono yourprogram.exe

```

Mono and the OS-X dynamic link editor ('man dyld' for details) does not require DYLD\_LIBRARY\_PATH to be set to the current directory ('.'). Note: Linux does require LD\_LIBRARY\_PATH to include the current directory, if that is your intention.


- Move those dll map files out of the way to remove them from the equation.
- Unset DYLD\_LIBRARY\_PATH
- cd in the directory that contains your CIL based exe, dlls and native dylib(s)
- MONO\_LOG\_LEVEL=debug MONO\_LOG\_MASK=dll mono yourprogram.exe


Using the native dll/shared library trace output you can track which library is not being found (or one of its dependancies) or if it is the wrong ARCH for your mono version.


If you are still having problems, we would need to know which SQLite library you are using the options that you are using to compile it (or the arch version if getting it via a Nuget). A posting your dll trace output would quickly solve things also.


Notes:


I am assuming you are using the System.Data.SQLite library and are compiling the the options ""/p:UseInteropDll=true /p:UseSqliteStandard=false"".


Mono includes a SQLite in it's default install, it is 32-bit on OS-X:



```
file /Library/Frameworks/Mono.framework/Versions/4.0.2/lib/libsqlite3.dylib
/Library/Frameworks/Mono.framework/Versions/4.0.2/lib/libsqlite3.dylib: Mach-O dynamically linked shared library i386

```

Assuming you are using the OS-X package installer from Mono, thus are getting the 32-bit version of Mono and thus need 32-bit versions of the native libraries.



```
>>file `which mono`
/usr/bin/mono: Mach-O executable i386

```

The /usr/lib/libsqlite3.0.dylib is a multi ARCH fat binary, so that library is not a problem, but your debug output might show another one that is a problem,



```
>>file /usr/lib/libsqlite3.0.dylib
libsqlite3.0.dylib: Mach-O universal binary with 3 architectures
libsqlite3.0.dylib (for architecture x86_64):   Mach-O 64-bit dynamically linked shared library x86_64
libsqlite3.0.dylib (for architecture i386): Mach-O dynamically linked shared library i386
libsqlite3.0.dylib (for architecture x86_64h):  Mach-O 64-bit dynamically linked shared library x86_64

```

"
"pre-selecting value in Combo box in extjs
I want to implement the following combobox in ExtJS. The question is, how to make the third option selected by default?



```
<select name=""meter_payment_option"" onChange=""smart_meter(this.value)""> 
    <option value=""1"">All Up-Front</option> 
    <option value=""2"">Reduced Up-Front</option> 
    <option value=""3"" selected=""selected"">No Up-Front</option> 
</select> 

```

What I currently have is:



```
var meter_payment_option_values = new Ext.data.SimpleStore({
    fields: ['id', 'value'],
    data: [
        ['1', 'All Up-Front'],
        ['2', 'Reduced Up-Front'],
        ['3', 'No Up-Front']]
});

var smart_meter_term = new Ext.form.ComboBox({
    name: 'smart_meter_term',
    editable: false,
    typeAhead: false,
    allowblank: false,
    triggerAction: 'all',
    hiddenName: 'my_dropdown',
    fieldLabel: 'SmartM.T',
    store: meter_payment_option_values,
    displayField: 'value',
    valueField: 'id',
    mode: 'local'
});

```

How do I make the 3rd option (No Up-Front) selected by default?


","You need to set the `value` config option to the id of the default value, e.g.:



```
var smart_meter_term = new Ext.form.ComboBox({
                  name:'smart_meter_term' ,
                  editable: false,
                  typeAhead: false, 
                  allowblank:false , 
                  triggerAction: 'all',
                  hiddenName: 'my_dropdown',
                  fieldLabel:'SmartM.T',
                  store:meter_payment_option_values,
                  displayField:'value',
                  valueField:'id',
                  mode:'local',
                  // default value is 3 (No Up-Front)
                  value: 3
                });

```

<http://www.sencha.com/deploy/dev/docs/index.html?class=Ext.form.ComboBox>


"
"Is it possible to remove an app's results from Modern UI search on the start screen?
How can I remove all search items (for example emails) from a particular pre-installed Modern UI app when searching by typing on the start screen?


","Open the Windows-8 Charm Bar by moving the mouse pointer to the top-right/lower-right corner of the screen and click on **Settings**. Alternatively you can directly press the `Windows`+`I` hotkey to open the `Settings` sidebar. Here click on the `Change PC Settings` link at the bottom to open Windows-8 Modern UI settings. 


![enter image description here](https://i.stack.imgur.com/zShGF.png)


Navigate to **Search** settings in `PC settings` and click on the button **Delete history** to delete all the searches you have performed.


To disable the search history tracker, toggle the `Let Windows save my searches as future search` suggestions settings from on to off. 


When you disable the history tracker, the Delete history button will be disabled as all your history will be deleted automatically.


If you want to disable Windows-8 search on any of the app, you can switch if off as well and omit the app from showing results in the start screen.


"
"Ioc/DI - Why do I have to reference all layers/assemblies in application's entry point?
(Related to this question, [EF4: Why does proxy creation have to be enabled when lazy loading is enabled?](https://stackoverflow.com/questions/9500642/ef4-mvc3-with-windsor-castle-ioc-di-navigation-properties-not-loading)).


I'm new to DI, so bear with me. I understand that the container is in charge of instantiating all of my registered types but in order to do so it requires a reference to all of the DLLs in my solution and their references.


If I weren't using a DI container, I wouldn't have to reference the EntityFramework library in my MVC3 app, only my business layer, which would reference my DAL/Repo layer. 


I know that at the end of the day all DLLs are included in the bin folder but my problem is having to reference it explicitly via ""add reference"" in VS in order to be able to publish a WAP with all necessary files.


","
> 
> If I wasn't using a DI container, I wouldn't have to reference EntityFramework library in my MVC3 app, only my business layer which would reference my DAL/Repo layer.
> 
> 
> 


Yes, that's exactly the situation DI works so hard to avoid :)


With tightly coupled code, each library may only have a few references, but these again have other references, creating a deep graph of dependencies, like this:


![Deep Graph](https://i.stack.imgur.com/6HBLN.png)


Because the dependency graph is deep, it means that most libraries drag along a lot of other dependencies - e.g. in the diagram, *Library C* drags along *Library H, Library E, Library J, Library M, Library K* and *Library N*. This makes it harder to reuse each library independently from the rest - for example [in unit testing](https://blog.ploeh.dk/2011/11/10/TDDImprovesReusability.aspx).


However, in a loosely coupled application, by moving all the references to the [Composition Root](https://blog.ploeh.dk/2011/07/28/CompositionRoot.aspx), the **dependency graph is severely flattened**:


![Shallow Graph](https://i.stack.imgur.com/y2xO2.png)


As illustrated by the green color, it's now possible to reuse *Library C* without dragging along any unwanted dependencies.


However, all that said, with many DI Containers, you don't *have* to add hard references to all required libraries. Instead, you can use **late binding** either in the form of convention-based assembly-scanning (preferred) or XML configuration.


When you do that, however, you must remember to copy the assemblies to the application's bin folder, because that no longer happens automatically. Personally, I rarely find it worth that extra effort.


A more elaborate version of this answer can be found in [this excerpt](https://freecontent.manning.com/dependency-injection-in-net-2nd-edition-understanding-the-composition-root/) from my book [Dependency Injection, Principles, Practices, Patterns](https://manning.com/seemann2/).


"
"Google Actions sdk not playing audio in ssml from firebase storage
Google Actions SDK is unable to play audio files in SSML Audio tag from firebase storage. Although I could play same .ogg format file from wikipedia. 
<https://upload.wikimedia.org/wikipedia/en/9/9f/Sample_of_%22Another_Day_in_Paradise%22.ogg>


**firebase file**:
<https://firebasestorage.googleapis.com/v0/b/assisto-skill.appspot.com/o/TIP103_converted.mp3?alt=media&token=d0d08f9d-e340-478c-af00-657109683136>


I'm using it by sending an SSML string that looks like:



```
<speak>
  <audio src='https://firebasestorage.googleapis.com/v0/b/assisto-skill.appspot.com/o/TIP103_converted.mp3?alt=media&token=d0d08f9d-e340-478c-af00-657109683136'>
  </audio>
</speak>

```

But I'm getting the following error:



> 
> expected\_inputs[0].input\_prompt.rich\_initial\_prompt.items[0].simple\_response: 'ssml' could not be parsed.
> 
> 
> 


What could be the reason...Is this a limitation in actions sdk that url needs to be parameter free?


","The problem isn't that it can't use parameters - it is that when parsing parameters in URLs that are embedded in SSML it doesn't parse the & separating the parameters the way most HTML parsers parse them these days.


There are a few potential solutions for this, depending on how you can format the URL.


RFC 1866 used to recommend using a ; in place of an &, and some servers allow this. Unfortunately, Firebase Storage is not one of them.


Firebase Storage doesn't require the final token parameter in most cases, so in this case you could omit it, but that isn't a general solution to the problem.


Finally, you could use SGML encoding and replace the & as `&amp;` (note the leading ampersand and final semicolon). So your SSML would look something like:



```
<speak>
  <audio src='https://firebasestorage.googleapis.com/v0/b/assisto-skill.appspot.com/o/TIP103_converted.mp3?alt=media&amp;token=d0d08f9d-e340-478c-af00-657109683136'>
  </audio>
</speak>

```

This last method is probably the best for general purposes.


"
"Why is cv2.resize transposing the requested dimensions?

```
            s1 = image.shape  # Image is an opencv2 image (ndarray)
            w,h,d = s1
            image_ = cv2.resize(image, (w*2, h*2), interpolation = cv2.INTER_CUBIC)
            s2 = image_.shape

```

In the above, I asked for output to resize to (960,1280), but I get (1280, 960).


What's up with that?



```
  image_ = cv2.resize(image, None, fx=2, fy=2, interpolation = cv2.INTER_CUBIC)

```

... works as expected. (No points for pointing this out).


[![enter image description here](https://i.stack.imgur.com/wJxnM.png)](https://i.stack.imgur.com/wJxnM.png)


Runnable example:



```
import cv2

def run():
    cap = cv2.VideoCapture(0)

    while cap.isOpened():
        success, image = cap.read()  # (480, 640, 3)
        if not success:
            print(""Ignoring empty camera frame."")
            continue

        s1 = image.shape
        w, h, d = s1
        s2 = (w * 2, h * 2)
        image_ = cv2.resize(image, s2, interpolation=cv2.INTER_CUBIC)
        s3 = image_.shape
        cv2.imshow('Camera',
                   image_
                   )

        key = cv2.waitKey(5)
        if key & 0xFF == 27:
            break
    cap.release()


if __name__ == '__main__':
    run()


```

","Based on OpenCV documentation:



> 
> img.shape returns (Height, Width, Number of Channels)
> where
> 
> 
> 1. **Height** represents the number of pixel rows in the image or the number of pixels in > each column of the image array
> 2. **Width** represents the number of pixel columns in the image or the number of pixels in each row of the image array.
> 3. Number of Channels represents the number of components used to represent each pixel.
> 
> 
> 


So resize in `opencv` should be happened like:



```
import cv2
 
cap = cv2.VideoCapture(0)
_, img = cap.read()
cap.release()
 
print('Original Dimensions : ',img.shape)
 
scale_percent = 60 # percent of original size
width = int(img.shape[1] * scale_percent / 100)
height = int(img.shape[0] * scale_percent / 100)
dim = (width, height)
  
# resize image
resized = cv2.resize(img, dim, interpolation = cv2.INTER_AREA)

```

And your question regard why `fx=2, fy=2` works as expected, based on [this link](http://www.opencv.org.cn/opencvdoc/2.3.2/html/modules/imgproc/doc/geometric_transformations.html?highlight=resize#cv.Resize), when no desired size if given, it computes the new dimensions based on `fx` and `fy` to compute the desired dsize:


- FX: Scale factor along the horizontal axis. When it is 0, it is computed as `(double)dsize.width/image.cols`
- FY: Scale factor along the vertical axis. When it is 0, it is computed as `(double)dsize.height/image.rows`


[![Desired_size_formula](https://i.stack.imgur.com/q7Gjn.png)](https://i.stack.imgur.com/q7Gjn.png)


To wrap up, I guess that's because in `OpenCV` C++ version, they have considered `X` in horizontal direction (columns) and `Y` in vertical direction (rows) and since `CV2` is just a wrapping over `OpenCV-C++` this mismatch happened for considering what is the first dimension and what is the second as we have in `Numpy` arrays.


**Update**
HansHirse comment:


`OpenCV` is a `C++` library with a dedicated `cv::Mat` class representing images. And, for example, for some `cv::Mat` image, `image.size` returns a `cv::Size` object, which has width and height in that order, such that dimension parameters (width, height) in several `OpenCV` functions are consistent to that. The Python API of `OpenCV` uses `NumPy` arrays, which uses row-major order by default, such that for some NumPy array image, `image.shape` will return height and width in that order. It's not an intentional design choice here.


"
"How to pretty-print json with long array in the same line
I want to store collected data in json, they are all arrays of integers with couple of thousand elements each. 


I want the file to have the elements in the same list to be in the same line, should look like



```
{
    ""foo"": [
        [0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0]
        [0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0]
    ],
    ""bar"": [
        [0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0]
        [0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0]
    ]
}

```

Right now, all I can find `json` in python can only be export with either 



```
{'foo': [[0, 0, 0, 0, 0, 0, 0, 0, 0, 0], [0, 0, 0, 0, 0, 0, 0, 0, 0, 0] ... , 'bar': [0 ...]}

```

Or



```
{
    ""foo"": [
        [
            0,
            0,
            0,
            0,
            0,
            0,
            0,
            0,
            0,
        ],
        [
            0,
            0,
            0,
            0,
            0,
            0,
            0,
            0,
            0,

```

But neither is desirable.


","Using [`jsbeautifier`](https://pypi.org/project/jsbeautifier/):



```
import json
import jsbeautifier

x = {'foo': [[0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0], [0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0,
 0, 0, 0, 0, 0, 0]], 'bar': [[0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0], [0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 
0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0]]}  


options = jsbeautifier.default_options()
options.indent_size = 2
print(jsbeautifier.beautify(json.dumps(x), options))

```

OUTPUT:



```
{
  ""foo"": [
    [0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0],
    [0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0]
  ],
  ""bar"": [
    [0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0],
    [0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0, 0]
  ]
}

```

"
"Why is Pandas.eval() with numexpr so slow?
Test code:



```
import numpy as np
import pandas as pd

COUNT = 1000000

df = pd.DataFrame({
    'y': np.random.normal(0, 1, COUNT),
    'z': np.random.gamma(50, 1, COUNT),
})

%timeit df.y[(10 < df.z) & (df.z < 50)].mean()
%timeit df.y.values[(10 < df.z.values) & (df.z.values < 50)].mean()
%timeit df.eval('y[(10 < z) & (z < 50)].mean()', engine='numexpr')

```

The output on my machine (a fairly fast x86-64 Linux desktop with Python 3.6) is:



```
17.8 ms ±  1.3 ms per loop (mean ± std. dev. of 7 runs, 100 loops each)
8.44 ms ±  502 µs per loop (mean ± std. dev. of 7 runs, 100 loops each)
46.4 ms ± 2.22 ms per loop (mean ± std. dev. of 7 runs,  10 loops each)

```

I understand why the second line is a bit faster (it ignores the Pandas index). But why is the `eval()` approach using `numexpr` so slow? Shouldn't it be faster than at least the first approach? The documentation sure makes it seem like it would be: <https://pandas.pydata.org/pandas-docs/stable/enhancingperf.html>


","From the investigation presented below, it looks like the unspectacular reason for the worse performance is ""overhead"".


Only a small part of the expression `y[(10 < z) & (z < 50)].mean()` is done via `numexpr`-module. `numexpr` [doesn't support indexing](https://github.com/pydata/numexpr/issues/60), thus we can only hope for `(10 < z) & (z < 50)` to be speed-up - anything else will be mapped to `pandas`-operations.


However, `(10 < z) & (z < 50)` is not the bottle-neck here, as can be easily seen:



```
%timeit df.y[(10 < df.z) & (df.z < 50)].mean()  # 16.7 ms
mask=(10 < df.z) & (df.z < 50)                  
%timeit df.y[mask].mean()                       # 13.7 ms
%timeit df.y[mask]                              # 13.2 ms

```

`df.y[mask]` -takes the lion's share of the running time.


We can compare the profiler output for `df.y[mask]` and `df.eval('y[mask]')` to see what makes the difference. 


When I use the following script:



```
import numpy as np
import pandas as pd

COUNT = 1000000

df = pd.DataFrame({
    'y': np.random.normal(0, 1, COUNT),
    'z': np.random.gamma(50, 1, COUNT),
})

mask = (10 < df.z) & (df.z < 50)
df['m']=mask

for _ in range(500):
   df.y[df.m] 
   # OR 
   #df.eval('y[m]', engine='numexpr')

```

and run it with `python -m cProfile -s cumulative run.py` (or `%prun -s cumulative <...>` in IPython), I can see the following profiles. 


For direct call of the pandas functionality:



```
   ncalls  tottime  percall  cumtime  percall filename:lineno(function)
    419/1    0.013    0.000    7.228    7.228 {built-in method builtins.exec}
        1    0.006    0.006    7.228    7.228 run.py:1(<module>)
      500    0.005    0.000    6.589    0.013 series.py:764(__getitem__)
      500    0.003    0.000    6.475    0.013 series.py:812(_get_with)
      500    0.003    0.000    6.468    0.013 series.py:875(_get_values)
      500    0.009    0.000    6.445    0.013 internals.py:4702(get_slice)
      500    0.006    0.000    3.246    0.006 range.py:491(__getitem__)
      505    3.146    0.006    3.236    0.006 base.py:2067(__getitem__)
      500    3.170    0.006    3.170    0.006 internals.py:310(_slice)
    635/2    0.003    0.000    0.414    0.207 <frozen importlib._bootstrap>:958(_find_and_load)

```

We can see that almost 100% of the time is spent in `series.__getitem__` without any overhead.


For the call via `df.eval(...)`, the situation is quite different:



```
   ncalls  tottime  percall  cumtime  percall filename:lineno(function)
    453/1    0.013    0.000   12.702   12.702 {built-in method builtins.exec}
        1    0.015    0.015   12.702   12.702 run.py:1(<module>)
      500    0.013    0.000   12.090    0.024 frame.py:2861(eval)
 1000/500    0.025    0.000   10.319    0.021 eval.py:153(eval)
 1000/500    0.007    0.000    9.247    0.018 expr.py:731(__init__)
 1000/500    0.004    0.000    9.236    0.018 expr.py:754(parse)
 4500/500    0.019    0.000    9.233    0.018 expr.py:307(visit)
 1000/500    0.003    0.000    9.105    0.018 expr.py:323(visit_Module)
 1000/500    0.002    0.000    9.102    0.018 expr.py:329(visit_Expr)
      500    0.011    0.000    9.096    0.018 expr.py:461(visit_Subscript)
      500    0.007    0.000    6.874    0.014 series.py:764(__getitem__)
      500    0.003    0.000    6.748    0.013 series.py:812(_get_with)
      500    0.004    0.000    6.742    0.013 series.py:875(_get_values)
      500    0.009    0.000    6.717    0.013 internals.py:4702(get_slice)
      500    0.006    0.000    3.404    0.007 range.py:491(__getitem__)
      506    3.289    0.007    3.391    0.007 base.py:2067(__getitem__)
      500    3.282    0.007    3.282    0.007 internals.py:310(_slice)
      500    0.003    0.000    1.730    0.003 generic.py:432(_get_index_resolvers)
     1000    0.014    0.000    1.725    0.002 generic.py:402(_get_axis_resolvers)
     2000    0.018    0.000    1.685    0.001 base.py:1179(to_series)
     1000    0.003    0.000    1.537    0.002 scope.py:21(_ensure_scope)
     1000    0.014    0.000    1.534    0.002 scope.py:102(__init__)
      500    0.005    0.000    1.476    0.003 scope.py:242(update)
      500    0.002    0.000    1.451    0.003 inspect.py:1489(stack)
      500    0.021    0.000    1.449    0.003 inspect.py:1461(getouterframes)
    11000    0.062    0.000    1.415    0.000 inspect.py:1422(getframeinfo)
     2000    0.008    0.000    1.276    0.001 base.py:1253(_to_embed)
     2035    1.261    0.001    1.261    0.001 {method 'copy' of 'numpy.ndarray' objects}
     1000    0.015    0.000    1.226    0.001 engines.py:61(evaluate)
    11000    0.081    0.000    1.081    0.000 inspect.py:757(findsource)

```

once again about 7 seconds are spent in `series.__getitem__`, but there are also about 6 seconds overhead - for example about 2 seconds in `frame.py:2861(eval)` and about 2 seconds in `expr.py:461(visit_Subscript)`.


I did only a superficial investigation (see more details further below), but this overhead doesn't seems to be just constant but at least linear in the number of element in the series. For example there is `method 'copy' of 'numpy.ndarray' objects` which means that data is copied (it is quite unclear, why this would be necessary per se).


My take-away from it: using `pd.eval` has advantages as long as the evaluated expression can be evaluated with `numexpr` alone. As soon as this is not the case, there might be no longer gains but losses due to quite large overhead.




---


Using `line_profiler` (here I use %lprun-magic (after loading it with `%load_ext line_profliler`) for the function `run()` which is more or less a copy from the script above) we can easily find where the time is lost in `Frame.eval`:



```
%lprun -f pd.core.frame.DataFrame.eval
       -f pd.core.frame.DataFrame._get_index_resolvers 
       -f pd.core.frame.DataFrame._get_axis_resolvers  
       -f pd.core.indexes.base.Index.to_series 
       -f pd.core.indexes.base.Index._to_embed
       run()

```

Here we can see were the additional 10% are spent:



```
Line #      Hits         Time  Per Hit   % Time  Line Contents
==============================================================
  2861                                               def eval(self, expr, 
....
  2951        10        206.0     20.6      0.0          from pandas.core.computation.eval import eval as _eval
  2952                                           
  2953        10        176.0     17.6      0.0          inplace = validate_bool_kwarg(inplace, 'inplace')
  2954        10         30.0      3.0      0.0          resolvers = kwargs.pop('resolvers', None)
  2955        10         37.0      3.7      0.0          kwargs['level'] = kwargs.pop('level', 0) + 1
  2956        10         17.0      1.7      0.0          if resolvers is None:
  2957        10     235850.0  23585.0      9.0              index_resolvers = self._get_index_resolvers()
  2958        10       2231.0    223.1      0.1              resolvers = dict(self.iteritems()), index_resolvers
  2959        10         29.0      2.9      0.0          if 'target' not in kwargs:
  2960        10         19.0      1.9      0.0              kwargs['target'] = self
  2961        10         46.0      4.6      0.0          kwargs['resolvers'] = kwargs.get('resolvers', ()) + tuple(resolvers)
  2962        10    2392725.0 239272.5     90.9          return _eval(expr, inplace=inplace, **kwargs)

```

and `_get_index_resolvers()` can be drilled down to `Index._to_embed`:



```
Line #      Hits         Time  Per Hit   % Time  Line Contents
==============================================================
  1253                                               def _to_embed(self, keep_tz=False, dtype=None):
  1254                                                   """"""
  1255                                                   *this is an internal non-public method*
  1256                                           
  1257                                                   return an array repr of this object, potentially casting to object
  1258                                           
  1259                                                   """"""
  1260        40         73.0      1.8      0.0          if dtype is not None:
  1261                                                       return self.astype(dtype)._to_embed(keep_tz=keep_tz)
  1262                                           
  1263        40     201490.0   5037.2    100.0          return self.values.copy()

```

Where the `O(n)`-copying happens.


"
"Zend - How to install on WAMP
If there is an easy to follow instruction or tutorial that I can use to learn how to install Zend on my machine that has a WAMP installation?


The video listed
<http://www.zendcasts.com/getting-started-with-zend-and-wamp-server/2009/06/>


is hard to read what he is writing.


Thank you


","Conceptually, the whole thing is:


1. Create virtual host pointing it to the sample app
2. Make sure that PHP `include_path` contains the path to the Zend library.


But the specifics can be tricky if you are not accustomed to it. So here is at least a little bit more color.


1. Create a folder for your app, something like `C:\apps\myapp`.
2. Copy a sample ZF app - like [this](http://framework.zend.com/demos/ZendFrameworkQuickstart.zip) or [this](http://github.com/harikt/zfbase) - into that space so that the `myapp` folder has the typical subfolders like `application`, `library`, `public`, `tests`, etc.
3. create a virtual host within your Apache. This is a two step process: 


3.1 Modify your `hosts` file - on my WinXP machine, it resides in the folder `C:\windows\system32\drivers\etc` to contain a line like 


`127.0.0.1 myvirtualapp`


I am intentionally choosing a virtual host name `myvirtualapp` that is different from the app folder name `myapp` to demonstrate that they are conceptually different creatures. One is a name that the OS and Apache recognize as an HTTP host; the other is a local filesystem path.


3.2 Add an entry into Apache's vhost file - typically in the Apache folder hierarchy at something like `conf/extra/httpd-vhosts.conf`. A minimal entry there will look something like this:


<VirtualHost \*:80>


DocumentRoot ""C:/apps/myapp/public""


ServerName myvirtualapp


</VirtualHost>
4. Restart Apache.
5. Make sure that the Zend library is copied into your `c:\apps\myapp\library` folder, so that there is a subfolder named `Zend` with the rest of the library contained inside.
6. Make sure that the folder `c:\apps\myapp\library` is on your PHP include path. There are many ways to do this, but typically this is done in `c:\apps\myapp\public\index.php`. Usually, that `library` folder is referenced in `index.php` as `realpath(APPLICATION_PATH . '/../library')`.
7. Browse to the url: `http://myvirtualapp/`


With any luck, you should see the app!


"
"Does Spring AOP do compile time weaving or load time weaving?
I am starting to use Spring AOP for a project and I am a little bit confused about weaving. I know that Spring AOP has a dependency on AspectJweaver.jar, but as the documentation says, this is not for the weaving, but just that it uses some of the classes from this jar.


But my question is, if it is not using AspectJ for weaving, does Spring AOP have its own weaving and is it performed at load time or compile time?


The relevant part of my Spring Configuration XML file is:



```
<context:annotation-config />

<tx:advice id=""txAdvice"" transaction-manager=""transactionManager"">
    <tx:attributes>
        <tx:method name=""get*"" read-only=""true"" />
        <tx:method name=""*"" />
    </tx:attributes>
</tx:advice>

<aop:config>
    <aop:pointcut id=""myaop"" expression=""execution(* my.package.*.*(..))"" />
    <aop:advisor advice-ref=""txAdvice"" pointcut-ref=""myaop"" />
</aop:config>

<bean id=""transactionManager"" class=""org.springframework.jdbc.datasource.DataSourceTransactionManager"">
    <property name=""dataSource"" ref=""dataSource"" />
</bean>

```

","<http://docs.spring.io/spring/docs/4.0.1.RELEASE/spring-framework-reference/htmlsingle/#aop-introduction-defn>


Under 8.1.1, item weaving, it says:



> 
> Weaving: linking aspects with other application types or objects to
>  create an advised object. This can be done at compile time (using the
>  AspectJ compiler, for example), load time, or at runtime. Spring AOP,
>  like other pure Java AOP frameworks, performs weaving at runtime.
> 
> 
> 


Spring doesn't do the same type of load-time weaving as AspectJ, but works on proxies, as described in another part of the documentation:


<http://docs.spring.io/spring/docs/4.0.1.RELEASE/spring-framework-reference/htmlsingle/#aop-understanding-aop-proxies>


Edit: Just saw your comment, you are correct in that assumption. The documentation gives a rather complete explanation of how it works exactly. :)


"
"Pandas datetimes with different formats in the same column
I have a pandas data frame which has datetimes with 2 different formats e.g.:



```
3/14/2019 5:15:32 AM
2019-08-03 05:15:35
2019-01-03 05:15:33
2019-01-03 05:15:33
2/28/2019 5:15:31 AM
2/27/2019 11:18:39 AM

```

...


I have tried various formats but get errors `like ValueError: unconverted data remains: AM`


I would like to get the format as 2019-02-28 and have the time removed


","You can use `pd.to_datetime().dt.strftime()` to efficienty convert the entire column to a datetime object and then to a string with Pandas intelligently guessing the date formatting:



```
df = pd.Series('''3/14/2019 5:15:32 AM
2019-08-03 05:15:35
2019-01-03 05:15:33
2019-01-03 05:15:33
2/28/2019 5:15:31 AM
2/27/2019 11:18:39 AM'''.split('\n'), name='date', dtype=str).to_frame()

print(pd.to_datetime(df.date).dt.strftime('%Y-%m-%d'))

```


```
0    2019-03-14
1    2019-08-03
2    2019-01-03
3    2019-01-03
4    2019-02-28
5    2019-02-27
Name: date, dtype: object

```

If that doesn't give you what you want, you will need to identify the different kinds of formats and apply different settings when you convert them to datetime objects:



```
# Classify date column by format type
df['format'] = 1
df.loc[df.date.str.contains('/'), 'format'] = 2
df['new_date'] = pd.to_datetime(df.date)

# Convert to datetime with two different format settings
df.loc[df.format == 1, 'new_date'] = pd.to_datetime(df.loc[df.format == 1, 'date'], format = '%Y-%d-%m %H:%M:%S').dt.strftime('%Y-%m-%d')
df.loc[df.format == 2, 'new_date'] = pd.to_datetime(df.loc[df.format == 2, 'date'], format = '%m/%d/%Y %H:%M:%S %p').dt.strftime('%Y-%m-%d')
print(df)

```


```
                    date  format    new_date
0   3/14/2019 5:15:32 AM       2  2019-03-14
1    2019-08-03 05:15:35       1  2019-03-08
2    2019-01-03 05:15:33       1  2019-03-01
3    2019-01-03 05:15:33       1  2019-03-01
4   2/28/2019 5:15:31 AM       2  2019-02-28
5  2/27/2019 11:18:39 AM       2  2019-02-27

```

"
"eval lines from a file
I am creating a program that stores options in a txt file, and I tried a few different lines but so far I think an `eval` is the best one.


I got this info in the file:



```
colordefondo = ""#abcdef""

```

and I got this `eval` in the program:



```
for line in open(""filename.txt"", ""r""):
  eval(line)

```

However, when I try the code it gives me this error:



```
Traceback (most recent call last):
File ""D:\project-sae\Sae.py"", line 25, in <module>
eval(line)
File ""<string>"", line 1
colordefondo = ""#abcdef""
             ^
SyntaxError: invalid syntax

```

My question is Why? And if anyone knows a better way to load and store the value of several options from a txt file it would be great. Nevertheless my question is about why that `eval` is failing, my knowledge about how `eval` works seems wrong and I don't know why, used it several times before, just never from a file.


","As the documentation says, ['eval'](http://docs.python.org/2/library/functions.html#eval) evaluates an *expression*. Not a statement. Assignments are statements.


You could use the `exec` statement. But if you want to exec all of the lines of a file, that's exactly what [`execfile`](http://docs.python.org/2/library/functions.html#execfile) does. (I'm assuming Python 2.x here; things are a little different in 3.x, but the idea is similar.)


However, this is a bad idea to do in the first place. A better way to load and store the value of several options from a txt file is to use *anything other than Python source* as your format. JSON, .ini files, .rc files, etc. Python comes with code built-in to handle a variety of such formats. See [File Formats](http://docs.python.org/2/library/fileformats.html) and [Internet Data Handling](http://docs.python.org/2/library/netdata.html) for a list of the relevant modules, read through them, and pick the one you want.


"
"use LIKE inside CASE/THEN in Oracle?
I have this part of the code



```
WHERE p.id = m.id AND (
      CASE
        WHEN (c.match_text IS NOT NULL)
        THEN (
          (lower(p.station) LIKE '%'||c.match_text||'%')
          OR
          (
            (lower(p.station) LIKE '%sealed in%')
            AND
            (lower(p.affected_units) LIKE '%'||c.match_text||'%')
          )
        )
        WHEN(u.name LIKE '%Memory%')
        THEN (
          (lower(p.affected_fuels) LIKE '%nuclear%')
          AND

```

I get error with LIKE in this (lower(p.station) LIKE '%'||c.match\_text||'%')!!


why is that?


I was thinking this approach will fix the issue:



```
WHEN (c.match_text IS NOT NULL)
THEN (
  CASE WHEN
    (lower(p.station) LIKE '%'||c.match_text||'%')
    OR
    (
      (lower(p.station) LIKE '%sealed in%')
      AND
      (lower(p.affected_units) LIKE '%'||c.match_text||'%')
    )
    THEN true
    ELSE false
    END
)

```

Thanks


","CASE() is basically a translator: it returns a value for a given set of conditions. The problem you have is that you are trying to force it to return a BOOLEAN value. This won't work because BOOLEAN is not a valid SQL data type.


So I think you'll have to implement your logic as nested AND/OR clauses. It's a bit gnarly but not much more complicated than your original statement. Something like this:



```
WHERE p.id = m.id 
AND (  
        ( c.match_text IS NOT NULL
         AND ( lower(p.station) LIKE '%'||c.match_text||'%'
              OR ( lower(p.station) LIKE '%sealed in%'
                    AND   lower(p.affected_units) LIKE '%'||c.match_text||'%'
                 )           
              ) 
         )        
     OR (u.name LIKE '%Memory%'
        AND lower(p.affected_fuels) LIKE '%nuclear%'
        AND ...

```

"
"Python: timezone.localize() not working
I am having some issues getting `timezone.localize()` to work correctly. My goal is to grab today's date and convert it from CST to EST. Then finally format the datetime before spitting it out. I am able to format the date correctly, but the datetime is not changing from CST to EST. Additionally when I format the date I don't see the text representation of the timezone included.


Below I have listed out a simple program I created to test this out:



```
#! /usr/bin/python
#Test script

import threading
import datetime
import pexpect
import pxssh
import threading
from pytz import timezone
import pytz

est = timezone('US/Eastern')
curtime = est.localize(datetime.datetime.now().strftime(""%a %b %d %H:%M:%S %Z %Y""))
#test time change
#curtime = datetime.datetime.now().strftime(""%a %b %d %H:%M:%S %Z %Y"")

class ThreadClass(threading.Thread):
  def run(self):
    #now = (datetime.datetime.now() + datetime.timedelta(0, 3600))
    now = (datetime.datetime.now())
    print ""%s says Hello World at time: %s"" % (self.getName(), curtime)

for i in range(3):
  t = ThreadClass()
  t.start()

```

","`.localize()` takes a naive datetime object and interprets it *as if* it is in that timezone. It does *not* move the time to another timezone. A naive datetime object has *no* timezone information to be able to make that move possible.


You want to interpret `now()` in your *local* timezone instead, then use `.astimezone()` to interpret the datetime in another timezone:



```
est = timezone('US/Eastern')
cst = timezone('US/Central')
curtime = cst.localize(datetime.datetime.now())
est_curtime = curtime.astimezone(est).strftime(""%a %b %d %H:%M:%S %Z %Y"")

def run(self):
    print(""%s says Hello World at time: %s"" % (self.getName(), est_curtime))

```

"
"Static variable resetting its value to 0 on it's own apparently
I have a base class called `Component` which has a number of classes derived from it. I want each class to have an integer associated with it (doesn't matter which component gets what value, as long as they start at 0 and are contiguous). I don't know how to do this directly, so in the same file as `Component` I added the following:



```
template <typename T>
class ComponentIdentifier
{
public:
    static unsigned int cid;
};

static unsigned int CIDCounter = 0;
template <typename T> unsigned int ComponentIdentifier<T> = CIDCounter++;

template <typename T> unsigned int ComponentID()
{
    return ComponentIdentifier<T>::cid;
}

unsigned int ComponentCount(); // Defined in .cpp file, just returns CIDCounter

```

Now I tested the ComponentID() function and it appears to work fine. Each component class that I tried ComponentID on returned a different integer just as I expected. However, whenever I call ComponentCount I get 0.


e.g. if I have the following lines of code:



```
std::cout << ComponentID<AAA>() << std::endl;
std::cout << ComponentID<BBB>() << std::endl;
std::cout << ComponentID<CCC>() << std::endl;
std::cout << ComponentCount() << std::endl;

```

then my output is:



```
0
1
2
0

```

I suspect what's happening is that CIDCounter is being set to 0 again after its used to set the cid's of each component, but I don't know for sure and that seems kind of weird. Is there a way to do what I want or am I crazy and this whole plan doomed to failure?


","You've declared your identifier as static:



```
static unsigned int CIDCounter = 0;

```

which means each compilation unit will get its own copy of the variable. Instead you should declare it as extern in the header file:



```
extern unsigned int CIDCounter;

```

and initialise it in the implementation file



```
unsigned int CIDCounter = 0;

```

One should note that without appropriate locks this won't be thread-safe.


To clarify further:


The `static` keyword in this context means that the variable or function is limited to the current compilation unit (usually a cpp file). I'm guessing you have, say, a `main.cpp` and an `idcounter.cpp` - so now we have two variables (one for each of the compilation units) `main_CIDCounter` and `idcounter_CIDCounter` (variable names are for explanation only!).


When you execute your test code in main.cpp the template function sees `main_CIDCounter` because that's the current compilation unit and you get the expected 1,2,3. However, when you call `ComponentCount()` code from `idcounter.cpp` is used, and that code sees `idcounter_CIDCounter` - which hasn't been modified at all. If you had other compilation units you would see similar behaviour, where each cpp file would appear to be maintaining its own ID counter.


The fix I've described is simply to only have one copy of `CIDCounter` in `idcounter.cpp`, and declare it as external in all the other compilation units.


"
"what is the difference between and in maven?
In my project I alway use `<dependency><dependency>` but I can see `<parent></parent>` in some project `pom.xml` like:



```
<parent>
    <groupId>org.springframework.boot</groupId>
    <artifactId>spring-boot-starter-parent</artifactId>
    <version>1.5.2.RELEASE</version>
</parent>

```

so I want to konw when to use it.


","**`<parent>` a superset of `<dependencies>`**


`<parent>` and `<dependencies>` elements are two distinct things but it exists all the same an important relation between them. 


Simply said the `parent` defines the parent pom of the current pom and `dependencies` defines the actual dependencies of the current pom.  

The parent pom can define `dependencies` but also many other things inherited by the children Maven project (and particularly the `dependencyManagement` element and the `build` element that allow to configure many things) can so be considered in a some way as a superset of the `dependencies` element.  

Here is the list of [elements inherited from the parent pom](https://maven.apache.org/pom.html#Inheritance_v) : 



```

groupId
version
description
url
inceptionYear
organization
licenses
developers
contributors
mailingLists
scm
issueManagement
ciManagement
properties
dependencyManagement
dependencies
repositories
pluginRepositories
build
plugin executions with matching ids
plugin configuration
etc.
reporting
profiles

```

**As use `dependencies` and as use `<parent>`?** 


We can use only the first, only the second or both.  

It depends really on the way which the Maven projects are designed.  

Trying to enumerate all possible configurations would be long and not necessary very helpful.  

So I think that you should really retain that `parent` is much more structuring as `dependencies` as it defines both more things for the children projects but it also allow not to repeat the actual configuration that you want to define in a set of projects.  

So you should favor `parent` as you want to make inherit some child Maven projects an overall configuration and not only a list of dependencies. 




---


Your example is perfect to illustrate the consequences on the client projects from using `<parent>` or `dependencies` as alternative.


**1) With parent inheriting**


Here the project inherits from the `spring-boot-starter-parent` pom :



```
<parent>
    <groupId>org.springframework.boot</groupId>
    <artifactId>spring-boot-starter-parent</artifactId>
    <version>1.5.2.RELEASE</version>
</parent>

```

As a consequence, the project will inherit any things defined in `dependencies` and `dependencyManagement` but it will also inherit from the `<build>` element defined in the super pom. 


For example you would have the Maven compiler plugin configured out of the box with Java 8 and UTF-8 (that you can of course redefined in your child project) :



```
<properties>
    <project.reporting.outputEncoding>UTF-8</project.reporting.outputEncoding>
    <java.version>1.8</java.version>
    <resource.delimiter>@</resource.delimiter>
    <maven.compiler.source>${java.version}</maven.compiler.source>
    <project.build.sourceEncoding>UTF-8</project.build.sourceEncoding>
    <maven.compiler.target>${java.version}</maven.compiler.target>
</properties>

```

Additionally, some other plugins potential useful by a Spring Boot project will also be defined in the super pom and be inherited by your project such as :



```
<pluginManagement>
  <plugins>
     ...
    <plugin>
        <groupId>org.springframework.boot</groupId>
        <artifactId>spring-boot-maven-plugin</artifactId>
        <executions>
            <execution>
                <goals>
                    <goal>repackage</goal>
                </goals>
            </execution>
        </executions>
        <configuration>
            <mainClass>${start-class}</mainClass>
        </configuration>
    </plugin>
    ...
  </plugins>
</pluginManagement>

```

Note that a parent pom may define `dependencies`, directly inherited by the child projects but not necessary.  

For example the `spring-boot-starter-parent` doesn't define any `dependency` directly inherited by child projects but instead of define `dependency` in `<dependencyManagement><dependencies>`.  

It means that children of this parent pom may use the dependencies but they have to explicitly state that in `dependencies`.  

For example :



```
<parent>
    <groupId>org.springframework.boot</groupId>
    <artifactId>spring-boot-starter-parent</artifactId>
    <version>1.5.2.RELEASE</version>
</parent>
<dependencies>
    <dependency>
        <groupId>org.springframework.boot</groupId>
        <artifactId>spring-boot-starter-web</artifactId>
    </dependency>
</dependencies>

```

Note that the version is not valued as inherited.


**2) Without parent inheriting**


You will have to define all required dependencies by your Spring Boot application in or more straightly use the [`spring-boot-dependencies`](https://docs.spring.io/spring-boot/docs/current/reference/html/using-boot-build-systems.html) dependency in `dependencyManagement` with a `import` scope to have a way to declare them thanks to the dependency management feature :



```
<dependencyManagement>
        <dependencies>
        <dependency>
            <!-- Import dependency management from Spring Boot -->
            <groupId>org.springframework.boot</groupId>
            <artifactId>spring-boot-dependencies</artifactId>
            <version>1.5.2.RELEASE</version>
            <type>pom</type>
            <scope>import</scope>
        </dependency>
    </dependencies>
</dependencyManagement>

```

But in any cases you will never inherit from the `plugins` configured out of the box by the parent as you don't have parent.  

So you should declare them explicitly in the pom.xml of your project. 


For example to define the compiler version, used encoding and configure the build to repackage the built component(to make it standalone executable), you will will have to specify more things :



```
<properties>
    <maven.compiler.source>1.8</maven.compiler.source>
    <maven.compiler.target>1.8</maven.compiler.target>
    <project.build.sourceEncoding>UTF-8</project.build.sourceEncoding>
    <springboot.version>1.5.2.RELEASE</springboot.version>
</properties>

<dependencyManagement>
    <dependencies>
        <dependency>
            <groupId>org.springframework.boot</groupId>
            <artifactId>spring-boot-dependencies</artifactId>
            <version>${springboot.version}</version>
            <type>pom</type>
            <scope>import</scope>
        </dependency>
    </dependencies>
</dependencyManagement>

<build>
   <plugins>
        <plugin>
            <groupId>org.springframework.boot</groupId>
            <artifactId>spring-boot-maven-plugin</artifactId>
             <version>${springboot.version}</version>
            <executions>
                <execution>
                    <goals>
                        <goal>repackage</goal>
                    </goals>
                </execution>
            </executions>
            <configuration>
                <mainClass>myClass</mainClass>
            </configuration>
        </plugin>      
   <plugins>
</build>

```

"
"Get SVG icons dynamically in Next.js
I wanted to get SVG icons dynamically. I found a method to do this but it seems I made some mistakes. Where am I doing it wrong?


**Icon.js**



```
import React from ""react"";
import { ReactComponent as Bollards } from ""./icons/bollards.svg"";
import { ReactComponent as Earthquake } from ""./icons/earthquake.svg"";
import { ReactComponent as Fire } from ""./icons/fire.svg"";
import { ReactComponent as Healthy } from ""./icons/heartbeat.svg"";
import { ReactComponent as Home } from ""./icons/home.svg"";
import { ReactComponent as Planting } from ""./icons/planting.svg"";
import { ReactComponent as Business } from ""./icons/suitcase.svg"";
import { ReactComponent as Travel } from ""./icons/airplane-around-earth.svg"";

const iconTypes = {
  bollards: Bollards,
  earthQuake: Earthquake,
  fire: Fire,
  healthy: Healthy,
  home: Home,
  planting: Planting,
  business: Business,
  travel: Travel
};

const IconComponent = ({ name, ...props }) => {
  let Icon = iconTypes[name];
  return <Icon {...props} />;
};

export default IconComponent;

```

**Feautures.js**



```
import React from ""react"";
import Icon from ""./icon"";

export default function Features() {
  return (
    <div>
      <Icon name=""bollards"" />
    </div>
  );
}

```

**I get this error when trying to export icons.**



```
error - ./components/icon.js
Attempted import error: 'ReactComponent' is not exported from './icons/bollards.svg' (imported as 'Bollards').

```

","You can use [SVGR](https://github.com/gregberge/svgr) that allows us to import SVGs into your React applications as components.


You need to add `@svgr/webpack` as a dependency and modify the `next.config.js` file like this.


`next.config.js`:



```
module.exports = {
  webpack(config) {
    config.module.rules.push({
      test: /\.svg$/,
      use: [""@svgr/webpack""]
    });
    return config;
  }
};

```

Then, you can simply import your icons without using `ReactComponent`.


`Icon.js`:



```
import React from ""react"";

import Bollards from './icons/bollards.svg';
import Earthquake from './icons/earthquake.svg';
import Fire from './icons/fire.svg';
import Healthy from './icons/heartbeat.svg';
import Home from './icons/home.svg';
import Planting from './icons/planting.svg';
import Business from './icons/suitcase.svg';
import Travel from './icons/airplane-around-earth.svg';

const iconTypes = {
  bollards: Bollards,
  earthQuake: Earthquake,
  fire: Fire,
  healthy: Healthy,
  home: Home,
  planting: Planting,
  business: Business,
  travel: Travel
};

const IconComponent = ({ name, ...props }) => {
  let Icon = iconTypes[name];
  return <Icon {...props} />;
};

export default IconComponent;

```

Working demo is [available on CodeSandbox](https://codesandbox.io/s/svg-icons-278et).


"
"Search Functionality for Listview
i'm new to android. I'm just trying to make one simple search functionality for my app. My app consists one `ListView` one `EditText` and one `Button` for search. My `ListView` contents are listed from database using custom adapter which is extends `BaseAdapter` 


Now, what i'm trying to do is, i want to search any records from `ListView` For example, if i've some records like 



> 
> Optimization, Operations, Data Mining, Computer Ethics, Computer Architecture and etc...
> 
> 
> 


So, when i type some record name like `op`


The listview should listed the records which is started from `op...` I've referred something for this, from i got [addTextChangedListener](http://developer.android.com/reference/android/widget/TextView.html#addTextChangedListener%28android.text.TextWatcher%29) But, i don't know how to do this?


And, Can we do this same functionality with click of `button` 


Has anyone having any idea on this? Thanks in advance.


","I had asked a similar kinda question a while back. Here: [Filtering a ListView with Baseadapter filters text not images](https://stackoverflow.com/q/10764675/450534). Although, my specific question concerned a GridView, the concept (and the code) can be substituted for a ListView.


NOTE: This will be a lengthy post but I think necessary for the sake of completeness (I am leaving out the `imports` though)


The main activity (*Friends.java*)



```
public class Friends extends SherlockActivity {

    // BUNDLE OBJECT TO GET DATA FROM EARLIER ACTIVITY
    Bundle extras;

    // INITIAL ALBUM ID AND NAME
    String initialUserID;

    // THE GRIDVIEW
    GridView gridOfFriends;

    // THE ADAPTER
    FriendsAdapter adapter;

    // ARRAYLIST TO HOLD DATA
    ArrayList<getFriends> arrFriends;

    // LINEARLAYOUT TO SHOW THE FOOTER PROGRESS BAR
    LinearLayout linlaProgressBar;

    // THE EDITTEXT TO FILTER USERS
    EditText filterText;

    @Override
    protected void onCreate(Bundle savedInstanceState) {
        super.onCreate(savedInstanceState);
        setContentView(R.layout.friends_grid_list);

        extras = getIntent().getExtras();

        if (extras.containsKey(""USER_ID""))  {
            initialUserID = extras.getString(""USER_ID"");
        } else {
            Toast.makeText(
                    getApplicationContext(), 
                    ""There was a problem getting your Friends Data. Please hit the back button and try again."", 
                    Toast.LENGTH_SHORT).show();
        }

        ActionBar actionBar = getSupportActionBar();
        actionBar.setDisplayHomeAsUpEnabled(true);
        actionBar.setDisplayShowHomeEnabled(true);
        actionBar.setDisplayShowTitleEnabled(true);
        actionBar.setTitle(""Your Friends"");

        // CAST THE GRIDVIEW
        gridOfFriends = (GridView) findViewById(R.id.gridFriends);

        // INSTANTIATE THE ARRAYLIST
        arrFriends = new ArrayList<getFriends>();

        // CAST THE ADAPTER
        adapter = new FriendsAdapter(Friends.this, arrFriends);

        // CAST THE LINEARLAYOUT THAT HOLDS THE PROGRESS BAR
        linlaProgressBar = (LinearLayout) findViewById(R.id.linlaProgressBar);
        linlaProgressBar.setVisibility(View.GONE);

        // GET THE LOGGED IN USERS FRIENDS DATA
        if (initialUserID != null)  {
            new getFriendsData().execute();
        } else {
            Toast.makeText(
                    getApplicationContext(), 
                    ""There was a problem getting your Friends Data. Please hit the back button and try again."", 
                    Toast.LENGTH_SHORT).show();
        }

        // CAST THE EDITTEXT AND SETUP FILTERING
        filterText = (EditText) findViewById(R.id.editFilterList);
        filterText.addTextChangedListener(filterTextWatcher);
    }

    private class getFriendsData extends AsyncTask<Void, Void, Void>    {

        @Override
        protected void onPreExecute() {

            // SHOW THE BOTTOM PROGRESS BAR (SPINNER) WHILE LOADING THE FRIENDS LIST
            linlaProgressBar.setVisibility(View.VISIBLE);
        }

        @Override
        protected Void doInBackground(Void... params) {

            try {
                String query = 
                        ""SELECT name, uid, pic_big "" +
                        ""FROM user "" +
                        ""WHERE uid in "" +
                        ""(SELECT uid2 FROM friend WHERE uid1=me()) "" +
                        ""order by name"";
                Bundle paramGetFriendsList = new Bundle();
                paramGetFriendsList.putString(""method"", ""fql.query"");
                paramGetFriendsList.putString(""query"", query);

                String resultFriendsList = Utility.mFacebook.request(paramGetFriendsList);

                JSONArray JAFriends = new JSONArray(resultFriendsList);

                getFriends friends;

                if (JAFriends.length() == 0)    {

                } else {
                    for (int i = 0; i < JAFriends.length(); i++) {
                        JSONObject JOFriends = JAFriends.getJSONObject(i);

                        friends = new getFriends();

                        // SET FRIENDS ID
                        if (JOFriends.has(""uid""))   {
                            friends.setFriendID(JOFriends.getString(""uid""));
                        } else {
                            friends.setFriendID(null);
                        }

                        // SET FRIENDS NAME
                        if (JOFriends.has(""name""))  {
                            friends.setFriendName(JOFriends.getString(""name""));
                        } else {
                            friends.setFriendName(null);
                        }

                        // SET FRIENDS PROFILE PICTURE
                        if (JOFriends.has(""pic_big""))   {
                            friends.setFriendProfile(JOFriends.getString(""pic_big""));
                        } else {
                            friends.setFriendProfile(null);
                        }

                        arrFriends.add(friends);

                    }
                }

            } catch (Exception e) {
                e.printStackTrace();
            }

            return null;
        }

        @Override
        protected void onPostExecute(Void result) {

            // SET THE ADAPTER TO THE GRIDVIEW
            gridOfFriends.setAdapter(adapter);

            // HIDE THE BOTTOM PROGRESS BAR (SPINNER) AFTER LOADING THE FRIENDS LIST
            linlaProgressBar.setVisibility(View.GONE);
        }

    }

    private TextWatcher filterTextWatcher = new TextWatcher() {

        @Override
        public void onTextChanged(CharSequence s, int start, int before, int count) {

            adapter.getFilter().filter(s.toString().toLowerCase());
            adapter.notifyDataSetChanged();
        }

        @Override
        public void beforeTextChanged(CharSequence s, int start, int count, int after) {

        }

        @Override
        public void afterTextChanged(Editable s) {

        }
    };
}

```

The *getFriends.java* class for the `ArrayList`:



```
public class getFriends {

    String friendID;
    String friendName;
    String friendProfile;

    // SET FRIENDS ID
    public void setFriendID(String friendID) {
        this.friendID = friendID;
    }

    // GET FRIENDS ID
    public String getFriendID() {
        return friendID;
    }

    // SET FRIENDS NAME
    public void setFriendName(String friendName) {
        this.friendName = friendName;
    }

    // GET FRIENDS NAME
    public String getFriendName() {
        return friendName;
    }

    // SET FRIENDS PROFILE
    public void setFriendProfile(String friendProfile) {
        this.friendProfile = friendProfile;
    }

    // GET FRIENDS PROFILE
    public String getFriendProfile() {
        return friendProfile;
    }
}

```

And finally, the adapter class (*FriendsAdapter.java*)



```
public class FriendsAdapter extends BaseAdapter implements Filterable {

    ProgressDialog dialog;

    Activity activity;

    LayoutInflater inflater = null;
    ImageLoader imageLoader;

    ArrayList<getFriends> arrayFriends;
    List<getFriends> mOriginalNames;

    FriendsAdapter(Activity a, ArrayList<getFriends> arrFriends) {

        activity = a;

        arrayFriends = arrFriends;

        inflater = (LayoutInflater) activity
                .getSystemService(Context.LAYOUT_INFLATER_SERVICE);
        imageLoader = new ImageLoader(activity.getApplicationContext());
    }

    public int getCount() {
        return arrayFriends.size();
    }

    public Object getItem(int position) {
        return arrayFriends.get(position);
    }

    public long getItemId(int position) {
        return position;
    }

    @Override
    public void notifyDataSetChanged() {
        super.notifyDataSetChanged();
    }

    public View getView(final int position, View convertView, ViewGroup parent) {
        View vi = convertView;
        if (convertView == null)
            vi = inflater.inflate(R.layout.friends_grid_items, null);

        ImageView imgProfilePicture = (ImageView) vi.findViewById(R.id.imgProfilePicture);
        TextView txtUserName = (TextView) vi.findViewById(R.id.txtUserName);
        FrameLayout mainContainer = (FrameLayout) vi.findViewById(R.id.mainContainer);


        txtUserName.setText(arrayFriends.get(position).getFriendName());

        if (arrayFriends.get(position).getFriendProfile() != null) {
            imageLoader.DisplayImage(arrayFriends.get(position).getFriendProfile(),imgProfilePicture);
        } else if (arrayFriends.get(position).getFriendProfile() == null) {
            imgProfilePicture.setVisibility(View.GONE);
        }

        mainContainer.setOnClickListener(new View.OnClickListener() {

            @Override
            public void onClick(View v) {
                Intent showFriendsProfile = new Intent(activity.getApplicationContext(), UserProfileNew.class);
                showFriendsProfile.putExtra(""USER_ID"", arrayFriends.get(position).getFriendID());
                showFriendsProfile.putExtra(""NAME"", arrayFriends.get(position).getFriendName());
                activity.startActivity(showFriendsProfile);
            }
        });

        return vi;
    }

    @Override
    public Filter getFilter() {

        Filter filter = new Filter() {

            @SuppressWarnings(""unchecked"")
            @Override
            protected void publishResults(CharSequence constraint,
                    FilterResults results) {

                arrayFriends = (ArrayList<getFriends>) results.values;
                notifyDataSetChanged();
            }

            @Override
            protected FilterResults performFiltering(CharSequence constraint) {

                FilterResults results = new FilterResults();
                ArrayList<getFriends> FilteredArrayNames = new ArrayList<getFriends>();

                if (mOriginalNames == null) {
                    mOriginalNames = new ArrayList<getFriends>(arrayFriends);
                }
                if (constraint == null || constraint.length() == 0) {
                    results.count = mOriginalNames.size();
                    results.values = mOriginalNames;
                } else {
                    constraint = constraint.toString().toLowerCase();
                    for (int i = 0; i < mOriginalNames.size(); i++) {
                        getFriends dataNames = mOriginalNames.get(i);
                        if (dataNames.getFriendName().toLowerCase()
                                .contains(constraint.toString())) {
                            FilteredArrayNames.add(dataNames);
                        }
                    }

                    results.count = FilteredArrayNames.size();
                    // System.out.println(results.count);

                    results.values = FilteredArrayNames;
                    // Log.e(""VALUES"", results.values.toString());
                }

                return results;
            }
        };

        return filter;
    }
}

```

You can use the concept here and substitute for your ListView. I have used this elsewhere for a ListView and works as it should. Unfortunately, I cannot give that code away on a public fora. Hope this helps you though. Again, a very lengthy post no doubt, but necessary I believe.


"
"C# equivalent to this VB6 string formatting?
I'm converting an ancient VB6 program to C# and I came across some VB6 code that looked like this . . . 



```
   Format(part(PI).far_xdev, "" #0.000;-#0.000"")

```

at first I was confused about the two format fields separated by a semicolon. But it turns out that in VB6 this means that it uses the first one if the value of the number being formatted is 0 or positive, and the second if it's negative. If there were three formatting fields it would be positive, negative, and zero; four would be positive, negative, zero and null.


What's the equivalent of this in C# string formatting? How do I say ""use this formating string for a positive number and that one for a negative number""?


(To whoever added the ""This question may already have an answer"": the problem with that link is that the linked question was asked in reference to some version of BASIC (based on the syntax) and did not explicitly say he was looking for an answer in C#, and neither of the two answers given specifically say they are in C#. We are left to surmise the languages involved based only on the tags. I think this new question and the resulting answers are much more clear, explicit and detailed) 


","Use the same separation by semi-column. Read more about that separator at [msdn](http://msdn.microsoft.com/en-us/library/0c899ak8(v=vs.110).aspx#SectionSeparator) (supports up to three sections)



```
Console.WriteLine(""{0:positive;negative;zero}"", +1); //prints positive
Console.WriteLine(""{0:positive;negative;zero}"", -1); //prints negative
Console.WriteLine(""{0:positive;negative;zero}"", -0); //prints zero

```

You can use `ToString` on numeric value and pass format there



```
string formatted = 1.ToString(""positive;negative;zero""); //will return ""positive""

```

or use `string.Format` as shown in the comment section. But still you need to pass order position `{0}` to it.



```
string formatted = string.Format(""{0:positive;negative;zero}"", 1); 

```

In order to check for null, you can use null coalescing operator (cast to object is required, since there is no implicit cast from `int?` to `string`). It becomes quite messy, so I would recommend to consider simple if statement.



```
int? v = null;
var formatted = string.Format(""{0:positive;negative;zero}"", (object) v ?? ""null"");

```

"
"Applying CSS rules based on input checkbox status
I'm currently developing a site that uses the :checked variable to apply rules to the CSS. I've got a checkbox input that when triggered, should accomplish two things. The first thing it should do is expand a div from a height of 0 to 200 when checked, and back to 0 when unchecked. I had no problems with triggering that first task. The second thing that *should* be accomplished upon checking the box, is a transform: rotate(45deg) of a div with a ""**+""** in it that should rotate 45 degrees into an ""**x**"" (not an actual x, but a rotated +).


I've currently got my code setup to display the animation on :hover, but that's just for illustrative purposes, that wouldn't be in my final code. So hover over the ""**+** to see what I'm trying to accomplish with the :checked input.


If you're willing to take a look at my [code](http://codepen.io/mtthwbckmn/pen/Actwn), and tell me what I'm doing wrong, I'd be greatly appreciative! Let me know if you have any questions.


**Note: Ideally I'm looking for a pure CSS solution without the need for JS. Let me know if this isn't possible.**


[Here's my code pen.](http://codepen.io/mtthwbckmn/pen/Actwn)


","I wrote a similar solution the other day, [here](https://stackoverflow.com/questions/21272354/css-selector-doing-something-wrong/21272419#21272419).


Basically, you are limited when using the `:checked` method. You are relying on the [adjacent](http://www.w3.org/TR/css3-selectors/#adjacent-sibling-combinators) and [general sibling combinators](http://www.w3.org/TR/css3-selectors/#general-sibling-combinators), [`+`](http://www.w3.org/TR/css3-selectors/#adjacent-sibling-combinators), [`~`](http://www.w3.org/TR/css3-selectors/#general-sibling-combinators). If the element isn't a general preceding sibling, it isn't going to work.


In this example, `.expand` was not a preceding sibling. Therefore the solution is to place the `input` element at the root of the document, and then use the selector `input[name='panel']:checked ~ label .rotate` to change the `.rotate` element. Note, that the general sibling combinator, `~` is now also being used as opposed to the adjacent sibling combinator, `+`.


No need for JS - [**UPDATED EXAMPLE**](http://codepen.io/anon/pen/lFmzj)


**Modified HTML:**



```
<input type=""checkbox"" name=""panel"" class=""hidden"" id=""panel""/>
<label for=""panel"">
  Click Me
  <div class=""rotate"">+</div>
</label>
<div class=""expand"">
  Content goes here.
</div>

```

**Updated CSS:**



```
input[name='panel']:checked ~ label .rotate {
    transform: rotate(45deg);
    -o-transform: rotate(45deg);
    -ms-transform: rotate(45deg);
    -webkit-transform: rotate(45deg);
}

```

It's worth noting that I moved the transition properties to the `.rotate` element too.


"
"Gnome3+: How do I remove favorites from Dash via terminal?
I'm guessing I need to edit one of the schemas available in `gsettings` but I don't know which one. and when I listed all the schemas, there's just too many of them.


","The key you want is `favorite-apps`, the schema ID is `org.gnome.shell`. Now to list your favorite apps you can simply run



```
gsettings get org.gnome.shell favorite-apps

```

or



```
dconf read /org/gnome/shell/favorite-apps

```

These will return an array of strings e.g.



```
['firefox.desktop', 'org.gnome.Terminal.desktop', 'org.gnome.Nautilus.desktop', 'org.gnome.gedit.desktop', 'gnome-calculator.desktop']

```

Now, to remove a value from that array you could use text processing tools like `sed`/`awk` to check if an item is in that list and remove it keeping the same format (not that trivial though definitely doable) and once you get it right just write the new settings to the database e.g. assuming you wanted to remove `org.gnome.Nautilus.desktop` you would run (note the double quotes):



```
gsettings set org.gnome.shell favorite-apps ""['firefox.desktop', 'org.gnome.Terminal.desktop', 'org.gnome.gedit.desktop', 'gnome-calculator.desktop']""

```

or



```
dconf write /org/gnome/shell/favorite-apps ""['firefox.desktop', 'org.gnome.Terminal.desktop', 'org.gnome.gedit.desktop', 'gnome-calculator.desktop']""

```

Still, it's easier to write your own utility ([using `gsettings` API](https://developer.gnome.org/gio/stable/GSettings.html)) that will accept one or more desktop file names as positional parameters and remove them from favorites; to get you started, here is a very basic example in `python` that accepts one param (run as `script.py firefox.desktop`):



```
#!/usr/bin/env python

from sys import argv
from gi.repository import Gio,GLib
item=argv[1]
gschema = Gio.Settings('org.gnome.shell')
gvalues=gschema.get_value('favorite-apps').unpack()
if item in gvalues: gvalues.remove(item)
gschema.set_value('favorite-apps', GLib.Variant('as', gvalues))

```

"
"How should I reason when I have to choose between a class, struct and enum in Swift?
Since classes, structs and enums all has constructors, properties and computed properties, how should I reason when choosing between them?


","ChristopheD's and Jack Wu's answers are good, but I feel they don't touch on enums, or miss their importance. Swift enums are (meant to be) a full implementation of algebraic data types. Classes and structs are traditionally used to model data in object-oriented languages, but enums are usually limited to being used as a convenient way to limit the value of a variable to a limited number of possibilities. E.g. (C++):



```
enum MaritalStatus { Unmarried, Married, Divorced, WidowedOrWidowered };
MaritalStatus m = Unmarried;

```

Swift enums can do the above but they can do a lot more. Of course the Language Guide has a pretty good [barcode modelling example](https://developer.apple.com/library/prerelease/ios/documentation/Swift/Conceptual/Swift_Programming_Language/Enumerations.html#//apple_ref/doc/uid/TP40014097-CH12-XID_189) but the best example I know of that really drives home the point of modelling data with algebraic data types is Scott Wlaschin's presentation: <http://www.slideshare.net/ScottWlaschin/ddd-with-fsharptypesystemlondonndc2013>


You would probably benefit from going through the whole presentation but really to 'get' the point all you need to see is slide 60 where he shows how to model a 'payment method' in a typical line-of-business app.


The examples in the presentation are in F# but F# isn't *that* far off from Swift and you can pretty easily map between them. E.g., the payment methods enum in Swift would look like:



```
enum PaymentMethod {
    case cash // No extra data needed.
    case cheque(Int) // Cheque #.
    case card(CardType, CardNumber) // 2 pieces of extra data.
}

```

The point of the above is that each order's payment method can be only one of the above three methods. Anything else will not be allowed by the compiler. This is a very succinct alternative to building entire class hierarchies to model these almost trivial things.


The presentation really takes off from there and the best part is Swift can do almost everything that F# can in terms of data modelling, using optional types, etc.


"
"window.open returns undefined in chrome extension
I have content script based Chrome extension. I initiate the sign in process through a popup window in the content script. 


I open a popup window using the below code and then wait till its closed.


However, I get an 'undefined' from `window.open` method. Does anybody know why this happens?


`loginwin` is `undefined` in below code although the popup window opens up fine with the specified `login_url`. The code below is called from my content script.



```
var loginWin = window.open(login_url, 'LoginWindow', ""width=655,height=490"");
console.log(loginWin);
// Check every 100 ms if the popup is closed.
var finishedInterval = setInterval(function() {
    console.log('checking if loginWin closed');
    if (loginWin.closed) {
        clearInterval(finishedInterval);
        console.log('popup is now closed');
        Backbone.history.navigate('index', true);
    }
}, 1000);

```

","
> 
> Note: This answer is obsolete. `window.open()` in a Chrome extension always returns either `null` (when the popup is blocked) or a `window` object. The information below only applies to very old (2012) versions of Chrome.
> 
> 
> 




---


Content scripts do not have any access to a page's global `window` object. For content scripts, the following applies:


- The **`window`** variable does not refer to the page's global object. Instead, it refers to a new context, a ""layer"" over the page. The page's DOM is fully accessible. [#execution-environment](http://code.google.com/chrome/extensions/content_scripts.html#execution-environment)


Given a document consisting of   `<iframe id=""frameName"" src=""http://domain/""></iframe>`:


- Access to the contents of a frame is restricted by the [Same origin policy](https://developer.mozilla.org/En/Same_origin_policy_for_JavaScript) of the page; the permissions of your extension does not relax the policy.
- `frames[0]` and `frames['frameName']`, (normally referring to the the frame's containing global `window` object) is **`undefined`**.
- `var iframe = document.getElementById('frameName');`
	- **`iframe.contentDocument`** returns a **`document`** object of the containing frame, because content scripts have access to the DOM of a page. This property is **`null`** when the Same origin policy applies.
	- **`iframe.contentDocument.defaultView`** (refers to the `window` object associated with the document) is **undefined**.
	- **`iframe.contentWindow`** is **undefined**.


As you can see, `window.open()` does not return a `Window` instance (neither does `window.opener`, and so forth).




---


### Alternatives


- [Inject the code in the page](https://stackoverflow.com/questions/9515704/building-a-chrome-extension-inject-code-in-a-page-using-a-content-script/9517879#9517879), so that it runs in the context of the page. Note: Only use this method if the page you're operating on can be trusted. To communicate between the injected script and the content script, you could use:



```
var login_url = 'http://example.com/';
var event_name = 'robwuniq' + Math.random().toString(16); // Unique name
document.addEventListener(event_name, function localName() {
    document.removeEventListener(event_name, localName); // Clean-up
    // Your logic:
    Backbone.history.navigate('index', true);
});
// Method 2b: Inject code which runs in the context of the page
var actualCode = '(' + function(login_url, event_name) {
    var loginWin = window.open(login_url, 'LoginWindow', ""width=655,height=490"");
    console.log(loginWin);
    // Check every 100 ms if the popup is closed.
    var finishedInterval = setInterval(function() {
        console.log('checking if loginWin closed');
        if (loginWin.closed) {
            clearInterval(finishedInterval);
            console.log('popup is now closed');
            // Notify content script
            var event = document.createEvent('Events');
            event.initEvent(event_name, false, false);
            document.dispatchEvent(event);
        }
    }, 1000);
} + ')(' + JSON.stringify(login_url+'') + ', ""' + event_name + '"")';
var script = document.createElement('script');
script.textContent = actualCode;
(document.head||document.documentElement).appendChild(script);
script.parentNode.removeChild(script);

```
- Launch the window from the background page using `window.open()`. This returns a `window` object which has a reliable `closed` property. See the next bullet point for more details on the communication flow.
- From the content script, [pass a message](http://code.google.com/chrome/extensions/messaging.html) to the [background page](http://code.google.com/chrome/extensions/background_pages.html). In the background page, use [`chrome.windows.create`](http://code.google.com/chrome/extensions/windows.html#method-create) to open a window. In the callback, assign an [`chrome.tabs.onRemoved`](http://code.google.com/chrome/extensions/tabs.html#event-onRemoved) and/or [`chrome.tabs.onUpdated`](http://code.google.com/chrome/extensions/tabs.html#event-onUpdated) event. When these event listeners are triggered, they should remove themselves, and notify the original caller (content script) using the `sendResponse` function of `chrome.extension.onMessage`.


"
"Connect two docker containers
I have two containers, the first one with a `django` and the second one with `postgresql`.


Well, in my first server I have running `django` and I'm trying to connect it with the second one.
The second container has the port `32770` exposed but internally running in the port `5432`.
In my local machine, I have the connection:
Server: 'Localhost'
Port: 32770
User: 'myuser'
Password: ''


And it's connecting, but with my `django` container, I'm getting this error:



```
could not connect to server: Connection refused
    Is the server running on host ""localhost"" (127.0.0.1) and accepting
    TCP/IP connections on port 32770?

```

The same happens for the port `5432`


How I can connect both servers?


","Since your are running the containers individually you have different options


**Run django on network of postgres container**



```
$ docker run -d ... postgres
<container id>

$ docker run -d ... --net container:<containerid> django

```

Then django can find postgres on `localhost:5432`


**Run django and postgres as named containers container**



```
$ docker run --name postgresdb -d ... postgres
<container id>

$ docker run -d ...  django

```

Now django can find db on `postgresdb:5432`


**Run both containers on host**



```
$ docker run --net host -d ... postgres
<container id>

$ docker run -d ... --net host django

```

Then django can find postgres on `localhost:5432`


**Run containers on same network**



```
$ docker network create mynet

$ docker run --name postgresdb --net mynet -d ... postgres
<container id>

$ docker run --net mynet -d ...  django

```

Now django can find db on `postgresdb:5432`


**Connect to the host IP and mapped port**



```
$ docker run -d -p 32770:5432 .... postgres

$ docker run -d .... django

```

Django can now connect to the DB on `<yourdockerhostip>:32770`. 


Better option is to run it using `docker-compose`. Learn more on <https://docs.docker.com/compose/>


"
"How do I maintain image size when using a colorbar?
I'm trying to plot two versions of the same image side by side. When I plot the figure without the color bar for one of the images, it seems to have the right sizes:


![without the color bar](https://i.stack.imgur.com/xfzWT.png)


But when I add a color bar to the image in the left, it scales the image down somehow:


![scales the image down](https://i.stack.imgur.com/IdhF8.png)


Here's the code where I have commented out the lines for color bar:



```
def plot_amaps(self, anisotropy_map, parallel):
        timepoint = self.t * self.timestep
        amap_directory = self.directory + ""amaps/""
        fig = plt.figure(facecolor='w', dpi=180)

        ax1 = fig.add_subplot(121)
        fig.subplots_adjust(top=0.85)
        ax1.grid(False)
        txt = ""Mean(r) = %.3f SD(r)= %.3f t=%dmin""
        txt = txt %(self.mean, self.sd, timepoint)
        ax1.set_title(txt)

        amap = ax1.imshow(anisotropy_map, cmap=""jet"", clim = self.clim)
        #divider = make_axes_locatable(ax1)
        #cax = divider.append_axes('right', size='5%', pad=0.05)
        #fig.colorbar(amap, cax=cax)

        ax2 = fig.add_subplot(122)
        ax2.set_title(""Intensity image"", fontsize=10)
        ax2.imshow(parallel, cmap=""gray"")
        ax2.grid(False)
        ax1.axis('off')
        ax2.axis('off')

        if self.save is True:
            self.make_plot_dir(amap_directory)
            name = self.cell + ""_time_""+str(timepoint)
            plt.savefig(amap_directory+name+self.saveformat, bbox_inches='tight')
        else:
            plt.show()
        plt.close('all')

```

What am I doing wrong, and how can I make sure that the two images are of the same size?


","When using 



```
divider = make_axes_locatable(ax1)
cax = divider.append_axes('right', size='5%', pad=0.05)

```

you explicitely ask for a 5% smaller axes. So if you don't want that, you should not create the axes for the colorbar using `make_axes_locatable`.


Instead you can simply create an axes at any point on the figure using



```
cax = fig.add_axes([left, bottom, width, height])

```

where `left, bottom, width, height` are in figure units ranging from 0 to 1. Then add the colorbar to it.  

If you want the colorbar in the middle, you could previously make some space using 



```
plt.subplots_adjust(wspace=0.3)

```

Of course you would have to experiment a bit with the numbers. 


"
"Confusing articles and documentation about the differences (if any) between System.Data.EntityState.Add & DbSet.Add
I am working on a C# ASP.NET MVC 5 web application with EF 5. Mapping of my database tables using EF generates a `DbContext` class and an `.edmx` file. Today, I was reading [a great article about creating generic DAL classes](https://blog.magnusmontin.net/2013/05/30/generic-dal-using-entity-framework/), but I stopped on the following sentence:



> 
> Note that using the Entry method to change the state of an entity will
>  only affect the actual entity that you pass in to the method. It won’t
>  cascade through a graph and set the state of all related objects,
>  unlike the DbSet.Add method.
> 
> 
> 


That contradicts what is mentioned in these questions:


- <http://forums.asp.net/p/2015170/5803192.aspx>
- <http://forums.asp.net/p/2060606/5943259.aspx>
- [Difference between DbSet.Add(entity) and entity.State = EntityState.Added](https://stackoverflow.com/questions/13548243/difference-between-dbset-addentity-and-entity-state-entitystate-added)
- [What is the difference between IDbSet.Add and DbEntityEntry.State = EntityState.Added?](https://stackoverflow.com/questions/9082922/what-is-the-difference-between-idbset-add-and-dbentityentry-state-entitystate)


In all the above questions’ answers, all users mentioned that using `System.Data.EntityState.Added` is exactly the same as using `DbSet.Add`. But the article I mentioned first states that using `System.Data.EntityState.Added` will not cascade through the graph.


Based on my test, I conclude that using `System.Data.EntityState.Added` will cascade through the graph same as in the `DBset.Add` case. Is the article wrong, or is it my test and the Q&A?


","Those methods are the same which you can verify by regular testing, or, if you want to be completely sure - by some exploration of EF 6 code.


1. `DbSet.Add` method (<http://entityframework.codeplex.com/SourceControl/latest#src/EntityFramework/DbSet.cs>)



```
public virtual TEntity Add(TEntity entity)
{
    Check.NotNull<TEntity>(entity, ""entity"");
    this.GetInternalSetWithCheck(""Add"").Add((object) entity);
    return entity;
}

```


This calls `InternalSet<T>.Add(object)` method.


2. `DbEntityEntry<T>.State` property (<http://entityframework.codeplex.com/SourceControl/latest#src/EntityFramework/Infrastructure/DbEntityEntry.cs>)



```
public EntityState State
{
    get { return _internalEntityEntry.State; }
    set { _internalEntityEntry.State = value; }
}

```


Where `_internalEntityEntry` is of `InternalEntityEntry` type.


`InternalEntityEntry.State` property (<http://entityframework.codeplex.com/SourceControl/latest#src/EntityFramework/Internal/EntityEntries/InternalEntityEntry.cs>)



```
    public virtual EntityState State
    {
        get { return IsDetached ? EntityState.Detached : _stateEntry.State; }
        set
        {
            if (!IsDetached)
            {
                if (_stateEntry.State == EntityState.Modified
                    && value == EntityState.Unchanged)
                {
                    // Special case modified to unchanged to be ""reject changes"" even
                    // ChangeState will do ""accept changes"".  This keeps the behavior consistent with
                    // setting modified to false at the property level (once that is supported).
                    CurrentValues.SetValues(OriginalValues);
                }
                _stateEntry.ChangeState(value);
            }
            else
            {
                switch (value)
                {
                    case EntityState.Added:
                        _internalContext.Set(_entityType).InternalSet.Add(_entity);
                        break;
                    case EntityState.Unchanged:
                        _internalContext.Set(_entityType).InternalSet.Attach(_entity);
                        break;
                    case EntityState.Modified:
                    case EntityState.Deleted:
                        _internalContext.Set(_entityType).InternalSet.Attach(_entity);
                        _stateEntry = _internalContext.GetStateEntry(_entity);
                        Debug.Assert(_stateEntry != null, ""_stateEntry should not be null after Attach."");
                        _stateEntry.ChangeState(value);
                        break;
                }
            }
        }
    }

```

You see that if entity is detached (your case) and state is Added - the same `InternalSet<T>.Add(object)` is called.


As for verification by testing:



```
using (var ctx = new TestDBEntities()) {
    // just some entity, details does not matter
    var code = new Code();
    // another entity
    var error = new Error();
    // Code has a collection of Errors
    code.Errors.Add(error);
    var codeEntry = ctx.Entry(code);
    // modify code entry and mark as added
    codeEntry.State = EntityState.Added;
    // note we did not do anything with Error
    var errorEntry = ctx.Entry(error);
    // but it is marked as Added too, because when marking Code as Added -
    // navigation properties were also explored and attached, just like when
    // you do DbSet.Add
    Debug.Assert(errorEntry.State == EntityState.Added);                
}

```

"
"Check if all elements of a list are of the same type
How can I check if the elements of a list are of the same type, without checking individually every element if possible?


For example, I would like to have a function to check that every element of this list is an integer (which is clearly false):



```
x = [1, 2.5, 'a']

def checkIntegers(x):
    # return True if all elements are integers, False otherwise

```

","Try using [`all`](http://docs.python.org/2/library/functions.html#all) in conjunction with [`isinstance`](http://docs.python.org/2/library/functions.html#isinstance):



```
all(isinstance(x, int) for x in lst)

```

You can even check for multiple types with `isinstance` if that is desireable:



```
all(isinstance(x, (int, long)) for x in lst)

```



---


Not that this will pick up inherited classes as well. e.g.:



```
class MyInt(int):
     pass

print(isinstance(MyInt('3'),int)) #True

```

If you *need* to restrict yourself to just integers, you could use `all(type(x) is int for x in lst)`. But that is a **VERY** rare scenario.




---


A fun function you could write with this is one which would return the type of the first element in a sequence if all the other elements are the same type:



```
def homogeneous_type(seq):
    iseq = iter(seq)
    first_type = type(next(iseq))
    return first_type if all( (type(x) is first_type) for x in iseq ) else False

```

This will work for any arbitrary iterable, but it will consume ""iterators"" in the process. 


Another fun function in the same vein which returns the set of common bases:



```
import inspect
def common_bases(seq):
    iseq = iter(seq)
    bases = set(inspect.getmro(type(next(iseq))))
    for item in iseq:
        bases = bases.intersection(inspect.getmro(type(item)))
        if not bases:
           break
    return bases

```



---


"
"Make nav dropdown show above all other divs
I have a page I'm working on where currently I have 2 items. Item 1 is a flexnav jQuery navigation menu with a dropdown. Item 2 is a slick jQuery div scroller. I am trying to position the slick scroller just below the flexnav menu. The problem I'm running into though is when you hover over one of the menu items the dropdown for the sub menu is covered up by the slick scroller divs. This only seems to be happening with a screen larger than 800px as the flexnav plugin changes to a mobile friendly navigation menu on small screens. 


I have tried changing the css position setting of both items but I just can't seem to figure out how to make the dropdown menus appear above the slick divs. Does anyone know what I'm doing wrong here or have any suggestions on how I could change things around to achieve what I am looking for? 


[Here is a example JSFiddle](http://jsfiddle.net/austin350s10/8b4yh3de/) 


The code I am using:



```
<header>
    <nav style=""background-color: #FAD10E; height:50px"">
    <div class=""menu-button"">Mobile Menu</div>
        <ul class=""flexnav"" data-breakpoint=""800"">
            <li><a href="""">Home</a></li>
            <li><a href="""">Stuff</a>

            <!-- THIS DROPDOWN IS COVERED BY THE AUTOPLAY DIV -->
              <ul>
                <li><a href="""">Stuff 1</a></li>
                <li><a href="""">Stuff 2</a></li>
                <li><a href="""">Stuff 3</a></li>
                <li><a href="""">Stuff 4</a></li>
                <li><a href="""">Stuff 5</a></li>
                <li><a href="""">Stuff 6</a></li>
              </ul>
            </li>
            <li><a href=""/"">Stuff 2</a></li>
            <li><a href="""">Stuff 3</a></li>
            <li><a href="""">Stuff 4</a></li>
            <li><a href="""">Stuff 5</a></li>
        </ul>
    </nav>
</header>


<div>
    <!-- THIS AUTOPLAY DIV SHOWS ON TOP OF THE MENU DROPDOWN ITEMS -->
    <div class=""autoplay"">
        <div><img src=""http://www.affordablehomecare.org/assets/images/fade/happy-home-care-client.jpg""></div>
        <div><img src=""http://www.affordablehomecare.org/assets/images/fade/helping-hands-home-care.jpg""></div>
        <div><img src=""http://www.affordablehomecare.org/assets/images/fade/loving-home-care-client.jpg""></div>
    </div>
</div>

```

","You only need to add two lines of CSS


# [Example fiddle](http://jsfiddle.net/8b4yh3de/1/)


**CSS**



```
.flexnav{
    -webkit-padding-start: 0px;
    -webkit-margin-before: 0px;
    -webkit-margin-after: 0px;
    margin-top: 0px;
    margin-right: auto;
    margin-bottom: 0px;
    margin-left: auto;
    width:90%;
    position: relative; /* <-- Added */
    z-index: 1; /* <-- Added */
}

```

The `position: relative` allows for the element to have a z-index applied (an element must not be statically positioned, relative positioning will allow the element to display in normal document flow without having a `static` positioning).


The `z-index: 1` provides a separate stacking context for the nav. Otherwise, because it precedes your carousel in document flow, will necessarily display beneath it when overlapped without a `z-index` given.


Stacking contexts apply *generally* only to elements which sit at the same hierarchical depth. So putting the flyout in your nav with a higher `z-index`won't work.


"
"Alternative for the ImageCapture API for better browser support
I use an OCR in order to recognize text from images. I take photos/snapshots from the user's camera/video input using the [ImageCapture API](https://developer.mozilla.org/en-US/docs/Web/API/ImageCapture) from a [MediaStream](https://developer.mozilla.org/en-US/docs/Web/API/MediaStream). Here is the code responsible for that:



```
function getBlobFromMediaStream() {
    const videoTrack = mediaStream.getVideoTracks()[0]

    const imageCapture = new ImageCapture(videoTrack);

    return imageCapture.takePhoto().then(blob => {
        if (!blob) throw ""Photo could not be taken"";

        return blob;
    })
}

```

Too bad that this API doesn't work on a lot of browsers (Firefox, IE, Safari). Is there an alternative that can be used?


","You could make a fallback with a `HTMLVideoElement` and a `HTMLCanvasElement`.  

First check if your browser supports the MediaStream Image Capture API by checking if the constructor exists in the window object.



```
if ('ImageCapture' in window) { 
  // Has support for API.
} else {
  // No support, use fallback.
}

```

Then based on that either use the API or use your fallback.


In the fallback create a video element, canvas element and a canvas context. Set the `MediaStreamTrack` as the `srcObject` of the video element. The video is able to decode the data and make an actual picture.


Use the image of the video to draw an image to the canvas element with the [`CanvasRenderingContext2D.drawImage()`](https://developer.mozilla.org/en-US/docs/Web/API/CanvasRenderingContext2D/drawImage) method. With this the canvas element will draw an image of the current frame in the video.


Now you can extract the data that has been drawn on the canvas and turn it into a `Blob` with the [`HTMLCanvasElement.toBlob()`](https://developer.mozilla.org/en-US/docs/Web/API/HTMLCanvasElement/toBlob) method.


**Wrap the function call inside a `Promise` for two reasons:**


1. You'll need to extract the blob from the callback, a Promise wrapper can help you do this.
2. The `ImageCapture.takePhoto()` method also returns a Promise. So you'll have the same behavior as you would have when the API is supported.


All put together it would look like this.



```
function getBlobFromMediaStream(stream) {

  if ('ImageCapture' in window) {

    const videoTrack = stream.getVideoTracks()[0];
    const imageCapture = new ImageCapture(videoTrack);
    return imageCapture.takePhoto();
    
  } else {

    const video = document.createElement('video');
    const canvas = document.createElement('canvas');
    const context = canvas.getContext('2d');

    video.srcObject = stream;

    return new Promise((resolve, reject) => {
      video.addEventListener('loadeddata', async () => {
        const { videoWidth, videoHeight } = video;
        canvas.width = videoWidth;
        canvas.height = videoHeight;

        try {
          await video.play();
          context.drawImage(video, 0, 0, videoWidth, videoHeight);
          canvas.toBlob(resolve, 'image/png');
        } catch (error) {
          reject(error);
        }
      });
    });

  }
  
}

```

Call your function like in the example below. I've discarded the `throw` statement so that all *rejected* states will be caught at the same point when you call the `getBlobFromMediaStream` function.


According to the docs of [`ImageCapture.takePhoto()`](https://developer.mozilla.org/en-US/docs/Web/API/ImageCapture/takePhoto), you either get a `Blob` when the promise has fulfilled and an error when it has rejected.


Same goes for the fallback. The only difference there is that the `HTMLCanvasElement.toBlob()` function throws a `SecurityError` if the bitmap of the canvas is not origin clean.



```
getBlobFromMediaStream(mediaStream).then(blob => {
  // Work with your blob.
}).catch(error => {
  console.log(`Photo could not be taken. ${error}`);
})

```

"
"Measure of power other than Turing completeness
I originally tried asking this on StackOverflow, but it was too subjective :-(. I am interested in methods of defining the power of programming languages. Turing completeness is one, but it is almost universally satisfied. What would be nice is to define a measure of power that discriminates among programming languages that are actually in used. For example, can anyone propose a non-subjective method that would discriminate between assembly and Java?


Turing completeness means that a language is maximally powerful in what it can output (which pretty much means it can do anything non-time based in the real world). So if we want to define a stronger measure of power, we need to take another approach. Shortness was suggested in the original question, but this is not easy to define at all. Does anyone have any other suggestions?


","The notion you are looking for is called **expressiveness** and Matthias Felleisen has a mathematically rigorous definition:



> 
> ""**[On the Expressive Power of Programming Languages](http://dx.doi.org/10.1016/0167-6423(91)90036-W)**""
> 
> 
> www.ccs.neu.edu/scheme/pubs/scp91-felleisen.ps.gz (Postscript version)
> 
> 
> 


The intuition behind the idea is that if you have two equivalent programs in two different languages-- say, program A in language X and program B in language Y-- and if you make a local change to A that requires a global change to B, then X is more expressive than Y.


One example Felleisen provides is assignment: In the Scheme programming languages you can remove the assignment operator and still have a Turing complete language. However, in such a restricted language, adding in a feature that would be localized if assignment was allowed would require a global change to the program without assignment.


My discussion has simplified some details, and you should read the paper itself for the full account.


To answer your other question: You can say that Java is more expressive than assembly because you can add a new class to your Java program, and then gain the benefits of polymorphism by having other parts of your program call its methods without global modification. Exception handling is another example where Java is more expressive than assembly: You simply need to write a single `throw` statement to transfer control up the stack. On a more elementary level, you can also add a new `case` statement near the beginning of a `switch` and you won't have to worry about recalculating any jump offsets by hand.


"
"jsTree - Last selected node highlighting issue
I'm using cookies plugin in my jsTree in order to remember and restore the last node the user clicked on.


The behavior of my tree is that on node's double-click i'm redirecting the client to appropriate page in my site.


I'm expecting the last double-clicked node to be highlighted after I was redirected but for some reason I must do the double-click again on the same node in order it to be highlighted, **it's seems like it's highlighting the previous selected node instead of the last selected node.**


Any idea?


","We're probably having the same problem. Add the following code to your jstree `cookies` config:


`cookies: { cookie_options: { path: '/' } }`


so you'll have a tree config similar to this:



```
$('.my-tree').jstree({
            plugins: [""html_data"", ""ui"", ""cookies""]
            cookies: { cookie_options: { path: '/' } }
    })

```

### What's going on?


The reason for this is that if [jQuery cookies](http://plugins.jquery.com/project/Cookie) (the library jstree uses to write its cookies) don't have their path set, the cookies will be stored relative to the current page you're on.


So, for example, if you have two nodes (also leaves/needles):


- `http://localhost/Home/` with node id `localhost_home`
- `http://localhost/Admin/` with node id `localhost_admin`


and you want to navigate from `http://localhost/Home/` to `http://localhost/Admin/` by clicking on the `localhost_admin` node, you're probably expecting that by the time you get to `http://localhost/Admin/` the `jstree_select` cookie will have `localhost_admin` selected right?


What really happened here is that 2 separate cookies now exist. One for `http://localhost/Home/` and another for `http://localhost/Admin/` containing 2 separate values.


By setting the path option to `/`, we're ensuring that only 1 cookie gets shared across all pages.


"
"Train Tensorflow model with estimator (from\_generator)
I am trying train an estimator with a generator, but I want to feed this estimator with a package of samples for each iteration. I show the code:



```
def _generator():
for i in range(100):
    feats  = np.random.rand(4,2)
    labels = np.random.rand(4,1)

    yield feats, labels


def input_func_gen():
    shapes = ((4,2),(4,1))
    dataset = tf.data.Dataset.from_generator(generator=_generator,
                                         output_types=(tf.float32, tf.float32),
                                         output_shapes=shapes)
dataset = dataset.batch(4)
# dataset = dataset.repeat(20)
iterator = dataset.make_one_shot_iterator()
features_tensors, labels = iterator.get_next()
features = {'x': features_tensors}
return features, labels


x_col = tf.feature_column.numeric_column(key='x', shape=(4,2))
es = tf.estimator.LinearRegressor(feature_columns=[x_col],model_dir=tf_data)
es = es.train(input_fn=input_func_gen,steps = None)

```

When I run this code, it raises this error:



```
    raise ValueError(err.message)
ValueError: Dimensions must be equal, but are 2 and 3 for 'linear/head/labels/assert_equal/Equal' (op: 'Equal') with input shapes: [2], [3].

```

How do I have to call to this structure??


thx!!!


","The batch size is automatically computed and added to the tensors shapes by Tensorflow, so it doesn't have to be done manually. Your generator should also be defined to output single samples.


Assuming the `4` in position 0 of your shapes are for the batch size, then:



```
import tensorflow as tf
import numpy

def _generator():
    for i in range(100):
        feats  = numpy.random.rand(2)
        labels = numpy.random.rand(1)

        yield feats, labels


def input_func_gen():
    shapes = ((2),(1))
    dataset = tf.data.Dataset.from_generator(generator=_generator,
                                         output_types=(tf.float32, tf.float32),
                                         output_shapes=shapes)
    dataset = dataset.batch(4)
    # dataset = dataset.repeat(20)
    iterator = dataset.make_one_shot_iterator()
    features_tensors, labels = iterator.get_next()
    features = {'x': features_tensors}
    return features, labels


x_col = tf.feature_column.numeric_column(key='x', shape=(2))
es = tf.estimator.LinearRegressor(feature_columns=[x_col])
es = es.train(input_fn=input_func_gen,steps = None)

```

"
"Passing an anonymous variable by reference
Standard C++ types such as int or char have ctors, so you can have expressions like:



```
int a = int(67); // create anonymous variable and assing it to variable a
int b(13);       // initialize variable b
int(77);         // create anonymous variable

```

User defined types (structures or classes) are able to do the same:



```
struct STRUCT
{
  STRUCT(int a){}
};

STRUCT c = STRUCT(67);
STRUCT d(13);
STRUCT(77);

```

The question is: why can we pass by a reference anonymous structure or class instances, but can not pass standard types?



```
struct STRUCT
{
  STRUCT(int a){}
};

void func1(int& i){}
void func2(STRUCT& s){}
void func3(int i){}
void func4(STRUCT s){}

void main()
{
  //func1(int(56));  // ERROR: C2664
  func2(STRUCT(65)); // OK: anonymous object is created then assigned to a reference
  func3(int(46));    // OK: anonymous int is created then assigned to a parameter
  func4(STRUCT(12)); // OK: anonymous object is created then assigned to a parameter
}

```

","If your compiler allows this, then it's not a standard compatible C++ compiler. You can not bind a temporary rvalue to a non-const lvalue reference. It's the rule. Both [clang](http://coliru.stacked-crooked.com/a/acfdf47d2edf574f) and [gcc](http://coliru.stacked-crooked.com/a/73db08a665558e1b) don't compile that code for `func2(STRUCT(65));`.


Instead you have alternatives:



```
void func1(int&& i){}

void func1(const int& i){}

```

Legacy from C++03: A (lvalue) reference to a non-const type (`int &i`) supposed to able to change the parameter then passing a temporary object such as `56` is not logical because it not changeable. A reference to a const type (`const int &i`) supposed to just observe the value as read-only, then passing a temporary value such as `52` is legal.


In C++11 you can reference to a non-const temporary object by `&&`.


"
"React / Redux: Should containers have any knowledge of state structure?
This [tutorial](https://egghead.io/lessons/javascript-redux-colocating-selectors-with-reducers) by Dan Abramov suggests that the advantage to using selectors that act on global state (rather than a slice of state) is that they allow containers to be decoupled from knowledge of the state structure.


If that's the case, shouldn't we also avoid directly mapping state values to props, and use selectors instead? Otherwise, our containers must still know about where those values exist in the state tree.


To illustrate using an example...


*Directly maps a nested state value to a prop:*



```
const mapStateToProps = (state) => ({
  isModalVisible: state.modal.isVisible,
});

```

**VS**


*Has no knowledge of state structure. Gets value using `isModalVisible()` selector:*



```
const mapStateToProps = (state) => ({
  isModalVisible: isModalVisible(state),
});

```

However, the problem with the latter approach is that for each value in the state tree, we have to write a selector. This seems like a lot of boilerplate code just to select a simple value. Is this considered best practice?


","The answer for your question is: ""it depends""


**Do you have a tiny app?** 


Maybe avoid using redux at all and stick to react components state.


      [**You Might Not Need Redux**](https://medium.com/@dan_abramov/you-might-not-need-redux-be46360cf367)


**Do you have a small app?**


Writing a bunch of selectors is a lot of boilerplate just like you said. Avoid writing them and stick to simply mapping, using `mapStateToProps`.


**Medium to large apps?**


This is where selectors and memoized selectors pay-off. You'll find yourself checking if modals is visible in many components. Currently in your state modal is under `state.modal.visible`. But tomorrow a modal might be a part of a parent modal, you'll have to change all mapping in all your components to `state.parentModal.modal.visible`. You can see how this can turn badly right?


**The pros of selectors:**


- Hides complexity when working in a team.
- Changing your state means **only** modifying your selector function.
- Avoid writing filters and reducing lists on each `mapPropsToState` function.


**Pros of Memoized selectors:**


- Performance.


**Cons**


- Boilerplate code and slower to get started.
- Additional libraries.


Hope it answers your question.


"
"wp\_rewrite in a WordPress Plugin
Ok, I've got this code that I've been using to spit out news to an application of mine. It was working until today. I've cutout all the logic in the following code to make it simpiler. But it should ""WORK"" Can someone help me fix this code to where it works, and is done right? I know it's hacked together , but it didn't seem to have any problems until today. I have not updated anything, don't know what the deal is.



```


 Plugin Name:   MyPlugin Example
 Version:       1.0.1


If ( ! class_exists(""MyPlugin"") )
{
    class MyPlugin
    {
        var $db_version = ""1.0""; //not used yet

        function init()
        {
   //Nothing as of now.
        }
        function activate()
        {
            global $wp_rewrite;
            $this->flush_rewrite_rules();
        }

        function pushoutput( $id )
        {
            $output->out =' The output worked!';
            $this->output( $output );

        }
        function output( $output )
        {
            ob_start();
            ob_end_clean();
            header( 'Cache-Control: no-cache, must-revalidate' );
            header( 'Expires: Mon, 26 Jul 1997 05:00:00 GMT' );
            header( 'Content-type: application/json' );

            echo json_encode( $output );
            //Must encode this...
        }

        function flush_rewrite_rules()
        {
            global $wp_rewrite;
            $wp_rewrite->flush_rules();
        }

        function createRewriteRules( $rewrite )
        {
            global $wp_rewrite;
            $new_rules = array( 'MyPlugin/(.+)' => 'index.php?MyPlugin=' . $wp_rewrite->preg_index(1) );
            if ( ! is_array($wp_rewrite->rules) )
            {
                $wp_rewrite->rules = array();
            }
            $wp_rewrite->rules = $new_rules + $wp_rewrite->rules;
            return $wp_rewrite;
        }


        function add_query_vars( $qvars )
        {
            $qvars[] = 'MyPlugin';
            return $qvars;
        }
        function template_redirect_intercept()
        {
            global $wp_query;
            if ( $wp_query->get('MyPlugin') )
            {
                $id = $wp_query->query_vars['MyPlugin'];
                $this->pushoutput( $id );


                exit;
            }
        }
    }
}
If ( class_exists(""MyPlugin"") )
{
    $MyPluginCode = new MyPlugin();
}
If ( isset($MyPluginCode) )
{
    register_activation_hook( __file__, array($MyPluginCode, 'activate') );
    add_action( 'admin-init', array(&$MyPluginCode, 'flush_rewrite_rules') );
    //add_action( 'init', array(&$MyPluginCode, 'init') );
    add_action( 'generate_rewrite_rules', array(&$MyPluginCode, 'createRewriteRules') );

    add_action( 'template_redirect', array(&$MyPluginCode, 'template_redirect_intercept') );
    // add_filter( 'query_vars', array(&$MyPluginCode, 'add_query_vars') );
}


```

","I changed a little bit of your code in the process, but this worked for me:



```
<?php

/**
* Plugin Name:   MyPlugin Example
* Version:       1.0.1
**/
class MyPlugin {

    function activate() {
        global $wp_rewrite;
        $this->flush_rewrite_rules();
    }

    // Took out the $wp_rewrite->rules replacement so the rewrite rules filter could handle this.
    function create_rewrite_rules($rules) {
        global $wp_rewrite;
        $newRule = array('MyPlugin/(.+)' => 'index.php?MyPlugin='.$wp_rewrite->preg_index(1));
        $newRules = $newRule + $rules;
        return $newRules;
    }

    function add_query_vars($qvars) {
        $qvars[] = 'MyPlugin';
        return $qvars;
    }

    function flush_rewrite_rules() {
        global $wp_rewrite;
        $wp_rewrite->flush_rules();
    }

    function template_redirect_intercept() {
        global $wp_query;
        if ($wp_query->get('MyPlugin')) {
            $this->pushoutput($wp_query->get('MyPlugin'));
            exit;
        }
    }

    function pushoutput($message) {
        $this->output($message);
    }

    function output( $output ) {
        header( 'Cache-Control: no-cache, must-revalidate' );
        header( 'Expires: Mon, 26 Jul 1997 05:00:00 GMT' );

        // Commented to display in browser.
        // header( 'Content-type: application/json' );

        echo json_encode( $output );
    }
}

$MyPluginCode = new MyPlugin();
register_activation_hook( __file__, array($MyPluginCode, 'activate') );

// Using a filter instead of an action to create the rewrite rules.
// Write rules -> Add query vars -> Recalculate rewrite rules
add_filter('rewrite_rules_array', array($MyPluginCode, 'create_rewrite_rules'));
add_filter('query_vars',array($MyPluginCode, 'add_query_vars'));

// Recalculates rewrite rules during admin init to save resourcees.
// Could probably run it once as long as it isn't going to change or check the
// $wp_rewrite rules to see if it's active.
add_filter('admin_init', array($MyPluginCode, 'flush_rewrite_rules'));
add_action( 'template_redirect', array($MyPluginCode, 'template_redirect_intercept') );

```

I've commented the important parts, but what I did was basically to move your hooks over to `use_filter` rather than `add_action`. I also moved the filters into the order that they are actually used in Wordpress. Seemed the thing to do at the time.


Finally, make sure that your permalinks are set to use pretty URLs. I had an issue where mine were set to default, which makes Wordpress ignore any rewrite conditions that it would otherwise need to parse. Change over to some pretty URLs and your conditions will refresh.


Let me know if that works for you. Hope that helps.


Thanks, Joe


"
"calculate vector after rotating it towards another by angle θ in 3D space
[![image](https://i887.photobucket.com/albums/ac80/nicos9pavlou/Untitled_zpsdd4280d3.png)](https://i887.photobucket.com/albums/ac80/nicos9pavlou/Untitled_zpsdd4280d3.png)
  



Unit vectors V and D lie on a 3D space. They have the same starting point. I want to rotate the vector V towards to vector D, but only by angle θ. 


Given that we know:


1. Vectors V and D are unit vectors.
2. We know V(x,y,z) and D(x,y,z).
3. We know the angle between the two vectors V and D, Δφ.
4. We also know the angle θ, where the vector V will rotate to approach vector D.
5. We know the starting point ""O"" where the three vectors start from.


We now want to calculate vector Z, which will also be a unit vector.
Is it possible to calculate the coordinates of vector Z, with this information given above?


Do you have any ideas how this problem can be solved?


","What you need to do is to define a third axis in the plane of the V and D vectors such that a 90 degree rotation from V points in this direction. I'll call the unit vector that points in this direction D'. With this, your Z vector is easy:



```
Z = cos(theta)*V + sin(theta)*D_tick;

```

So how to compute D'? That, too is easy. First compute a vector orthogonal to V and D using the cross product. Call this W: W = V×D. Next compute a vector orthogonal to W and V: D' = W×V = (V×D)×V. This points in the right direction, but it will only be a unit vector if your V and D are orthogonal. So normalize: D' = D'/||D'||, where ||D'|| is the magnitude of the vector D'. If you have a vector math package, you can do this via



```
D_tick = ((V.cross(D)).cross(V)).normalize();

```

One caveat: What if ||D'|| is zero? That happens if and only if your Δφ is a multiple of pi radians (or 180 degrees). Alternatively, it happens when V and D are parallel or anti-parallel to one another. Your question is ill posed in this special case. You should check for this special case.


  





---


**Addendum**  

My (V×D)×V and comingstorm's D - V\*((D·V)/(V·V)) are one and the same for vectors in three dimensional space. Because V is a unit vector, his D-V\*((D·V)/(V·V)) reduces to D-V(D·V). My (V×D)×V is equal to D(V·V)-V(D·V) per the vector triple product identity (<http://mathworld.wolfram.com/VectorTripleProduct.html>), and this reduces to D-V(D·V), again because V is a unit vector.


"
"PHP SharePoint API Document Subdirectory Listings
I've just implemented the SharePointAPI for PHP which can be found <https://github.com/thybag/PHP-SharePoint-Lists-API>


I was able to retrieve all of the SharePoint Lists from the WSDL, and was able to use the ->read($list) function to retrieve a List which holds Document folders and files.


Within the list, if the item is a file, I'm able to download it fine.
However, if the item is another folder (a subdirectory), how do I access that directory's folders and files?


The ->read($list) function doesn't work, and I'm kind of at a loss. Does anybody know how I can read a List item's folders and files as it were a List?


Thanks alot


","Although part of me suspects you are probably the person who ended up managing to answer their own question over on the github issues; I thought I may as well repost that answer here in case others have the same problem.


Essentially (although its a bit of a work around) you can get a listing of the contents of a folder in a sharepoint list, by calling the following method:



```
$sp->read($listName,
          NULL,//$limit 
          NULL,//$query
          NULL,//$view 
          NULL,//$sort 
          ""<Folder>"" . $listName . ""\\"" . $folderName . ""</Folder>""//$options 
)

```

Hopefully once manage to find a bit of spare time, I'll add a dedciated method for doing this to the library, just to make it a little more straight forward in future.


"
"Counting different values using Data.Map leaks memory
The following program uses 100+ MB RAM when counting different line lengths in a 250 MB file. How do I fix it to use less RAM? I suppose I misused lazy IO, `foldr` and laziness of `Data.Map` in values.



```
import Control.Applicative
import qualified Data.Map as M
import Data.List

main = do
  content <- readFile ""output.csv""
  print $ (foldr count M.empty . map length . lines) content

count a b = M.insertWith (+) a 1 b

```

","The first big mistake in



```
main = do
  content <- readFile ""output.csv""
  print $ (foldr count M.empty . map length . lines) content

count a b = M.insertWith (+) a 1 b

```

is using `foldr`. That constructs an expression of the form



```
length firstLine `count` length secondLine `count` ... `count` length lastLine `count` M.empty

```

traversing the entire list of lines constructing the thunk - at that time not even evaluating the `length` calls due to laziness - before it is then evaluated right to left. So the entire file contents is in memory in addition to the thunk for building the `Map`.


If you build up a map from a list of things, **always** use a strict left fold (well, if the list is short, and the things not huge, it doesn't matter) unless the semantics require a right fold (if you're combining values using a non-commutative function, that might be the case, but even then it is often preferable to use a left fold and `reverse` the list before building the map).


`Data.Map`s (or `Data.IntMap`s) are spine-strict, that alone makes it impossible to generate partial output before the entire list has been traversed, so the strengths of `foldr` cannot be used here.


The next (possible) problem is (again laziness) that you don't evaluate the mapped-to values when putting them in the `Map`, so if there is a line length that occurs particularly often, that value becomes a huge thunk



```
((...((1+1)+1)...+1)+1)

```

Make it



```
main = do
  content <- readFile ""output.csv""
  print $ (foldl' count M.empty . map length . lines) content

count mp a = M.insertWith' (+) a 1 mp

```

so that the lines can be garbage collected as soon as they have been read in, and no thunks can build up in the values. That way you never need more than one line of the file in memory at once, and even that need not be in memory entirely, since the `length` is evaluated before it is recorded in the `Map`.


If your `containers` package is recent enough, you could also



```
import Data.Map.Strict

```

and leave `count` using `insertWith` (without prime, the `Data.Map.Strict` module always evaluates values put into the map).


"
"List comprehension in pure BASH?
Is it possible to do LC like in python and other languages but only using BASH constructs?


What I would like to be able to do, as an example is this:



```
function ignoreSpecialFiles()
{
    for options in ""-L"" ""-e"" ""-b"" ""-c"" ""-p"" ""-S"" ""! -r"" ""! -w""; do
        if [[ $options ""$1"" -o $options ""$2"" ]];then
            return $IGNORED
        fi
    done
}

```

instead of using code like this:



```
if [[ -L ""$1"" -o -e ""$1"" -o -b ""$1"" -o -c ""$1"" -o -p ""$1"" -o -S ""$1"" -o\
! -r ""$1"" -o ! -w ""$1"" ]]

```

Do you know of any recipes for simulating LC like this??


# Edit:


A more LC specific example:



```
M = [x for x in S if x % 2 == 0] #is python

```

what is the most pythonic way to do the same in bash?



```
for x in S; do if x % 2 == 0;then (HERE MY HEAD EXPLODES) fi done

```

","Here's a semi-general format for doing the equivalent of LC in bash with arrays:



```
outarray=()
for x in ""${inarray[@]}""; do
    if SOMECONDITION; then
        outarray+=(SOMEFUNCTIONOFx)
    fi
done

```

Here's your second example in this format:



```
s=({1..10})
echo ""${s[@]}""
# Prints: 1 2 3 4 5 6 7 8 9 10
m=()
for x in ""${s[@]}""; do
    if (( x % 2 == 0 )); then
        m+=($x)
    fi
done
echo ""${m[@]}""
# Prints: 2 4 6 8 10

```

Here's another example, with a less trivial ""function"" but without the conditional:



```
paths=(""/path/to/file 1"" ""/path/somewhere/else"" ""/this/that/the other"" ""/here/there/everywhere"")
filenames=()
for x in ""${paths[@]}""; do
    filenames+=( ""$(basename ""$x"")"" )
done
printf ""'%s' "" ""${filenames[@]}""
# Prints: 'file 1' 'else' 'the other' 'everywhere' 

```

"
"what are the changes in mysql 8 result rowset case?
when running 



```
SELECT maxlen FROM `information_schema`.`CHARACTER_SETS`;

```

mysql 5.7 and mysql 8 produce different results:


- on mysql 5.7 the results row names are lower cased,
- on mysql 8 the results row names are upper cased.


NB : in the `CHARACTER_SETS` table, the comumn name is `MAXLEN` (upper cased).


Since I can't find a resource documenting it, my question is : 



> 
> what are the changes in mysql 8 result rowset case ?
> 
> 
> 


","[MySQL 8.0 did change](https://dev.mysql.com/blog-archive/mysql-8-0-improvements-to-information_schema/) the implementation of some views in the INFORMATION\_SCHEMA.



> 
> Now that the metadata of all database tables is stored in transactional data dictionary tables, it enables us to design an INFORMATION\_SCHEMA table as a database VIEW over the data dictionary tables. This eliminates costs such as the creation of temporary tables for each INFORMATION\_SCHEMA query during execution on-the-fly, and also scanning file-system directories to find FRM files. It is also now possible to utilize the full power of the MySQL optimizer to prepare better query execution plans using indexes on data dictionary tables.
> 
> 
> 


So it's being done for good reasons, but I understand that it has upset some of your queries when you fetch results in associative arrays based on column name.


You can see the definition of the view declares the column name explicitly in uppercase:



```
mysql 8.0.14> SHOW CREATE VIEW CHARACTER_SETS\G
*************************** 1. row ***************************
                View: CHARACTER_SETS
         Create View: CREATE ALGORITHM=UNDEFINED DEFINER=`mysql.infoschema`@`localhost` SQL SECURITY DEFINER VIEW `CHARACTER_SETS` AS 
  select 
    `cs`.`name` AS `CHARACTER_SET_NAME`,
    `col`.`name` AS `DEFAULT_COLLATE_NAME`,
    `cs`.`comment` AS `DESCRIPTION`,
    `cs`.`mb_max_length` AS `MAXLEN` -- delimited column explicitly uppercase
  from (`mysql`.`character_sets` `cs` 
  join `mysql`.`collations` `col` on((`cs`.`default_collation_id` = `col`.`id`)))

character_set_client: utf8
collation_connection: utf8_general_ci

```

You can work around the change in a couple of ways:


You can declare your own column aliases in the case you want when you query a view:



```
mysql 8.0.14> SELECT MAXLEN AS `maxlen` 
  FROM `information_schema`.`CHARACTER_SETS` LIMIT 2;
+--------+
| maxlen |
+--------+
|      2 |
|      1 |
+--------+

```

You could start a habit of querying columns in uppercase prior to 8.0. Here's a test showing results in my 5.7 sandbox:



```
mysql 5.7.24> SELECT MAXLEN 
  FROM `information_schema`.`CHARACTER_SETS` LIMIT 2;
+--------+
| MAXLEN |
+--------+
|      2 |
|      1 |
+--------+

```

Or you could fetch results into a non-associative array, and reference columns by column number, instead of by name.


"
"How can I protect data in a SQLite database in Android?
I'm currently developing an Android game which saves data into a SQLite database. This is not really ""sensitive"" data, but I don't want users to be able to **modify** it (for obvious reasons of game balance, as it would be cheating). And it's quite easy to access and modify a SQLite db **when your phone is rooted** (there are plenty of applications for that in the market).


So should I even worry about that, or consider users with a rooted phone can do whatever they want including cheating and that's their choice? Or could I somehow **encrypt** the data I don't want them to modify, or add a **MD5 checksum** or something similar?


Another approach would be to abandon SQLite altogether and use some kind of binary files with game data.


Please let me know if some of you already encountered similar issues, and what are the approaches you followed, as well as the ""good practices"" in the Android gaming development community.


","Root access for everybody and security are mutually exclusive.


Any application or user with root permissions can read *and modify* each and every file on your system, as well as all of the main memory. That doesn't leave many places to store a potential encryption key for the database.


You could hide parts of the key in the executables, configuration files etc, but everything you could come up with would be nothing more than obfuscation and security by obscurity.


If a user opts to grant root access to everybody, that's their decision, and it's not your job as an app developer to prevent any harm that might be caused.


Update:
[Storing API keys in Android, is obfustication enough?](https://stackoverflow.com/questions/4671859/storing-api-keys-in-android-is-obfustication-enough) is a pretty similar issue - it's about protecting API keys, but it's the same situation with regards to your options.


"
"QR code scan from image file
Tried to use several libraries like ZXing, ZBar and their forks but didn't find way to scan barcode not from camera but from file.


Can someone point me to right direction? Preferably I'm looking into ZXing: how to scan image from file (not from camera).


Please.


","In the end I've found solution. Code is (originated from [here](https://stackoverflow.com/questions/14861553/zxing-convert-bitmap-to-binarybitmap)):



```
import com.google.zxing.*;

public static String scanQRImage(Bitmap bMap) {
    String contents = null;

    int[] intArray = new int[bMap.getWidth()*bMap.getHeight()];
    //copy pixel data from the Bitmap into the 'intArray' array
    bMap.getPixels(intArray, 0, bMap.getWidth(), 0, 0, bMap.getWidth(), bMap.getHeight());

    LuminanceSource source = new RGBLuminanceSource(bMap.getWidth(), bMap.getHeight(), intArray);
    BinaryBitmap bitmap = new BinaryBitmap(new HybridBinarizer(source));

    Reader reader = new MultiFormatReader();
    try {
        Result result = reader.decode(bitmap);
        contents = result.getText();
    }
    catch (Exception e) {
        Log.e(""QrTest"", ""Error decoding barcode"", e);
    }
    return contents;
}

```

Gradle referencing as:



```
dependencies {
    compile 'com.google.zxing:core:3.2.1'
}

```

Usage:



```
InputStream is = new BufferedInputStream(new FileInputStream(file));
Bitmap bitmap = BitmapFactory.decodeStream(is);
String decoded=scanQRImage(bitmap);
Log.i(""QrTest"", ""Decoded string=""+decoded);

```

"
"How do I know to which core my Python process has been bound?
How do I know to which process my Python process has been bound?
Alone these same lines, are child processes going to execute on the same core (i.e. CPU) that the parent is currently executing?


","Processes and native OS threads are only bound to specific processors if somebody specifically requests for that to happen. By default, processes and threads can (and will) be scheduled on any available processor.


Modern operating systems use pre-emptive multi-threading and can interrupt a thread's execution at any moment. When that thread is next scheduled to run, it can be executed on a different processor. This is known as a context switch. The thread's entire execution context is stored away by the operating system and then when the thread is re-scheduled, the execution context is restored.


Because of all this, it makes no real sense to ask what processor your thread is executing on since the answer can change at any moment. Even during the execution of the function that queried which the current thread's processor.


Again, by default, there's no relationship between the processors that two separate processes execute on. The two processes could execute on the same processor, or different processors. It all depends on how the OS decides to schedule the different threads.


In the comments you state:



> 
> The Python process will execute on only one core due to the GIL lock.
> 
> 
> 


That statement is simply incorrect. For example, a section of Python code would claim the GIL, get context switched around all the available processors, and then release the GIL.


Right at the start of the answer I said alluded to the possibility of binding a process or thread to a particular processor. For example, on Windows you can use `SetProcessAffinityMask` and `SetThreadAffinityMask` to do this. However, it is unusual to do this. I can only recall ever doing this once, and that was to ensure that an execution of `CPUID` run on a specific processor. In the normal run of things, processes and threads have affinity with all processors.


In another comment you say:



> 
> I am creating the child processes to use multi cores of the CPU.
> 
> 
> 


In which case you have nothing to worry about. Typically you would create as many processes as there are logical processors. The OS scheduler is sensible and will schedule each different process to run on a different processor. And thus make the optimal use of the available hardware resources.


"
"How to encrypt data in core data (sqllite) in OS-X application
I have found that if I use transformable type of attributes and NSXMLStoreType my data is encrypted, that is attributes that has been of transformable type, are not readable. There is no need of doing anything else, no code is required. **Please note that I am working on OS-X application that uses core data.**


However, if I change my store type to NSSQLiteStoreType, that is not the case. 


I can open the database with sqllitebrowser, select the transformable field, and If I click on export button, in the generated text file, I can read the value normally, that is the value (data) is not encrypted.


I have asked [the same question](https://stackoverflow.com/questions/34169270/how-to-encrypt-and-decrypt-transformable-properties-on-os-x-application) about 4 months ago and I get no answer. 


Also, I have found [this post](https://stackoverflow.com/questions/1645007/how-can-i-encrypt-coredata-contents-on-an-iphone) here on stackoverflow.



> 
> You can encrypt individual properties in your Core Data model entities
>  by making them transformable properties, then creating an
>  NSValueTransformer subclass which will encrypt and decrypt the data
>  for that property.
> 
> 
> 


Unlucky for me, the author of the answer, @Brad Larson, didn't provide an simple example of how this can be done.


Can anyone provide any sample code of how I can encrypt transformable properties so that It' cant be readable in any way? 


","you could do something like shown here  [Cross-platform-AES-encryption](https://github.com/Pakhee/Cross-platform-AES-encryption)


add a new Objective-C file to project select category and NSData class call it Additions


**NSData+Additions.h**



```
#import <Foundation/Foundation.h>
#import <CommonCrypto/CommonDigest.h>
#import <CommonCrypto/CommonCryptor.h>

@interface NSData (Additions)

#pragma mark - data encryption

+ (NSData *)encrypt:(NSData *)plainText key:(NSData *)key iv:(NSData *)iv;
+ (NSData *)decrypt:(NSData *)encryptedText key:(NSData *)key iv:(NSData *)iv;

+ (NSData *)dataFromHexString:(NSString *)string;
+ (NSData *)sha256forData:(id)input;

+ (NSData *)generateRandomIV:(size_t)length;

@end

```

**NSData+Additions.m**



```
#import ""NSData+Additions.h""

@implementation NSData (Additions)

+ (NSData *)encrypt:(NSData *)dataToEncrypt key:(NSData *)key iv:(NSData *)iv {

    NSUInteger dataLength = [dataToEncrypt length];

    size_t buffSize = dataLength + kCCBlockSizeAES128;
    void *buff = malloc(buffSize);

    size_t numBytesEncrypted = 0;

    CCCryptorStatus status = CCCrypt(kCCEncrypt,
                                     kCCAlgorithmAES128,
                                     kCCOptionPKCS7Padding,
                                     [key bytes], kCCKeySizeAES256,
                                     [iv bytes],
                                     [dataToEncrypt bytes], [dataToEncrypt length],
                                     buff, buffSize,
                                     &numBytesEncrypted);

    if (status == kCCSuccess) {
        return [NSData dataWithBytesNoCopy:buff length:numBytesEncrypted];
    }

    free(buff);
    return nil;
}

+ (NSData *)decrypt:(NSData *)encryptedData key:(NSData *)key iv:(NSData *)iv {

    NSUInteger dataLength = [encryptedData length];

    size_t buffSize = dataLength + kCCBlockSizeAES128;

    void *buff = malloc(buffSize);

    size_t numBytesEncrypted = 0;
    CCCryptorStatus status = CCCrypt(kCCDecrypt,
                                     kCCAlgorithmAES128,
                                     kCCOptionPKCS7Padding,
                                     [key bytes], kCCKeySizeAES256,
                                     [iv bytes],
                                     [encryptedData bytes], [encryptedData length],
                                     buff, buffSize,
                                     &numBytesEncrypted);
    if (status == kCCSuccess) {
        return [NSData dataWithBytesNoCopy:buff length:numBytesEncrypted];
    }

    free(buff);
    return nil;
}

+ (NSData *)dataFromHexString:(NSString *)string {

    NSMutableData *stringData = [[NSMutableData alloc] init];
    unsigned char whole_byte;
    char byte_chars[3] = {'\0','\0','\0'};

    for (int counter = 0; counter < [string length] / 2; counter++) {
        byte_chars[0] = [string characterAtIndex:counter * 2];
        byte_chars[1] = [string characterAtIndex:counter * 2 + 1];
        whole_byte = strtol(byte_chars, NULL, 16);
        [stringData appendBytes:&whole_byte length:1];
    }

    return stringData;
}

+ (NSData *)sha256forData:(id)input {
    NSData *dataIn;

    if ([input isKindOfClass:[NSString class]]) {
        dataIn = [input dataUsingEncoding:NSUTF8StringEncoding];
    } else if ([input isKindOfClass:[NSData class]]) {

        NSUInteger dataLength = [input length];
        NSMutableString *string = [NSMutableString stringWithCapacity:dataLength * 2];
        const unsigned char *dataBytes = [input bytes];

        for (NSInteger idx = 0; idx < dataLength; ++idx)
            [string appendFormat:@""%02x"", dataBytes[idx]];

        dataIn = [string dataUsingEncoding:NSUTF8StringEncoding];
    }

    NSMutableData *macOut = [NSMutableData dataWithLength:CC_SHA256_DIGEST_LENGTH];

    CC_SHA256(dataIn.bytes, (CC_LONG)[dataIn length],  [macOut mutableBytes]);

    return macOut;
}

+ (NSData *)generateRandomIV:(size_t)length {
    NSMutableData *data = [NSMutableData dataWithLength:length];

    SecRandomCopyBytes(kSecRandomDefault, length, [data mutableBytes]);

    return data;
}

@end

```

then you are going to need username, password, iVector and some random salt. in the salt i have used replace # with random characters but make sure that if u use % do not forget to double it %% otherwise it will be incomplete format specifier warning. 


in your class use it like this



```
#define CC_USERNAME         @""secretName""
#define CC_PASSWORD         @""secretPassword""
#define CC_SALTED_STRING    [NSString stringWithFormat:@""####################%@#####################"", CC_PASSWORD]

```

then create NSData representation of your salted string ran through SHA256



```
NSData *hash = [NSData sha256forData:CC_SALTED_STRING];

```

next step is to generate 16 bytes of random generated iVector data



```
NSData *iVector = [NSData generateRandomIV:16];

```

and use these objects to encrypt your string. create a NSMutableData object with first 16 bytes of iVector data (make sure u use iVector object and do not generate new random or you will be unable to decrypt).



```
NSString *message = @""my secret message to the world"";
NSData *messageData = [message dataUsingEncoding:NSUTF8StringEncoding];

NSMutableData *encryptedData = [[NSMutableData alloc] initWithData:iVector];

NSData *payLoad = [NSData encrypt:messageData key:hash iv:iVector];

[encryptedData appendData:payLoad];

```

to decrypt, separate first 16 bytes and the rest of data and use it with the hash.



```
NSData *pureData = [encryptedData subdataWithRange:NSMakeRange(16, [encryptedData length] - 16)];
NSData *extractedVector = [encryptedData subdataWithRange:NSMakeRange(0, 16)];

NSData *decryptedData = [NSData decrypt:pureData key:hash iv:extractedVector];

NSString *decryptedMessage = [[NSString alloc] initWithData:decryptedData encoding:NSUTF8StringEncoding];

```

you can do some extra md5 on the hash or even pack encrypted data with zlib before storing it. 


enjoy your custom made crypto.


"
"Sql Server - Constraint - Allow to set column A only if column B is null and vice-versa
Is there a way to add a constraint in SQL Server 2008 that will verify that if a user try to enter a value in a column A, it can only if column B is null and vice-versa


Example


- if A is NULL, B can have a value
- if B iS NULL, A can have a value
- A and B can't have value at the same time


","Something like this:



```
ALTER TABLE foo WITH CHECK ADD
    CONSTRAINT CK_Foo_reason CHECK (
        ColA IS NOT NULL AND ColB IS NULL
        OR
        ColA IS NULL AND ColB IS NOT NULL
        )

```

Edit: after question update


It depends if both columns are allowed NULL in which case



```
        ColA IS NULL OR ColB IS NULL

```


Edit 2: For 3 columns, exactly one of which must be NOT NULL


No smarty-pants answer I'm afraid. I've added spaces to hopefully make it clearer



```
ALTER TABLE foo WITH CHECK ADD
    CONSTRAINT CK_Foo_reason CHECK (
        ColA IS NOT NULL AND ColB IS NULL     AND ColC IS NULL
        OR
        ColA IS NULL     AND ColB IS NOT NULL AND ColC IS NULL
        OR
        ColA IS NULL     AND ColB IS NULL     AND ColC IS NOT NULL
        )

```

"
"Using list item contents as string name
I'm using C# and making lists from sqlite query responses.


List1 contains a field with the name of a field in list2



```
List1[i].field = fieldname

```

I want to use the fieldname stored in `List1[i].field` to access an item in `List2` something like:



```
List2[ii].+List1[i].field+

```

I have no idea if this is possible in this way in C# i'm relatively new to it and essentially going off my knowledge of PHP where something like this does work.


","For this type of data structure, you usually use a dictionary (key/value pairs) to dynamically select properties of an object. For your specific example, this could be accomplished with reflection, but it is not very commonly used since other structures make it much easier.



```
List2[ii].GetType().GetProperty(List1[i].field).GetValue(List2[ii], null);

```

Reflection isn't the fastest type of solution typically, and it's usually not great to come in behind and maintain code with a lot of reflection if it can be avoided.


Alternatively, you could use Linq to transfer List2 from a list of your object types into a list of dictionary objects so you could do something like `List2[ii][List1[i].field]` to accomplish the same thing.



```
var listDictionaries = List2
    .Select(l => new Dictionary<string, object>()
    {
        new { ""Prop1"", l.Prop1 },
        new { ""Prop2"", l.Prop2 },
        new { ""Prop3"", l.Prop3 },
    }).ToArray();
listDictionaries[ii][List1[i].field];

```

"
"Checking correctness of function call expression
`std::result_of` computes the return type of a calling expression at compile time.


As [the reference](http://en.cppreference.com/w/cpp/types/result_of) says, if the call is ill-formed the instantation of `std::result_of` results in a compilation error. But suppose that what we need is to check if the call is well formed, before getting the result type.


**Is there any way to write a trait that checks if a calling expression is well-formed?**


Something like:



```
template<typename F , typename... ARGS>
struct is_valid_call : public impl::is_valid_call<F,typelist<ARGS...>>
{};

namespace impl
{
    struct sfinae_result{};

    template<typename F , typename ARGS , typename ENABLED = sfinae_result>
    struct is_valid_call;

    template<typename F , typename... ARGS>
    struct is_valid_call<F,typelist<ARGS...>,
                         decltype( std::declval<F>()(std::declval<ARGS>()...) )
                        > : 
        public std::true_type
    {};

    template<typename F , typename... ARGS>
    struct is_valid_call<F,typelist<ARGS...>,sfinae_result> : 
        public std::false_type
    {};
}

```

**EDIT:** Of course the solution posted [doesn't work](http://ideone.com/yuXO8R)


","Here's something that works:



```
#include <type_traits>
#include <utility>

template<typename F, typename... Args>
struct is_valid_call {
private:
    template<typename FF, typename... AA>
    static constexpr auto check(int) ->
    decltype( std::declval<FF>()(std::declval<AA>()...), std::true_type());

    template<typename FF, typename... AA>
    static constexpr std::false_type check(...);
public:
    static constexpr bool value = decltype(check<F, Args...>(0)){};
};

#include <cstdio>

int main()
{
   printf(""%d"", int (is_valid_call<decltype(&puts), const char*>::value));
   printf(""%d"", int (is_valid_call<decltype(&puts), double>::value));
}

```

Output: `10`


"
"Need Jersey client api way for posting a webrequest with json payload and headers
I am writing a client for one of my REST API using jersey(org.glassfish.jersey.client.\*).


api url is : `http://localhost:5676/searchws/search/getresults` (POST)


this api returns a json response. i need to provide a payload using jersey client and thats where i am stuck. FOllowing is a sample extract of payload which i need to provide (preferably as string)


Question is how can i provide a payload (XML/JSON) as string or entity to my webtarget.


I saw the answer to providing payload mentioned by calden [How to send Request payload to REST API in java?](https://stackoverflow.com/questions/15570656/how-to-send-request-payload-to-rest-api-in-java) but i am looking for a way to do it in jersey client.


Here is my code till now which does not work fully for post requests.



```
public class RequestGenerator 
{

    private WebTarget target;
    private ClientConfig config;
    private Client client;
    private Response response;

    public RequestGenerator(Method RequestSendingMethod) throws Exception
    {
        switch (RequestSendingMethod)
        {
            case POST :
                config = new ClientConfig();
                client = ClientBuilder.newClient(config);
                target = client.target(""http://localhost:5676/searchws"").path(""search"").path(""getresults"");
                String payload = ""{\""query\"":\""(filter:(\\\""google\\\"")) AND (count_options_availbale:[1 TO *])\""}""; //This is just a sample json payload actual one is pretty large
                response = target.request(MediaType.APPLICATION_JSON_TYPE).post(Entity.json("""")); // What to do here
                String jsonLine = response.readEntity(String.class);
                System.out.println(jsonLine);
}

}

```

","I got this working using following code, Salil's code works fine as well(+1 with thanks to him), thanks everyone who contributed to this problem, loving stackoverflow:



```
public class RequestGenerator 
{

    private WebTarget target;
    private ClientConfig config;
    private Client client;
    private Response response;

    public RequestGenerator(Method RequestSendingMethod) throws Exception
    {
               switch (RequestSendingMethod)
            {
                case POST :             
                    String payload = ""\r\n{\r\n\""query\"": \""google \"",\r\n\""rows\"": 50,\r\n\""return_docs\"": true,\r\n\""is_facet\"": true\r\n}""; //this is escapped json string in single line
                    config = new ClientConfig();
                    client = ClientBuilder.newClient(config);
                    target = client.target(""http://localhost:7400/searchws/search/getresults"");
                    response = target.request().accept(MediaType.APPLICATION_JSON).post(Entity.entity(payload, MediaType.APPLICATION_JSON), Response.class);
                    processresponse(response); //This could be any method which processes your json response and gets you your desired data.
                    System.out.println(response.readEntity(String.class));
                    break;
                case GET :
                    config = new ClientConfig();
                    client = ClientBuilder.newClient(config);
                    target = client.target(""http://localhost:7400/search-service/searchservice"").path(""search"").path(""results"").path(""tiger"");
                    response = target.request().accept(MediaType.APPLICATION_JSON).get();
                    processresponse(response); //This could be any method which processes your json response and gets you your desired data.
                    System.out.println(response.readEntity(String.class));
            }

}

```

"
"Spin UIImageView continuously
I am having a problem while trying to rotate `UIImageview` continuously with a ball's image inside. I would like this ball to spin continuously on its center axis.
I have tried using `CGAffineTransform` but it didn't work.


Please help!


","This may be an old Q but it's near the top of search results for the topic. Here's a more cut and dry solution: (make sure to import QuartzCore/QuartzCore.h) 



```
CABasicAnimation *rotation;
rotation = [CABasicAnimation animationWithKeyPath:@""transform.rotation""];
rotation.fromValue = [NSNumber numberWithFloat:0];
rotation.toValue = [NSNumber numberWithFloat:(2*M_PI)];
rotation.duration = 1.1; // Speed
rotation.repeatCount = HUGE_VALF; // Repeat forever. Can be a finite number.
[yourView.layer addAnimation:rotation forKey:@""Spin""];

```

Then, to ~~stop~~ remove/reset the animation: (see comments for how to stop-in-place)



```
[yourView.layer removeAnimationForKey:@""Spin""];

```

In swift:



```
let rotation = CABasicAnimation(keyPath: ""transform.rotation"")
rotation.fromValue = 0
rotation.toValue = 2 * M_PI
rotation.duration = 1.1
rotation.repeatCount = Float.infinity
view.layer.addAnimation(rotation, forKey: ""Spin"")

```

"
"Return a partial view from a controller?
So, we can return a partial view from a controller like this:



```
using System;
using System.Collections.Generic;
using System.Linq;
using System.Web;
using System.Web.Mvc;
using MvcApplication1.Models;

namespace MvcApplication1.Controllers
{
    public class HomeController : Controller
    {
        public ActionResult Index()
        {
            ViewBag.Message = ""Modify this template to jump-start your ASP.NET MVC application."";

            return View();
        }

        public ActionResult About()
        {
            ViewBag.Message = ""Your app description page."";

            return View();
        }

        public ActionResult Contact()
        {
            ViewBag.Message = ""Your contact page."";

            return View();
        }

        public PartialViewResult Address()
        {
            Address a = new Address 
            { 
                Line1 = ""111 First Ave N."", 
                Line2 = ""APT 222"", 
                City = ""Miami"", 
                State = ""FL"", 
                Zip = ""33133"" 
            };

            return PartialView(@""~/Views/Home/_Address.cshtml"", a);
        }
    }
}

```

But, how am I supposed to use the returned partial view? I created \_Address.cshtml under Views/Home, like this:



```
@model MvcApplication1.Models.Address

<p>
    This is a partial view of address.
</p>
<p>
  @Model.City
</p>

```

And, at the end of Views/Home/Contact.cshtml, I added this line:



```
@Html.Partial(@""~/Views/Home/_Address.cshtml"")

```

And I expect to see the City of my address, but it doesn't show up. I am confused.


","When the partial takes a different model than the method you're including it in you need to use the overload that takes a model parameter and supply the model for the view. By default it uses the same model as the including view. Typically you only need the path if it's in a different, non-shared folder. If it's in the same controller's folder, using just the name should do the trick.



```
@Html.Partial(""_Address"", Model.Address)

```

On the other hand, if you're asking how do I get the partial view from an action included in my page, then you want to use the `Action` method instead of the `Partial` method.



```
@Html.Action(""Address"")

```

**EDIT**


To make the partial work you need to pass a `Contact` model to the contact view.



```
public ActionResult Contact()
{
     var contact = new Contact
     {
        Address = new Address
                  { 
                       Line1 = ""111 First Ave N."",
                       Line2 = ""APT 222"",
                       City = ""Miami"",
                       State = ""FL"",
                       Zip = ""33133""
                  }
     }

     return View(contact);
}

```

"
"ActionServlet - Synchronized block - working of threads
if there is a `synchronized` block in the `service()` method of say for example in `ActionServlet` of struts, how will multiple requests/threads work if it is a busy site having large number of hits.


Will each thread wait for the next other one to release lock from the synchronized block? Will that create delay in response


","**Don't synchronize the service method of a servlet.**


If you synchronize the `service` method of a servlet, you are in fact making a ""reservation on access"" for a thread at a time for that instance of the Servlet.


The Struts `ActionServlet` class is a `HttpServlet` and basically the `doGet` and `doPost` methods are of interest here. If we are to speak about Struts, the [process](http://struts.apache.org/1.x/apidocs/org/apache/struts/action/ActionServlet.html#process%28javax.servlet.http.HttpServletRequest,%20javax.servlet.http.HttpServletResponse%29) method is the main entry point, but the same principle applies to all methods as it does for the general `service` method.


The idea is this.


When you declare a servlet in your `web.app`, the servlet container (e.g. Tomcat) will create only one instance of that servlet. This means **there is only one instance to serve all requests**.


If more requests arrive at the same time, each request thread gets it chance at the `service` method because there is no synchronization enforced. 


If you have 10 request threads, each will execute concurrently in the `service` method. This is normally safe because processing done in the `service` method does not involve any state related to the current request it is handling. You go into issues if you add state to your servlets. [Here is an article with more details on the subject](http://www.javaworld.com/javaworld/jw-07-2004/jw-0712-threadsafe.html). 


**Now back to Struts**.


Struts uses a pattern called a [Front Controller](http://java.sun.com/blueprints/patterns/FrontController.html) with the `ActionServlet` being that controller. This will in turn delegate specific requests to specific `Action` classes as specified in its configuration (a.k.a `struts-config.xml`).


All incoming request pass though here. If you place synchronization at this point (the Struts `process` method or the servlet `service` method higher up) you are reserving the servlet for a thread at a time. **In case of struts, you are reserving all processing of a request to a single thread at a time**. 


That means that if 10 request arrive simultaneousely, in the case without synchronization all can execute side by side, while in the case with synchronization request 2 will have to wait until request 1 is done, 3 waits for 2 and so on (i.e. requests are sequentially processed). And this means poor performance. 


Maybe for the lucky user that made request 1 there will be no performance impact but number 10 will have to wait. Then how about number 100? 200?


There is no need to synchronize the entry point if you program your application with thread safety in mind. If your application is of the sort that you just can't avoid synchronization, then synchronizing the entry point will reduce performance.


P.S. Just one other thing. If you are thinking into moving the synchronization lower in the process, namely the `Action` classes, note that they are not thread safe either and there is [only one instance of it inside the Struts framework](https://stackoverflow.com/questions/5650207/struts-action-singleton/5655287#5655287). 


"
"How do I get first three characters from an edittext?
I want to get the first three characters from an edittext and then turn them onto a string, but I cant find anything about that. Any ideas?


","To get text from an EditText, or a TextView, Button, etc. (pretty much any View that has text), you call getText(). This returns a CharSequence, which is *almost* a String, but not quite, so to turn it into a String object, call toString() on it. And then to get the first 3 letters, use the substring() method, where the first argument is the index of the character to start, and the second is *one past the last* character you want. So you want the first 3 characters, which are indices 0,1,2, so we must have 3.



```
EditText yourEditText = (EditText) findViewById(R.id.your_edit_text);
CharSequence foo = yourEditText.getText();
String bar = foo.toString();
String desiredString = bar.substring(0,3);

```

In addition, you will probably want to make sure that the user has actually put something in the EditText before assuming that there is and getting a NullPointerException when you try to use the string. So i usually use EditTexts in the following way.



```
EditText yourEditText = (EditText) findViewById(R.id.your_edit_text);
String foo = yourEditText.getText().toString();
if(foo.length() > 0) { //just checks that there is something. You may want to check that length is greater than or equal to 3
    String bar = foo.substring(0, 3);
    //do what you need with it
}

```

"
"Reading struct values using pointer
I have a struct that contains a float var. I am trying to read the value using a pointer to a struct. Here's the code:



```
struct mas {
    float m;
};

int main(void)
{
    struct mas *ms;
    ms=(struct mas*)malloc(sizeof(struct mas));
    scanf(""%f"",&(ms->m));
    printf(""%f"",ms->m);
    return 0;
}

```

But running the program produces the following error:



```
scanf floating point formats not linked

```

The compiler used is Borland Turbo C++ (3.0) on a Windows PC. Why is this so?


","This might be helpful: <http://www.faqs.org/faqs/msdos-programmer-faq/part2/section-5.html>


From the article: 



> 
> Borland's compilers try to be smart and not
>  link in the floating- point (f-p) library unless you need it. Alas, they
>  all get the decision wrong. ... (To force them to link it) define this function somewhere in a source file but don't
>  call it:
> 
> 
> 



```
static void forcefloat(float *p)
{
  float f = *p;
  forcefloat(&f);
}

```

Also:



> 
> If you have Borland C++ 3.0, the README file documents a slightly less
>  ugly work-around. Insert these statements in your program:
> 
> 
> 



```
 extern unsigned _floatconvert;
 #pragma extref _floatconvert

```

"
"Upload File in Nativescript Webview with iframe
My problem is I have one webview Containing Dynamic Form or form which is created dynamic with WP Builders and that is included in iframe and i want to select file from application in nativescript this will not allow me select file.


So I want to try 


[How to load files into webview in native script](https://stackoverflow.com/questions/39994187/how-to-load-files-into-webview-in-native-script)


<http://shripalsoni.com/blog/nativescript-webview-native-bi-directional-communication/>


But problem is form is dynamic is there any script which will detect click event of file input inside frame and also put file path to it when file is selected if we can detect event for file input click.


another issue is what if there are more than one file input in dynamic form


","I Got Answer by Implementing WebChromeClient In android



```
 let webview: WebView = this.webViewRef.nativeElement;
 let myWebChromeClient: MyWebChromeClient = new MyWebChromeClient();
 webview.android.setWebChromeClient(myWebChromeClient);

```

and class MyWebChromeClient is here



```
import * as imagepicker from ""nativescript-imagepicker"";
import * as fs from ""file-system"";

export class MyWebChromeClient extends android.webkit.WebChromeClient {
    file_path:string;
    constructor() {
        super();

        return global.__native(this);
    }

    onShowFileChooser(webView,filePathCallback,fileChooserParams) {

         this.filepathCallback(filePathCallback);

        return true;
    }

    filepathCallback(filePathCallback)
    {
        let context = imagepicker.create({
            mode: ""single"",
            mediaType: imagepicker.ImagePickerMediaType.Any
        });
       this.startSelection(context,filePathCallback);
    }

    startSelection(context,filePathCallback) {
        context.authorize().then(() => {
                                    return context.present();
                                  })
            .then((selection) => {
                selection.forEach((selected) => {
                    let path = selected.android;
                    let file = fs.File.fromPath(path);
                    this.file_path = file.path;
                    this.file_path = ""file://"" + this.file_path;
                    let results = Array.create(android.net.Uri, 1);
                    results[0] = android.net.Uri.parse(this.file_path);

                    filePathCallback.onReceiveValue(results);
                    return this.file_path;
                });
            }).catch(function (e) {
                console.log(e);
            });
    }
}

```

"
"Oracle Java code conventions
Does someone know where I can find the document by Oracle which describes Java code conventions?


This URL is not available anymore, for that reason I created new question for this topic.


<http://www.oracle.com/technetwork/java/codeconv-138413.html> 


","One option is to use wayback machine, which [seems to contain the document](http://web.archive.org/web/20140222045352/http://www.oracle.com/technetwork/java/codeconv-138413.html) ([here's direct link to pdf version](http://web.archive.org/web/20140228225807/http://www.oracle.com/technetwork/java/codeconventions-150003.pdf)). However I would also be interested to find out what Oracle did with it and do they intend to get rid of it.


There's a [discussion on Oracle forums](https://community.oracle.com/thread/3544233) about this being raised to OTN support, but no response is mentioned.




---


Update on 17.6.2014: There's now a [forum posting](https://community.oracle.com/message/12472357#12472357) added to the site explaining the following:



> 
> Those Java Code Conventions were written in 1999 and have not been
>  maintained since.
> 
> 
> The information might not be up to date; links within the documents
>  might not work. That is why the pages were removed. Unfortunately
>  there are other sites that point to that document which were not
>  updated.
> 
> 
> To avoid confusions we have re-posted the original document –with an
>  appropriate disclaimer about the information not being up to date-
>  while we clean up those other sites.
> 
> 
> Try [Code Conventions for the Java Programming Language](http://www.oracle.com/technetwork/java/index-135089.html)
> 
> 
> 


And indeed, seems the page has been restored.


"
"running my sample app using express.js and node.js

```
var sys = require(""sys""),  
my_http = require(""http"");  
my_http.createServer(function(request,response){  
    sys.puts(""I got kicked"");  
    response.writeHeader(200, {""Content-Type"": ""text/plain""});  
    response.write(""Hello World"");  
    response.end();  
}).listen(8080);  
sys.puts(""Server Running on 8080"");

```

The above is **my basic web-server**, now i want to run my application which contains a HTML and a JS file. Where would i place those files so that i can access it through my port. 


I use `Apache and Xampp`, so i place my files in `htdocs` directory and access it via my browser, but in terms of `node.js` i am totally confused?.


","Let's get step by step here. 


**Identify the location for your application.**


Firs identify the location where for your application. Let's take it as `C:\your_app`. The path doesn’t matter, so feel free to locate the directory wherever is best for you.


**Installing Node.js**


Here is where we will setup `Node.js and Express`. Node.js is a framework and Express provides a web server. The web server we need does not need to do anything fancy. The only feature that the web server needs is the ability to provide static files.


To get started download and install Node.JS: <http://nodejs.org/>


**Install Express**


Express is a package that execute within Node.js. To install express, in the Command Prompt navigate to your directory for the application which is c:\your\_app.


Now lets `install Express as a package for Node.js.` At the command prompt type `“npm install express”`. That installed Express and should have created a directory called `“node_modules”`. 


**server.js**


Now that Express is installed, we need to configure it to execute as a webserver. Create another file in the c:\your\_app directory call “server.js”.



```
var express = require('express');
var app = express();
port = process.argv[2] || 8000;

app.configure(function () {
    app.use(
        ""/"", //the URL throught which you want to access to you static content
        express.static(__dirname) //where your static content is located in your filesystem
    );
});
app.listen(port); //the port you want to use
console.log(""Express server running"");

```

**Start Express Web Server in Node.js**


In the Command Prompt confirm you are at the c:\your\_app directory and execute the following commend. 



```
node server.js 8000

```

Now the web server `should be running on port 8000` and your index.html page should be **displayed in the browser**.


"
"Is there a way to control the number of partitions when reading a text file in PySpark
I am reading a text file using the following command in PySpark



```
rating_data_raw = sc.textFile(""/<path_to_csv_file>.csv"")

```

Is there a way to specify the number of partitions that RDD rating\_data\_raw should be split into? I want to specify a large number of partitions for greater concurrency.


","As other user said you can set the minimal number of partitions that will be created, while reading the file, by setting it in the optional parameter `minPartitions` of [textFile](http://spark.apache.org/docs/latest/api/python/pyspark.html#pyspark.SparkContext.textFile).



```
rating_data_raw = sc.textFile(""/<path_to_csv_file>.csv"", minPartitions=128)

```

Another way to achieve this is by using [repartition](https://spark.apache.org/docs/latest/api/python/pyspark.html?highlight=rdd#pyspark.RDD.repartition) or [coalesce](https://spark.apache.org/docs/latest/api/python/pyspark.html?highlight=rdd#pyspark.RDD.repartition), if you need to reduce the number of partition you may use `coalesce`, otherwise you can use `repartition`.



```
rating_data_raw = sc.textFile(""/<path_to_csv_file>.csv"").repartition(128)

```

"
"What is the purpose of liftIO?
I have the following code snippet from internet: 



```
calculateLength :: LengthMonad Int
calculateLength = do
  -- all the IO operations have to be lifted to the IO monad in the monad stack
  liftIO $ putStrLn ""Please enter a non-empty string: ""
  s <- liftIO getLine
  if null s
    then throwError ""The string was empty!""
    else return $ length s

```

and could not understand, why the author use `liftIO`? 


What is the purpose of `liftIO`? 


It is defined as follows:



```
class (Monad m) => MonadIO m where
    -- | Lift a computation from the 'IO' monad.
    liftIO :: IO a -> m a  

```

Is it possible to lift `IO a -> [a]`? It looks like natural transformation. 


","`IO` operations like `getLine, putStrLn ""...""` only work inside the `IO` monad. Using them inside any other monad will trigger a type error.


Still, there are many monads `M` which are defined in terms of `IO` (e.g. `StateT Int IO`, and apparently your `LengthMonad` as well) and because of that they allow `IO` actions to be converted into `M`-actions, and executed as such.


However, we need a conversion for each `M`:



```
convertIOintoM1 :: IO a -> M1 a 
convertIOintoM2 :: IO a -> M2 a
convertIOintoM3 :: IO a -> M3 a
...

```

Since this is cumbersome, the library defines a typeclass `MonadIO` having such conversion function, so that all the functions above can be named `liftIO` instead.


In practice, `liftIO` is used each time one wants to run `IO` actions in another monad, provided such monad allows for it.


"
"Gitlab Rack Attack Insallation
I was installing gitlab and followed all the instructions of the installation guide of Gitlab. Everything works fine but then at Step ""Initialize Database and Activate Advanced Features""
it failed! 


<https://github.com/gitlabhq/gitlabhq/blob/master/doc/install/installation.md>


I got this error Message.



```
rake aborted!
uninitialized constant Rack::Attack
/home/git/gitlab/config/application.rb:82:in `<class:Application>'
/home/git/gitlab/config/application.rb:13:in `<module:Gitlab>'
/home/git/gitlab/config/application.rb:12:in `<top (required)>'
/home/git/gitlab/Rakefile:5:in `require'
/home/git/gitlab/Rakefile:5:in `<top (required)>'
(See full trace by running task with --trace)

```

In the application.rb file I uncommented the line as it says in the instruction.



```
# Enable rack attack middleware
# Find and uncomment the line 'config.middleware.use Rack::Attack' 

```

","Note that, from [issue 5245](https://github.com/gitlabhq/gitlabhq/issues/5245):



> 
> Rack Attack is new for 6-2 stable which will be released on 22rd of october.
> 
> 
> 


That being said, read ""[`doc/security/rack_attack.md`](https://github.com/gitlabhq/gitlabhq/blob/master/doc/security/rack_attack.md)"": 


uncomment the line `config.middleware.use Rack::Attack` in the [`config/application.rb`](https://github.com/gitlabhq/gitlabhq/blob/master/config/application.rb) find **isn't enough**.


The other steps are:



> 
> - Rename [`config/initializers/rack_attack.rb.example`](https://github.com/gitlabhq/gitlabhq/blob/master/config/initializers/rack_attack.rb.example) to `config/initializers/rack_attack.rb`
> - Review the `paths_to_be_protected` and add any other path you need protecting
> - Restart GitLab instance
> 
> 
> 


"
"In R, transform a character string into an operation
If I have a character vector, how can I tell R to execute it as a command?


Simple example:



```
> a <- paste(""2"", ""+"", ""3"")

>a

>[1] ""2 + 3""

```

I would like R to actually do 2+3, and not just print the content of ""a"".


","The way to do it is



```
a <- paste(""2"", ""+"", ""3"")
eval(parse(text = a))

```

But, in general, it's not a great idea. There's usually a better way to do what you're trying to do.



```
> require(fortunes)
> fortune(""parse"")

If the answer is parse() you should usually rethink the question.
   -- Thomas Lumley
      R-help (February 2005)

```

There's lots of info on why `eval(parse())` is considered a ""bad practice"" in the answers to [What specifically are the dangers of eval(parse())?](https://stackoverflow.com/q/13649979/903061); in summary it is 


- bug-prone
- harder to debug
- usually there's a more readable way to do whatever you're trying to do
- can pose a security risk


(though that last one probably doesn't apply for most R use cases)


"
"CentOS 7: What is /bin/sh? It looks like Bash but seems to be something else
When I run a Centos 7 Docker Image like this



```
docker run -it centos:7 bash

```

Running something which is uses *Process Substitution* is fine (as expected, as Bash supports Process Substitution since the beginning of time - Bash 1.4.x actually).


For example:



```
while IFS= read -r test; do echo $test; done < <(cat anaconda-post.log)

```

But when I switch to */bin/sh* the same code doesn't work anymore



```
/bin/sh
while IFS= read -r test; do echo $test; done < <(cat anaconda-post.log)
sh: syntax error near unexpected token `<'

```

Although */bin/sh* seems to be Bash



```
/bin/sh --version
GNU bash, version 4.2.46(2)-release (x86_64-redhat-linux-gnu)
Copyright (C) 2011 Free Software Foundation, Inc.
License GPLv3+: GNU GPL version 3 or later <http://gnu.org/licenses/gpl.html>

```

But then why doesn't process substitution work anymore? Other non-POSIX features seems to work, though



```
echo ${PATH//:/ }
/usr/local/sbin /usr/local/bin /usr/sbin /usr/bin /sbin /bin

```

","Yes, `bash`, when called as `sh`, runs in POSIX mode, disabling all its bash only features. From the manual - [Invoked with name sh](https://www.gnu.org/software/bash/manual/html_node/Bash-Startup-Files.html#Bash-Startup-Files)



> 
> If Bash is invoked with the name sh, it tries to mimic the startup behavior of historical versions of sh as closely as possible, while conforming to the POSIX standard as well
> 
> 
> 



```
$ ls -lrth /bin/sh
lrwxrwxrwx. 1 root root 4 Aug 26  2018 /bin/sh -> bash
$ /bin/bash -c 'set -o | grep posix'
posix           off
$ /bin/sh -c 'set -o | grep posix'
posix           on

```

with `posix` mode enabled, non-standard features like process substitution won't be enabled. See [Bash POSIX Mode](https://www.gnu.org/software/bash/manual/html_node/Bash-POSIX-Mode.html) to see its complete behavior running in the mode.


From release 5.1 of the shell, process substitutions are available in POSIX mode.


"
"How do I handle uninitialized data in a dynamic predicate in Ada?
This is some simplified code I haven't tested as is (so it may contain errors) that demonstrates the problem I'm experiencing:



```
type Space is private;

--Depending on members of Space, determines whether Outer fully contains Inner
function Contains(Outer : Space; Inner : Space);

--Outer should fully contain Inner
type Nested_Space is
record
    Inner : Space;
    Outer : Space;
end record
with Dynamic_Predicate => Contains(Outer, Inner);

```

I haven't been able to find a convenient way to initialize a Nested\_Space without failing the assert defined by the predicate. If I try to set the members of Inner first, the members of Outer are still wherever they defaulted. But if I try to set the members out Outer first, the members of Inner are still wherever they defaulted. Even if I tried forcing a default on either type, there's still no way to pick a default that will certainly be within the bounds of any arbitrary Nested\_Space.


Even trying to initialize with something like



```
declare
    My_Inner : Space := (...);
    My_Outer : Space := (...);
    My_NS : Nested_Space := (Inner => My_Inner, Outer => My_Outer);
begin
    ....
end;

```

I can't seem to keep it from failing the assert. I can come up with some pretty clunky ideas (such as adding an Initialized : Boolean to Nested\_Space specifically to check in the predicate, or alternatively setting members of the two different Spaces) but I was hoping there might be a solution that doesn't affect the structure of the record for something not required for the use case.


GNAT solutions are welcome if there's no solution in the ARM.


Thanks in advance!


","
> 
> I haven't been able to find a convenient way to initialize a Nested\_Space without failing the assert defined by the predicate. If I try to set the members of Inner first, the members of Outer are still wherever they defaulted. But if I try to set the members out Outer first, the members of Inner are still wherever they defaulted.
> 
> 
> 


[ARM 3.2.4 (35/3)](http://www.ada-auth.org/standards/rm12_w_tc1/html/RM-3-2-4.html#I1439) says, ""A Static\_Predicate, like a constraint, always remains True for all objects of the subtype, except in the case of uninitialized variables and other invalid values. A Dynamic\_Predicate, on the other hand, is checked as specified above, but can become False at other times. For example, the predicate of a record subtype is not checked when a subcomponent is modified."" You seem to be saying that this is not followed, and the record predicate is checked when you assign to a record component. If so, then you have found a compiler error. That it fails for an aggregate seems to support this idea.


However, unless you post a compilable example that demonstrates your problem, we cannot be sure that this is a compiler error.


"
"How to set correct format of DateTime?
I would like to print DateTime in format `HH:mm:ss`. 
This is current code



```
DateTime.Now.ToString(""T"");

```

this give me format: `HH:mm:ss:fff`


I found that `CultureInfo` can help, so I transform code like



```
DateTime.Now.ToString(""T"", new CultureInfo(fr-FR))

```

or



```
DateTime.Now.ToString(""T"", new CultureInfo(hr-HR))

```

with no change. Can you help me?


","[The `""T""` Format Specifier](https://msdn.microsoft.com/en-us/library/az4se3k1%28v=vs.110%29.aspx#LongTime) uses [`LongTimePattern`](https://msdn.microsoft.com/en-us/library/system.globalization.datetimeformatinfo.longtimepattern) property of the culture settings you used. And this property changes over the culture settings.


For example; it can be `h:mm:ss tt`, `H:mm:ss` or `hh:mm:ss tt` etc..


But [`InvariantCulture`](https://msdn.microsoft.com/en-us/library/system.globalization.cultureinfo.invariantculture) uses `HH:mm:ss` format which is exactly what you want. Just pass `InvariantCulture` as a second parameter in your `ToString` method and you will be fine.



```
DateTime.Now.ToString(""T"", CultureInfo.InvariantCulture);

```

Or you can use [custom date and time formatting](https://msdn.microsoft.com/en-us/library/8kb3ddd4.aspx) like;



```
DateTime.Now.ToString(""HH:mm:ss"", CultureInfo.InvariantCulture);

```

By the way, `fr-FR` culture *also* use `HH:mm:ss` as a `LongTimePattern` but `hr-HR` culture use `H:mm:ss` format.


"
"Is there a subset of assembler language which is platform-independent?
I've recently come across the `asm` reserved word in C. I want to exploit it to avoid some C semantic rules: in particular I would like to modify a variable out of a function:



```
int a = 5;
asm(""asm code to change a from 5 to, like, 6"")

int main() {
    printf(""a equal to %d\n"", a);
}

```

Now, from my computer science course I know that assembly is platform dependent, so its use should be very limited. However, I've seen some recurrent instructions, like MOV or ADD.


My questions are:


1. Although assembly is globally platform dependent, is there a subset of it that is understood by all assembler? I mean, if I were to write `asm(""MOV something somethingelse"")` and assuming MOV is inside this hypothetical subset, would the code be correct regardless of the platform?
2. If that platform independent subset exists indeed, can someone link it?
3. If there is no platform-independent subset, which are the most common assembler syntaxes?


","
> 
> 1) Although assembly is globally platform dependent, is there a subset
>  of it that is understood by all assembler? I mean, if I were to write
>  asm(""MOV something somethingelse"") and assuming MOV is inside this
>  hypothetical subset, would the code be correct regardless of the
>  platform?
> 
> 
> 


Nope. While all versions of assembly language might have a command like MOV, syntax would be different, registers would be different, and when it was assembled into machine language, the actual binary code representing that instruction would probably be different.


Think of it this way - a platform-independent subset would have to assemble down to the exact same set of bits, which each computer would have to interpret the same, (or assemblers would have to be written for all different architectures for the same assembly language). If this was the case, platform independence would be a rather trivial task.



> 
> 2) If that platform independent subset exists indeed, can someone link
>  it?
> 
> 
> 


It doesn't, so nope.



> 
> 3) If there is no platform-independent subset, which are the most
>  common assembler syntaxes?
> 
> 
> 


MASM, NASM, and GAS (especially for ARM architectures)


"
"JavaScript bitwise & operator appends 1 on unintended position
When I run `0x00000000C17E000F & 0x00000000C17E0000` on javascript, it returns `-1048707072` instead of `3246260224`.


Binary of each is


- `0x00000000C17E000F` : (32 **zeros** omitted) 1100 0001 0111 1110 0000 0000 0000 1111
- `0x00000000C17E0000` : (32 **zeros** omitted) 1100 0001 0111 1110 0000 0000 0000 0000


so I expected the result of `0x00000000C17E000F & 0x00000000C17E0000` as `3246260224`, but it results in `-1048707072`.


- `3246260224` : (32 of **0** omitted) 1100 0001 0111 1110 0000 0000 0000 0000
- `-1048707072` : (32 of **1** omitted) 1100 0001 0111 1110 0000 0000 0000 0000


Why does Javascript calculates `0 & 0` as `1` at there (in 33 to 64th bit)?


(I heard that Javascript Number consists of 64bit.)


","From [the MDN on bitwise operators](https://developer.mozilla.org/en-US/docs/Web/JavaScript/Reference/Operators/Bitwise_Operators#Bitwise_AND):



> 
> The operands are converted to 32-bit integers and expressed by a series of bits (zeroes and ones). Numbers with more than 32 bits get their most significant bits discarded.
> 
> 
> 


Your numbers aren't between -2 ^ 31 and 2 ^ 31 so they're changed by the conversion.


Note that this isn't really overflowing nor a limitation of the storage: Numbers in Javascript are 64 bits (IEEE754 doubles) which means they can store all integers between - 2 ^53 and 2 ^53. The limitation at 32 bits is just with the bitwise operators, by design.


A consequence of this remark is that you may very well design your own functions to do bitwise operations on bigger integers:



```
function and(a, b){
    let abits = Array.from(a.toString(2));
    let bbits = Array.from(b.toString(2));
    return parseInt(abits.map((abit, i)=>abit & bbits[i]).join(""""), 2);
}
console.log(and(0x00000000C17E000F, 0x00000000C17E0000)); // 3246260224

```

"
"How to verify a method inside a method is called in mockito
I was doing some unit testing in flutter with mockito, and I feels unable to verify a method is called within another method. The code I've written so far as follows,


The class I want to test



```
class A {
  void doSomething() {
    callMe();
  }

  void callMe() {}
}

```

Mocked class



```
class MockA extends Mock implements A {}

```

The test I wrote,



```
test(""Test method is called"", () {
    A a = new MockA();
    a.doSomething();
    verify(a.callMe()).called(1);
  });

```

When I run the above test I am getting an error



```
No matching calls. All calls: MockA.doSomething()
(If you called `verify(...).called(0);`, please instead use `verifyNever(...);`.)

```

If i verify doSomething is called it works, but for a call on callMe within doSomething doesn't work. Is this the default behavior or am I doing something wrong? Please note I need to verify the callMe() method is called when doSomething() is called.


","You mocked `A` and replaced it with `MockA`. Mocks have no implementation. `MockA.doSomething()` *does nothing* and does not and cannot call `MockA.callMe()`.


That `A.doSomething()` calls `A.callMe()` should be considered an implementation detail of of `doSomething()`; making a test rely on that would tightly couple the test to the specific implementation and would be brittle.


You can't use a mock to verify the implementation of the thing being mocked. If you want to verify the implementation of `A.doSomething()`, you instead should use an actual object and verify observable properties on that object.


But if you still really want to do this, then you would need to modify `A` to not call methods on itself and to instead call methods on a provided object (i.e., ""dependency injection""). For example:



```
class A {
  final late A a;

  A({A? a}) {
    this.a = a ?? this;
  }

  void doSomething() {
    a.callMe();
  }

  void callMe() {}
}

test(""Test method is called"", () {
    var mockA = MockA();
    var actualA = A(a: mockA);
    actualA.doSomething();
    verify(mockA.callMe()).called(1);
  });

```

It's a bit unusual for a class to depend on a mock of itself, however, and it would not scale if you then want to verify calls made by `callMe()`.


Another approach that would scale better (but with significantly more work) would be to create your own fake class that tracks method calls:



```
class TrackedA implements A {
  int doSomethingCallCount = 0;
  int callMeCallCount = 0;

  @override
  void doSomething() {
    doSomethingCallCount += 1;
    super.doSomething();
  }

  @override
  void callMe() {
    callMeCallCount += 1;
    super.callMe();
  }
}

```

But again, that's very brittle, and I would not recommend it.


"
"[iOS]How to make a ""snapshot"" of the current view's state
I need to make a snapshot or a screenshot - call it as you like- of the current view and then display it in the modal view. 


Because if I just write this in the ModalView's View controller



```
MyAppViewController *viewController = [[MyAppViewController alloc] init];
self.view = viewController.view;

```

All the methods of the MyAppViewController are called as well, but I don't need it, I just need to ""save"" everything that was on the screen when the ModalView appeared and show it in the ModalView's view.


How can I do it?
Thanks in advance!


","I would suggest doing this: 


Have a method that creates an image out of the contents of the view.



```
-(UIImage*) makeImage {

UIGraphicsBeginImageContext(self.view.bounds.size);

[self.view.layer renderInContext:UIGraphicsGetCurrentContext()];

UIImage *viewImage = UIGraphicsGetImageFromCurrentImageContext();
UIGraphicsEndImageContext();

return viewImage;
}

```

Create a custom init method for your modal view, and also give your modalView an instance variable that can hold a UIImage something like...



```
- (id)initWithImage:(UIImage *)temp {
   myImage = temp;
}

```

Then in your modalView, perhaps in the viewDidLoad method, create a UIImageView and set the image to `myImage`


Hopefully this achieves what you are trying to do.


"
"Creating Monoids for every subclass using Scalaz (or Shapeless)
Is it possible to create monoids for every sub-class? For example, 



```
package currency 

final case class GBP[A: Monoid](amount: A)

object Implicits {
  implicit class CurrencyOps[A: Monoid](a: A) {
    def GBP = currency.GBP(a)
  }

  implicit def gbpMonoid[A: Monoid]: Monoid[GBP[A]] = new Monoid[GBP[A]] {
    override def zero =
      GBP(Monoid[A].zero)

    override def append(f1: GBP[A], f2: => GBP[A]): GBP[A] =
      GBP(Semigroup[A].append(f1.amount, f2.amount))
  }
}

test(""GBP support plus"") {
  1.GBP |+| 2.GBP shouldBe 3.GBP // passed
}

```

I would like to add more case classes representing *currency* (e.g `USD`, `EUR`, ..)



```
sealed trait Currency
final case class GBP[A: Monoid](amount: A) extends Currency
final case class USD[A: Monoid](amount: A) extends Currency
final case class EUR[A: Monoid](amount: A) extends Currency

```

As a result, I have to implement monoids for new case classes. It's somewhat boilerplate.



```
implicit class CurrencyOps[A: Monoid](a: A) {
  def GBP = currency.GBP(a)
  def EUR = currency.EUR(a)
  def USD = currency.USD(a)
}

implicit def gbpMonoid[A: Monoid]: Monoid[GBP[A]] = new Monoid[GBP[A]] {
  override def zero =
    GBP(Monoid[A].zero)

  override def append(f1: GBP[A], f2: => GBP[A]): GBP[A] =
    GBP(Semigroup[A].append(f1.amount, f2.amount))
}

implicit def usdMonoid[A: Monoid]: Monoid[USD[A]] = new Monoid[USD[A]] {
  override def zero =
    USD(Monoid[A].zero)

  override def append(f1: USD[A], f2: => USD[A]): USD[A] =
    USD(Semigroup[A].append(f1.amount, f2.amount))
}

implicit def eurMonoid[A: Monoid]: Monoid[EUR[A]] = new Monoid[EUR[A]] {
  override def zero =
    EUR(Monoid[A].zero)

  override def append(f1: EUR[A], f2: => EUR[A]): EUR[A] =
    EUR(Semigroup[A].append(f1.amount, f2.amount))
}

```

","## Small suggestions


First I'd like to propose remove `Monoid` requirement from case classes, as they will carry implicit value in each `Currency` instance. Without this requirement your wrappers could be much more efficient and even implemented as [value classes](http://docs.scala-lang.org/overviews/core/value-classes.html):



```
  sealed trait Currency extends Any
  final case class GBP[A](amount: A) extends AnyVal with Currency
  final case class USD[A](amount: A) extends AnyVal with Currency
  final case class EUR[A](amount: A) extends AnyVal with Currency

```

## Shapeless Implementation


From here you can build simple implementation via `shapeless` as required:



```
import scalaz._
import shapeless._

implicit def monoidCurrency[A, C[_] <: Currency]
(implicit monoid: Monoid[A], gen: Generic.Aux[C[A], A :: HNil]) =
  new Monoid[C[A]] {
    def zero: C[A] = gen.from(monoid.zero :: HNil)
    def append(f1: C[A], f2: => C[A]): C[A] = {
      val x = gen.to(f1).head
      val y = gen.to(f2).head
      gen.from(monoid.append(x, y) :: HNil)
    }
  }

```

and verify it 



```
import scalaz.syntax.monoid._
import scalaz.std.anyVal._

println(2.USD |+| 3.USD) // USD(5)

```

## Further improvements


You can get rid of shapeless at all. Consider such implementations:



```
trait CurrencyUnit{
  def show(amounts: String) = s""$amounts $this""
}

final case class Currency[A, U <: CurrencyUnit](amount: A) extends AnyVal

```

`CurrencyUnit` now is not a matter of class, it's just compile-time type-tag



```
implicit case object GBP extends CurrencyUnit
implicit case object USD extends CurrencyUnit{
  override def show(amounts: String) = s""$$$amounts ""
}
implicit case object EUR extends CurrencyUnit

implicit class CurrencyOps[A](a: A) {
  def GBP = Currency[A, GBP.type](a)
  def EUR = Currency[A, EUR.type](a)
  def USD = Currency[A, USD.type](a)
}

```

which you can configure for your needs 



```
import scalaz.syntax.show._

implicit def currencyShow[A: Show, U <: CurrencyUnit](implicit unit: U) =
  new Show[Currency[A, U]] {
    override def shows(f: Currency[A, U]) = unit.show(f.amount.shows)
  }

```

And most important easily derive typeclasses via `scalaz.Isomorphism.Iso` functionality:



```
import Isomorphism._

implicit def currencyIso[A, U <: CurrencyUnit] = new (Currency[A, U] <=> A) {
  def to: (Currency[A, U]) => A = _.amount
  def from: (A) => Currency[A, U] = Currency[A, U]
}

implicit def currencyMonoid[A: Monoid, U <: CurrencyUnit] =
  new IsomorphismMonoid[Currency[A, U], A] {
    def G: Monoid[A] = implicitly
    def iso: Currency[A, U] <=> A = implicitly
  }

```

Finally you can verify this solution too



```
import scalaz.syntax.monoid._
import scalaz.std.anyVal._

println((2.USD |+| 3.USD).shows) // $5

```

"
"Order varchar string as numeric
Is it possible to order result rows by a `varchar` column cast to `integer` in Postgres 8.3?


","It's absolutely possible.



```
ORDER BY varchar_col::int

```

Be sure to have valid integer literals in your `varchar` column for each entry or you get an exception `invalid input syntax for integer`. Leading and trailing white space is ok - that's trimmed automatically.


If that's the case, though, then why not convert the column to `integer` to begin with? Smaller, faster, cleaner, simpler.


### How to avoid exceptions?


#### Postgres 16 or newer


**If** the input might not be valid integer literals, use the dedicated function [`pg_input_is_valid()`](https://www.postgresql.org/docs/16/functions-info.html#FUNCTIONS-INFO-VALIDITY) to test without raising an exception:



```
ORDER BY CASE WHEN pg_input_is_valid(varchar_col, 'integer') THEN varchar_col::int END

```

SQL `CASE` defaults to `null` in the absence of `ELSE` - which sorts last in default ascending order. See:


- [Sort by column ASC, but NULL values first?](https://stackoverflow.com/questions/9510509/sort-by-column-asc-but-null-values-first/9511492#9511492)


The new built-in function is faster and more reliable than below manual solution. Also different in that it does not try to fix broken input.


#### Postgres 15 or older


To remove non-digit characters before the cast and thereby avoid possible exceptions:



```
ORDER BY NULLIF(regexp_replace(varchar_col, '\D', '', 'g'), '')::int

```

- The [`regexp_replace()`](https://www.postgresql.org/docs/current/functions-string.html#FUNCTIONS-STRING-OTHER) expression effectively removes all non-digits, so only digits remain or an empty string. (See below.)
- `\D` is shorthand for the character class `[^[:digit:]]`, meaning all non-digits (`[^0-9]`).  

In old Postgres versions with the outdated setting `standard_conforming_strings = off`, you have to use Posix escape string syntax `E'\\D'` to escape the backslash `\`. This was default in Postgres 8.3, so you'll need that for your outdated version.
- The 4th parameter `g` is for *""globally""*, instructing to replace *all* occurrences, not just the first.
- You *may* want to allow a leading dash (`-`) for negative numbers.
- If the the string has no digits at all, the result is an empty string which is not valid for a cast to `integer`. Convert empty strings to `NULL` with [`NULLIF`](https://www.postgresql.org/docs/current/functions-conditional.html#FUNCTIONS-NULLIF). (You might consider `0` instead.)


The result is guaranteed to be valid. This procedure is for a cast to **`integer`** as requested in the body of the question, *not for `numeric`* as the title mentions.


### How to make it fast?


One way is an [index on an expression](https://www.postgresql.org/docs/current/indexes-expressional.html).



```
CREATE INDEX tbl_varchar_col2int_idx ON tbl
(cast(NULLIF(regexp_replace(varchar_col, '\D', '', 'g'), '') AS integer));

```

Then use the same expression in the `ORDER BY` clause:



```
ORDER BY
cast(NULLIF(regexp_replace(varchar_col, '\D', '', 'g'), '') AS integer)

```

Test with `EXPLAIN ANALYZE` whether the functional index actually gets used.


"
"App\_Offline in MSBuild Remote Web Deploy
I have the following task in my MSBuild script to deploy to a remote server using Web Deploy (MSDeploy service):



```
  <Target Name=""Deploy"">
    <MSBuild
            Projects=""$(SolutionFile)""
            Properties=""Configuration=Release; DeployOnBuild=True; 
              DeployTarget=MsDeployPublish; MSDeployPublishMethod=WMSvc; 
              MsDeployServiceUrl=$(DeployServiceUrl); 
              DeployIisAppPath=$(DeployIisAppPath); 
              UserName=$(DeployUserName); Password=$(DeployPassword); 
              CreatePackageOnPublish=True; AllowUntrustedCertificate=True"" />
  </Target>

```

It works fine. However, I would like to place an app\_offline.htm file (on the remote server) before the app is deployed and remove the app\_offline.htm file after deployment (or on error). Is there a MSBuild property or any other scripting tweak to make it happen?


Thanks in advance.


","I just recently blogged about this at <http://sedodream.com/2012/01/08/HowToTakeYourWebAppOfflineDuringPublishing.aspx>. It's more difficult than it should be and I'm working on simplifying that for a later version. In any case I've pasted all the content here for you.


I received a customer email asking how they can take their web application/site offline for the entire duration that a publish is happening from Visual Studio. An easy way to take your site offline is to drop an app\_offline.htm file in the sites root directory. For more info on that you can read ScottGu’s post, link in below in resources section. Unfortunately Web Deploy itself doesn’t support this . If you want Web Deploy (aka MSDeploy) to natively support this feature please vote on it at <http://aspnet.uservoice.com/forums/41199-general/suggestions/2499911-take-my-site-app-offline-during-publishing>.


Since Web Deploy doesn’t support this it’s going to be a bit more difficult and it requires us to perform the following steps:


1. Publish app\_offline.htm
2. Publish the app, and ensure that app\_offline.htm is contained inside the payload being published
3. Delete app\_offline.htm


**1** will take the app offline before the publish process begins. 


**2** will ensure that when we publish that app\_offline.htm is not deleted (and therefore keep the app offline) 


**3** will delete the app\_offline.htm and bring the site back online


Now that we know what needs to be done let’s look at the implementation. First for the easy part. Create a file in your Web Application Project (WAP) named app\_offline-template.htm. This will be the file which will end up being the app\_offline.htm file on your target server. If you leave it blank your users will get a generic message stating that the app is offline, but it would be better for you to place **static HTML** (no ASP.NET markup) inside of that file letting users know that the site will come back up and whatever other info you think is relevant to your users. When you add this file you should change the Build Action to None in the Properties grid. This will make sure that this file itself is not published/packaged. Since the file ends in .htm it will by default be published. See the image below.


![enter image description here](https://i.stack.imgur.com/IzAgq.png)


Now for the hard part. For Web Application Projects we have a hook into the publish/package process which we refer to as “wpp.targets”. If you want to extend your publish/package process you can create a file named {ProjectName}.wpp.targets in the same folder as the project file itself. Here is the file which I created you can copy and paste the content into your wpp.targets file. I will explain the significant parts but wanted to post the entire file for your convince. Note: you can grab my latest version of this file from my github repo, the link is in the resource section below.



```
<?xml version=""1.0"" encoding=""utf-8""?>
<Project xmlns=""http://schemas.microsoft.com/developer/msbuild/2003"">
  <Target Name=""InitalizeAppOffline"">
    <!-- 
    This property needs to be declared inside of target because this is imported before
    the MSDeployPath property is defined as well as others -->
    <PropertyGroup>
      <MSDeployExe Condition="" '$(MSDeployExe)'=='' "">$(MSDeployPath)msdeploy.exe</MSDeployExe>
    </PropertyGroup>    
  </Target>

  <PropertyGroup>
    <PublishAppOfflineToDest>
      InitalizeAppOffline;
    </PublishAppOfflineToDest>
  </PropertyGroup>

  <!--
    %msdeploy% 
      -verb:sync 
      -source:contentPath=""C:\path\to\app_offline-template.htm"" 
      -dest:contentPath=""Default Web Site/AppOfflineDemo/app_offline.htm""
  -->

  <!--***********************************************************************
  Make sure app_offline-template.htm gets published as app_offline.htm
  ***************************************************************************-->
  <Target Name=""PublishAppOfflineToDest"" 
          BeforeTargets=""MSDeployPublish"" 
          DependsOnTargets=""$(PublishAppOfflineToDest)"">
    <ItemGroup>
      <_AoPubAppOfflineSourceProviderSetting Include=""contentPath"">
        <Path>$(MSBuildProjectDirectory)\app_offline-template.htm</Path>
        <EncryptPassword>$(DeployEncryptKey)</EncryptPassword>
        <WebServerAppHostConfigDirectory>$(_MSDeploySourceWebServerAppHostConfigDirectory)</WebServerAppHostConfigDirectory>
        <WebServerManifest>$(_MSDeploySourceWebServerManifest)</WebServerManifest>
        <WebServerDirectory>$(_MSDeploySourceWebServerDirectory)</WebServerDirectory>
      </_AoPubAppOfflineSourceProviderSetting>

      <_AoPubAppOfflineDestProviderSetting Include=""contentPath"">
        <Path>""$(DeployIisAppPath)/app_offline.htm""</Path>
        <ComputerName>$(_PublishMsDeployServiceUrl)</ComputerName>
        <UserName>$(UserName)</UserName>
        <Password>$(Password)</Password>
        <EncryptPassword>$(DeployEncryptKey)</EncryptPassword>
        <IncludeAcls>False</IncludeAcls>
        <AuthType>$(AuthType)</AuthType>
        <WebServerAppHostConfigDirectory>$(_MSDeployDestinationWebServerAppHostConfigDirectory)</WebServerAppHostConfigDirectory>
        <WebServerManifest>$(_MSDeployDestinationWebServerManifest)</WebServerManifest>
        <WebServerDirectory>$(_MSDeployDestinationWebServerDirectory)</WebServerDirectory>
      </_AoPubAppOfflineDestProviderSetting>
    </ItemGroup>

    <MSdeploy
          MSDeployVersionsToTry=""$(_MSDeployVersionsToTry)""
          Verb=""sync""
          Source=""@(_AoPubAppOfflineSourceProviderSetting)""
          Destination=""@(_AoPubAppOfflineDestProviderSetting)""
          EnableRule=""DoNotDeleteRule""
          AllowUntrusted=""$(AllowUntrustedCertificate)""
          RetryAttempts=""$(RetryAttemptsForDeployment)""
          SimpleSetParameterItems=""@(_AoArchivePublishSetParam)""
          ExePath=""$(MSDeployPath)"" />
  </Target>

  <!--***********************************************************************
  Make sure app_offline-template.htm gets published as app_offline.htm
  ***************************************************************************-->
  <!-- We need to create a replace rule for app_offline-template.htm->app_offline.htm for when the app get's published -->
  <ItemGroup>
    <!-- Make sure not to include this file if a package is being created, so condition this on publishing -->
    <FilesForPackagingFromProject Include=""app_offline-template.htm"" Condition="" '$(DeployTarget)'=='MSDeployPublish' "">
      <DestinationRelativePath>app_offline.htm</DestinationRelativePath>
    </FilesForPackagingFromProject>

    <!-- This will prevent app_offline-template.htm from being published -->
    <MsDeploySkipRules Include=""SkipAppOfflineTemplate"">
      <ObjectName>filePath</ObjectName>
      <AbsolutePath>app_offline-template.htm</AbsolutePath>
    </MsDeploySkipRules>
  </ItemGroup>

  <!--***********************************************************************
  When publish is completed we need to delete the app_offline.htm
  ***************************************************************************-->
  <Target Name=""DeleteAppOffline"" AfterTargets=""MSDeployPublish"">
    <!--
    %msdeploy% 
      -verb:delete 
      -dest:contentPath=""{IIS-Path}/app_offline.htm"",computerName=""..."",username=""..."",password=""...""
    -->
    <Message Text=""************************************************************************"" />
    <Message Text=""Calling MSDeploy to delete the app_offline.htm file"" Importance=""high"" />
    <Message Text=""************************************************************************"" />

    <ItemGroup>
      <_AoDeleteAppOfflineDestProviderSetting Include=""contentPath"">
        <Path>$(DeployIisAppPath)/app_offline.htm</Path>
        <ComputerName>$(_PublishMsDeployServiceUrl)</ComputerName>
        <UserName>$(UserName)</UserName>
        <Password>$(Password)</Password>
        <EncryptPassword>$(DeployEncryptKey)</EncryptPassword>
        <AuthType>$(AuthType)</AuthType>
        <WebServerAppHostConfigDirectory>$(_MSDeployDestinationWebServerAppHostConfigDirectory)</WebServerAppHostConfigDirectory>
        <WebServerManifest>$(_MSDeployDestinationWebServerManifest)</WebServerManifest>
        <WebServerDirectory>$(_MSDeployDestinationWebServerDirectory)</WebServerDirectory>
      </_AoDeleteAppOfflineDestProviderSetting>
    </ItemGroup>

    <!-- 
    We cannot use the MSDeploy/VSMSDeploy tasks for delete so we have to call msdeploy.exe directly.
    When they support delete we can just pass in @(_AoDeleteAppOfflineDestProviderSetting) as the dest
    -->
    <PropertyGroup>
      <_Cmd>""$(MSDeployExe)"" -verb:delete -dest:contentPath=""%(_AoDeleteAppOfflineDestProviderSetting.Path)""</_Cmd>
      <_Cmd Condition="" '%(_AoDeleteAppOfflineDestProviderSetting.ComputerName)' != '' "">$(_Cmd),computerName=""%(_AoDeleteAppOfflineDestProviderSetting.ComputerName)""</_Cmd>
      <_Cmd Condition="" '%(_AoDeleteAppOfflineDestProviderSetting.UserName)' != '' "">$(_Cmd),username=""%(_AoDeleteAppOfflineDestProviderSetting.UserName)""</_Cmd>
      <_Cmd Condition="" '%(_AoDeleteAppOfflineDestProviderSetting.Password)' != ''"">$(_Cmd),password=$(Password)</_Cmd>
      <_Cmd Condition="" '%(_AoDeleteAppOfflineDestProviderSetting.AuthType)' != ''"">$(_Cmd),authType=""%(_AoDeleteAppOfflineDestProviderSetting.AuthType)""</_Cmd>
    </PropertyGroup>

    <Exec Command=""$(_Cmd)""/>
  </Target>  
</Project>

```

## 1 Publish app\_offline.htm


The implementation for #1 is contained inside the target PublishAppOfflineToDest. The msdeploy.exe command that we need to get executed is.



```
msdeploy.exe 
    -source:contentPath='C:\Data\Personal\My Repo\sayed-samples\AppOfflineDemo01\AppOfflineDemo01\app_offline-template.htm' 
    -dest:contentPath='""Default Web Site/AppOfflineDemo/app_offline.htm""',UserName='sayedha',Password='password-here',ComputerName='computername-here',IncludeAcls='False',AuthType='NTLM' -verb:sync -enableRule:DoNotDeleteRule

```

In order to do this I will leverage the MSDeploy task. Inside of the PublishAppOfflineToDest target you can see how this is accomplished by creating an item for both the source and destination.


## 2 Publish the app, and ensure that app\_offline.htm is contained inside the payload being published


This part is accomplished by the fragment



```
<!--***********************************************************************
Make sure app_offline-template.htm gets published as app_offline.htm
***************************************************************************-->
<!-- We need to create a replace rule for app_offline-template.htm->app_offline.htm for when the app get's published -->
<ItemGroup>
  <!-- Make sure not to include this file if a package is being created, so condition this on publishing -->
  <FilesForPackagingFromProject Include=""app_offline-template.htm"" Condition="" '$(DeployTarget)'=='MSDeployPublish' "">
    <DestinationRelativePath>app_offline.htm</DestinationRelativePath>
  </FilesForPackagingFromProject>

  <!-- This will prevent app_offline-template.htm from being published -->
  <MsDeploySkipRules Include=""SkipAppOfflineTemplate"">
    <ObjectName>filePath</ObjectName>
    <AbsolutePath>app_offline-template.htm</AbsolutePath>
  </MsDeploySkipRules>
</ItemGroup>

```

The item value for FilesForPackagingFromProject here will convert your app\_offline-template.htm to app\_offline.htm in the folder from where the publish will be processed. Also there is a condition on it so that it only happens during publish and not packaging. We do not want app\_offline-template.htm to be in the package (but it’s not the end of the world if it does either).


The element for MsDeploySkiprules will make sure that app\_offline-template.htm itself doesn’t get published. This may not be required but it shouldn’t hurt.


## 3 Delete app\_offline.htm


Now that our app is published we need to delete the app\_offline.htm file from the dest web app. The msdeploy.exe command would be:


%msdeploy% 
 -verb:delete 
 -dest:contentPath=""{IIS-Path}/app\_offline.htm"",computerName=""..."",username=""..."",password=""..."" 


This is implemented inside of the DeleteAppOffline target. This target will automatically get executed after the publish because I have included the attribute AfterTargets=”MSDeployPublish”. In that target you can see that I am building up the msdeploy.exe command directly, it looks like the MSDeploy task doesn’t support the delete verb.


If you do try this out please let me know if you run into any issues. I am thinking to create a Nuget package from this so that you can just install that package. That would take a bit of work so please let me know if you are interested in that.


## Resources


1. [The latest version of my AppOffline wpp.targets file](https://github.com/sayedihashimi/sayed-samples/blob/master/AppOfflineDemo01/AppOfflineDemo01/AppOfflineDemo01.wpp.targets)
2. [ScottGu’s blog on app\_offline.htm](http://weblogs.asp.net/scottgu/archive/2006/04/09/442332.aspx)


"
"re matches none: Differences in Python implementation of regex?
I'm having some trouble matching a specific pattern using the python regex library (`re`). I'm trying to match lines with a number (of up to 3 digits), followed a collection of words (with no space between the first word and the number) which are terminated by exactly two spaces. Some examples, with the matching string enclosed in parentheses:


`test(58your own becoming )Adapted from Pyramid Text utterance 81.`


`(46ancestral fires )In Sumerian, a language recently supplanted by`


`(45lap of God )Ginzberg, Legends of the Bible, p. 1.`


`(9Island of the Egg )The symbolism of the cosmic egg is an integral aspect of almost every mythological tradition. In the`


I'm using the following expression:



```
(\d+).+(  )

```

The relevant python code is as follows:



```
# the search string is `tmp`
pattern = re.compile(""(\d+).+(  )"")
footnotes = pattern.finditer(tmp)
for footnote in footnotes:
    # do something with each match

```

When I use a testing site like regexr, all the above examples match exactly as intended. However, python matches none. Is there something simple I'm missing? I've also tried passing the expression to `re` as a raw string. I can't seem to find anything else to try in the documentation. Any help would be greatly appreciated!


# EDIT


The full string can be found [here](https://raw.githubusercontent.com/rosslaird/stone/master/stone.md?token=AA8QrJsY8fTR7zxEcL-U-lyJh8jGKaNlks5ZkmxFwA%3D%3D).


At this point, I'm fairly certain it has something to do with how I'm handling the string. If I read from a text file, and execute the following code, the output is empty:



```
with open(""stone.md"", ""r+"") as f:
    tmp = f.read()
    pattern = re.compile(r""(\d+).+  "")
    footnotes = pattern.finditer(tmp)
    for footnote in footnotes:
        print tmp[footnote.start():footnote.end()]

```

But, If I run:



```
tmp = """"""test58your own becoming  Adapted from Pyramid Text utterance 81.""""""
pattern = re.compile(r""(\d+).+  "")
footnotes = pattern.finditer(tmp)
for footnote in footnotes:
    print tmp[footnote.start():footnote.end()]

```

I get `58your own becoming`


","## You've fallen victim to Unicode homoglyphs.


Your regex contains ASCII-encoded space characters (the regular spaces you are used to). However, the full text that you are operating on contain non-breaking spaces, which in HTML is `&nbsp;` and in Unicode `U+00A0`. It looks exactly like a regular space to the human eye, but it isn't an ASCII space.



```
Python 3.6.2 (default, Jul 20 2017, 03:52:27) 
[GCC 7.1.1 20170630] on linux
Type ""help"", ""copyright"", ""credits"" or ""license"" for more information.
>>> '  '.encode('ascii')
b'  '
>>> '  '.encode('ascii')
Traceback (most recent call last):
  File ""<stdin>"", line 1, in <module>
UnicodeEncodeError: 'ascii' codec can't encode characters in position 0-1: ordinal not in range(128)
>>> '  '.encode('utf-8')
b'\xc2\xa0\xc2\xa0'

```

The following regex will give you what you want:


`pattern = re.compile(b'(\d+).+(\xc2\xa0)'.decode('utf-8'))`


What this is doing is constructing a bytes object then decoding it to a utf-8 string so `re` can work with it.


Or, even better, you can use `\s`, which matches any whitespace character (covers Unicode) in the regex flavor you're using:


`pattern = re.compile('(\d+).+(\s\s)')`


## Why then, did the regex in your answer appear to work?


**Because browsers render the non-breaking space as an ASCII space**, which propagates to the browser copy paste buffer as an ASCII space.


I was only able to discover this once you had disclosed the original text file you were working on. I downloaded the raw format with `wget` on the URL, which preserved the Unicode spaces in the original file, something that would not have happened had I copy pasted your large text file in the browser to a file on my local computer.


Wow. This was a really fun puzzle to solve. Thanks for the question.


"
"How do I check if a string has only alphanumeric characters in guava?
Actually, I would like to check if it has only alphanumeric plus these: .-\_ [space]


If using an external library I would like to use Guava since it's already included in my project...


","Regexp-free and somewhat more readable (?) solution could be using Guava's [`CharMatcher`](https://google.github.io/guava/releases/snapshot/api/docs/com/google/common/base/CharMatcher.html) class:



```
boolean matched = CharMatcher.JAVA_LETTER_OR_DIGIT.or(CharMatcher.WHITESPACE)
    .or(CharMatcher.anyOf(""_.-"")).matchesAllOf(yourString);

```

Maybe ASCII chars are OK for your use case? If so, use:



```
boolean matched = CharMatcher.ASCII.matchesAllOf(yourString);

```

See [wiki page](https://code.google.com/p/guava-libraries/wiki/StringsExplained#CharMatcher) for examples and description. 


Also, you may want to extract your matcher to constant and precompute it:



```
private static final CharMatcher CHAR_MATCHER = CharMatcher.JAVA_LETTER_OR_DIGIT
    .or(CharMatcher.WHITESPACE)
    .or(CharMatcher.anyOf(""_.-""))
    .precomputed();

```

What's more interesting, if you read `CharMatcher`'s documentation you may find that [""digit""](https://google.github.io/guava/releases/snapshot/api/docs/com/google/common/base/CharMatcher.html#JAVA_DIGIT), [""letter""](https://google.github.io/guava/releases/snapshot/api/docs/com/google/common/base/CharMatcher.html#JAVA_LETTER) and [""whitespace""](https://google.github.io/guava/releases/snapshot/api/docs/com/google/common/base/CharMatcher.html#WHITESPACE) in Java are quite ambigious terms:



> 
> Determines whether a character is a digit according to Java's definition. If you only care to match ASCII digits, you can use `inRange('0', '9')`.
> 
> 
> 


or 



> 
> Determines whether a character is a letter according to Java's definition. If you only care to match letters of the Latin alphabet, you can use `inRange('a', 'z').or(inRange('A', 'Z'))`.
> 
> 
> 


so you may want use explicit char ranges:



```
private static final CharMatcher CHAR_MATCHER_ASCII = 
    CharMatcher.inRange('0', '9')
        .or(CharMatcher.inRange('a', 'z'))
        .or(CharMatcher.inRange('A', 'Z'))
        .or(CharMatcher.anyOf("" _.-"")) // note space here
        .precomputed();

```

"
"Combine two potentially null lists with linq
I have two lists:



```
    var myIds = new List<int>()
        {
            1, 2, 3
        };

    var yourIds = new List<int>()
        {
            2, 3, 4
        };

```

The two lists can be combined into one like this:



```
myIds.Union(yourIds)
.Select(x => new 
             { 
                 Id = x,
                 Mine = myIds.Contains(x), 
                 Yours = yourIds.Contains(x)
             });

```

The new list would look like this:



```
Id    Mine     Yours
---------------------
1      T         F
2      T         T
3      T         T
4      F         T

```

This works great when the lists contain elements. However, what if the lists have the possibility of being null. How could I handle the null lists?


","If the lists could be `null`, then you have to check for `null`. You can also use null coalescing operator and if any of the list is `null` then return an empty list *(internally [it](http://msdn.microsoft.com/en-us/library/ms173224.aspx) will also check for null)*. 


You can do:



```
(mIds ?? Enumerable.Empty<int>()).Union(yourIds ?? Enumerable.Empty<int>())

```

*(thanks to @phoog [comment](https://stackoverflow.com/questions/25789121/combine-two-potentially-null-lists-with-linq/25789206?noredirect=1#comment40335029_25789206))*


**But IMO, it is better to check against `null`**


But if any of the list is empty, then you don't have to do any check, There will be no exception. Even if both the lists are empty , you will get an empty result back. The above code would return an empty list if any of the list is `null`. 


"
"ASP.NET WebAPI Conditional Serialization based on User Role
I have an ORM (NHibernate) that maps to POCOs which will be returned in ApiControllers. I realize that JSON.NET allows me to put conditional serialization methods (ShouldSerialize\*) on my models; however, these models and their methods have no knowledge of anything about their environment, nor should they. What I would like to do is conditionally serialize a model or one or more of its properties based on the user's role when they're signed into my website. I can conceptually perceive how this can be done but I'm lost at one part. Here's a sample model:



```
public class SomeModel
{
    public string SomeProperty { get; set; }

    [Sensitive]
    public string SomeOtherProperty { get; set; }
}

```

I would like to be able to put an attribute on a property to flag it as ""Sensitive"". Then in my WebApi when it is serializing it for output, I would like for it to check the model for this attribute and, if it exists, it should check the user's role. If the user is in the specified role, then the serializer should serialize the attribute, otherwise it will either mask it out or simply not serialize it. So would I have to write my own custom formatter to handle this or is there a way to hook into the built-in ones to perform this check? Or am I too limited in my thinking, and there's another way to handle this?


I did think that another way this could be handled would be at the ORM level but couldn't find good examples online.


Much appreciated!


EDIT: I did find another similar question here:
[Contextual serialization from WebApi endpoint based on permissions](https://stackoverflow.com/questions/16172076/contextual-serialization-from-webapi-endpoint-based-on-permissions)
but there was no solution. Also, I don't like the idea of setting role-based access in the models via attributes. I believe that should be handled in the web application.


","As mentioned in the question I had raised the question ""[Contextual serialization from WebApi endpoint based on permissions](https://stackoverflow.com/questions/16172076/contextual-serialization-from-webapi-endpoint-based-on-permissions)"" which I have [answered](https://stackoverflow.com/a/17472174/227200) myself. I did initially look at using a [`MediaFormatter`](http://msdn.microsoft.com/en-us/library/system.net.http.formatting.mediatypeformatter%28v=vs.108%29.aspx) but I believe this would then restrict you to what kind of response you could return. If you wanted to return JSON and XML you would need to implement two formatters. To only have to implement the filter in one place you need to go higher up the stack to the[`DelegatingHandler`](http://msdn.microsoft.com/en-us/library/system.net.http.delegatinghandler.aspx).


In my implementation I wanted to look up which fields the client had access to and remove any from the response that the client should not see. This is quite similar to what you want to do.


In your senario you would need to reflect over the object and pick out any fields that contain your attribute and set those values to null. If returning JSON or XML then the properties are not included in the response if the values are null so you will not even leak the property name.


## Example


Here is an example implementation of a `DelegatingHandler` that uses reflection to filter out properties on the response object content that implement the `Sensitive` attribute. It assumes the object hierarchy is flat so if you have nested objects you will need to navigate the object graph and do the same for each object in the hierarchy.



```
public class ResponseDataFilterHandler : DelegatingHandler
{
    protected override System.Threading.Tasks.Task<HttpResponseMessage> SendAsync(HttpRequestMessage request, CancellationToken cancellationToken)
    {
        return base.SendAsync(request, cancellationToken)
            .ContinueWith(task =>
            {
                var response = task.Result;

                //Manipulate content here
                var content = response.Content as ObjectContent;
                if (content != null && content.Value != null)
                {
                    FilterFields(content.Value);
                }

                return response;
            });
    }

    private void FilterFields(object objectToFilter)
    {
        var properties = objectToFilter
                             .GetType()
                             .GetProperties(
                                 BindingFlags.IgnoreCase |
                                 BindingFlags.GetProperty |
                                 BindingFlags.Instance |
                                 BindingFlags.Public);

        foreach (var propertyInfo in properties)
        {
            if (propertyInfo.GetCustomAttributes(typeof(SensitiveAttribute), true).Any())
            {
                propertyInfo.SetValue(objectToFilter, null, new object[0]);
            }
        }   
    }
}

```

"
"Stemming words using tm package in R does not work properly?
I am doing some text mining (PCA, HC, K-Means) and so far I have managed to code everything right. However, there is a small flaw I'd like to fix.


When I try to stem my Corpus it does not work properly as there are different words with the same radical which aren't identified in the correct way. 
These are the words I am particularly interested in (it's in Spanish and they mean ""kids"" or related):



```
niñera, niños, niñas, niña, niño

```

But when I run the code I get that these words are still the same except for 



```
niña, niño --> niñ 

```

But the other remain the same so I end up only stemming for niña/niño but not for the others.


This is my code for creating the corpus:



```
corp <- Corpus(DataframeSource(data.frame(x$service_name)))
docs <- tm_map(corp, removePunctuation)
docs <- tm_map(docs, removeNumbers) 
docs <- tm_map(docs, tolower)
docs <- tm_map(docs, removeWords, stopwords(""spanish""))
docs <- tm_map(docs, stemDocument, language = ""spanish"") 
docs <- tm_map(docs, PlainTextDocument) 
dtm <- DocumentTermMatrix(docs)   
dtm  

```

I'd really appreciate some suggestions!
Thank you


","It seems that the stemming transform can only be applied to PlainTextDocument types. See `? stemDocument`. 



```
sp.corpus = Corpus(VectorSource(c(""la niñera. los niños. las niñas. la niña. el niño."")))
docs <- tm_map(sp.corpus, removePunctuation)
docs <- tm_map(docs, removeNumbers) 
docs <- tm_map(docs, tolower)
docs <- tm_map(docs, removeWords, stopwords(""spanish""))
docs <- tm_map(docs, PlainTextDocument)  # needs to come before stemming
docs <- tm_map(docs, stemDocument, ""spanish"")
print(docs[[1]]$content)

# "" niñer  niñ  niñ  niñ  niñ""

```

versus



```
# WRONG
sp.corpus = Corpus(VectorSource(c(""la niñera. los niños. las niñas. la niña. el niño."")))
docs <- tm_map(sp.corpus, removePunctuation)
docs <- tm_map(docs, removeNumbers) 
docs <- tm_map(docs, tolower)
docs <- tm_map(docs, removeWords, stopwords(""spanish""))
docs <- tm_map(docs, stemDocument, ""spanish"")  # WRONG: apply PlainTextDocument first
docs <- tm_map(docs, PlainTextDocument)  
print(docs[[1]]$content)

# "" niñera  niños  niñas  niña  niñ""

```

In my opinion, this detail is not obvious and it'd be nice to get at least a warning when stemDocument is invoked on a non-PlainTextDocument.


"
"Reinstalled Firefox. It's not a fresh install. Why?
My firefox started to get really sluggish, crashing sometimes and CPU usage was way too high. It was at 30-55% the last time I checked. I have a Thinkpad T450s, so there shouldn't be any issues with performance and resources. The T450s got an i7, 12GB RAM and a 512GB SSD. I just uninstalled Firefox, downloaded the newest version and installed it. The first thing to greet me were my old tabs. Everything was like nothing happened! Every bookmark, the history and all cookies are still present. What happened?


","Your personal information is stored within the following location: 



> 
> %AppData%\Mozilla\Firefox\Profiles\ 
> 
> 
> 


When you remove and reinstall Firefox, those files and folders with your profile settings are left intact. Although you actually *did* successfully replace the Firefox program files, the things that you see are still retained within your Windows user profile.


I saw the same thing happen when the 64-bit version of Firefox was first released. I removed the 32-bit version of Firefox from my Windows 10 Pro x64 machine, and when I installed the 64-bit version of Mozilla's browser it made for a very smooth transition. The following information was still there: 



> 
> 1. Bookmarks
> 2. Browsing and download history
> 3. Passwords
> 4. Open windows and tabs
> 5. Cookies
> 6. Web form auto-fill information
> 7. Personal dictionary
> 
> 
> 


For future reference, should you need to do this again it would probably be a bit faster and easier to utilize the built-in **Refresh Firefox** feature, instead. In your Firefox address bar enter `about:support` and then select the `Refresh Firefox` button on the upper-right: 


[![Refresh Firefox](https://i.stack.imgur.com/f60em.jpg)](https://i.stack.imgur.com/f60em.jpg)


**Sources**:  

[Profiles - Where Firefox stores your bookmarks, passwords and other user data](https://support.mozilla.org/en-US/kb/profiles-where-firefox-stores-user-data)  

[Refresh Firefox - reset add-ons and settings](https://support.mozilla.org/en-US/kb/refresh-firefox-reset-add-ons-and-settings)


"
"Create an array or List of all dates between two dates
I am generating multi-series graphs with the date along the X-Axis. 


The problem is that not all of the series in the graph have the same dates in the date range. Meaning that if I choose 1 Feb through 30 Apr that one series may have data that starts at 1 Feb but only goes through the end of March but another series may have data for the entire date range. 


This skews the charts I need to create. Go, given the date range taken at the begining of the query I'd like to generate a list of dates and populate the data to be graphed, padding those series with 0's for those dates that have no data.


","**LINQ:**



```
Enumerable.Range(0, 1 + end.Subtract(start).Days)
          .Select(offset => start.AddDays(offset))
          .ToArray(); 

```

**For loop:**



```
var dates = new List<DateTime>();

for (var dt = start; dt <= end; dt = dt.AddDays(1))
{
   dates.Add(dt);
}

```

EDIT:
As for padding values with defaults in a time-series, you could enumerate all the dates in the full date-range, and pick the value for a date directly from the series if it exists, or the default otherwise. For example:



```
var paddedSeries = fullDates.ToDictionary(date => date, date => timeSeries.ContainsDate(date) 
                                               ? timeSeries[date] : defaultValue);

```

"
"How to construct different object according to their types in a template function?
Here is my question, I have a base class `Base`, two derived classes `ClassA` and `ClassB` form `Base` taking only two arguments, and two classes `ClassB0` and `ClassB1` derived from `ClassB` taking three arguments. I want to create class intance accroding to the type passed to the function `create`, if the type is derived from `ClassB`, the first argument will be filled with 10. The compiler always warn that there is no matching constructors.



```
#include <iostream>
#include <type_traits>

class Base {};
class ClassA : public Base {
public:
  ClassA(int a, int b) : Base() {
    std::cout << ""ClassA: "" << a << "" "" << b << ""\n\n"";
  }
};

class ClassB : public Base {
public:
  ClassB(int a, int b, int c) : Base() {
    std::cout << ""ClassB: "" << a << "" "" << b << "" "" << c << ""\n"";
  }
};

class ClassB0 : public ClassB {
public:
  ClassB0(int a, int b, int c) : ClassB(a, b, c) { 
    std::cout << ""ClassB0: "" << a << "" "" << b << "" "" << c << ""\n\n"";
  }
};
class ClassB1 : public ClassB {
public:
  ClassB1(int a, int b, int c) : ClassB(a, b, c) {
    std::cout << ""ClassB1: "" << a << "" "" << b << "" "" << c << ""\n\n"";
  }
};

template <typename T, typename ...Args>
T* create(Args&&... args) {
  T* comp = nullptr;

  if (std::is_base_of<ClassB, T>::value) {
    std::cout << ""True "";
    comp = new T(10, std::forward<Args>(args)...);
  } else {
    std::cout << ""False "";
    comp = new T(std::forward<Args>(args)...);
  }

  return comp;
}

int main() {
  create<ClassA>(1, 2);
  create<ClassB0>(2, 3);
  create<ClassB1>(2, 3);
}

```

","The problem is that both the branch of `if` and `else` need to be evaluated at compile-time, despite of which one would be evaluated at run-time.


You can apply [constexpr if statement](https://en.cppreference.com/w/cpp/language/if#Constexpr_If) (since C++17), for which the condition must be known at compile-time, and either the `if` or `else` branch would be discarded and won't be evaluated at compile-time again.



> 
> If the value is `true`, then *statement-false* is discarded (if present), otherwise, *statement-true* is discarded.
> 
> 
> 



```
template <typename T, typename ...Args>
T* create(Args&&... args) {
  T* comp = nullptr;

  if constexpr (std::is_base_of<ClassB, T>::value) {
  // ^^^^^^^^^
    std::cout << ""True "";
    comp = new T(10, std::forward<Args>(args)...);
  } else {
    std::cout << ""False "";
    comp = new T(std::forward<Args>(args)...);
  }

  return comp;
}

```

[LIVE](https://wandbox.org/permlink/tPwI3vswtNIcHxH3)


Before C++17, you can apply overloading with [SFINAE](https://en.cppreference.com/w/cpp/language/sfinae). e.g.



```
template <typename T, typename ...Args>
typename std::enable_if<std::is_base_of<ClassB, T>::value, T*>::type
create(Args&&... args) {
  T* comp = nullptr;
  std::cout << ""True "";
  comp = new T(10, std::forward<Args>(args)...);
  return comp;
}
template <typename T, typename ...Args>
typename std::enable_if<!std::is_base_of<ClassB, T>::value, T*>::type
create(Args&&... args) {
  T* comp = nullptr;
  std::cout << ""False "";
  comp = new T(std::forward<Args>(args)...);
  return comp;
}

```

[LIVE](https://wandbox.org/permlink/59BBD4cKznUCe36t)


"
"Java: is there a map function?
I need a [map](http://docs.python.org/library/functions.html#map) function. Is there something like this in Java already?


(For those who wonder: I of course know how to implement this trivial function myself...)


","There is no notion of a function in the JDK as of java 6.


[Guava](https://github.com/google/guava) has a [Function](http://google.github.io/guava/releases/19.0/api/docs/com/google/common/base/Function.html) interface though and the  

[`Collections2.transform(Collection<E>, Function<E,E2>)`](http://google.github.io/guava/releases/19.0/api/docs/com/google/common/collect/Collections2.html#transform%28java.util.Collection,%20com.google.common.base.Function%29)  

method provides the functionality you require.


**Example:**



```
// example, converts a collection of integers to their
// hexadecimal string representations
final Collection<Integer> input = Arrays.asList(10, 20, 30, 40, 50);
final Collection<String> output =
    Collections2.transform(input, new Function<Integer, String>(){

        @Override
        public String apply(final Integer input){
            return Integer.toHexString(input.intValue());
        }
    });
System.out.println(output);

```

**Output:**



```
[a, 14, 1e, 28, 32]

```



---


These days, with Java 8, there is actually a map function, so I'd probably write the code in a more concise way:



```
Collection<String> hex = input.stream()
                              .map(Integer::toHexString)
                              .collect(Collectors::toList);

```

"
"PHP: Check if it has been one week since timestamp
Let's assume:



```
$time = '2010-05-17 02:49:30' // (retrieved from MySQL TIMESTAMP field)

```

How do I do the following in PHP:


1) Check if it has been more than one week since this time has passed?


2) Assuming ""false"" on (1), find out how much more time until the one week mark, rounded to days and hours remaining.


I know this is pretty straightforward, but it uses a very specific syntax. Having never played with time calculations before, I'd appreciate some guidance.


Thanks!


","
```
$time = strtotime('2010-05-10 02:49:30');
$one_week_ago = strtotime('-1 week');

if( $time > $one_week_ago ) { 
    // it's sooner than one week ago
    $time_left = $time - $one_week_ago;
    $days_left = floor($time_left / 86400); // 86400 = seconds per day
    $hours_left = floor(($time_left - $days_left * 86400) / 3600); // 3600 = seconds per hour
    echo ""Still $days_left day(s), $hours_left hour(s) to go."";
}

```

"
"Changing layout of view in Joomla 2.5
I know there are several similar topics around but I read and tried most of them but still can't figure out how to do this.


I have a written a component in Joomla 2.5 and it works so far. I have different views and I can load the views using the controller.php. 
One of the views shows a table out of my data base (data about teams). 


Now I'd like to have another layout of the same view which would display the data base table as a form so can change the content.


That's the file structure:


views/  

- teams/  

- - tmpl/  

- - - default.php  

- - - modify.php  

- - view.html.php 


That's out of the view.html.php file:



```
...
// Overwriting JView display method
function display($tpl = null) {

    ...

    $this->setLayout('modify');
    echo $this->getLayout();
    // Display the view
    parent::display($tpl);
}

```

I tried different combinations of setLayout, $tpl = ..., default\_modify.php, etc.
but I always either get the default layout or some error like 'can't find layout modify'


I load the site with .../index.php?option=com\_test&task=updateTeams


And the controller.php looks like this:



```
function updateTeams(){
    $model = $this->getModel('teams');
    $view = $this->getView('teams','html');
    $view->setModel($model);

    $view->display();
}

```

","I had a similar problem, I created some kind of user profile view and wanted them to be able to edit the fields without having to create a new model for it (would have similar functions, hate redundancy...). What worked for me is to simply call the layout like this:


index.php?option=com\_mycomponent&view=myview&**layout=edit** (""edit"" would be ""modify"" in your case)


To do this I didn't touch the view.html.php (well I did at first but I didn't have to.). And you don't need to use the controller either. If you want to load the modify view, just add a button to your regular view linking to the modify layout. No need to change anything else.


I happen to have written a blog article about it, check it out if you want: <http://violetfortytwo.blogspot.de/2012/11/joomla-25-multiple-views-one-model.html> 


Hope this helps.


"
"What are the appropriate NGINX configs if you are only sending JSON objects?
I'm an iOS developer and my back end is all written in Django. I use gunicorn as my HTTP server. I have three workers running on a small EC2 instance. 


My iOS app does not require any images or static content. At most, I am sending 1-20 JSON objects at a time per request. Each JSON object has at most about 5-10 fields. 


I'm quite new to NGINX. I heard it can do proxy buffering. I would like to add proxy buffering for slow clients, but I don't know the appropriate specific settings to use for the following modules:



```
proxy_buffers
Syntax:     proxy_buffers number size
Default:    8 4k|8k
Context:    http server location
Reference:  proxy_buffers


proxy_busy_buffers_size
Syntax:     proxy_busy_buffers_size size
Default:    8k|16k
Context:    http server location
Reference:  proxy_busy_buffers_size

proxy_buffer_size
Syntax:     proxy_buffer_size size
Default:    4k|8k
Context:    http server location
Reference:  proxy_buffer_size

```

The only setting which I know how to use (which is pretty sad) is the one below:



```
proxy_buffering
Syntax:     proxy_buffering on | off
Default:    on
Context:    http server location
Reference:  proxy_buffering

```

Your expertise in this area would be greatly appreciated by this kind lost soul!


","## `proxy_buffers`


The `number` defines how many buffers nginx will create and the `size` how big each buffer will be. When nginx starts receiving data from the upstream it starts filling up those buffers, either until the buffers are full or the upstream sends `EOF` or `EOT`. If any of those two conditions is met, nginx will send the contents of the buffers to the client.


If the client isn’t reading the buffers quickly enough, it will attempt to write the contents of them to disk and send them when the client is able to receive.


Take the average size of your JSON responses. Modern disks and file systems can handle even huge buffer sizes, but you should use something to the power of two and by creating a good balance between the number and size you can speed up the buffering process.


## `proxy_busy_buffers_size`


These are buffers that were already passed downstream but not yet completely send and therefor they can’t be reused. This directive limits the maximum total size of such buffers and thus allows remaining buffers to be used to read upstream responses.


## `proxy_buffer_size`


The main buffer that is always in use. Even if you disable `proxy_buffering`, nginx will still fill up this buffer and flush it as soon as it’s full or `EOF`/`EOT`.


## `proxy_max_temp_file_size`


This directive controls how many data might be written to disk if the buffers are full (still utilizing in-memory buffers, because you need them to communicate with the disk). If all buffers and this file are full, nginx stops reading from upstream and has to wait for downstream to fetch the data before it can continue with the same procedure.


"
"Character Casing in Silverlight 4 TextBox
I am writing a Silverlight 4 business application and have run into an issue. I need the text input in TextBoxes to be forced to UpperCase. What I understand from various forums is Silverlight does not Implement CharacterCasing and CSS Styling. 


Is there any other way to achieve this? 


","You can achieve this by creating a behavior, like this:



```
public class UpperCaseAction : TriggerAction<TextBox>
{

    protected override void Invoke(object parameter)
    {
        var selectionStart = AssociatedObject.SelectionStart;
        var selectionLenght = AssociatedObject.SelectionLength;
        AssociatedObject.Text = AssociatedObject.Text.ToUpper();
        AssociatedObject.SelectionStart = selectionStart;
        AssociatedObject.SelectionLength = selectionLenght;
    }
}

```

Then, use it in your TextBox, like this:



```
<Grid x:Name=""LayoutRoot"" Background=""White"">
    <TextBox TextWrapping=""Wrap"" VerticalAlignment=""Top"" Margin=""10"">
        <i:Interaction.Triggers>
            <i:EventTrigger EventName=""TextChanged"">
                <ASD_Answer009_Behaviors:UpperCaseAction/>
            </i:EventTrigger>
        </i:Interaction.Triggers>
    </TextBox>
</Grid>

```

Where `i:` is a namespace for 


`clr-namespace:System.Windows.Interactivity;assembly=System.Windows.Interactivity`


Code behind:



```
System.Windows.Interactivity.EventTrigger eventTrigger = new System.Windows.Interactivity.EventTrigger(""TextChanged"");
eventTrigger.Actions.Add(new UpperCaseAction());   
System.Windows.Interactivity.Interaction.GetTriggers(myTextBox).Add(eventTrigger);

```

In order to create and use behaviors, you need to download and install the [Expression Blend SDK for Silverlight 4](http://www.microsoft.com/downloads/en/details.aspx?FamilyID=d197f51a-de07-4edf-9cba-1f1b4a22110d&displaylang=en) and add a reference to [System.Windows.Interactivity.dll](http://msdn.microsoft.com/en-us/library/system.windows.interactivity(Expression.40).aspx).


"
"Logistic regression diagnostic plots in R
For testing purposes I made up some correlated data in R like this:



```
mydata = data.frame(
  outcome   = c(1, 0, 1, 0, 0, 1, 1, 0, 1, 1),
  predictor = c(0.1, -0.2, 0, 0.1, -0.3, 0.3, 0.2, -0.1, 0.1, 0.1)
)

```

Then I did this in order to create a logistic model that modeled this data:



```
model1 = glm(family = binomial, formula = outcome ~ predictor, data = mydata)

```

Running `plot(model1)` yields the following plots:


[![enter image description here](https://i.stack.imgur.com/tbdpD.png)](https://i.stack.imgur.com/tbdpD.png)


[![enter image description here](https://i.stack.imgur.com/4O3jW.png)](https://i.stack.imgur.com/4O3jW.png)


I need answers to some questions in order to understand how to perform diagnostics on such a logistic model. As someone with only an introductory course in statistics I'm having trouble gathering knowledge on how to interpret the plots.


1. What do the ""Predicted Values"" in the first plot represent?
2. What does residual mean in the context of logistic regression?
3. Which of these plots can in any way be useful for model diagnostics based on real data? How?


","This question is related to: [Interpretation of plot(glm.model)](https://stats.stackexchange.com/q/121490/7290), which it may benefit you to read. Regarding your specific questions: 


1. What constitutes a *predicted value* in logistic regression is a tricky subject. That's because the prediction can be made on several different scales. I think the most intuitive predicted value is the fitted probability of 'success' for the given observation. However, you could also use the fitted *odds*, or the fitted *log odds*. The fitted model equation / coefficients that is returned by statistical software will be on the scale of the linear predictor, that is, on the log odds scale. As a result, the fitted log odds of 'success' is typically used as the default. In `R`, for example, [?predict.glm](https://stat.ethz.ch/R-manual/R-devel/library/stats/html/predict.glm.html) will default to `type=""link""` (the log odds); since your predicted values extend below $0$, it is clear that the log odds of success is what is being plotted. 


Here are some additional resources that might help you: 


	- [Interpretation of simple predictions to odds ratios in logistic regression](https://stats.stackexchange.com/a/34638/7290)
2. Likewise, what constitutes a *residual* in logistic regression is even more tricky. There are *lots* of ways to compute residuals for a generalized linear model. In my opinion, the most intuitive residual would be the raw residual ($r\_i = y\_i - \hat y\_i$), but they are actually hard to use, so you may well never see them. By default, [?residuals.glm](https://stat.ethz.ch/R-manual/R-devel/library/stats/html/glm.summaries.html) defaults to `type=""deviance""`. Deviance residuals reflect a datum's contribution to the model's total deviance. Deviance residuals (and some other common types) are briefly discussed in the [lecture notes](http://data.princeton.edu/wws509/notes/c3s8.html) for Germán Rodríguez's GLM class. 


Suggested reading: 


	- [What do the residuals in a logistic regression mean?](https://stats.stackexchange.com/q/1432/)
3. I have argued, in my answer to the thread linked at the top, that it is best not to use these to examine a fitted logistic regression model. 


Further reading: 


	- [Diagnostics for logistic regression?](https://stats.stackexchange.com/q/45050/)


"
"Adding Security Provider multiple times in Java application
We have a Java application where a job is scheduled to run every 5 minutes. In that job, there is a security component that does the following every time it is executed:



```
java.security.Security
            .addProvider(new org.bouncycastle.jce.provider.BouncyCastleProvider());

```

My questions are:


1. Is it required to add the security provider multiple times in the application? Does it serve any purpose? To me, it doesn't make sense and adding it once should suffice.
2. Is it a candidate for a potential memory leak in the application?


To clarify, I want to add Bouncy Castle security provider programmatically in my application and not statically via JRE. IMO, adding Bouncy Castle security provider once in the application is enough and I need not do in multiple times.


","According to [`addProvider`'s javadoc](https://docs.oracle.com/javase/8/docs/api/java/security/Security.html#addProvider-java.security.Provider-):



> 
> **Returns:**  
> 
>  the preference position in which the provider was added, or -1 **if the provider was not added because it is already installed**
> 
> 
> 


`addProvider` already checks if the provider is installed, so, even if you have multiple calls across your application, it will just be added once. And after you add it, it will remain there until the JVM stops (or if someone calls [`removeProvider`](https://docs.oracle.com/javase/8/docs/api/java/security/Security.html#removeProvider-java.lang.String-)).


Of course you can optimize it and put just a single call in a main class (some class that you know it's always loaded at application startup), but I wouldn't worry much about that.


I had worked with systems that had more than one call to `addProvider` in different parts (because they could be called in any order and were independent from each other), all running in the same JVM, and it never got any memory leaks. Of course it's just my case, but I'm not aware of this causing leaks.




---


If you want to call `addProvider` only if the provider is not added yet, you can call `Security.getProvider()` to check that. If the provider is not in the JVM, it returns `null`:



```
// add provider only if it's not in the JVM
if (Security.getProvider(BouncyCastleProvider.PROVIDER_NAME) == null) {
    Security.addProvider(new BouncyCastleProvider());
}

```

"
"Workaround for missing whitelist in phonegap for windows phone
In my [other question](https://stackoverflow.com/questions/9197222/whitelist-external-hosts-in-phonegap-for-windows-phone) I found out that there is no whitelist for windows phones.


Now I am looking for a native code workaround but I have never written a line of native code for windows phones. So it's not easy for me. I think I can download a page like this:



```
void GetAirportData()
{
  var url = new Uri(""http://server.example.com/data.php"", UriKind.Absolute);
  var webClient = new WebClient();
  webClient.OpenReadCompleted += new OpenReadCompletedEventHandler(webClient_OpenReadCompleted);
  webClient.OpenReadAsync(url, url);
}

```

But how can a get this data to my javascript app?


","Here is a workaround. The following code is a Phonegap command that implements Cross Domain Call functionality. 



```
using System;
using System.IO;
using System.Net;
using System.Runtime.Serialization;
using WP7CordovaClassLib.Cordova;
using WP7CordovaClassLib.Cordova.Commands;
using WP7CordovaClassLib.Cordova.JSON;

namespace Cordova.Extension.Commands //namespace is predefined, don't change it!
{
    public class Cdc : BaseCommand //Cross domain call
    {
        [DataContract]
        public class CdcOptions
        {
            [DataMember(Name = ""path"")]
            public string Path { get; set; }
        }

        public void Call(string args)
        {
            CdcOptions options = JsonHelper.Deserialize<CdcOptions>(args);

            var url = new Uri(options.Path, UriKind.Absolute);

            var webClient = new WebClient();

            webClient.OpenReadCompleted += (s, e) =>
            {
                if (e.Error != null)
                {
                    DispatchCommandResult(new PluginResult(PluginResult.Status.ERROR, ""Error""));
                    return;
                }

                //Stream -> string
                var sr = new StreamReader(e.Result);
                var result = sr.ReadToEnd();

                DispatchCommandResult(
                    new PluginResult(PluginResult.Status.OK, result));
            };

            webClient.OpenReadAsync(url, url);

        }
    }
}

```

Test on the client side:



  





```
      <script type=""text/javascript"">

          function cdc(path, success, fail) {

              PhoneGap.exec(
                            success, //success
                            fail, //fail
                            ""Cdc"", //service
                            ""Call"", //action
                             path //args
                           );
          };

          function onDeviceReady(e) {

                cdc(
                    {
                        path: ""http://stackoverflow.com/questions/9291809/workaround-for-missing-whitelist-in-phonegap-for-windows-phone""
                    },
                    function (arg) {
                        document.getElementById('test').innerHTML = arg;
                    }, function (arg) {
                        document.getElementById('test').innerHTML = arg;
                    });

            }

            document.addEventListener(""deviceready"", onDeviceReady, false);


      </script>
  </head>
    <body>
        <div id=""test""></div>
    </body>
</html>

```

"
"How to pass local variable to remote using ssh and bash script?

```
ssh remotecluster 'bash -s' << EOF
> export TEST=""sdfsd""
> echo $TEST
> EOF

```

This prints nothing.


Also it still does not work even if I store the variable into file and copy it to remote.



```
TEST=""sdfsdf""
echo $TEST > temp.par
scp temp.par remotecluster
ssh remotecluster 'bash -s' << EOF
> export test2=`cat temp.par`
> echo $test2
> EOF

```

Still prints nothing. 


So my question is how to pass local variable to the remote machine as a variable ?


Answers have been give in [this](https://stackoverflow.com/questions/15775911/how-to-properly-escaping-qsub-command-with-long-input-args-within-ssh-command#15776073)


","The variable assignment `TEST=""sdfsd""` given in the here document is no real variable assignment, i. e. the variable assignment will actually not be performed directly in the declaration / definition of the here document (but later when the here document gets evaluated by a shell).


In addition, the `$TEST` variable contained in an unescaped or unquoted here document will be expanded by the **local shell** before the local shell executes the `ssh` command.
The result is that `$TEST` will get resolved to the empty string if it is not defined in the local shell **before** the `ssh` command or the here document respectively.


As a result, the variable assignment `export TEST=""sdfsd""` in the here document will not take effect in the local shell, but first be sent to the shell of the remote host and only there be expanded, hence your *prints nothing* experience.


The solution is to use an escaped or single-quoted here document, `<<\EOF` or `<<'EOF'`; or only escape the `\$TEST` variable in the here document; or just define the `$TEST` variable before the `ssh` command (and here document).



```
# prints sdfsd
export TEST=""sdfsd""
ssh localhost 'bash -s' << EOF
echo $TEST
EOF

# prints sdfsd
ssh localhost 'bash -s' << EOF
export TEST=""sdfsd""
echo \$TEST
EOF

# prints sdfsd
ssh localhost 'bash -s' <<'EOF'
export TEST=""sdfsd""
echo $TEST
EOF

```

"
"How to hide navigation portlet after logout?
How to hide a navigation portlet after logging out? What I exactly want is, if a logged-in admin visits a site, the navigation portlet should appear and be visible to him until he logs out from the plone site. But if a normal user visits the plone site, the portlet should not appear.


","A nice feature of Plone is the option to assign portlets not only to locations, but also to contenttypes and groups. To solve your request, simply assign the portlet to the group 'Administrators'.


To do that, go to your Plonesite's controlpanel, click 'Users and Groups', click on green tab 'Groups', click on 'Administrators', click on the tab 'Group portlets', add a portlet.


To reproduce this programtically, use Genericsetup (export portlets.xml of the site and include the relevant parts in your product).


For completeness: Similiar for assigning portlets to a contenttype, you go to the controlpanel, click 'Types', choose the wanted type (f.e. 'Event') of the dropdownlist and click on 'Manage portlets assigned to this content type', assign portlet. Progragramtically reproducable via exporting 'Types' and including it to your products.


"
"Express.js 4 - use middleware for authentication before static files
In my `express` app I've set static files to be served from the `/public` directory with this line:



```
app.use(express.static(__dirname + '/public'));

```

Now I need to add a middleware for authentication before serving the static content and if the user is not authenticated to be redirected to a route for authentication (e.g., `/login`).  

I'm not really sure how I have to do it. Any ideas?


","Since you didn't specify it, I'm going to assume that you already have some kind of authentication system. 


In Express, the order of the middlewares in the code matters: if you want to have middleware 1 executed before middleware 2, you should place them accordingly in your code. Since `express.static` *is* a middleware, if you want authentication before serving your static files you can simply write your authentication middleware before the call to `express.static`



```
app.use(function (req, res, next) {
    if (!userAuthenticated(req)) {
        return res.redirect('/login');
    }
    next();    
});

app.use(express.static(__dirname + '/public'));

```

I am assuming you have a `userAuthenticated` function which is for instance checking if the HTTP requests contains a valid access-token. 


[Read more about middlewares.](http://expressjs.com/en/guide/using-middleware.html)


"
"Flutter BottomNavigationBar not working with more than three items
I´ve a problem with the **BottomNavigationBar** in Flutter (0.6). As soon as I add **more then three BottomNavigationBarItems** as children, the buttons in the bar have **white icons and they are messed up**. When I use only three or less items, everything´s fine.


Here´s the widget code I use and which breaks the bar:



```
bottomNavigationBar: BottomNavigationBar(
          currentIndex: 0,
          iconSize: 20.0,
          items: [
          BottomNavigationBarItem(
              title: Text('Home'), icon: Icon(Icons.accessibility)),
          BottomNavigationBarItem(
              title: Text('Preise'), icon: Icon(Icons.account_box)),
          BottomNavigationBarItem(
              title: Text('Test'), icon: Icon(Icons.adb)),
          BottomNavigationBarItem(
              title: Text('Mehr'), icon: Icon(Icons.menu))
        ])

```

What's wrong here?


","For 4 or more items, set the `type` to fixed.



```
bottomNavigationBar: BottomNavigationBar(
  type: BottomNavigationBarType.fixed, // This is all you need!
  items: // ...,
)

```



---


From <https://github.com/flutter/flutter/issues/13642#issuecomment-371875044>



> 
> When more than 3 BottomNavigationBar items are provided the type, if
> unspecified, changes to `BottomNavigationBarType.shifting` per
> <https://docs.flutter.io/flutter/material/BottomNavigationBar/BottomNavigationBar.html>.
> This bit of information should be highlighted in the class's doc. It's
> easy to overlook where it is (I overlooked it).
> 
> 
> When the BottomNavigationBar's type is
> `BottomNavigationBarType.shifting` the items text and icons are
> rendered in white, via DefaultTextStyle and IconTheme. It's assumed
> that their`BottomNavigationBarItem.backgroundColor` will be specified
> as a contrasting color. This is obviously confusing.
> 
> 
> The overall idea with `shifting` type bottom navigation bars is that
> each item will have a different background color (that contrasts with
> white), since that color will become the color of the entire
> navigation bar, when the item is selected.
> 
> 
> The doc for BottomNavigationBar, and NavigationBarItem needs to be
> improved.
> 
> 
> 


"
"matplotlib candlestick chart error xy = (t-OFFSET, lower), TypeError: unsupported operand type(s) for -: 'datetime.date' and 'float'
I am trying to plot a candlestick chart using matplotplib. I have queried my database, returned the relevant data and appended in to an anrray named candleAr, in the required format (date,open,close,high,low)


My code is as follows:



```
import matplotlib.pyplot as plt
import numpy as np
import matplotlib.ticker as mticker
import matplotlib.dates as mdates
from matplotlib.finance import candlestick


candleAr=[]

cursor = conx.cursor()
query= 'SELECT ticker,date,time,open,low,high,close FROM eurusd WHERE date > ""2013-02-28""'
cursor.execute(query)
for line in cursor:
    #appendLine in correct format for candlesticks - date,open,close,high,low
    appendLine = line[1],line[3],line[6],line[5],line[4]
    candleAr.append(appendLine)


fig = plt.figure()
ax1 = plt.subplot(1,1,1)
candlestick(ax1, candleAr, width=1, colorup='g', colordown='r')

ax1.grid(True)


plt.xlabel('Date')
plt.ylabel('Price')
plt.show()

```

However I am getting the following error message:



```
Traceback (most recent call last):
  File ""C:\Users\Stuart\Desktop\Python Programming\Apache\Liclipse\Andres-Apache\FX\fx2.py"", line 53, in <module>
candlestick(ax1, candleAr, width=1, colorup='g', colordown='r')
  File ""C:\Users\Stuart\AppData\Local\Enthought\Canopy32\User\lib\site-packages\matplotlib\finance.py"", line 359, in candlestick
xy    = (t-OFFSET, lower),
TypeError: unsupported operand type(s) for -: 'datetime.date' and 'float'

```

When I print out my candleAr, I get the following (I have just included the first couple of results for an example):



```
[(datetime.date(2013, 3, 1), 1.306, 1.305, 1.306, 1.305), (datetime.date(2013, 3, 1), 1.305, 1.306, 1.306, 1.305)

```

So there are the datetime and floats relating to the error message no doubt, but surely the candlestick import from matplotlib can handle datetimes and floats - I mean what else can be passed to it to build a chart related to prices and dates, if not dates/datetimes and floats!??


Would anyone be able to point out what i am doing wrong?


","Matplotlib uses floating point numbers representing ordinals (counting days from 0001-01-01), with decimals representing fractions of a day.


This is documented in the [`candlestick` docstring](https://github.com/matplotlib/matplotlib/blob/master/lib/matplotlib/finance.py#L822-L863):



> 
> *time* must be in `float` days format - see `date2num`
> 
> 
> 


The [`matplotlib.dates` module](http://matplotlib.org/api/dates_api.html) gives you tools to convert `datetime` objects to such numbers:



```
>>> from datetime import date
>>> from matplotlib.dates import date2num
>>> date2num(date(2013, 3, 1))
734928.0

```

For your code that looks like:



```
from matplotlib.dates import date2num

appendLine = date2num(line[1]), line[3], line[6], line[5], line[4]

```

For `datetime.date()` this basically comes down to `datetime.date.toordinal()` as a float instead of an integer.


"
"Google Drive Rest API: Daily Limit for Unauthenticated Use Exceeded. Continued use requires signup
I'm migrating my app from deprecated Google Drive Android API to Drive Rest API. I implemented required behavior using new packages and everything works fine in emulator - app asks for access to user's Drive and getting it uploads files. When I started to test it on real device I get this error:



> 
> Daily Limit for Unauthenticated Use Exceeded. Continued use requires signup
> 
> 
> 


I already had configured project in <https://console.developers.google.com> that worked fine with old library (it has required ../auth/drive.file scope). According to [migration documentation](https://developers.google.com/drive/android/deprecation) - old configuration should work with new library without changes.


What other reasons could be for this error?


","Funny, I've been dealing with this the past so many days and waiting 24 hours to see if something fixed it. Its a really crappy problem.


**PRIMARY ANSWER**


I have a feeling you're using ProGuard same as I am assuming by ""real device"" that you're building signed apk as you do not specify if you're using signed or debug apk. ProGuard can cause this error to happen. I fixed it using the following.



```
# Fix OAuth Drive API failure for release builds
-keep class * extends com.google.api.client.json.GenericJson { *; }
-keep class com.google.api.services.drive.** { *; }
-keepclassmembers class * { @com.google.api.client.util.Key <fields>; }

```

**SECONDARY ANSWER**


Note that you **DO NOT** need to use keys/tokens using the Drive rest API with Android like you may find from other solutions (it may not hurt either, but it can). It doesn't match up here to what people talk about elsewhere (here they don't know what they're talking about).


See my notes here for further info: [Google Drive via OAuth release version receives dailyLimitExceededUnreg](https://stackoverflow.com/questions/58378312/google-drive-via-oauth-release-version-receives-dailylimitexceededunreg/58383292#58383292)


**IF you have the problem however in debug build** then you did not do everything correctly. My notes should get you on the right track.


If you need further help, I might assist because of how nuts it is.


"
"TFile.ReadAllText with TEncoding.UTF8 omits first 3 chars
I have a UTF-8 text file which starts with this line:



```
<HEAD><META name=GENERATOR content=""MSHTML 10.00.9200.16521""><body>

```

When I read this file with `TFile.ReadAllText` with TEncoding.UTF8:



```
MyStr := TFile.ReadAllText(ThisFileNamePath, TEncoding.UTF8);

```

then the first 3 characters of the text file are omitted, so MyStr results in:



```
'AD><META name=GENERATOR content=""MSHTML 10.00.9200.16521""><body>...'

```

However, when I read this file with `TFile.ReadAllText` without TEncoding.UTF8:



```
MyStr := TFile.ReadAllText(ThisFileNamePath);

```

then the file is read completely and correctly:



```
<HEAD><META name=GENERATOR content=""MSHTML 10.00.9200.16521""><body>...

```

Does `TFile.ReadAllText` have a bug?


","The first three bytes are skipped because the RTL code assumes that the file contains a UTF-8 BOM. Clearly your file does not. 


The `TUTF8Encoding` class implements a `GetPreamble` method that specifies the `UTF-8` BOM. And `ReadAllBytes` skips the preamble specified by the encoding that you pass.


One simple solution would be to read the file into a byte array and then use `TEncoding.UTF8.GetString` to decode it into a string.



```
var
  Bytes: TBytes;
  Str: string;
....
Bytes := TFile.ReadAllBytes(FileName);
Str := TEncoding.UTF8.GetString(Bytes);

```

An more comprehensive alternative would be to make a `TEncoding` instance that ignored the UTF-8 BOM.



```
type
  TUTF8EncodingWithoutBOM = class(TUTF8Encoding)
  public
    function Clone: TEncoding; override;
    function GetPreamble: TBytes; override;
  end;

function TUTF8EncodingWithoutBOM.Clone: TEncoding;
begin
  Result := TUTF8EncodingWithoutBOM.Create;
end;

function TUTF8EncodingWithoutBOM.GetPreamble: TBytes;
begin
  Result := nil;
end;

```

Instantiate one of these (you only need one instance per process) and pass it to `TFile.ReadAllText`.


The advantage of using a singleton instance of `TUTF8EncodingWithoutBOM` is that you can use it anywhere that expects a `TEncoding`.


"
"Empirical rejection rate of a wrong null hypothesis -is it the ""power of the test""?
I guess this is basic, but it forms part of the holes in my statistics education.  

What I know is that the power of a hypothesis test is not a fixed magnitude, but a function of various factors (significance level, distance of the alternative hypothesis from the null...). Assuming that this is correct: 


We perform a Monte Carlo study and we consider a two-tail t-test of statistical significance on an unknown parameter -so the null hypothesis is that this parameter is zero, and the alternative is that it is not zero.


The data are generated using a non-zero parameter. So our null is a false hypothesis. For a given significance level, we observe that the test rejects the false null XX% of the times it is executed. This looks like an empirically measured ""power of the test"" -but it is just a number, not a function. Granted, the sample size is given, the number of samples generated is given, the significance level is given. Changing any one of these will, I guess, give us different empirical rejection rates (and so an empirically determined range of the power function w.r.t. these factors) -but where and how does the alternative hypothesis comes in? Shouldn't it also co-determine the ""power of the test""?


**In general, how the empirical rejection rate of a wrong null hypothesis by a two-tailed test of statistical significance is to be interpreted?**


","The usual way to investigate the power properties is via a *power curve* (or sometimes, power surface, if we want to investigate the response to varying two things at once).


On these curves, the y-variable is the rejection rate and the x-value has the particular value of the thing we're varying.


The most common type of power curve is one we produce as we vary the parameter that is the subject of the test (e.g. in a test of means, as the true mean changes from the hypothesized value). Here's an example of a power curve for a two-sample t-test under a particular set of conditions:


$\hspace{1cm}$![enter image description here](https://i.stack.imgur.com/2ajgO.png)


(that one was not generated empirically, but by calling a function)


Here's a power comparison of paired t-test (curve) and Signed Rank test (points) for 4 pairs of normal observations (it's actually two-sided, but the left half isn't shown as it's a mirror image of the right half):  

![enter image description here](https://i.stack.imgur.com/Wz2Ao.png)


The t-test was carried out at the exact significance level of the signed rank test (since it can only take a few significance levels).


Here's a pair of (one-sided) power curves for a power comparison in a test of normality, where the alternatives are gamma distributed (which include the normal as a limiting case, under appropriate standardization):


![enter image description here](https://i.stack.imgur.com/dFYEt.png)


(this one was generated empirically in essentially the manner you describe)


As you suggest, at some specified value of the alternative, you can compute the power, and then as you vary that, you obtain a function that varies with the parameter you change, giving a power curve (or more strictly a curve of rejection rates, since at the null that's not power, but significance level).


I've generated quite a few of these curves in various of my answers. See [here](https://stats.stackexchange.com/a/83942/805) for another example that compares power from an ""algebraic"" calculation (/function call) and an empirical calculation (i.e. simulation).




---


Some general advice relating to empirical power:


1) since these are empirical rejection rates (i.e. binomial proportions), we can compute standard errors, confidence intervals and so on. So you know how accurate they are. If I can spare the time I usually simulate enough samples so the standard error is on the order of a pixel in my image (or even somewhat less), at least if it's not a large image (if you're doing vector graphics, think maybe half a percent or so of the height-dimension of the plot instead).


2) power curves are typically going to be smooth. Very smooth. As such, we can, with a bit of cleverness, avoid calculating power at a huge number of values (and indeed we can use this fact to reduce the number of simulations required at each point). One thing I do is to take a transformation of the power that will ""straighten"" the power curve, at least when we're away from 0 (inverse normal cdf is often a good choice), do cubic spline smoothing, and then transform back (beware doing that anywhere your rejection rates are exactly 0 or 1; you may want to leave those alone). If you do that well, you should be able to get away with 10-20 points or so. 


If you already have so many simulations your points are accurate to the pixel, after transforming to approximate local linearity, *linear* interpolation will usually be sufficient, and produces smooth, highly accurate curves after you transform back. If in doubt, produce a few more points and see whether the curves are generally within a couple of standard errors of those simulated values (since if those standard errors are only on the order of a pixel, you can't actually see the difference... so the miniscule bias this might introduce really doesn't matter).


You can also exploit obvious symmetries and so on. (In the t-test vs signed rank test power curve above, we exploited a relationship between the within-pair correlation ($\rho$) and the standard error of the difference to give different x-axes (above and below the plot) that then have the same power curves). 


Sometimes a little fiddling is required to get it just so, but you should get very smooth, more accurate estimates of the power with such smoothing. (On the other hand, sometimes its faster just to do more points - but in any case I would rarely do more than about 30 points because the eye happily fills the rest in.)


3) Since we're doing Monte Carlo simulation we can exploit various variance reduction techniques (though keep in mind the impact on calculated standard errors; at the worst, if you can't calculate it any more, the unreduced variance will be an upper bound). For example, we can use control variates - one thing I did when comparing power of a nonparametric test with a t-test was to compute the empirical rate for both tests and then use the error in the power for the t-test to help reduce the error in the other test (again, smoothing the result a little) ... but it works better if you do it on the right scale. A number of other variance-reduction techniques can also be used. (If I remember rightly, I might have used a control variate on a transformed-scale for the one-sample-t vs signed-rank test comparison above.)


But often simple brute force will suffice and requires little brain effort. If all it takes is going for a cup of coffee while the full simulation runs, might as well let it go. (No point spending half an hour working out some clever computation to save 15 minutes of a half hour run time.)


"
"Can you please explain Simpson's paradox with equations, instead of contingency tables?
I probably don't have a clear understanding of the [Simpson's paradox](https://en.wikipedia.org/wiki/Simpson%27s_paradox). Informally I know that the average of response Y1, grouped over all possible levels of factor A, can be higher than the average of response Y2 over all levels of A, even if the average of Y1 for each level of A (each group) is always less than the corresponding average of Y2. I've read examples, but I still get surprised each time I see it, maybe because I don't learn well by specific examples: I have troubles generalizing them. I learn best, and would rather see, an explanation in formulas. Can you please explain the paradox relying on equations, instead than count tables? 


Also, I think the reason for my surprise is that I could unconsciously be making some assumptions about the averages involved in the paradox, which may not be true in general. Maybe I forget to weight by the number of samples in each group? But then, I would like to see an equation which shows me that the estimate of the total average is more accurate if I weight each group average by the number of samples in each group, because (if this is true) it isn't obvious to me in general. Naively I would think that the estimate of $\mathbf{E}[Y\_1]$ has a lower standard error when I have more samples, irrespective of weighting.


","Here is a general approach to understanding Simpson's Paradox algebraically for count data.


Suppose that we have survival data for an exposure and we create a 2x2 contingency table. To keep things simple we will have the same counts in each cell. We could relax this, but it would make the algebra quite messy.


\begin{array}{|c|c|c|c|}
\hline
& \text{Died} & \text{Survived} & \text{Death Rate} \\ \hline
\text{Exposed} & X & X & 0.5 \\ \hline
 \text{Unexposed}& X & X & 0.5\\ \hline
\end{array}


In this case, the Death Rate is the same in both the Exposed and Unexposed groups. 


Now, if we split the data, say into one group for females and another group for males, we obtain 2 tables, with the following counts:


Males:
\begin{array}{|c|c|c|c|}
\hline
& \text{Died} & \text{Survived} & \text{Death Rate} \\ \hline
\text{Exposed} & Xa & Xb & \frac{a}{a+b} \\ \hline
\text{Unexposed}& Xc & Xd & \frac{c}{c+d}\\ \hline
\end{array}


and for females:
\begin{array}{|c|c|c|c|}
\hline
& \text{Died} & \text{Survived} & \text{Death Rate} \\ \hline
\text{Exposed} & X(a-1) & X(b-1) & \frac{a-1}{a+b-2} \\ \hline
 \text{Unexposed}& X(c-1) & X(d-1) & \frac{c-1}{c+d-2}\\ \hline
\end{array}


where $a,b,c,d \in [0,1]$ are the proportions of each cell in the aggregated data table that are male.


Simpson's Paradox will occur when the death rates for exposed males is greater than the death rate for unexposed males *AND* the death rate for exposed females is greater than the death rate for unexposed females. Alternatively, it will also occur when the death rates for exposed males is **less** than the death rate for unexposed males *AND* the death rate for exposed females is **less** than the death rate for unexposed females.
That is, when


$$\left(\frac{a}{a+b} < \frac{c}{c+d}\right) \text{ and } \left(\frac{a-1}{a+b-2} < \frac{c-1}{c+d-2}\right)$$


$$ \text{Or }$$


$$\left(\frac{a}{a+b} > \frac{c}{c+d}\right) \text{ and } \left(\frac{a-1}{a+b-2} > \frac{c-1}{c+d-2}\right)$$


As a concrete example, let $X=100$, and $a=0.5, b=0.8, c=0.9$. Then we will have Simpson's paradox when:


$$\left(\frac{0.5}{0.8+0.9} < \frac{0.9}{0.9+d}\right) \text{ and } \left(\frac{0.5-1}{0.5+0.8-2} < \frac{0.9-1}{0.9+d-2}\right)$$


$$ (-9 < d < 1.44) \text{ and } (0.96 < d < 1.1) $$


From which we conclude that d must lie in $(0.96,1]$


The 2nd set of inequalities gives:


$$\left(\frac{0.5}{0.8+0.9} > \frac{0.9}{0.9+d}\right) \text{ and } \left(\frac{0.5-1}{0.5+0.8-2} > \frac{0.9-1}{0.9+d-2}\right)$$


$$ (d < -0.9 \text{ or } d>1.44) \text{ and } (0.96 < d \text{ or } d > 1.44) $$


which has no solution for $d \in [0,1]$


So for the three values that we chose for $a,b,$ and $c$, to invoke Simpson's paradox, $d$ must be greater than 0.96. In the case where the value was $0.99$ then we would obtain a Death Rate for Males of


$$ 0.5/ (0.5+0.8) = 38 \text{% in the exposed group} $$
$$ 0.9/ (0.9+0.99) = 48 \text{% in the unexposed group} $$


and for Females:


$$ (0.5-1)/ (0.5+0.8-2) = 71 \text{% in the exposed group} $$
$$ (0.9-1)/ (0.9+0.99-2) = 91 \text{% in the unexposed group} $$


So, males have a higher death rate in the unexposed group than in the exposed group, and females *also* have a higher death rate in the unexposed group than the exposed group, yet the death rates in the aggregated data are the same for exposed and unexposed.


"
"Oracle Sql Developer ""string literal too long"" error
I have the following SQL that I would like to run in Oracle SQL Developer against an Oracle 10g server: 



```
WITH openedXml AS (
  SELECT extractvalue(column_value, '/theRow/First') FIRST,
         extractvalue(column_value, '/theRow/Last') LAST,
         to_number(extractvalue(column_value, '/theRow/Age')) Age
    FROM TABLE(XMLSequence(XMLTYPE('
  <theRange>
    <theRow><First>Bob</First><Last>Smith</Last><Age>30</Age></theRow>
    <theRow><First>Sue</First><Last>Jones</Last><Age>34</Age></theRow>
...
...
...
    <theRow><First>Tom</First><Last>Anderson</Last><Age>39</Age></theRow>
    <theRow><First>Ali</First><Last>Grady</Last><Age>45</Age></theRow>
  </theRange>
  ').extract('/theRange/theRow')))
)
SELECT *
FROM openedxml
WHERE age BETWEEN 30 AND 35;

```

When I attempt to run it I get the following error:



```
Error at Command Line:1 Column:0 Error report: SQL Error: ORA-01704: string literal too long
01704. 00000 -  ""string literal too long""
*Cause:    The string literal is longer than 4000 characters.
*Action:   Use a string literal of at most 4000 characters.
           Longer values may only be entered using bind variables.

```

My strings will occasionally be much longer than 4000 characters. Any ideas about how I can get around this problem?


","You will need to use a CLOB as the input to XMLTYPE() instead of a VARCHAR.


Using either `dbms_lob.loadclobfromfile` to load the xml from a file, or by breaking up the xml into 32000 character chunks and appending to the CLOB.



```
DECLARE
   xmlClob CLOB;
BEGIN
/* Build Clob here */

WITH openedXml AS (
  SELECT extractvalue(column_value, '/theRow/First') FIRST,
         extractvalue(column_value, '/theRow/Last') LAST,
         to_number(extractvalue(column_value, '/theRow/Age')) Age
    FROM TABLE(XMLSequence(XMLTYPE(xmlClob).extract('/theRange/theRow')))
)
SELECT *
FROM openedxml
WHERE age BETWEEN 30 AND 35;
END;

```

"
"How to keep a big and complex software product maintainable over the years?
I have been working as a software developer for many years now. It has been my experience that projects get more complex and unmaintainable as more developers get involved in the development of the product. 


It seems that software at a certain stage of development has the tendency to get ""hackier"" and ""hackier"" especially when none of the team members that defined the architecture work at the company any more.


I find it frustrating that a developer who has to change something has a hard time getting the big picture of the architecture. Therefore, there is a tendency to fix problems or make changes in a way that works against the original architecture. The result is code that gets more and more complex and even harder to understand.


Is there any helpful advice on how to keep source code really maintainable over the years?


","The only real solution to avoid code rot is to code well! 


How to code well is another question. It's hard enough even if you're an excellent programmer working alone. In a heterogeneous team, it becomes much harder still. In outsourced (sub)projects... just pray.


The usual good practices may help:


1. Keep it simple.
2. Keep it simple. This applies especially to the architecture, the ""big picture"". If developers are having hard time to get the big picture, they *are* going to code against it. So make the architecture simple so that all the developers get it. If the architecture has to be less than simple, then the developers must be trained to understand that architecture. If they don't internalize it, then they shouldn't code in it.
3. Aim for [low coupling](http://en.wikipedia.org/wiki/Coupling_%28computer_programming%29) and [high cohesion](http://en.wikipedia.org/wiki/Cohesion_%28computer_science%29). Make sure everyone in the team understands this idea. In a project consisting of loosely coupled, cohesive parts, if some of the parts becomes unmaintainable mess, you can simply unplug and rewrite that part. It's harder or near impossible if the coupling is tight.
4. Be consistent. *Which* standards to follow matters little, but please do follow *some* standards. In a team, everyone should follow the same standards of course. On the other hand, it's easy to become too attached with standards and forget the rest: please do understand that while standards are useful, they are only a small part of making good code. Don't make a big number of it.
5. [Code reviews](http://en.wikipedia.org/wiki/Code_review) may be useful to get a team to work consistently.
6. Make sure that all tools - IDEs, compilers, version control, build systems, documentation generators, libraries, *computers*, *chairs*, *overall environment* etc. etc. - are well maintained so that developers don't have to waste their time with secondary issues such as fighting project file version conflicts, Windows updates, noise and whatever banal but irritating stuff. Having to repeatedly waste considerable time with such uninteresting stuff lowers the morale, which at least won't improve code quality. In a large team, there could be one or more guys whose main job is to maintain the developer tools.
7. When making technological decisions, think what it would take to switch the technology; which decisions are irreversible and which are not. Evaluate the irreversible decisions extra carefully. For example, if you decide to write the project in [Java](http://en.wikipedia.org/wiki/Java_%28programming_language%29), that's a pretty much irreversible decision. If you decide to use some self-boiled binary format for data files, that's also a fairly irreversible decision (once the code is out in the wild and you have to keep supporting that format). But colors of the GUI can easily be adjusted, features initially left out can be added later on, so stress less about such issues.


"
"Spring Boot Kafka Configure DefaultErrorHandler?
I created a batch-consumer following the Spring Kafka docs:



```
@SpringBootApplication
public class ApplicationConsumer {
  private static final Logger LOGGER = LoggerFactory.getLogger(ApplicationConsumer.class);
  private static final String TOPIC = ""foo"";

  public static void main(String[] args) {
    ConfigurableApplicationContext context = SpringApplication.run(ApplicationConsumer.class, args);
  }

  @Bean
  public RecordMessageConverter converter() {
    return new JsonMessageConverter();
  }

  @Bean
  public BatchMessagingMessageConverter batchConverter() {
    return new BatchMessagingMessageConverter(converter());
  }

  @KafkaListener(topics = TOPIC)
  public void listen(List<Name> ps) {
    LOGGER.info(""received name beans: {}"", Arrays.toString(ps.toArray()));
  }
}

```

I was able to successfully get the consumer running by defining the following additional configuration env variables, that Spring automatically picks up:



```
export SPRING_KAFKA_BOOTSTRAP-SERVERS=...
export SPRING_KAFKA_CONSUMER_GROUP-ID=...

```

So the above code works. But now I want to customize the default error handler to use exponential backoff. From the ref docs I tried adding the following to ApplicationConsumer class:



```
@Bean
public ConcurrentKafkaListenerContainerFactory<?, ?> kafkaListenerContainerFactory() {
    ConcurrentKafkaListenerContainerFactory<String, Object> factory = new ConcurrentKafkaListenerContainerFactory<>();
    factory.setCommonErrorHandler(new DefaultErrorHandler(new ExponentialBackOffWithMaxRetries(10)));
    factory.setConsumerFactory(consumerFactory());
    return factory;
}

@Bean
public ConsumerFactory<String, Object> consumerFactory() {
    return new DefaultKafkaConsumerFactory<>(consumerConfigs());
}

@Bean
public Map<String, Object> consumerConfigs() {
    Map<String, Object> props = new HashMap<>();
    return props;
}

```

But now I get errors saying that it can't find some of the configuration. It looks like I'm stuck having to redefine all of the properties in consumerConfigs() that were already being automatically defined before. This includes everything from bootstrap server uris to the json-deserialization config.


Is there a good way to update my first version of the code to just override the default-error handler?


","Just define the error handler as a `@Bean` and Boot will automatically wire it into its auto configured container factory.


**EDIT**


This works as expected for me:



```
@SpringBootApplication
public class So70884203Application {

    public static void main(String[] args) {
        SpringApplication.run(So70884203Application.class, args);
    }

    @Bean
    DefaultErrorHandler eh() {
        return new DefaultErrorHandler((rec, ex) -> {
            System.out.println(""Recovered: "" + rec);
        }, new FixedBackOff(0L, 0L));
    }

    @KafkaListener(id = ""so70884203"", topics = ""so70884203"")
    void listen(String in) {
        System.out.println(in);
        throw new RuntimeException(""test"");
    }

    @Bean
    NewTopic topic() {
        return TopicBuilder.name(""so70884203"").partitions(1).replicas(1).build();
    }

}

```


```
foo
Recovered: ConsumerRecord(topic = so70884203, partition = 0, leaderEpoch = 0, offset = 0, CreateTime = 1643316625291, serialized key size = -1, serialized value size = 3, headers = RecordHeaders(headers = [], isReadOnly = false), key = null, value = foo)

```

"
"Count of Non Formula Cells in a column
lets say you had formulas in each cell between A1 & A9 but A10 had a hard coded number. Within VBA is there a formula you could enter in A11 that would give you the number 1 due to one record in the column having a value instead of a formula. (Or even a ""False"" would work in ""A11"" showing that column had 1 fail)?


","Try 1 of the array formulas below (Validated with `Ctrl+Shift+Enter`):


• Number of cells containing a formula:



```
=COUNT(IF(ISFORMULA(A1:A10),A1:A10))

```

• Number of cells NOT containing a formula (including empty cells):



```
=COUNT(IF(NOT(ISFORMULA(A1:A10)),A1:A10))

```

• Number of cells NOT containing a formula (excluding empty cells):



```
=COUNT(IF((NOT(ISBLANK(A1:A10))*NOT(ISFORMULA(A1:A10)))>0,A1:A10))

```

Note you can also use the previous formula and `COUNTBLANK`



```
=COUNT(IF(NOT(ISFORMULA(A1:A10)),A1:A10)) - COUNTBLANK(A1:A10)

```

• First row with a value instead of a formula:



```
=MIN(IF((NOT(ISBLANK(A1:A10))*NOT(ISFORMULA(A1:A10)))>0,ROW(A1:A10)))

```

"
"How to use AuthorizationServerSecurityConfigurer?
I am looking at a Spring boot project which has this code:



```
public void configure(AuthorizationServerSecurityConfigurer oauthServer) throws Exception {
    oauthServer
        .tokenKeyAccess(""permitAll()"")
        .checkTokenAccess(""isAuthenticated()"");
}

```

Unfortunately, I am not able to find any resources anywhere (i.e. Google, Spring docs, Spring oauth docs) that explains to me how to actually use `AuthorizationServerSecurityConfigurer`. Moreover, I do not understand exactly what `tokenKeyAccess(""permitAll()"")` or `checkTokenAccess(""isAuthenticated()"")` do.


Other than helping me understand what those two functions do, please help me learn where to look for these types of information in the future.


","Spring Security OAuth exposes two endpoints for checking tokens ([`/oauth/check_token`](https://github.com/spring-projects/spring-security-oauth/blob/master/spring-security-oauth2/src/main/java/org/springframework/security/oauth2/provider/endpoint/CheckTokenEndpoint.java) and [`/oauth/token_key`](https://github.com/spring-projects/spring-security-oauth/blob/master/spring-security-oauth2/src/main/java/org/springframework/security/oauth2/provider/endpoint/TokenKeyEndpoint.java)). Those endpoints are not exposed by default (have access ""denyAll()"").


So if you want to verify the tokens with this endpoint you'll have to add this to your authorization servers' config:



```
@Override
public void configure(AuthorizationServerSecurityConfigurer oauthServer) throws Exception {
    oauthServer.tokenKeyAccess(""isAnonymous() || hasAuthority('ROLE_TRUSTED_CLIENT')"")
               .checkTokenAccess(""hasAuthority('ROLE_TRUSTED_CLIENT')"");
}

```

Some more details can be found in the [""Resource Server Configuration"" section of the Spring Security OAuth2 documentation](http://projects.spring.io/spring-security-oauth/docs/oauth2.html#resource-server-configuration).


"
"what does ${pasteargs// /- } parameter expansion mean?
context: <https://stackoverflow.com/a/47348104/15603477>



```
printf -v pasteargs %*s 16
paste -d\  ${pasteargs// /- } < <(seq 1 42)
1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16
17 18 19 20 21 22 23 24 25 26 27 28 29 30 31 32
33 34 35 36 37 38 39 40 41 42

```


> 
> paste -d, --delimiters=LIST reuse characters from LIST instead of
> TABs
> 
> 
> 


${parameter/pattern/string}



> 
> The pattern is expanded to produce a pattern just as in filename
> expansion. Parameter is expanded and the longest match of pattern
> against its value is replaced with string. The match is performed
> according to the rules described below (see Pattern Matching). If
> pattern b**egins with ‘/’, all matches of pattern are replaced with**
> **string.**
> 
> 
> 


after checked with manual.


1. what does `${pasteargs// /- }` do.
2. I do know `%s` refers to `printf` argument. But `%*s 16` I don't know.
3. Even I quoted manual, still not sure `paste -d\`


","1. `printf %*s 16` means: print 16 spaces. See [this answer](https://stackoverflow.com/a/1000574/8529284) for further explanation. So now `pasteargs` is a variable with a value of 16 spaces.
2. `${pasteargs// /- }` means: Replace **all** occurrences of  (space) in the variable with `-`  (in other words: add a hyphen before each space in the variable). As you quoted from the manual:



> 
> If pattern begins with ‘/’, all matches of pattern are replaced with string.
> 
> 
> 


And the pattern here is `/` , which means: **all** matches of space. So now the value of `pasteargs` is 16 hyphens separated by spaces.
3. Regarding the `paste` command, you first need to understand that it's followed by 16 hyphens, meaning 16 streams. Basically it will merge every 16 consecutive lines into one line. By default, when those lines are merged, they are delimited by **tabs**. So `paste -d\`  (notice the trailing space after the backslash) means to separate the lines by **spaces** (`\` ) instead of **tabs**.


To summarize, this command (as advertised) just merges each 16 consecutive lines from the input into one line separated by spaces.


"
"Redirecting Servlet to Unicode-Domains
I'm having problems with sending redirections to servlet with Unicode-URLs.


i.e. consider the following url in Turkish



> 
> [http://türkçeisimtescil.com](http://t%C3%BCrk%C3%A7eisimtescil.com)
> 
> 
> 


It works if you paste it into your browser's address bar. However it is translated to



> 
> <http://xn--trkeisimtescil-ijb74a.com>
> 
> 
> 


by your browser upon your request.


Let's say I have first URL with UTF8-specific chars and I get it successfully from DB. I want to redirect my servlet to that URL. 


However when I just do `response.sendRedirect(url);` (according to headers) it redirects me to `www.t%1frk%e7eisimtescil.com`


I tried even `response.sendRedirect(""http://www.t\u011Frk\u00E7eisimtescil.com"");` (inline encoding) and the response is exactly the same.


Maybe if I obtain türkçeisimtescil.com on the headers, browser will convert it to `xn--..` format and it will succeed.


I could not figure out where the encoding got broken. Any helps are appreciated.


","That's an [Internationalized Domain Name](http://en.wikipedia.org/wiki/Internationalized_domain_name) (IDN). Its conversion between [ASCII and Unicode](http://en.wikipedia.org/wiki/Internationalized_domain_name#ToASCII_and_ToUnicode) is specified in [RFC 3490](http://en.wikipedia.org/wiki/Internationalized_domain_name#ToASCII_and_ToUnicode). In Java, you can use [`java.net.IDN`](http://download.oracle.com/javase/6/docs/api/java/net/IDN.html) to convert between the one and other. You can use [`java.net.URL`](http://download.oracle.com/javase/6/docs/api/java/net/URL.html) to obtain the host part from the URL.



```
String host = new URL(""http://türkçeisimtescil.com"").getHost();
String idn = IDN.toASCII(host);
String newURL = ""http://"" + idn;

```

"
"Where is the list of supported CPU profiles in VirtualBox?
I thought I found a list of CPU profiles for VirtualBox. I just can't find it again.


I can see my current Host CPU information with the following command:



```
vboxmanage list hostcpuids

```

but I wanted to test various profiles with one of my VMs which is not currently booting. Something like this:



```
VBoxManage modifyvm ""myVM"" --cpu-profile ""Intel Core i7–6700K""

```

Only if the profile doesn't exist, the VM will fail (it won't start at all). So I'm wondering how can I find a list of CPU profiles that will work on my host system?


Update:


From `/proc/cpuinfo` (one entry of 64):



```
processor   : 63
vendor_id   : GenuineIntel
cpu family  : 6
model       : 85
model name  : Intel(R) Xeon(R) Silver 4216 CPU @ 2.10GHz
stepping    : 7
microcode   : 0x500002c
cpu MHz     : 800.882
cache size  : 22528 KB
physical id : 1
siblings    : 32
core id     : 15
cpu cores   : 16
apicid      : 63
initial apicid  : 63
fpu     : yes
fpu_exception   : yes
cpuid level : 22
wp      : yes
flags       : fpu vme de pse tsc msr pae mce cx8 apic sep mtrr pge mca cmov pat pse36 clflush dts acpi mmx fxsr sse sse2 ss ht tm pbe syscall nx pdpe1gb rdtscp lm constant_tsc art arch_perfmon pebs bts rep_good nopl xtopology nonstop_tsc cpuid aperfmperf pni pclmulqdq dtes64 monitor ds_cpl vmx smx est tm2 ssse3 sdbg fma cx16 xtpr pdcm pcid dca sse4_1 sse4_2 x2apic movbe popcnt tsc_deadline_timer aes xsave avx f16c rdrand lahf_lm abm 3dnowprefetch cpuid_fault epb cat_l3 cdp_l3 invpcid_single intel_ppin ssbd mba ibrs ibpb stibp ibrs_enhanced tpr_shadow vnmi flexpriority ept vpid fsgsbase tsc_adjust bmi1 avx2 smep bmi2 erms invpcid cqm mpx rdt_a avx512f avx512dq rdseed adx smap clflushopt clwb intel_pt avx512cd avx512bw avx512vl xsaveopt xsavec xgetbv1 xsaves cqm_llc cqm_occup_llc cqm_mbm_total cqm_mbm_local dtherm ida arat pln pts pku ospke avx512_vnni md_clear flush_l1d arch_capabilities
bugs        : spectre_v1 spectre_v2 spec_store_bypass swapgs taa itlb_multihit
bogomips    : 4201.65
clflush size    : 64
cache_alignment : 64
address sizes   : 46 bits physical, 48 bits virtual
power management:

```

","I did some research myself and found something in their .dll (on a Windows host).


In the `.rdata` section of `Oracle\VirtualBox\VBoxVMM.dll` you can find some types which might work (I did not test all of them). I reformatted it a little to make it more readable:



```
AMD Athlon 64 3200+
AMD Athlon(tm) 64 X2 Dual Core Processor 4200+  AMD Athlon 64 X2 Dual Core 4200+
 Quad-Core AMD Opteron(tm) Processor 2384
 Quad-Core AMD Opteron 2384
 AMD Phenom(tm) II X6 1100T Processor    AMD Phenom II X6 1100T  AMD FX(tm)-8150 Eight-Core Processor    AMD FX-8150 Eight-Core  Intel(R) 8086   Intel 8086
 Intel(R) 80186  Intel 80186
Intel(R) 80286  Intel 80286
Intel(R) 80386  Intel 80386
Intel(R) 80486DX2
Intel 80486
Intel(R) Atom(TM) CPU  330   @ 1.60GHz  Intel Atom 330 1.60GHz  Intel(R) Pentium(R) CPU  N3530  @ 2.16GHz
Intel Pentium N3530 2.16GHz
Intel(R) Pentium(R) 4 CPU 3.00GHz
Intel Pentium 4 3.00GHz Intel(R) Pentium(R) M processor 2.00GHz Intel Pentium M processor 2.00GHz
Genuine Intel(R) CPU
    T2600  @ 2.16GHz Intel Core Duo T2600 2.16GHz    Intel(R) Core(TM)2 CPU
  T7600  @ 2.33GHz Intel Core2 T7600 2.33GHz
Intel(R) Core(TM)2 CPU
  X6800  @ 2.93GHz Intel Core2 X6800 2.93GHz
Intel(R) Xeon(R) CPU
    X5482  @ 3.20GHz Intel Xeon X5482 3.20GHz
 Intel(R) Core(TM) i7-2635QM CPU @ 2.00GHz
Intel Core i7-2635QM    Intel(R) Core(TM) i5-3570 CPU @ 3.40GHz Intel Core i5-3570
 Intel(R) Core(TM) i7-3960X CPU @ 3.30GHz
 Intel Core i7-3960X
Intel(R) Core(TM) i7-5600U CPU @ 2.60GHz
 Intel Core i7-5600U
Intel(R) Core(TM) i7-6700K CPU @ 4.00GHz
 Intel Core i7-6700K 

```

"
"Java read file and remove line, Cant figure out error
I am working on a project that has a method that reads a file, finds a a string that is passed in writes a new file called temp (that doesnt have the passed in line in it) and then deletes the original and renames temp to the opriginal name. I can run it without errors but it outputs nothing into the new file. I have done the usually debugging and found that the error lies in the lines where it writes out the line to the new file. I feel like I have made a small error calling something wrong. Any help solving it would be great...


Here is the code



```
public static void LineDelete(String Filename, String Content) throws IOException {
    try {
        File flights = new File(""AppData/"" + Filename);
        File temp;
        temp = new File(""AppData/temp.txt"");
        FileWriter fstream;
        BufferedWriter out;
        try (Scanner sc = new Scanner(flights)) {

            fstream = new FileWriter(""AppData/temp.txt"", true);
            out = new BufferedWriter(fstream);
            boolean exis = temp.exists();
            if (exis) {
                temp.delete();
                temp = new File(""AppData/temp.txt"");
                boolean createNewFile = temp.createNewFile();
            } else {
                boolean creatNewFile = temp.createNewFile();
            }
            String f;
            while (sc.hasNextLine()) {
                f = sc.nextLine();
                if (!f.equals(Content)) {

                    out.newLine();
                    out.write(f);

                }

            }
        }
        fstream.close();
        //out.close();
        flights.delete();
        File flightsn = new File(""AppData/"" + Filename);
        temp.renameTo(flightsn);
    } catch (FileNotFoundException ex) {
        Logger.getLogger(FileWrite.class.getName()).log(Level.SEVERE, null, ex);
    }
}

```

}


","You should call close on `out` (BufferedReader). 


You should also close it in the finally clause in try-catch-finally.


Your code should be more or less



```
import java.io.BufferedWriter;
import java.io.File;
import java.io.FileWriter;
import java.io.IOException;
import java.util.Scanner;
import java.util.logging.Level;
import java.util.logging.Logger;

public class FileWrite {
    public static void LineDelete(String Filename, String Content)
            throws IOException {
        BufferedWriter out = null;
        File flights = new File(""AppData/"" + Filename);
        File temp = new File(""AppData/temp.txt"");
        FileWriter fstream = null;

        try {
            Scanner sc = new Scanner(flights);
            fstream = new FileWriter(""AppData/temp.txt"", true);
            out = new BufferedWriter(fstream);
            boolean exis = temp.exists();
            if (exis) {
                temp.delete();
                temp = new File(""AppData/temp.txt"");
                boolean createNewFile = temp.createNewFile();
            } else {
                boolean creatNewFile = temp.createNewFile();
            }
            String f;
            while (sc.hasNextLine()) {
                f = sc.nextLine();
                if (!f.equals(Content)) {
                    out.newLine();
                    out.write(f);
                }

            }
        } catch (IOException ex) {
            Logger.getLogger(FileWrite.class.getName()).log(Level.SEVERE, null, ex);
        } finally {
            try {
                out.close();
            } catch (Exception e) {
                Logger.getLogger(FileWrite.class.getName()).log(Level.SEVERE, null, e);
            }
        }

        if(flights.exists()){           
            flights.delete();
            File flightsn = new File(""AppData/"" + Filename);
            temp.renameTo(flightsn);
        }
    }
}

```

"
"""Payload"" option on Facebook Bots buttons
[Facebook Send API](https://developers.facebook.com/docs/messenger-platform/send-api-reference) mentions a ""payload"" type you can set for buttons on the Generic Response Template. However, they don't say how it works, other than:



> 
> For postback buttons, this data will be sent back to you via webhook
> 
> 
> 


But how is it sent back? I don't receive any messages when I click the payload button. Has anyone successfully used it?


","I tested it and it works for me. The payload of a button acts like the value on a html button. This means it's not visible to the user but it's the value that's send back to you. 


If you create a button like that:



```
'attachment': {
    'type': 'template',
    'payload': {
        'template_type': 'button',
        'text': 'This is the description',
        'buttons': [
             {
                 'type': 'postback',
                 'title': 'This is the visible text',
                 'payload': 'This is the value you get back'
             }
        ]
 }

```

A callback including a payload looks like that:



```
{'timestamp': 1461697120850, 'postback': {'payload': 'this is the value you get back'}, 'recipient': {'id': xxxxxxxx}, 'sender': {'id': xxxxxxxx}}

```

"
"Is it guaranteed for two (or more) versions of the .NET Framework to be installed side by side?
Based on many articles I have read, I came to the conclusion that it is not guaranteed for a .NET Framework to be 100% backward compatible with the previous versions. So my question is: is it guaranteed for two (or more) versions of the .NET Framework to be installed side by side without any problems (is this documented somewhere by Microsoft)?


For example, can .NET 1.1 be installed side by side with .NET 4.5, or will I have to re-compile my project to work on .NET 4.5? 


","There's a difference between the Framework version and the runtime version. You can have different runtimes operate side-by-side without interfering if your application is properly marked with the expected runtime in its configuration file.


Programs written for a particular runtime (e.g. 1.1, 2.0, 4.0) can run independently against their given framework with no side effects from the other frameworks. That being said, you must [modify your config file](https://www.google.com/url?sa=t&rct=j&q=&esrc=s&source=web&cd=1&cad=rja&uact=8&ved=0CB4QFjAA&url=https%3A%2F%2Fmsdn.microsoft.com%2Fen-us%2Flibrary%2Fw4atty68%2528v%3Dvs.110%2529.aspx&ei=itPmVI2EK4m_ggT9ooT4Cw&usg=AFQjCNHCJy_OIA5FV5XO2g0PCMsuCq4F3A&bvm=bv.86475890,d.eXY) to ensure that a .NET 2.0 application runs with the .NET 2.0 CLR when the 4.0 CLR is installed on the machine


The CLR's are affected by the version of the Framework installed on the machine. For example, 4.0 and 4.5 both run with the 4.0 CLR and having 4.5 installed introduces some incompatibilities. That being said, these issues are usually very specific and most applications will not be affected by them.


Here's [Microsoft's page on compatibility](https://msdn.microsoft.com/en-us/library/ff602939%28v=vs.110%29.aspx) between Framework versions.


"
"Are ""(function ( ) { } ) ( )"" and ""(function ( ) { } ( ) )"" functionally equal in JavaScript?
Both of these code blocks below alert `foo` then `bar`. The only difference is `})()` and `}())`.


Code 1:



```
(function()
{
    bar = 'bar';
    alert('foo');
})();

alert(bar);

```

Code 2:



```
(function()
{
    bar = 'bar';
    alert('foo');
}());

alert(bar);

```

So is there any difference, apart from the syntax?


","No; they are identical




---


However, if you add `new` beforehand and `.something` afterwards, they will be different.


# Code 1



```
new (function() {
    this.prop = 4;
}) ().prop;

```

This code creates a new instance of this function's class, then gets the `prop` property of the new instance.  

It returns `4`.


It's equivalent to



```
function MyClass() {
    this.prop = 4;
}
new MyClass().prop;

```



---


# Code 2



```
new ( function() {
    return { Class: function() { } }; 
}() ).Class;

```

This code calls `new` on the `Class` property.  

Since the parentheses for the function call are inside the outer set of parentheses, they aren't picked up by the `new` expression, and instead call the function normally, returning its return value.  

The `new` expression parses up to the `.Class` and instantiates that. (the parentheses after `new` are optional)


It's equivalent to



```
var namespace = { Class: function() { } };

function getNamespace() { return namespace; }

new ( getNamespace() ).Class;
//Or,
new namespace.Class;

```

Without the parentheses around the call to `getNamespace()`, this would be parsed as `(new getNamespace()).Class` — it would call instantiate the `getNamespace` class and return the `Class` property of the new instance.


"
"Recursive function calls in python in list comprehension gives different result than assigning the result of the call to a variable first
Why doesn't this:



```
def subsets(x):
    if not x:
        return [[]]
    else:
        return [x + y for x in [[], [x[0]]] for y in subsets(x[1:])]

print subsets(['a', 'b'])

```

produce the same output as this:



```
def subsets(x):
    if not x:
        return [[]]
    else:
        z = subsets(x[1:])
        return [x + y for x in [[], [x[0]]] for y in z]

print subsets(['a', 'b'])

```

","The problem is that you are using the name `x` for two different things. When you introduce `x` as a variable inside a list comprehension it hides the `x` that was defined in the function. As a result, the final `x` in your list comprehension isn't refering to the `x` that you had hoped for.


Renaming one of the `x` to something else fixes the problem:



```
def subsets(original_x):
    if not original_x:
        return [[]]
    else:
        return [x + y for x in [[], [original_x[0]]]
                      for y in subsets(original_x[1:])]

```

I would also suggest that you try to find more descriptive names for your variables. This will reduce the likelihood of accidentally reusing a name.


"
"UPDATE an array using PDO
I'm creating a multi-step form for my users. They will be allowed to update any or all the fields. So, I need to send the values, check if they are set and if so, run an `UPDATE`. Here is what I have so far:



```
public function updateUser($firstName, $lastName, $streetAddress, $city, $state, $zip, $emailAddress, $industry, $password, $public = 1, 
    $phone1, $phone2, $website,){

    $updates = array(
        'firstName'             => $firstName,
        'lastName'              => $lastName,
        'streetAddress'         => $streetAddress,
        'city'                  => $city,
        'state'                 => $state,
        'zip'                   => $zip,
        'emailAddress'          => $emailAddress,
        'industry'              => $industry,
        'password'              => $password,
        'public'                => $public,
        'phone1'                => $phone1,
        'phone2'                => $phone2,
        'website'               => $website,

);

```

Here is my PDO (well, the beginning attempt)



```
    $sth = $this->dbh->prepare(""UPDATE user SET firstName = ""); //<---Stuck here
    $sth->execute();
    $result = $sth->fetchAll(PDO::FETCH_ASSOC);
    return $result; 

```

Basically, how can I create the `UPDATE` statement so it only updates the items in the array that are not `NULL`?


I thought about running a `foreach` loop like this:



```
    foreach($updates as $key => $value) {
        if($value == NULL) {
            unset($updates[$key]);
        }
    }

```

but how would I write the `prepare` statement if I'm unsure of the values?


If I'm going about this completely wrong, please point me in the right direction. Thanks.


","First of all, use [`array_filter`](http://php.net/filter) to remove all NULL values:



```
$updates = array_filter($updates, function ($value) {
    return null !== $value;
});

```

Secondly, bind parameters, that makes your live a lot easier:



```
$query = 'UPDATE table SET';
$values = array();

foreach ($updates as $name => $value) {
    $query .= ' '.$name.' = :'.$name.','; // the :$name part is the placeholder, e.g. :zip
    $values[':'.$name] = $value; // save the placeholder
}

$query = substr($query, 0, -1).';'; // remove last , and add a ;

$sth = $this->dbh->prepare($query);

$sth->execute($values); // bind placeholder array to the query and execute everything

// ... do something nice :)

```

"
"Using integer values and String identifier in an Enum
So I currently have two enums:



```
public enum AuthorizationLevel 
    {   
        FULL,
        HALF,
        NONE;
    };


public enum DatabaseLoggingLevel
{
    HIGH,
    MED,
    LOW,
    AUTH_ONLY,
    NONE
}

```

I want to be able to associate integers with the values in the enums so that I can have code like so:



```
if(databaseLoggingLevel < ClassName.DatabaseLoggingLevel.HIGH) return;

```

This is just so that certain logging is disabled when the level is less than `HIGH`. I thought about making a helper function that returns an integer value associated with each enum value with a switch statement, but that seems hacky. Is there something I am missing?


","You can add an integer instance field to your enum, and add a parameterized constructor:



```
enum AuthorizationLevel {

    FULL(1), HALF(2), NONE(3);

    private final int level;

    private AuthorizationLevel(int level) {
        this.level = level;
    }

    public int getLevel() {
        return level;
    }
};

```

Note that there is an `ordinal()` method which will give you position of your enum constants, but you should avoid using it. Again, quoting [Effective Java Item - 31 (Use instance fields instead of ordinal)](http://books.google.co.in/books?id=ka2VUBqHiWkC&pg=PR10&lpg=PR10&dq=Effective+Java+Item+31&source=bl&ots=yYJnMen5O_&sig=8VQVA0z44gWmg4zOJRoY9oyp6EE&hl=en&sa=X&ei=N7ooUrmmDMSQrQe6sIDwDA&ved=0CD8Q6AEwAw#v=onepage&q=Effective%20Java%20Item%2031&f=false):



> 
> If the constants are reordered, the numberOfMusicians method will
>  break. If you want to **add a second enum constant associated with an
>  int value that you’ve already used**, you’re out of luck.
> 
> 
> Also, you can’t add a constant for an int value without adding
>  constants for all intervening int values. For example, suppose you
>  want to add a constant representing a triple quartet, which consists
>  of twelve musicians. There is no standard term for an ensemble
>  consisting of eleven musicians, so you are forced to add a dummy
>  constant for the unused int value (11). At best, this is ugly. If many
>  int values are unused, it’s impractical.
> 
> 
> 


"
"Convert a simple mono drawing image to a 2d text array
I am trying to develop an algorithm that converts simple mono line images ie Maze, to a text 2d array.


For example, the image below, it would be converted to the following text array.


[![enter image description here](https://i.stack.imgur.com/HJVPe.jpg)](https://i.stack.imgur.com/HJVPe.jpg)



```
[|------------         |]
[|     |               |]
[|                     |]
[|  |------|      ---- |]
[|         |      |    |]
[|         |     ---   |] 
[|---      |        |  |]
[|         |---     |  |] 
[|     |      |        |]
[|    ---------------  |]
[|                     |]
[|  -------------------|]
        

```

and finally, like this, where 0=obstacle and 1=free passage



```
[0000000000000111111110]
[0111110111111111111110]
[0111111111111111111110]
[0110000000011111100000]
[0111111111011111011110]
[0111111111011111000110] 
[0000111111011111111010]
[0111111111000011111010] 
[0111110111111011111110]
[0111100000000000000010]
[0111111111111111111110]
[0110000000000000000000]

```

I am thinking to use an Image to Line Art Text like algorithms, ie
<https://www.text-image.com/convert/pic2ascii.cgi>


What do you think about this approach?


","Interseting problem its basically vector form of [Image to ASCII art conversion](https://stackoverflow.com/a/32987834/2521214)... I managed to do this with this algorithm:


1. **preprocess image**


You gave us JPG which has lossy compresion meaning your image contain much more than just 2 colors. So there are shades and artifacts which will screw things up. So first we must get rid of those by thresholding and recoloring. So we can have 2D BW image (no grayscales)
2. **vectorize**


Your maze is axis aligned so it contains only horizontal and vertical (`h`,`v`) lines. So simply scan each line of image find first starting wall pixel then its ending pixel and store somewhere... repeat until whole line is processed and do this for all lines. Again do the same for rows of image. As your image has thick walls ignore lines sorter than thickness threshold and remove adjacent (duplicates) line that are (almost) the same.
3. **get list of possible grid coordinates from h,v lines**


simply make a list of all x and y (separately) coordinates from lines start and end points. Then sort them and remove too close coordinates (duplicates).


Now the min and max values gives you AABB of your maze and GCD of all the `coordinate-lowest coordinate` will give you grid size.
4. **align h,v lines to grid**


simply round all start/end points to nearest grid position ...
5. **create text buffer for maze**


AABB along with grid size will give you resolution of your maz in cells so simply create 2D text buffer where each cell has `NxN` characters. I am using `6x3` cells which looks nice enough (square and with enough space inside).
6. **renmder h,v lines into text**


simply loop through all lines and render `-` or `|` instead of pixels... I am using also `+` if the target position does not contain ' '.
7. **convert 2D text array into wanted text output**


simply copy the lines into single text ... or if you clever enough you can have 1D and 2D at the same memory place with `eol` encoded between lines.


**Here simple example in C++/VCL I made from the exampe in the link above:**



```
//---------------------------------------------------------------------------
#include <vcl.h>
#include <jpeg.hpp>
#pragma hdrstop

#include ""win_main.h""
#include ""List.h""
//---------------------------------------------------------------------------
#pragma package(smart_init)
#pragma resource ""*.dfm""
TForm1 *Form1;
Graphics::TBitmap *bmp=new Graphics::TBitmap;
int txt_xs=0,txt_ys=0,txt_xf=0;
//---------------------------------------------------------------------------
template <class T> void sort_asc_bubble(T *a,int n)
    {
    int i,e; T a0,a1;
    for (e=1;e;n--)                                     // loop until no swap occurs
     for (e=0,a0=a[0],a1=a[1],i=1;i<n;a0=a1,i++,a1=a[i])// proces unsorted part of array
      if (a0>a1)                                        // condition if swap needed
      { a[i-1]=a1; a[i]=a0; a1=a0; e=1; }               // swap and allow to process array again
    }
//---------------------------------------------------------------------------
AnsiString bmp2lintxt(Graphics::TBitmap *bmp)
    {
    bool debug=false;
    const int cx=6;             // cell size
    const int cy=3;
    const int thr_bw=400;       // BW threshold
    const int thr_thickness=10; // wall thikness threshold
    char a;
    AnsiString txt="""",eol=""\r\n"";
    int x,y,x0,y0,x1,y1,xs,ys,gx,gy,nx,ny,i,i0,i1,j;
    union { BYTE db[4]; DWORD dd; } c; DWORD **pyx;
    List<int> h,v;  // horizontal and vertical lines (x,y,size)
    List<int> tx,ty;// temp lists for grid GCD computation
    // [init stuff]
    bmp->HandleType=bmDIB;
    bmp->PixelFormat=pf32bit;
    xs=bmp->Width ;
    ys=bmp->Height;
    if (xs<=0) return txt;
    if (ys<=0) return txt;
    pyx=new DWORD*[ys];
    for (y=0;y<ys;y++) pyx[y]=(DWORD*)bmp->ScanLine[y];
    i=xs; if (i<ys) i=ys;
    // threshold bmp to B&W
    x0=xs; x1=0; y0=xs; y1=0;
    for (y=0;y<ys;y++)
     for (x=0;x<xs;x++)
        {
        c.dd=pyx[y][x];
        i =c.db[0];
        i+=c.db[1];
        i+=c.db[2];
        if (i>=thr_bw) c.dd=0x00FFFFFF;
        else           c.dd=0x00000000;
        pyx[y][x]=c.dd;
        }
    if (debug) bmp->SaveToFile(""out0_bw.bmp"");
    // [vectorize]
    // get horizontal lines
    i0=0; i1=0; h.num=0;
    for (y0=0;y0<ys;y0++)
        {
        for (x0=0;x0<xs;)
            {
            for (     ;x0<xs;x0++) if (!pyx[y0][x0]) break;
            for (x1=x0;x1<xs;x1++) if ( pyx[y0][x1]){ x1--; break; }
            i=x1-x0;
            if (i>thr_thickness)
                {
                h.add(x0);
                h.add(y0);
                h.add(i);
                }
            x0=x1+1;
            }
        // remove duplicate lines
        for (i=i0;i<i1;i+=3)
         for (j=i1;j<h.num;j+=3)
          if ((abs(h[i+0]-h[j+0])<thr_thickness)&&(abs(h[i+2]-h[j+2])<thr_thickness))
            {
            h.del(i);
            h.del(i);
            h.del(i);
            i1-=3; i-=3; break;
            }
        i0=i1; i1=h.num;
        }
    // get vertical lines
    i0=0; i1=0; v.num=0;
    for (x0=0;x0<xs;x0++)
        {
        for (y0=0;y0<ys;)
            {
            for (     ;y0<ys;y0++) if (!pyx[y0][x0]) break;
            for (y1=y0;y1<ys;y1++) if ( pyx[y1][x0]){ y1--; break; }
            i=y1-y0;
            if (i>thr_thickness)
                {
                v.add(x0);
                v.add(y0);
                v.add(i);
                }
            y0=y1+1;
            }
        // remove duplicate lines
        for (i=i0;i<i1;i+=3)
         for (j=i1;j<v.num;j+=3)
          if ((abs(v[i+1]-v[j+1])<thr_thickness)&&(abs(v[i+2]-v[j+2])<thr_thickness))
            {
            v.del(i);
            v.del(i);
            v.del(i);
            i1-=3; i-=3; break;
            }
        i0=i1; i1=v.num;
        }
    // [compute grid]
    x0=xs; y0=ys; x1=0; y1=0;   // AABB
    gx=10; gy=10;               // grid cell size
    nx=0; ny=0;                 // grid cells
    tx.num=0; ty.num=0;         // clear possible x,y coordinates
    for (i=0;i<h.num;i+=3)
        {
        x =h[i+0];
        y =h[i+1];
        if (x0>x) x0=x; if (x1<x) x1=x; for (j=0;j<tx.num;j++) if (tx[j]==x){ j=-1; break; } if (j>=0) tx.add(x);
        if (y0>y) y0=y; if (y1<y) y1=y; for (j=0;j<ty.num;j++) if (ty[j]==y){ j=-1; break; } if (j>=0) ty.add(y);
        x+=h[i+2];
        if (x0>x) x0=x; if (x1<x) x1=x; for (j=0;j<tx.num;j++) if (tx[j]==x){ j=-1; break; } if (j>=0) tx.add(x);
        }
    for (i=0;i<v.num;i+=3)
        {
        x =v[i+0];
        y =v[i+1];
        if (x0>x) x0=x; if (x1<x) x1=x; for (j=0;j<tx.num;j++) if (tx[j]==x){ j=-1; break; } if (j>=0) tx.add(x);
        if (y0>y) y0=y; if (y1<y) y1=y; for (j=0;j<ty.num;j++) if (ty[j]==y){ j=-1; break; } if (j>=0) ty.add(y);
        y+=v[i+2];
        if (y0>y) y0=y; if (y1<y) y1=y; for (j=0;j<ty.num;j++) if (ty[j]==y){ j=-1; break; } if (j>=0) ty.add(y);
        }
    // order tx,ty
    sort_asc_bubble(tx.dat,tx.num);
    sort_asc_bubble(ty.dat,ty.num);
    // remove too close coordinates
    for (i=1;i<tx.num;i++) if (tx[i]-tx[i-1]<=thr_thickness){ tx.del(i); i--; }
    for (i=1;i<ty.num;i++) if (ty[i]-ty[i-1]<=thr_thickness){ ty.del(i); i--; }
    // estimate gx,gy
    for (gx=x1-x0,i=1;i<tx.num;i++){ x=tx[i]-tx[i-1]; if (gx>x) gx=x; } nx=(x1-x0+1)/gx; gx=(x1-x0+1)/nx; x1=x0+nx*gx;
    for (gy=y1-y0,i=1;i<ty.num;i++){ y=ty[i]-ty[i-1]; if (gy>y) gy=y; } ny=(y1-y0+1)/gy; gy=(y1-y0+1)/ny; y1=y0+ny*gy;
    // align x,y to grid: multiplicate nx,ny by cx,cy to form boxes and enlarge by 1 for final border lines
    nx=(cx*nx)+1;
    ny=(cy*ny)+1;
    // align h,v lines to grid
    for (i=0;i<h.num;i+=3)
        {
        x=h[i+0]-x0; x=((x+(gx>>1))/gx)*gx; h[i+0]=x+x0;
        y=h[i+1]-y0; y=((y+(gy>>1))/gy)*gy; h[i+1]=y+y0;
        j=h[i+2];    j=((j+(gx>>1))/gx)*gx; h[i+2]=j;
        }
    for (i=0;i<v.num;i+=3)
        {
        x=v[i+0]-x0; x=((x+(gx>>1))/gx)*gx; v[i+0]=x+x0;
        y=v[i+1]-y0; y=((y+(gy>>1))/gy)*gy; v[i+1]=y+y0;
        j=v[i+2];    j=((j+(gy>>1))/gy)*gy; v[i+2]=j;
        }
    // [h,v lines -> ASCII Art]
    char *text=new char[nx*ny];
    char **tyx=new char*[ny];
    for (y=0;y<ny;y++)
     for (tyx[y]=text+(nx*y),x=0;x<nx;x++)
      tyx[y][x]=' ';
    // h lines
    for (i=0;i<h.num;i+=3)
        {
        x=(h[i+0]-x0)/gx;
        y=(h[i+1]-y0)/gy;
        j=(h[i+2]   )/gx; j+=x;
        x*=cx; y*=cy; j*=cx;
        for (;x<=j;x++) tyx[y][x]='-';
        }
    // v lines
    for (i=0;i<v.num;i+=3)
        {
        x=(v[i+0]-x0)/gx;
        y=(v[i+1]-y0)/gy;
        j=(v[i+2]   )/gy; j+=y;
        x*=cx; y*=cy; j*=cy;
        for (;y<=j;y++)
         if (tyx[y][x]=='-') tyx[y][x]='+';
          else               tyx[y][x]='|';
        }
    // convert char[ny][nx] to AnsiString
    for (txt="""",y=0;y<ny;y++,txt+=eol)
     for (x=0;x<nx;x++) txt+=tyx[y][x];
    txt_xs=nx;  // just remember the text size for window resize
    txt_ys=ny;
    delete[] text;
    delete[] tyx;
    // [debug draw]
    // grid
    bmp->Canvas->Pen->Color=TColor(0x000000FF);
    for (i=1,x=x0;i;x+=gx)
        {
        if (x>=x1){ x=x1; i=0; }
        bmp->Canvas->MoveTo(x,y0);
        bmp->Canvas->LineTo(x,y1);
        }
    for (i=1,y=y0;i;y+=gy)
        {
        if (y>=y1){ y=y1; i=0; }
        bmp->Canvas->MoveTo(x0,y);
        bmp->Canvas->LineTo(x1,y);
        }
    if (debug) bmp->SaveToFile(""out1_grid.bmp"");
    // h,v lines
    bmp->Canvas->Pen->Color=TColor(0x00FF0000);
    bmp->Canvas->Pen->Width=2;
    for (i=0;i<h.num;)
        {
        x=h[i]; i++;
        y=h[i]; i++;
        j=h[i]; i++;
        bmp->Canvas->MoveTo(x,y);
        bmp->Canvas->LineTo(x+j,y);
        }
    for (i=0;i<v.num;)
        {
        x=v[i]; i++;
        y=v[i]; i++;
        j=v[i]; i++;
        bmp->Canvas->MoveTo(x,y);
        bmp->Canvas->LineTo(x,y+j);
        }
    bmp->Canvas->Pen->Width=1;
    if (debug) bmp->SaveToFile(""out2_maze.bmp"");

    delete[] pyx;
    return txt;
    }
//---------------------------------------------------------------------------
void update()
    {
    int x0,x1,y0,y1,i,l;
    x0=bmp->Width;
    y0=bmp->Height;
    // Font size
    Form1->mm_txt->Font->Size=Form1->cb_font->ItemIndex+4;
    txt_xf=abs(Form1->mm_txt->Font->Size);
    // mode
    Form1->mm_txt->Text=bmp2lintxt(bmp);
    // output
    Form1->mm_txt->Lines->SaveToFile(""pic.txt"");
    x1=txt_xs*txt_xf;
    y1=txt_ys*abs(Form1->mm_txt->Font->Height);
    if (y0<y1) y0=y1;
    x0+=x1+16+Form1->flb_pic->Width;
    y0+=Form1->pan_top->Height;
    if (x0<340) x0=340;
    if (y0<128) y0=128;
    Form1->ClientWidth=x0;
    Form1->ClientHeight=y0;
    Form1->Caption=AnsiString().sprintf(""Picture -> Text ( Font %ix%i )"",abs(Form1->mm_txt->Font->Size),abs(Form1->mm_txt->Font->Height));
    }
//---------------------------------------------------------------------------
void draw()
    {
    Form1->ptb_gfx->Canvas->Draw(0,0,bmp);
    }
//---------------------------------------------------------------------------
void load(AnsiString name)
    {
    if (name=="""") return;
    AnsiString ext=ExtractFileExt(name).LowerCase();
    if (ext=="".bmp"")
        {
        bmp->LoadFromFile(name);
        }
    if (ext=="".jpg"")
        {
        TJPEGImage *jpg=new TJPEGImage;
        jpg->LoadFromFile(name);
        bmp->Assign(jpg);
        delete jpg;
        }
    bmp->HandleType=bmDIB;
    bmp->PixelFormat=pf32bit;
    Form1->ptb_gfx->Width=bmp->Width;
    Form1->ClientHeight=bmp->Height;
    Form1->ClientWidth=(bmp->Width<<1)+32;
    }
//---------------------------------------------------------------------------
__fastcall TForm1::TForm1(TComponent* Owner):TForm(Owner)
    {
    }
//---------------------------------------------------------------------------
void __fastcall TForm1::FormDestroy(TObject *Sender)
    {
    delete bmp;
    }
//---------------------------------------------------------------------------
void __fastcall TForm1::FormPaint(TObject *Sender)
    {
    draw();
    }
//---------------------------------------------------------------------------
void __fastcall TForm1::flb_picChange(TObject *Sender)
    {
    load(flb_pic->FileName);
    update();
    }
//---------------------------------------------------------------------------
void __fastcall TForm1::FormActivate(TObject *Sender)
    {
    flb_pic->SetFocus();
    flb_pic->Update();
    if (flb_pic->ItemIndex==-1)
     if (flb_pic->Items->Count>0)
        {
        flb_pic->ItemIndex=0;
        flb_picChange(this);
        }
    }
//---------------------------------------------------------------------------

```

Just ignore the VCL stuff and convert the resulting text into whatever you have at disposal. I also use mine dynamic list template so:


  
`List<double> xxx;` is the same as `double xxx[];`
  
`xxx.add(5);` adds `5` to end of the list
  
`xxx[7]` access array element (safe)
  
`xxx.dat[7]` access array element (unsafe but fast direct access)
  
`xxx.num` is the actual used size of the array
  
`xxx.reset()` clears the array and set `xxx.num=0`
  
`xxx.allocate(100)` preallocate space for `100` items


So use whatever list you got or recode or use `std::vector` instead...


I edited out the texts from your image:


[![input](https://i.stack.imgur.com/5SFpO.jpg)](https://i.stack.imgur.com/5SFpO.jpg)


And this is the result using that as input:



```
+-----------+------------------     |
|           |                       |
|           |                       |
|     |     |     |     +-----------+
|     |           |     |           |
|     |           |     |           |
|     +-----------+     +-----+     |
|                 |           |     |
|                 |           |     |
+------     |     +-----+     |     |
|           |           |           |
|           |           |           |
|     ------+-----------+------     |
|                                   |
|                                   |
|     ------------------------------+

```

And here the saved debug bitmaps (from left to right: BW,Grid,Maze):


[![BW](https://i.stack.imgur.com/Rg8lH.png)](https://i.stack.imgur.com/Rg8lH.png) [![Grid](https://i.stack.imgur.com/QMEoF.png)](https://i.stack.imgur.com/QMEoF.png) [![Maze](https://i.stack.imgur.com/tZjUP.png)](https://i.stack.imgur.com/tZjUP.png)


The only important stuff from the code is function:



```
AnsiString bmp2lintxt(Graphics::TBitmap *bmp);

```

Which returns text from VCL (GDI based) bitmap.


"
"How to find why control loses focus
I have bunch of controls, views and editors. I found that editor is loosing focus sometimes.
This issue hard to reproduce in debug because when I switch to breakpoint in debugger and back to application, editor is never loses focus.


Can you recommend tips & tricks how to find why control loses focus ?


","Adding some context to @david's answer:


You can add a `FocusListener` to all your `Widget`s. Within this listener you can output some information about the widget that lost/gained the focus.


To make things easier, you can add meta-data to the `Widget`s using `Widget#setData(Object)`.




---


Here is a code example that should help you figure things out:



```
public static void main(String[] args)
{
    Display display = new Display();
    final Shell shell = new Shell(display);
    shell.setText(""StackOverflow"");
    shell.setLayout(new FillLayout());

    FocusListener focusListener = new FocusListener()
    {
        @Override
        public void focusLost(FocusEvent e)
        {
            System.out.println(""Focus out: "" + e.widget.getData());
        }

        @Override
        public void focusGained(FocusEvent e)
        {
            System.out.println(""Focus in: "" + e.widget.getData());
        }
    };

    Text text = new Text(shell, SWT.BORDER);
    text.setText(""Text"");
    text.setData(""Text"");
    text.addFocusListener(focusListener);

    Button button = new Button(shell, SWT.PUSH);
    button.setText(""Button"");
    button.setData(""Button"");
    button.addFocusListener(focusListener);

    shell.pack();
    shell.open();
    while (!shell.isDisposed())
    {
        if (!display.readAndDispatch())
            display.sleep();
    }
    display.dispose();
}

```

Notice the utilization of `setData()` and `getData()`...


"
"Cannot find function name
I'm trying to do my first TypeScript/React project and I'm running into issues.


Using [this answer](https://stackoverflow.com/a/27852403), I have managed to read and play sound from my mic and also display some sample analysis data in the console. Now i'm trying to translate it into TS. Going step by step, I've arrived at this:



```
export class Processor {
    readonly BUFFER_SIZE = 16384;

    audioContext: AudioContext;
    gainNode: GainNode;
    microphoneStream: MediaElementAudioSourceNode;

    constructor() {
        this.audioContext = new AudioContext();

        console.log('audio is starting up ...');

        if (navigator.getUserMedia) {
            navigator.getUserMedia(
                { audio: true },
                function (stream) {
                    startMicrophone(stream);
                },
                function (e) {
                    alert('Error capturing audio.');
                });
        } else {
            alert('Seems like this browser might not be supported.');
        }
    }

    private startMicrophone(stream: MediaStream) {
        this.gainNode = this.audioContext.createGain();
        this.gainNode.connect(this.audioContext.destination);

        this.microphoneStream = 
this.audioContext.createMediaStreamSource(stream);
    }
}

```

Except the call to startMicrophone gives me



```
'Cannot find name 'startMicrophone'.'

```

I also tried to refer to it with `this`, which results in a different error:



```
''this' implicitly has type 'any' because it does not have a type annotation.'

```

I don't know what I'm doing wrong and could really use a bit of guidance.


","**Recommended:** You have to use arrow function if you want to use this because if you write this inside function block it refers current function this not parent this.



```
export class Processor {
    readonly BUFFER_SIZE = 16384;

    audioContext: AudioContext;
    gainNode: GainNode;
    microphoneStream: MediaElementAudioSourceNode;

    constructor() {
        this.audioContext = new AudioContext();

        console.log('audio is starting up ...');

        if (navigator.getUserMedia) {
            navigator.getUserMedia({
                    audio: true
                },
                (stream) => {
                    this.startMicrophone(stream);
                },
                (e) => {
                    alert('Error capturing audio.');
                });
        } else {
            alert('Seems like this browser might not be supported.');
        }
    }

    private startMicrophone(stream: MediaStream) {
        this.gainNode = this.audioContext.createGain();
        this.gainNode.connect(this.audioContext.destination);

        this.microphoneStream =
            this.audioContext.createMediaStreamSource(stream);
    }
}

```

Another way is you can assign this to some other variable and use `const self= this;` use self inside the function.



```
constructor() {
    const self = this;
    this.audioContext = new AudioContext();
    if (navigator.getUserMedia) {
        navigator.getUserMedia({
                audio: true
            },
            function (stream) {
                self.startMicrophone(stream);
            },
            function (e) {
                alert('Error capturing audio.');
            });
    } else {
        alert('Seems like this browser might not be supported.');
    }
}

```

"
"Dynamic allocation of classes
I have a question about object array allocation,
the problem is this:
I have a class called person :



```
class Person {
    char* name;
    int ID;


public:
    Person(char* name = NULL, int ID = 0) :name(name), ID(ID) {};
}

```

and I'm trying to make an array of persons like so:



```
Person *pArr = new Person[size];

```

and after that I extract the data from a file (char\* for string and int for ID)
and I use for loop and the constructors to place the persons inside the loop like so:



```
for (int j = 0; j < size3; j++) {
pArr[j] = Person(Name, id);
}

```

after I finish my program I want to use the Destructor and free the allocated string that's stored inside the char\* name, but when I add the Destructor it gets triggered right after the loops ends,
every Person that's being created is destroyed immediately,
I know its possible to make an array of pointers and allocated the person aswell, but in this assignment I'm suppose to do it like this, 
is there a right way to do it without triggering the destructor right away? 


","On this line:



```
pArr[j] = Person(Name, id);

```

You are creating a *temporary* `Person` object, and then assigning it to the `pArr[j]` object. `Person` has no explicit copy assignment operator defined, so the compiler auto-generates one for you, which simply performs a member-by-member copy of values from one object to another.


When the temporary goes out of scope on the `;`, it is automatically destroyed. The compiler-generated assignment operator copied the `name` pointer as-is to the `pArr[j]` object, so if `Person` has a destructor that frees its `name`, the `pArr[j]` object will be left with a dangling `name` pointer to freed memory.


Your `Person` class is not following the [Rule of Three](https://en.wikipedia.org/wiki/Rule_of_three_(C%2B%2B_programming)):



> 
> The rule of three (also known as the Law of The Big Three or The Big Three) is a rule of thumb in C++ (prior to C++11) that claims that if a class defines one (or more) of the following it should probably explicitly define all three:
> 
> 
> - destructor
> - copy constructor
> - copy assignment operator
> 
> 
> 


Since you want `Person` to have a destructor that frees `name`, it also needs a copy constructor and a copy assignment operator so it can make copies of `name` for the destructor to free, eg:



```
class Person
{
    char* name;
    int id;

public:
    Person(const char* Name = NULL, int ID = 0)
        : name(new char[std::strlen(Name)+1]), id(ID)
    {
        std::strcpy(name, Name);
    }

    Person(const Person &src)
        : name(new char[std::strlen(src.name)+1]), id(src.id)
    {
        std::strcpy(name, src.name);
    }

    ~Person()
    {
        delete[] name;
    }

    Person& operator=(const Person &rhs)
    {
        if (&rhs != this)
        {
            Person tmp(rhs);

            //std::swap(tmp.name, name);
            char *p = tmp.name;
            tmp.name = name;
            name = p;

            id = tmp.id;
        }
        return *this;
    }
};

```

Now, back on this line:



```
pArr[j] = Person(Name, id);

```

The temporary will be copied to `pArr[j]` safely and properly. Just note that if you dynamically allocated `Name` beforehand, you will now have to free it afterwards, since `Person` makes its own internal copy.




---


If you can change `char*` to `std::string`, then you don't have to worry about the compiler making default copies, since `std::string` is compliant with the Rule of Three and will be copied properly:



```
class Person
{
    std::string name;
    int id;

public:
    Person(const std::string &name = std::string(), int ID = 0)
        : name(name), id(ID)
    {
    }

    // no explicit copy constructor, destructor, or
    // assignment operator is needed, as compiler-generated
    // defaults will suffice...
};

```

"
"Should I put default values of attributes on the prototype to save space?
Suppose I've got a JavaScript function that I'm treating like a class - that is, I want to make many instances of it:



```
function Blerg() {
  this._a = 5;
}

Blerg.prototype.getA = function() {
  return this._a;
}

Blerg.prototype.setA = function(val) {
  this._a = val;
}

```

This class has one attribute, `a`, which the constructor instantiates to a default value of `5`. It is accessed with the getter and setter.


Now suppose I have such a class but it has 30+ attributes (`a`, `b`, `c`, etc), and suppose like `a` that these all have unique default values, but also that it is *uncommon for them to be changed from the default.*


Suppose also that I need to make 10,000 or more `Blerg` instances, and so a design goal is to save space.


I'm wondering if it is a good idea to put all of the *default values* for my 30+ attributes on the `prototype` of my class instead. That way, when I create 10,000+ instances of my class, none of them have `a`, `b`, `c`, etc attributes, but calling `getA()` will still return the correct default value.


So I present this modified `Blerg` function, `Blerg2`:



```
function Blerg2() {
  // nothing!
}

Blerg2.prototype._a = 5;

Blerg2.prototype.getA = function() {
  return this._a;
}

Blerg2.prototype.setA = function(val) {
  this._a = val;
}

```

Are there downsides to taking this approach?




---


### Some notes


The prototype way seems faster to create, see: <http://jsperf.com/blergs>


For that matter, the worst-case scenario does not look that bad: <http://jsperf.com/blergs/2>


And creating many of them using the code in the jsperf test (in its own html page) and adding:



```
var a = [];
for (var i = 0; i < 400000; i++) {
  a.push(new Blerg()); // or Blerg2
}

```

Suggests that the heap size for the objects in question is cut in half by using Blerg2.


","related jsperf <http://jsperf.com/12312412354>


Well this is a really bad idea. Objects are always considered having different class if they don't have exactly the same set of properties in the same order. So
a function that accepts these objects will in best case be polymorphic and in worst case megamorphic all the while you are thinking
you are passing it same class of objects. This is fundamental to all JS engines although the specifics that follow focus on V8.


Consider:



```
function monomorphic( a ) {
    return a.prop + a.prop;
}

var obj = {prop: 3};
while( true ) {
    monomorphic( obj );
}

```

Now, since the passed object `a` always has the same class, we will get really good code:



```
    ; load from stack to eax
15633697    23  8b4508         mov eax,[ebp+0x8] 
    ;; test that `a` is an object and not a small integer
1563369A    26  f7c001000000   test eax,0x1 
156336A0    32  0f8485000000   jz 171  (1563372B) ;;deoptimize if it is not
    ;; test that `a`'s class is as expected
156336A6    38  8178ffb9f7902f cmp [eax+0xff],0x2f90f7b9
156336AD    45  0f857d000000   jnz 176  (15633730) ;;deoptimize if it is not
    ;; load a.prop into ecx, as you can see it's like doing struct->field in C
156336B3    51  8b480b         mov ecx,[eax+0xb]
    ;; this will untag the tagged pointer so that integer arithmetic can be done to it
156336B6    54  d1f9           sar ecx,1
    ;; perform a.prop + a.prop
    ;; note that if it was a.prop + a.prop2
    ;; then it wouldn't need to do all those checks again
    ;; so for one time check inside the function we can load all the properties
    ;; quickly
156336B8    56  03c9           add ecx,ecx

```

Notice what happened here, V8 saw that we always pass the same class of object to the function
`monomoprhic` and generated really tight code that assumes we will always get that class of object
in the future as well.


Now let's do:



```
function polymorphic( a ) {
    return a.prop + a.prop;
}

var obj = {prop: 3};
var obj2 = {prop: 3, prop2: 4};
while( true ) {
    polymorphic( Math.random() < 0.5 ? obj : obj2 );
}

```

Now the function must consider 2 different classes of objects. The classes are different but similar enough
that the client code can stay as it is as both classes contain a field `prop`.


Let's see:



```
    ; load from stack to eax
04C33E17    23  8b4508         mov eax,[ebp+0x8]
    ;; test that `a` is an object and not a small integer
04C33E1A    26  f7c001000000   test eax,0x1
04C33E20    32  0f8492000000   jz 184  (04C33EB8) ;; deoptimize if not
    ;; test that `a`'s class is one of the expected classes
04C33E26    38  8178ffb9f7401c cmp [eax+0xff],0x1c40f7b9
04C33E2D    45  0f840d000000   jz 64  (04C33E40) ;; if it is, skip the second check and go to the addition code
    ;; otherwise check that `a`'s class is the second one of the expected classes
04C33E33    51  8178ff31f8401c cmp [eax+0xff],0x1c40f831
04C33E3A    58  0f857d000000   jnz 189  (04C33EBD) ;; deoptimize if not
    ;; load a.prop into ecx
    ;; if you are still reading this you will probably notice that this
    ;; is actually relying on the fact that both classes declared the prop field
    ;; first
04C33E40    64  8b480b         mov ecx,[eax+0xb]
    ;; this will untag the tagged pointer so that integer arithmetic can be done to it
04C33E43    67  d1f9           sar ecx,1
    ;; do the addition
04C33E45    69  03c9           add ecx,ecx

```

Ok so the situation is still pretty good but here we are relying on the fact that properties are in same order
and that there are only 2 different classes.


Let's do the same in different order so that V8 can't use the same instruction (`mov ecx,[eax+0xb]`) for both objects:



```
function polymorphic( a ) {
    return a.prop + a.prop;
}

var obj = {prop: 3};
var obj2 = {prop2: 4, prop: 3};
while( true ) {
    polymorphic( Math.random() < 0.5 ? obj : obj2 );
}

```

And:



```
06C33E84    36  8b4508         mov eax,[ebp+0x8]
    ;; small integer check
06C33E87    39  f7c001000000   test eax,0x1
06C33E8D    45  0f84d3000000   jz 262  (06C33F66)
    ;; class check 1
06C33E93    51  8178ffb9f75037 cmp [eax+0xff],0x3750f7b9
06C33E9A    58  7505           jnz 65  (06C33EA1)
06C33E9C    60  8b480b         mov ecx,[eax+0xb]
06C33E9F    63  eb10           jmp 81  (06C33EB1)
    ;; class check 2
06C33EA1    65  8178ff31f85037 cmp [eax+0xff],0x3750f831
06C33EA8    72  0f85bd000000   jnz 267  (06C33F6B)
06C33EAE    78  8b480f         mov ecx,[eax+0xf]

06C33EB1    81  f6c101         test_b cl,0x1
06C33EB4    84  0f851e000000   jnz 120  (06C33ED8)
06C33EBA    90  d1f9           sar ecx,1
06C33EBC    92  89ca           mov edx,ecx
06C33EBE    94  03d1           add edx,ecx

```

Ok just as expected, just using different offsets depending on the class.


So you can see where this is going, if you end up with 10 different classes then you will
just get 30 instructions (instead of 3) whenever a function will need to lookup a property. Which is still
much better than a hash table (100s of instructions?) lookup.


Well no, turns out there is [a limit of 4 different classes](https://github.com/v8/v8/blob/e6848c0cd37a2e6dbcaff5f7e13df53ded0165a9/src/ic.cc#L1024)
and then you go into megamorphic mode.


So with this, we should see radically different code output if we use 5 or more different classes:



```
function megamorphic( a ) {
    return a.prop + a.prop;
}

var objs = [
    {prop: 3},
    {prop3: 4, prop: 3, prop2: 4},
    {prop4: 4, prop2: 4, prop: 3, prop5: 6},
    {prop: 3, prop12: 6},
    {prop7: 15, prop30: 12, prop314: 4, prop34: 15, prop: 3}
];

while( true ) {
    var index = Math.random() * objs.length | 0;
    megamorphic( objs[index] );
}

```

Indeed:



```
3D3342E7    39  8b5508         mov edx,[ebp+0x8]
3D3342EA    42  b9f5e2a115     mov ecx,15A1E2F5
3D3342EF    47  e84c6ffeff     call LoadIC_Initialize  (3D31B240)
3D3342F4    52  8945ec         mov [ebp+0xec],eax
3D3342F7    55  8b75f0         mov esi,[ebp+0xf0]
3D3342FA    58  8b5508         mov edx,[ebp+0x8]
3D3342FD    61  b9f5e2a115     mov ecx,15A1E2F5
3D334302    66  e8396ffeff     call LoadIC_Initialize  (3D31B240)
3D334307    71  8b4dec         mov ecx,[ebp+0xec]
3D33430A    74  f6c101         test_b cl,0x1
3D33430D    77  0f8527000000   jnz 122  (3D33433A)
3D334313    83  d1f9           sar ecx,1
3D334315    85  89c2           mov edx,eax
3D334317    87  f6c201         test_b dl,0x1
3D33431A    90  0f8546000000   jnz 166  (3D334366)
3D334320    96  d1fa           sar edx,1
3D334322    98  03d1           add edx,ecx

```

That actually looks a lot like code that loads properties from objects that are in hash table mode. Is it fast? Unfortunately jsperf is down now
so you have to run it yourself with code [here](http://pastebin.com/TJqZDUHV):



```
Megamorphic 492
Monomorphic 30

```

So the same function body ran 15 times faster because the object passed to it always had the same class. It is also hard to predict if you will actually save memory
if all those 10000 objects allocate a different class for instance.


It is also complicated to really look at all the downsides but having optional properties is one of those things that it is easy to say is very bad.


"
"Angular 2 Forms: binding html select to collection of objects
I would like to display a collection of selectable objects in an angular 2 application (RC 5, forms 0.3.0)



```
<select [(ngModel)]=""selectedItem"">       
  <option *ngFor=""let item of selectableItems""
          [value]=""item"">
    {{ item }}
  </option>
</select>

<div> {{ selectedItem }} </div>

```

The list itself is correctly displayed.


But all that is displayed for 'selectedItem' is [object Object]. When accessing the item in code, I get the corresponding string ""[object Object]"". I tried switching to ngValue but that produces the same results.


The whole thing works when I use primitive values instead of objects. But I suspect that I am missing some crucial point here.


Thanks for your help. I've been wasting several hours on Internet search and trial and error. Maybe somebody has encountered the same problem.


**Edit** (24-08-16): I came across this tutorial which explains how to create common forms widgets today. Maybe it's useful for somebody who lands on this page: <https://scotch.io/tutorials/how-to-deal-with-different-form-controls-in-angular-2>


","From your code and explanation, I assume ""item"" is a complex object, and not a simple object. So to display it in HTML you would use:



```
<div> {{ selectedItem | json }} </div>

```

In code you have to access the ""item"" properties (as in item.value or whatever the item's properties are. If you want to work with simple values you can do something like this:



```
<select [(ngModel)]=""selectedItem"">       
  <option *ngFor=""let item of selectableItems""
          [value]=""item.id"">
    {{ item }}
  </option>
</select>

```

This is of course assuming item is a complex type.


**Edit**


After some digging (because this has plagued me as well), I found the solution I really wanted! To use complex object you just need to use [ngValue] when binding the value! The fix is described at <https://github.com/angular/angular/issues/4843>. So the above will look like this:



```
<select [(ngModel)]=""selectedItem"">       
     <option *ngFor=""let item of selectableItems""
             [ngValue]=""item"">
             {{item}}
     </option>
</select>

```

And then selectedItem will the the correct item from the list, including all it's properties.


"
"How do I pass the this context into an event handler?
I know this question doesn't make much sense, but let me try and clarify a bit.


I have a class, called `ScrollBanner`, and it looks somewhat as follows (a lot omitted for brevity):



```
function ScrollBanner() {
    this.initialize = function(selector) {
        $('span#banner1-nav').click(this._onClickNavigation);
    }

    this._onClickNavigation = function(event) {
        this.restartTimer(); // this == span#banner1-nav element from this.initialize
        //...
    }

    this.restartTimer() {
        //...
    }
}

```

As you can see `this.initialize` sets a click handler to `this._onClickNavigation`. Some might expect the `this` inside the event handler to refer to the `ScrollBanner` instance, but sadly it doesn't. It refers to the element that triggered the click event, in this case `span#banner1-nav`.


What would be the best way to get `this` to refer to the `ScrollBanner` class instance?


","**The old/traditional way:**


Capture `this` in a variable:



```
this.initialize = function(selector) {
    var that = this;
    $('span#banner1-nav').click(function(event) {
       that._onClickNavigation(event);
    });
}

```

You could also assign `this` to a variable e.g. `instance`:



```
function ScrollBanner() {
    var instance = this;
    // ...
}

```

and refer to `instance` instead of `this` in all the calls.


The overall idea is to store `this` in a variable in a higher scope.




---


**The ECMAScript5 way:**


ECMAScript5 introduces a new property of functions: [`.bind()`](http://ecma262-5.com/ELS5_HTML.htm#Section_15.3.4.5). [MDC's documentation](http://developer.mozilla.org/en/JavaScript/Reference/Global_Objects/Function/bind) shows an implementation for browsers that don't support it. With it you can bind a certain context to a function:



```
this.initialize = function(selector) {
    $('span#banner1-nav').click(this._onClickNavigation.bind(this));
}

```

but behind the scenes it is doing the same thing. The advantage is that you make use of built-in functionality in browser that support is.


Note that this is different from [`apply`](https://developer.mozilla.org/en/JavaScript/Reference/Global_Objects/function/apply) or [`call`](https://developer.mozilla.org/en/JavaScript/Reference/Global_Objects/function/call). Both of these set the context **and** execute the function, whereas `bind` only sets the context without executing the function.




---


**The jQuery way:**


jQuery provides a method [`$.proxy()`](http://api.jquery.com/jQuery.proxy/) that is doing the same:



```
$('span#banner1-nav').click($.proxy(this._onClickNavigation, this));

```

"
"IntelliJ says the package does not exist, But I can access the package
I get the error below, but I can find the package in rt.jar.
I can see the JDK being used from Project Structure.
I'm not sure what's missing.



```
Error:(6, -1) Play 2 Compiler: 
 C:\user\projects\portal\app\com\example\security\cert\X509Cert.java:6: package sun.security.pkcs10 does not exist
 import sun.security.pkcs10.*;
 C:\user\projects\portal\app\com\v\security\cert\GenerateCSR.java:75: cannot find symbol

```

","This is not an intellij issue.


You need to compile with the `-XDignore.symbol.file` option. Some internal packages
are hidden by default unless you add this option to javac.


<https://bugs.openjdk.java.net/browse/JDK-7141442>


[sun.security.x509.CertAndKeyGen and sun.security.pkcs.PKCS10 - missing in JDK8. Implementation suggestions](https://stackoverflow.com/questions/29060064/sun-security-x509-certandkeygen-and-sun-security-pkcs-pkcs10-missing-in-jdk8)


Also the comment of user @user180100 is relevant and important.



> 
> The sun.\* packages are not part of the supported, public interface.  
> 
>  A
>  Java program that directly calls into sun.\* packages is not guaranteed
>  to work on all Java-compatible platforms. In fact, such a program is
>  not guaranteed to work even in future versions on the same platform.
> 
> 
> 


"
"Kotlin null checks(!!) when nullability is already checked
I have a question about Kotlin and its null checks warnings.


Let's assume I created an object called ""user"" with some attribute like name, surname, etc etc. The following code is an example:



```
if(user != null) {
    val name = user!!.name
    val surname = user.surname
    val phoneNumber = user.phoneNumber
} else 
    // Something else

```

Why, even if I checked that user is not null, Kotlin wants me to use !! the first time I call user? It cannot be null at this point.


I know that I can use the following block but I don't understand this behavior.



```
user?.let{
    // Block when user is not null
}?:run{
    // Block when user is null
}

```

","There's a reason for that behavior. Basically, it's because the compiler can't ensure that the value of `user` doesn't become null after the `if` check. 


This behavior is only applicable for `var user`, not for `val user`. So for example,



```
val user: User? = null;
if (user != null) {
  // user not null
  val name = user.name // won't show any errors
}

```


```
var user: User? = null;
if (user != null) {
  // user might be null
  // Since the value can be changed at any point inside the if block (or from another thread).
  val name = user.name // will show an error
}

```

`let` allows you to ensure immutability even for `var` variables. `let` creates a new final value separate from the original variable.



```
var user: User? = null
user?.let {
  //it == final non null user
  //If you try to access 'user' directly here, it will show error message,
  //since only 'it' is assured to be non null, 'user' is still volatile.
  val name = it.name // won't show any errors
  val surname = user.surname // will show an error
}

```

"
"Capture logs from Chrome during test is running. Python
I'm trying to capture logs from chrome, during automated test is running. Code is below:



```
 binary_path = 'Chromium'
chromedriver_path = 'chromedriver'

opts = Options()
opts.binary_location = binary_path
d = DesiredCapabilities.CHROME
d['goog:loggingPrefs'] = { 'browser':'ALL',
                           'driver': 'ALL'}

driver = webdriver.Chrome(chromedriver_path, desired_capabilities=d, options=opts)

#EXAMPLE TO GET LOGS
driver.get('https://www.google.com/')
driver.find_element_by_name('q').send_keys('abc')

print('browser = ', driver.get_log('browser'))
print('driver = ', driver.get_log('driver'))
for entry in driver.get_log('browser'):
    print('entry = ', entry)

driver.quit()

```

output from print:



```
browser =  []
driver =  []

```

My question is, why entry isn't printing and why there is no output from chrome logs


","Try using [ChromeOptions](https://chromedriver.chromium.org/capabilities) instead of Options. 


Also, not sure why you use Chromium, this might affect the result. 
I modified your code and following code worked for me with Chrome browser: 



```
from selenium.webdriver.common.desired_capabilities import DesiredCapabilities
from selenium.webdriver import ChromeOptions

opts = ChromeOptions()
d = DesiredCapabilities.CHROME
d['goog:loggingPrefs'] = {'browser': 'ALL',
                           'driver': 'ALL'}

driver = webdriver.Chrome(desired_capabilities=d, options=opts)

#EXAMPLE TO GET LOGS
driver.get('https://www.google.com/')
driver.find_element_by_name('q').send_keys('abc')

print('browser = ', driver.get_log('browser'))
print('driver = ', driver.get_log('driver'))
for entry in driver.get_log('browser'):
    print('entry = ', entry)

driver.quit()

```

it printed out => 



```
browser =  []
driver =  [{'level': 'INFO', 'message': 'Populating Preferences file: {\n   ""alternate_error_pages"": {\n      ""enabled"": false\n   },\n   ""autofill"": {\n      ""enabled"": false\n   }, ... ]

```

I hope this helps, good luck!


"
"What is the difference between a smartcard and SIM card?
I'm aware that a SIM card is electrically just a smart card but to focus my question more, can I use any smartcard with contacts that I've cut down to size as a SIM card if programmed correctly. Is software the only thing that makes a SIM card a smartcard? I know I can just cut the plastic away and that shouldn't affect the chip itself. Is there proprietary code loaded on first? 


I'm trying to write my own SIM cards for a private LTE network I'm developing. Smartcards are cheap, but ""SIM"" cards from gemalto are not. I can use pysim well enough, can I just plug a smart card into that and go to town? There's a lot of security concerns as well, do regular smartcards fulfill the SIM spec?


Thanks,
Ryan


","SIM card is a smart card (similarly a web server is a ""[server](https://en.wikipedia.org/wiki/Server_(computing))"") it just implements SIM card specific APIs (like a web server implements HTTP).


Mobile network operators use available SIM card products which they personalize/customize to fit their needs (just like internet projects use existing web server software and configure it to serve their content).


Technically nothing prevents you from implementing your own SIM card on a compatible smart card (in the same way you would implement a web server on any ""server"" supporting TCP/IP communication) as SIM card API requirements are publicly available /e.g. [GSM 11.11](http://www.etsi.org/deliver/etsi_gts/11/1111/05.03.00_60/gsmts_1111v050300p.pdf),[3GPP TS 51.011](https://portal.3gpp.org/desktopmodules/Specifications/SpecificationDetails.aspx?specificationId=2793).../ (just like HTTP is specified in [RFC 7230](https://www.rfc-editor.org/rfc/rfc7230)).


It would be much easier to implement it on some Java Card /which is open and quite well supported/ than on some ""bare chip"" /which is quite complicated to get and develop for/ (similarly it is much easier to develop a web server for linux than for [Cisco IOS](https://en.wikipedia.org/wiki/Cisco_IOS)).


For a proof-of-concept SIM card implementation have a look at [SSIM](https://github.com/vlp/ssim) which I wrote some long time ago for PIC16F877+24LC256 Silver Wafer Card (unfortunately in assembly).


You might find [Sysmocom SIM/USIM cards for sale](http://shop.sysmocom.de/t/sim-card-related/sim-cards) useful (I am not sure about LTE support thought).


Disclaimer: My hobbyist GSM experience ended several years ago (before 2.5G) so please do validate my thoughts...


Good luck!


"
"Is a URL starting with // valid?
We can see many HTML page using `src=""//example.com/myjavascript.js""` to include a **.js** and let the browser use `http://` / `https://` depending on the current scheme of the page URL.


Is this a valid URI ?


","**Yes, it is definitely valid.** It is a ""scheme relative"" or ""protocol relative"" URI. In the spec since the beginning. Quite helpful to deal with http/https issues.


Read a much better description and caveats:
[Is it valid to replace http:// with // in a <script src=""http://..."">?](https://stackoverflow.com/questions/550038/is-it-valid-to-replace-http-with-in-a-script-src-http)


A few points to keep in mind:


- There are some minor issues on IE 7 & 8 with double downloads
- If you are viewing an HTML page from a file the browser will replace the scheme with `file://`, and not load your JS file from the server like it would with a full URL starting with `http://` or `https://`.


**Edit for modern webdev practices:**


While the URL is still valid, it is now recommended to use `https` for third party resources (and, serve those resources from secure pages). The performance or compatibility issues from years ago are largely resolved with updated protocols and browsers.


"
"Using a base class for an integration test
My base class for an integration test:



```
[TestClass]
public class BaseIntegrationTest
{
    protected readonly ApplicationDbContext _dbContext;

    public BaseIntegrationTest()
    {
        _dbContext = new ApplicationDbContext(""DefaultConnectionTest"");
    }

    [ClassInitialize]
    public void Initialize()
    {
        InsertTestData();
    }

    public void InsertTestData()
    {
        _dbContext.Users.Add(new User { Name = ""John Doe"" });
    }
}

```

Here's a test class that inherits from `BaseIntegrationTest`:



```
[TestClass]
public class UserRepositoryTests : BaseIntegrationTest
{
    // how to lose this method here and use the one in the BaseIntegrationTest?
    //[ClassInitialize]
    //public static void Setup(TestContext context)
    //{
    //  BaseIntegrationTest test = new BaseIntegrationTest();
    //  test.InsertTestData();
    //}

    [TestMethod]
    public void GetUser()
    {
        Filter f = new Filter()
        {
            Name = ""John Doe""
        };

        UserRepository fr = new UserRepository(_dbContext);
        var result = fr.GetUser(f);

        Assert.IsTrue(result.Any(x => x.Name == ""John Doe""));
    }
}

```

The method decorated with the `[ClassInitialize]` attribute in the `BaseIntegrationTest` never gets executed, and because of that, my tests fail. No data gets inserted into the database.


But if I uncomment the `public static void Setup(TestContext context)` in the `UserRepositoryTests` it works.


I have a few more repositories I want to test. And I want the data being inserted into the database being done in the base class so that I can run all the tests without recreating the database for every repository I want to test.


Converting `public void Initialize()` into `public static void Initialize(TestContext context)` doesn't work either. The `Initialize()` wont get executed/called.


Am I missing something? Or are these limitations of unit test/integration test? I'm pretty new to unit testing/integration testing.


","From MSDN:



> 
> Will the same approach that was used for `[TestInitialize]` work with
>  `[ClassInitialize]` in a base class?
> 
> 
> Not exactly, if you create a `[ClassInitialize]` attributed method in
>  the base class it won’t ever get called unless you explicitly call it
>  at the beginning of your derived test classes `ClassInitialize`
>  method; which of course is nowhere near as nice as the above approach.
> 
> 
> 


<http://blogs.msdn.com/b/densto/archive/2008/05/16/using-a-base-class-for-your-unit-test-classes.aspx>


Later on it describes how to do initialisation before any methods in your inheriting class:



> 
> How can I create an initialization method that will run before any
>  class initialization methods in my test project?
> 
> 
> Building on the common base class approach described above you could
>  simply add a static constructor to your base class and either perform
>  the initialization there or call the method that will perform the
>  desired initialization. The resulting base class might look like
>  this:
> 
> 
> 



```
[TestClass]
public class TestBase
{
    static TestBase()
    {
        s_log = new StringBuilder();
        Log.AppendLine(""TestBase.ctor()"");
    }

    [TestInitialize]
    public void BaseTestInit()
    {
        Log.AppendLine(""TestBase.BaseTestInit()"");
    }

    [TestCleanup]
    public void BaseTestCleanup()
    {
        Console.WriteLine(Log.ToString());
    }

    public static StringBuilder Log
    {
        get { return s_log; }
    }

    static StringBuilder s_log;
} 

```

"
"PowerShell functions and performance
I've been wondering about the performance impact of functions in PowerShell.
Let's say we want to generate 100.000 random numbers using System.Random.



```
$ranGen = New-Object System.Random

```

Executing



```
for ($i = 0; $i -lt 100000; $i++) {
    $void = $ranGen.Next()
}

```

finishes within 0.19 seconds.
I put the call inside a function 



```
Get-RandomNumber {
    param( $ranGen )

    $ranGen.Next()
}

```

Executing



```
for ($i = 0; $i -lt 100000; $i++) {
    $void = Get-RandomNumber $ranGen
}

```

takes about 4 seconds.


Why is there such a huge performance impact?


Is there a way I can use functions and still get the performance I have with the direct call?


Are there better (more performant) ways of code encapsulation in PowerShell?


","a function call is *expensive*. The way to get around that is to put as much as you can IN the function. take a look at the following ... 



```
$ranGen = New-Object System.Random
$RepeatCount = 1e4

'Basic for loop = {0}' -f (Measure-Command -Expression {
    for ($i = 0; $i -lt $RepeatCount; $i++) {
        $Null = $ranGen.Next()
    }
    }).TotalMilliseconds

'Core in function = {0}' -f (Measure-Command -Expression {
function Get-RandNum_Core {
    param ($ranGen)
    $ranGen.Next()
    }

    for ($i = 0; $i -lt $RepeatCount; $i++) {
        $Null = Get-RandNum_Core $ranGen
    }
    }).TotalMilliseconds

'All in function = {0}' -f (Measure-Command -Expression {
    function Get-RandNum_All {
        param ($ranGen)
        for ($i = 0; $i -lt $RepeatCount; $i++) {
            $Null = $ranGen.Next()
            }
        }
    Get-RandNum_All $ranGen
    }).TotalMilliseconds

```

output ... 



```
Basic for loop = 49.4918
Core in function = 701.5473
All in function = 19.5579

```

from what i vaguely recall [and can't find again], after a certain number of repeats, the function scriptblock gets JIT-ed ... that seems to be where the speed comes from. 


"
"How to really target table at runtime in Entity Framework Core
I have `KPI` tables and now it has more than 500M rows. I would like to split this table into `KPI_Year2011`, `KPI_Year2012`...`KPI_2020` (will be created on 1st January 2020).


PS: All tables have same structure and Entity type is `KPI`.


How would you target table at runtime like this 



```
public IActionResult GetKpis(int year) {
    /*var kpis = context.KPI.where(k => k.year == year).ToList(); */
    var kpis = context.SetTargetTable($""KPI_Year{year}"").ToList();
    return View(kpis);
}

```

","For you and your program, logically, it's one big table. And the fact that the database needs to split it into multiple smaller ones to make access faster should be none of your business.


This is called **Partitioning** and should really be a job of the database.


For example:


[Partitioning in SQL Server](https://learn.microsoft.com/en-us/sql/relational-databases/partitions/create-partitioned-tables-and-indexes?view=sql-server-2017)


[Partitioning in Oracle](https://www.oracle.com/technetwork/database/options/partitioning/overview/index.html)


[Partitioning in Postgres](https://www.postgresql.org/docs/10/ddl-partitioning.html)


[Partitioning in MariaDB](https://mariadb.com/kb/en/library/partitioning-overview/)


You could manually implement it in your program logic, but that would just mean you reinvent the wheel. In fact, probably a much crappier and tinier wheel that breaks more often and does not really fit onto the rest of your car.


Once you leave this job to the database, your question has a solution: do nothing. Your program does not even know it's multiple tables in the database. It's one *partitioned* table.


"
"How do I keep the smaller value in a bar chart in front (when I cannot place the bars side by side)?
I am plotting two sets of data on a chart that has the amino acid position along the x-axis. Sometimes assay 1 as a smaller value and sometimes assay 2 has the smaller value. How do I ensure that the lower value is always in front? 


I am using python and matplotlib / pyplot.


Right now I am adding two bar charts as:



```
p1 = plot.bar(x1, y1, color='r', alpha=1, edgecolor='none',linewidth=0,width=0.5, log=False)
p2 = plot.bar(x2, y2, color='b', alpha=1, edgecolor='none', linewidth=0,width=0.5, log=False)

```

Thanks!


","You can use the [`numpy.ma`](http://docs.scipy.org/doc/numpy/reference/maskedarray.generic.html#the-numpy-ma-module) module to define masks:



```
import numpy.ma as ma

mask1 = ma.where(y1>=y2)
mask2 = ma.where(y2>=y1)

p1 = plt.bar(x1[mask1], y1[mask1], color='r', alpha=1, edgecolor='none',linewidth=0,width=0.5, log=False)
p2 = plt.bar(x2, y2, color='b', alpha=1, edgecolor='none', linewidth=0,width=0.5, log=False)
p3 = plt.bar(x1[mask2], y1[mask2], color='r', alpha=1, edgecolor='none',linewidth=0,width=0.5, log=False)

```

Example:



```
import matplotlib.pyplot as plt
import numpy as np
import numpy.ma as ma

x1 = x2 = np.arange(5)
y1 = np.array([1,4,25,2,4])
y2 = np.array([4,2,3,32,6])

mask1 = ma.where(y1>=y2)
mask2 = ma.where(y2>=y1)

p1 = plt.bar(x1[mask1], y1[mask1], color='r', alpha=1, edgecolor='none',linewidth=0,width=0.5, log=False)
p2 = plt.bar(x2, y2, color='b', alpha=1, edgecolor='none', linewidth=0,width=0.5, log=False)
p3 = plt.bar(x1[mask2], y1[mask2], color='r', alpha=1, edgecolor='none',linewidth=0,width=0.5, log=False)

```

![enter image description here](https://i.stack.imgur.com/0EpRi.png)


"
"Fitting a distribution to data - MATLAB
I am trying to fit a distribution to some data I've collected from microscopy images. We know that the peak at about 152 is due to a Poisson process. I'd like to fit a distribution to the large density in the center of the image, while ignoring the high intensity data. I know how to fit a Normal distribution to the data (red curve), but it doesn't do a good job of capturing the heavy tail on the right. Although the Poisson distribution should be able to model the tail to the right, it doesn't do a very good job either (green curve), because the mode of the distribution is at 152. 



```
PD = fitdist(data, 'poisson');

```

The Poisson distribution with lambda = 152 looks very Gaussian-like. 


Does anyone have an idea how to fit a distribution that will do a good job of capturing the right-tail of the data?


![enter image description here](https://i.stack.imgur.com/pPiX4.png)


[Link to an image showing the data and my attempts at distribution fitting.](https://i.stack.imgur.com/2GwvF.jpg) 


","The distribution looks a bit like an [Ex-Gaussian](http://en.wikipedia.org/wiki/Exponentially_modified_Gaussian_distribution) (see the green line in the first wikipedia figure), that is, a mixture model of a normal and an exponential random variable.


On a side note, are you aware that, although the events of a poisson process are poisson distributed, the waiting times between the events are exponentially distributed? Given that a gaussian noise added to your measurement, an ex-gaussian distribution could be theoretically possible. (Of course this does not mean that this is also plausible.)


A tutorial on fitting the ex-gaussian with MatLab can be found in



> 
> Lacouture Y, Cousineau D. (2008)
>  How to use MATLAB to fit the ex‐Gaussian and other probability functions to a distribution of response times. 
>  Tutorials in Quantitative Methods for Psychology 4 (1), p. 35‐45.
>  <http://www.tqmp.org/Content/vol04-1/p035/p035.pdf>
> 
> 
> 


"
"How do I set up NUnit to run my project's unit tests?
I'm starting to use NUnit to write test cases in C# with Visual Studio 2010 on .NET 4.0. I want to use NUnit to test against a .dll (a C# class library project)'s public functions. How do I set up NUnit to work with my project?


1. Should I add NUnit code to the same class library project to test against, or should I add a separate project in the same solution for NUnit test cases? Which is the best practice?
2. If I need to create a separate project for the NUnit test cases, should I make it a class library project or an executable? If I made it a class library project, how do I run it?
3. If I need to test against an executable and *not* a class library project, are there any changes to the process and/or projects?


","Re 2


Generally, keep your [test case dll] separate from your [business logic dll]. Your business logic dll shouldn't include any knowledge of NUnit, to separate concerns and simplify deployment/maintenance.


Your test case dll should include a reference to NUnit, and to your business logic dll.


You do not need to share a namespace. You can expose internal members of your business logic dll to your test case dll by modifying the business logic dll's AssemblyInfo.cs file to expose internals to the test case dll. This lets you preserve your desired visibility in the business logic dll.


Re 3


Your test cases should go in a dll (i.e. class library project). You can load this directly into NUnit's UI, or use NUnit's console runner in an integration environment to run your tests automatically.


How I do it:


- open test case class library project
properties, Debug tab
- set to open with external program:
point this to your nunit.exe
- set command line arguments to the
exact name of your test case dll:
MyTests.dll
- working directory: just click the
ellipsis button and it will
pre-select your test dll output
directory for the current build
config
- set the test project to be the
default start project in the
solution; this way, whenever you hit
F5 (or the ""Play"" button), NUnit will
come right up with your updated tests
preloaded - very convenient and
quick.


Best of luck - also try out the Test project type avl. in Visual Studio, it's very similar to NUnit. I still prefer NUnit but while learning it's good to try some various options.


"
"CentOS is ignoring my public key
I am trying to connect to my own CentOS server using SSH; I am able to connect fine with a password; I am trying to connect with my public key, but the server does not attempt public key authentication. I have followed all instructions on [this page](http://wiki.centos.org/HowTos/Network/SecuringSSH#head-9c5717fe7f9bb26332c9d67571200f8c1e4324bc) in the CentOS wiki, and verified the following things:


- I am running CentOS 6.0 or later
- My sshd\_config has the following:



```
PermitRootLogin no
PubkeyAuthentication Yes
RSAAuthentication yes
AuthorizedKeysFile .ssh/authorized_keys

```
- I have created a private key in the .ssh/ folder in my home directory on the client
- The full text of the public key (generated by `ssh-keygen`) takes up a single line in the .ssh/authorized\_keys file in my home directory on the server.
- I have the following permission modes set on both client and server:



```
chmod 700 .ssh
chmod 600 .ssh/*

```


","According this [issue report](http://bugs.centos.org/view.php?id=4959) on the CentOS web site, this is caused by an upstream ""feature"" in SELinux; if SELinux is enforcing, it will prevent the server from using pubkey authentication. There is no indication that this behavior will change in the near future. To see if this problem applies to you, type `sestatus` and see if it reports ""Current mode"" as `enforcing`.


I know of two workarounds.


- You can set SELinux to permissive mode. Type `setenforce 0` to change the policy immediately, and then open `/etc/selinux/config` and make sure `SELINUX=permissive`.
- Restore the file to the correct security context:



```
restorecon -R -v ~$USER/.ssh

```


"
"Kernel in a logistic regression model LogisticRegression scikit-learn sklearn
How can I use a kernel in a logistic regression model using the sklearn library? 



```
logreg = LogisticRegression()

logreg.fit(X_train, y_train)

y_pred = logreg.predict(X_test)
print(y_pred)

print(confusion_matrix(y_test,y_pred))
print(classification_report(y_test,y_pred))
predicted= logreg.predict(predict)
print(""Accuracy:"",metrics.accuracy_score(y_test, y_pred))

```

","Very nice question but `scikit-learn` currently does not support neither kernel logistic regression nor the ANOVA kernel.


You can implement it though.


**Example 1 for the ANOVA kernel:**



```
import numpy as np
from sklearn.metrics.pairwise import check_pairwise_arrays
from scipy.linalg import cholesky
from sklearn.linear_model import LogisticRegression

def anova_kernel(X, Y=None, gamma=None, p=1):
    X, Y = check_pairwise_arrays(X, Y)
    if gamma is None:
        gamma = 1. / X.shape[1]

    diff = X[:, None, :] - Y[None, :, :]
    diff **= 2
    diff *= -gamma
    np.exp(diff, out=diff)
    K = diff.sum(axis=2)
    K **= p
    return K

# Kernel matrix based on X matrix of all data points
K = anova_kernel(X)
R = cholesky(K, lower=False)

# Define the model
clf = LogisticRegression()

# Here, I assume that you have split the data and here, train are the indices for the training set
clf.fit(R[train], y_train)
preds = clf.predict(R[test])¨

```



---


**Example 2 for Nyström:**



```
from sklearn.kernel_approximation import Nystroem
from sklearn.linear_model import LogisticRegression
from sklearn.pipeline import Pipeline

K_train = anova_kernel(X_train)
clf = Pipeline([
    ('nys', Nystroem(kernel='precomputed', n_components=100)),
    ('lr', LogisticRegression())])
clf.fit(K_train, y_train)

K_test = anova_kernel(X_test, X_train)
preds = clf.predict(K_test)

```

"
"Entity Framework many-to-many question
Please help an EF n00b design his database.
I have several companies that produce several products, so there's a many-to-many relationship between companies and products. I have an intermediate table, Company\_Product, that relates them.


Each company/product combination has a unique SKU. For example Acme widgets have SKU 123, but Omega widgets have SKU 456. I added the SKU as a field in the Company\_Product intermediate table.


EF generated a model with a 1:\* relationship between the company and Company\_Product tables, and a 1:\* relationship between the product and Company\_Product tables. I really want a *:* relationship between company and product. But, most importantly, there's no way to access the SKU directly from the model.


Do I need to put the SKU in its own table and write a join, or is there a better way?


","I just tested this in a new VS2010 project (EFv4) to be sure, and here's what I found:


When your associative table in the middle (Company\_Product) has ONLY the 2 foreign keys to the other tables (CompanyID and ProductID), then adding all 3 tables to the designer ends up modeling the many to many relationship. It doesn't even generate a class for the Company\_Product table. Each Company has a Products collection, and each Product has a Companies collection.


However, if your associative table (Company\_Product) has other fields (such as SKU, it's own Primary Key, or other descriptive fields like dates, descriptions, etc), then the EF modeler will create a separate class, and it does what you've already seen.


Having the class in the middle with 1:\* relationships out to Company and Product is not a bad thing, and you can still get the data you want with some easy queries.



```
// Get all products for Company with ID = 1
var q =
    from compProd in context.Company_Product
    where compProd.CompanyID == 1
    select compProd.Product;

```

True, it's not as easy to just navigate the relationships of the model, when you already have your entity objects loaded, for instance, but that's what a data layer is for. Encapsulate the queries that get the data you want. If you really want to get rid of that middle Company\_Product class, and have the many-to-many directly represented in the class model, then you'll have to strip down the Company\_Product table to contain only the 2 foreign keys, and get rid of the SKU.


Actually, I shouldn't say you HAVE to do that...you might be able to do some edits in the designer and set it up this way anyway. I'll give it a try and report back.


**UPDATE**


Keeping the SKU in the Company\_Product table (meaning my EF model had 3 classes, not 2; it created the Company\_Payload class, with a 1:\* to the other 2 tables), I tried to add an association directly between Company and Product. The steps I followed were:


- Right click on the Company class in the designer
- Add > Association
- Set ""End"" on the left to be Company (it should be already)
- Set ""End"" on the right to Product
- Change both multiplicities to ""\* (Many)""
- The navigation properties should be named ""Products"" and ""Companies""
- Hit OK.
- Right Click on the association in the model > click ""Table Mapping""
- Under ""Add a table or view"" select ""Company\_Product""
- Map Company -> ID (on left) to CompanyID (on right)
- Map Product -> ID (on left) to ProductID (on right)


But, it doesn't work. It gives this error:
Error 3025: Problem in mapping fragments starting at line 175:Must specify mapping for all key properties (Company\_Product.SKU) of table Company\_Product.


So that particular association is invalid, because it uses Company\_Product as the table, but doesn't map the SKU field to anything.


Also, while I was researching this, I came across this ""Best Practice"" tidbit from the book Entity Framework 4.0 Recipies (note that for an association table with extra fields, besides to 2 FKs, they refer to the extra fields as the ""payload"". In your case, SKU is the payload in Company\_Product).



> 
> **Best Practice**
> 
> 
> Unfortunately, a project
>  that starts out with several,
>  payload-free, many-to-many
>  relationships often ends up with
>  several, payload-rich, many-to-many
>  relationships. Refactoring a model,
>  especially late in the development
>  cycle, to accommodate payloads in the
>  many-to-many relationships can be
>  tedious. Not only are additional
>  entities introduced, but the queries
>  and navigation patterns through the
>  relationships change as well. Some
>  developers argue that every
>  many-to-many relationship should start
>  off with some payload, typically a
>  synthetic key, so the inevitable
>  addition of more payload has
>  significantly less impact on the
>  project.
> 
> 
> So here's the best practice.
>  If you have a payload-free,
>  many-to-many relationship and you
>  think there is some chance that it may
>  change over time to include a payload,
>  start with an extra identity column in
>  the link table. When you import the
>  tables into your model, you will get
>  two one-to-many relationships, which
>  means the code you write and the model
>  you have will be ready for any number
>  of additional payload columns that
>  come along as the project matures. The
>  cost of an additional integer identity
>  column is usually a pretty small price
>  to pay to keep the model more
>  flexible.
> 
> 
> 


(From Chapter 2. Entity Data Modeling Fundamentals, 2.4. Modeling a Many-to-Many Relationship with a Payload)


Sounds like good advice. Especially since you already have a payload (SKU).


"
"Best practices for architecturing data validation in a Django multi sided project
I know it's a broad question.
But let's imagine you have a Django project with an API used by mobile app clients.


There are many places where you can add validation logic:


- Models fields themselves
- Models, by overriding `save()` method
- Django forms and model forms.
- API Serializers (generic serializers or model serializers)


When architecturing your project, what rules/elements help you to choose whether validation should be put here or there?


","This is an opinion and style question. I think that that models validation is an absolute requirement, the last defence line against mistakes. Once models validations are in place, you can go on with forms, client and API facing validation. 


The specific model validation, fields or model methods (or custom managers etc) is a choice of style, and what works for the specific app, but keep models validation. It's the one place to catch it all. You may have many forms, many APIs, but one point where data goes into the DB.


One of django advantages is that in many cases models validation propagates, e.g. ModelForms. So it's also a convenience, but this is a nice to have: models validation is essentially the last gate to the DB and should be treated properly.


Side note: ""fields"" validation is actually ""DB validation"", and ""django field validation"".


To keep with the DRY principle as much as possible:


1. Use the django approach of generic validators. So even if you validate in several places, you use the same validation. If this validation changes, it changes everywhere, see <https://docs.djangoproject.com/en/1.9/ref/validators/>
2. When possible, all client facing objects should pull the model validations, and use them, or map them to the correct client side validation, similarly to ModelForms.
3. When possible, use the model validations error messages to create a meaningful feedback to client.


But validations may repeat steps after all, mainly because validation has (at least) two roles: clean the data, and provide feedback to user. The best example is client side validation. It's useful to provide immediate feedback,w/o roundtrip to the server, but it repeats the server side validation. The model validation just makes sure that as you add more forms, API, options, if you forget to implement the correct validation - the model will prevent it.


"
"How can I get a reference to the key and value immediately after inserting into a `HashMap`?

```
use std::collections::HashMap;
use std::collections::hash_map::Entry::*;

fn hook(k: &str, v: &str) {}

fn tt(k: String, v: String) -> Option<String> {
    let mut a: HashMap<String, String> = HashMap::new();
    match a.entry(k) {
        Occupied(mut occupied) => {
            let old = occupied.insert(v);
            //hook(&k, &old);
            Some(old)
        }
        Vacant(vacant) => {
            let v = vacant.insert(v);
            let k = vacant.key(); // Why doesn't it work?
            //hook(&k, v);
            None
        }
    }
}

```

I would like to call `hook` immediately after a key is inserted into the `HashMap`. It seems I have to use `Entry`. However, I am unable to call `vacant.key()` right after `vacant.insert`.


","TL;DR: You cannot (right now?)




---


The compiler *tells you why it doesn't work*. Don't hesitate to read Rust compiler messages; they aren't scary and a lot of effort has gone into them!



```
error[E0382]: use of moved value: `vacant`
  --> src/main.rs:16:21
   |
15 |             let v = vacant.insert(v);
   |                     ------ value moved here
16 |             let k = vacant.key();
   |                     ^^^^^^ value used here after move
   |
   = note: move occurs because `vacant` has type `std::collections::hash_map::VacantEntry<'_, std::string::String, std::string::String>`, which does not implement the `Copy` trait

```

This is just normal Rust code. [`VacantEntry::insert`](https://doc.rust-lang.org/std/collections/hash_map/struct.VacantEntry.html#method.insert) consumes `self`by value, returning a reference to the inserted value: 



```
fn insert(self, value: V) -> &'a mut V

```

There are [over 100 other question/answer pairs](https://stackoverflow.com/search?q=%5Brust%5D+use+of+moved+value+is%3Aq) that ask about this error message, so I'll assume you've read enough of them to understand the problem; that's why we answer questions on Stack Overflow after all!




---


So *why* is it this way? That's tougher to answer. When you insert into a `HashMap`, it takes ownership of the key and the value. When you use the `entry` API, there's an intermediate step - you've given ownership of the key to the entry. The entry *also* has a mutable reference to the `HashMap`; a ""bookmark"" of where the value is.


When the value is missing and then you insert a value, the key is transferred out of the entry and into the `HashMap`. This means that the reference to the key inside the entry would be invalidated. That's why you can't reorder the two lines.


However, thinking about it a bit deeper, the *value* returned from `insert` refers to the underlying `HashMap`, after the value has been inserted. I can't see any reason preventing a function from being added that returns a reference to the key and the value. However, such a function doesn't exist now.


See also [How can I keep a reference to a key after it has been inserted into a HashMap?](https://stackoverflow.com/q/32401857/155423)




---


I'm pretty sure in *this* case you don't need the functionality though.


If you want to call it for the *new* value, just do all of this before the insertion:



```
hook(&k, &v);
a.insert(k, v);

```

If you want to do it only for the old value, doing nothing when there wasn't previously a value, you can:



```
Occupied(mut occupied) => {
    let old = occupied.insert(v);
    hook(occupied.key(), &old);
    Some(old)
}

```

If you want to call the hook with the **old** value if there was one and the **new** value if inserting (which seems inconsistent), you can call it *before* adding, the `hook` function will be none the wiser as the arguments are just references:



```
Vacant(vacant) => {
    hook(vacant.key(), &v);
    let v = vacant.insert(v);
    None
}

```

"
"What is string lexicographically? Java
The compareTo() method in Java compares two strings ""lexicographically"". Can someone please simply explain how the lexicographic comparison works in java?


I found **[this post](https://stackoverflow.com/questions/4064633/string-comparison-in-java)** that explains the three cases of <0 , ==0, and >0 ; However, I am still confused...


Does this mean that the int returned is the number of places away the strings are form one another if they were to be sorted alphabetically like a dictionary? 


Also, how does the method deal with case sensitivity? Are lower case letters first in line before uppercase? Is there a chart for this?


For example, the below code produces an output of -31. Does this mean that the string Dog is -31 places away from the string cat?



```
public static void main(String[] args) {
     Scanner keyboard = new Scanner(System.in);   

     String str1 = ""Dog"";

     String str2 = ""cat"";

     int result = str1.compareTo(str2);
     System.out.println(result);

```

","The value returned does not really matter as the `compareTo` contract is to return negative, positive or 0 (as you already know).


However, if really you want to understand why `-31` is returned when comparing `Dog` with `cat` (or any other string) then you could simply look at the method directly in `String` class :



```
public int compareTo(String anotherString) {
    int len1 = value.length;
    int len2 = anotherString.value.length;
    int lim = Math.min(len1, len2);
    char v1[] = value;
    char v2[] = anotherString.value;

    int k = 0;
    while (k < lim) {
        char c1 = v1[k];
        char c2 = v2[k];
        if (c1 != c2) {
            return c1 - c2;
        }
        k++;
    }
    return len1 - len2;
}

```

Keep in mind that `value` is the `char` array backing the string.



```
private final char value[];

```



---


So how does this method proceed ?


- You retrieve the minimum of both string length in a variable `lim`.
- You create a copy of both string char array.
- You loop over each characters (verifying if they are equals) until reaching the lowest limit.
- If two characters at same index are not equals, you return the result of substracting the second one to the first. The `char` can be represented as `int` value (which take their ascii value) and are already ordered. Thus when substracting a negative number will be returned if the second char is ""higher"" then the first one. A positive will be returned if the second char is ""lower"" then the first one. 0 will be returned if both are equals.
- If all characters were equals while looping for the lowest string length, you return a substraction of both length.


In your example, first letter of both words are not equals so you get to compare `D` with `c` which are respectively represented as `68` and `99`. Substract 99 to 68 and you get -31. 


So to answer this question :



> 
> Does this mean that the int returned is the number of places away the
>  strings are form one another if they were to be sorted alphabetically
>  like a dictionary?
> 
> 
> 


No, it is actually either the difference between two non-matching char's ascii value or the difference of both length.



> 
> Also, how does the method deal with case sensitivity? Are lower case
>  letters first in line before uppercase? Is there a chart for this?
> 
> 
> 


If you want to ignore the case when comparing, you can use [`String#compareToIgnoreCase`](http://docs.oracle.com/javase/8/docs/api/java/lang/String.html#compareToIgnoreCase-java.lang.String-).


Also you can check [this chart](http://www.asciitable.com/) for ascii values (upper and lower case).


"
"SELinux: Letting Apache talk to MySQL on CentOS
So I have chrooted MySQL into /opt/chroot/mysql, and everything runs fine. Apache is configured normally, and also runs fine.


I created a small PHP script which connects to the MySQL Daemon, and configured PHP so that the default socket used is the chrooted one. When SELinux is disabled, the script is able to connect fine. However, when SELinux is enabled, the script fails to connect (with error number 13), and the audit log tells me that SELinux denied the request.


Every tutorial I've found that deals with this problem tells me to disable SELinux. This is not what I want to do, so please don't suggest it! I want a solution that works whilst SELinux is enabled...I assume one must exist...


I'd guess that I probably have to change the SELinux contexts for the mysql files, but I'm not sure what to change them into to make it work.


","**UPDATE 2**



> 
> type=AVC msg=audit(1318863312.959:435): avc: denied { connectto } for
>  pid=12472 comm=""httpd""
>  path=""/opt/chroot/mysql/var/lib/mysql/mysql.sock""
>  scontext=unconfined\_u:system\_r:httpd\_t:s0
>  tcontext=unconfined\_u:unconfined\_r:unconfined\_t:s0-s0:c0.c1023
>  tclass=unix\_stream\_socket
> 
> 
> 


You can build the custom SELinux policy module by following steps:



```
# grep httpd_t audit.log | audit2allow -m httpd > httpd.te
# checkmodule -M -m -o httpd.mod httpd.te
# semodule_package -m httpd.mod -o httpd.pp 
# semodule -i httpd.pp

```

Refer to [this](https://serverfault.com/questions/321301/how-do-i-view-the-contents-of-a-selinux-policy-package) topic for more details.




---


**UPDATE**


1. Run `semanage` command to add a context mapping for `/opt/chroot/mysql/var/lib/mysql/`:



```
# semanage fcontext -a -t mysqld_db_t ""/opt/chroot/mysql/var/lib/mysql(/.*)?""

```
2. And use `restorecon` command to apply this context mapping:



```
# restorecon -Rv /opt/chroot/mysql/var/lib/mysql

```




---


If you are connecting via TCP/IP, try this:



```
# setsebool -P httpd_can_network_connect 1

```

"
"How could i create a TreeView like Microsoft apps
Hi im training to develop an app but i want to add a tree with a specific appearance i succeed in doing that but the appearance of what i have is not what my clients want the app is developed using MFC and C++.


**My app looks like this:**


[![enter image description here](https://i.stack.imgur.com/RdpKh.png)](https://i.stack.imgur.com/RdpKh.png)


**I want it to look like this:**


[![enter image description here](https://i.stack.imgur.com/KakRD.png)](https://i.stack.imgur.com/KakRD.png)


[![enter image description here](https://i.stack.imgur.com/qnweT.png)](https://i.stack.imgur.com/qnweT.png)


[![enter image description here](https://i.stack.imgur.com/ogyjW.png)](https://i.stack.imgur.com/ogyjW.png)


**The code I'm using:**



```
CRect rectDummy(0,0,200,600);
//rectDummy.SetRectEmpty();

CTreeCtrl m_wndClassView;

const DWORD dwViewStyle = WS_CHILD | WS_VISIBLE | TVS_HASLINES | TVS_LINESATROOT | TVS_HASBUTTONS | WS_CLIPSIBLINGS | WS_CLIPCHILDREN;

if (!m_wndClassView.Create(dwViewStyle, rectDummy, &m_wndView, 2))
{
    TRACE0(""Failed to create Class View\n"");
    return -1;      // fail to create
}


HTREEITEM hRoot = m_wndClassView.InsertItem(_T(""System Summary""), 0, 0);
m_wndClassView.SetItemState(hRoot, TVIS_BOLD, TVIS_BOLD);

HTREEITEM hClass = m_wndClassView.InsertItem(_T(""Hardwre Resources""), 1, 1, hRoot);
hClass = m_wndClassView.InsertItem(_T(""Components""), 1, 1, hRoot);
hClass = m_wndClassView.InsertItem(_T(""Software Enviroment""), 1, 1, hRoot);

```


> 
> I prefer the solution to be in MFC/C++ if possible and not using a third party libraries.
> 
> 
> 


","Add dependency to the `uxtheme.lib` library, include the `<Uxtheme.h>` header, call the [SetWindowTheme()](https://msdn.microsoft.com/en-us/library/windows/desktop/bb759827(v=vs.85).aspx) function after you've created the control:



```
SetWindowTheme(hYourTreeviewHandle, L""Explorer"", NULL);

```

which will give your treeview control a Windows Explorer look and feel. Things are much more involved as you need to supply your own image list and also utilize the [TreeView\_SetImageList()](https://msdn.microsoft.com/en-us/library/windows/desktop/bb760056(v=vs.85).aspx) macro to attach the image lists to the treeview:



```
TreeView_SetImageList(hYourTreeviewHandle, hYourImagesHandle, TVSIL_NORMAL);

```

"
"How to automate ""run asp.net website"" from command line?
Right now, I do the following manual steps to run an ASP.NET website on my PC:


1. Open Visual Studio and the project inside it
2. Press Ctrl+F5 which:
	1. Builds the solution
	2. Runs IIS express
	3. Opens a browser


How to write a batch file that does the same thing? The last step (opening a browser) is optional but at least I need to build the project and start it on IIS express (or whatever is configured in the project file).


","From the visual studio command line you could do the following:



```
devenv ""C:\path\FooSolution.sln"" /run

```

[**MSDN Devenv Command Line Switches Reference**](https://msdn.microsoft.com/en-us/library/xee0c8y7.aspx)


Wrapping this all up into a batch file (assuming VS 2012) it would become:



```
call ""C:\Program Files\Microsoft Visual Studio 11.0\Common7\Tools\VsDevCmd.bat"" 
devenv ""C:\path\FooSolution.sln"" /Run

```

**Update**


To run this outside of Visual Studio with IIS Express you would use the following commands:



```
msbuild.exe ""C:\path\FooSolution.sln"" 
iisexpress /path:c:\path\fooapp\ /port:666
start """" http://localhost:666

```

Please note there are many configuration options for both msbuild and iisexpress command. You will need to tailor them to suite your needs.


**[Running IIS Express from the Command Line](http://www.iis.net/learn/extensions/using-iis-express/running-iis-express-from-the-command-line)**


"
"Where does hugging face's transformers save models?
Running the below code downloads a model - does anyone know what folder it downloads it to?



```
!pip install -q transformers
from transformers import pipeline
model = pipeline('fill-mask')

```

","**Update 2023-05-02:** The cache location has changed again, and is now **`~/.cache/huggingface/hub/`**, as reported by @Victor Yan. Notably, the sub folders in the `hub/` directory are also named similar to the cloned model path, instead of having a SHA hash, as in previous versions.




---


**Update 2021-03-11:** The cache location has now changed, and is located in **`~/.cache/huggingface/transformers`**, as it is also detailed in the answer by @victorx.




---


[This post](https://github.com/huggingface/transformers/issues/2157) should shed some light on it (plus some investigation of my own, since it is already a bit older).


As mentioned, the default location in a Linux system is **`~/.cache/torch/transformers/`** (I'm using transformers v 2.7, currently, but it is unlikely to change anytime soon.). The cryptic folder names in this directory seemingly correspond to the Amazon S3 hashes.


Also note that the pipeline tasks are just a ""rerouting"" to other models. To know which one you are currently loading, see [here](https://github.com/huggingface/transformers/blob/master/src/transformers/pipelines.py#L1513). For your specific model, `pipeline(fill-mask)` actually utilizes a `distillroberta-base` model.


"
"How to detect installed Chrome version from my Android App?
I am working on a feature where I need to transition user from Android Native APP to Chrome custom tab only if chrome version installed on the device is greater than version 65. So my question is, is there a way we can detect chrome version from Android Native App ?


","
```
private boolean isChromeInstalledAndVersionGreaterThan65() {
    PackageInfo pInfo;
    try {
        pInfo = getPackageManager().getPackageInfo(""com.android.chrome"", 0);
    } catch (PackageManager.NameNotFoundException e) {
        //chrome is not installed on the device
        return false;
    }
    if (pInfo != null) {
        //Chrome has versions like 68.0.3440.91, we need to find the major version
        //using the first dot we find in the string
        int firstDotIndex = pInfo.versionName.indexOf(""."");
        //take only the number before the first dot excluding the dot itself
        String majorVersion = pInfo.versionName.substring(0, firstDotIndex);
        return Integer.parseInt(majorVersion) > 65;
    }
    return false;
}

```


> 
> Obviously this will work until Chrome will be versioned how they did until now, if they will decide to change the versioning the logic should be updated as well. (I don't think this will happen but for max safety put everything in a try-catch)
> 
> 
> 


"
"Can someone explain how to use FastTags
There are two ways to create customs tags with the play framework.


1. By defining a groovy template in app/view/tags
2. Directly in pure java by having a class extend FastTags


The latest is NOT documented. 


","So, similar to how JavaExtensions work by extending the JavaExtensions class, to create a FastTag you need to create a class that extends FastTags. Each method that you want to execute as a tag needs to conform to the following method structure.



```
public static void _tagName(Map<?, ?> args, Closure body, PrintWriter out, ExecutableTemplate template, int fromLine)

```

**Note the underscore before the name of the tag.**


To understand how to build an actual tag, the easiest way is to look at the source code for a FastTag, and see one in action.


Here is the source straight from git hub.
<https://github.com/playframework/play/blob/master/framework/src/play/templates/FastTags.java>


Below are a few I have copied, so that I can explain how this works.



```
public static void _verbatim(Map<?, ?> args, Closure body, PrintWriter out, ExecutableTemplate template, int fromLine) {
    out.println(JavaExtensions.toString(body));
}

```

So, this first method is the **verbatim** tag, and simply calls the toString method on the JavaExtensions, and passes in the body of the tag. The body of the tag would be anything between the open and close tag. So



```
<verbatim>My verbatim</verbatim>

```

The body value would be 



```
My verbatim

```

The second example, is slightly more complex. It is a tag that relies on a parent tag to function.



```
public static void _option(Map<?, ?> args, Closure body, PrintWriter out, ExecutableTemplate template, int fromLine) {
    Object value = args.get(""arg"");
    Object selectedValue = TagContext.parent(""select"").data.get(""selected"");
    boolean selected = selectedValue != null && value != null && selectedValue.equals(value);
    out.print(""<option value=\"""" + (value == null ? """" : value) + ""\"" "" + (selected ? ""selected=\""selected\"""" : """") + """" + serialize(args, ""selected"", ""value"") + "">"");
    out.println(JavaExtensions.toString(body));
    out.print(""</option>"");
}

```

This code works by outputting an HTML option tag, and sets the selected value, by checking which value is selected from the parent tag. The first 3 lines just get data, and set up the data ready to output. Then, the final 3 lines outputs the result of the tag.


There are many more examples in the source code I have linked to, with varying degrees of complexity, but hopefully this will be a good starting point for you.


To ensure that your tags do not conflict between projects, or with the core Play tags, you can set up namespaces, using the class level annotation @FastTags.Namespace.


So, for a hello tag, in a namespace of my.tags, you would do the following



```
@FastTags.Namespace(""my.tags"") 
public class MyFastTag extends FastTags {
    public static void _hello (Map<?, ?> args, Closure body, PrintWriter out, ExecutableTemplate template, int fromLine) {
        ...
    }
}

```

and then in your templates, you would reference the hello tag as



```
#{my.tags.hello/}

```

"
"How to have the same background image on each each side?
I when looking at the options for background images I cant find an option of the 
`CSS` `background-repeat` property which fits my desired design. My desired design is like this:


[![enter image description here](https://i.stack.imgur.com/KPLNJ.png)](https://i.stack.imgur.com/KPLNJ.png)


## Question:


My desired design is exactly 2 images like this which are a similar distance the window.
Is there any way to achieve this in CSS and still keeping the images as a background and not putting them in `<img>` tags?


Here is a code snippet you can use to apply the `CSS` to:





```
.content{
height: 300px;
width: 1000px;
background-color: grey;
background-image: url(""https://lh4.ggpht.com/mJDgTDUOtIyHcrb69WM0cpaxFwCNW6f0VQ2ExA7dMKpMDrZ0A6ta64OCX3H-NMdRd20=w300-rw"")
}
```


```
<div class=""content"">
 </div>
```






","You can simply use the image twice in background and adjust position/size as you need (you will have **multiple** backgrounds). You may also put the background-color on the same property.


Pay attention to order. The **first** background will be on **the top**, so the background color should be the last one so it doesn't cover the images.





```
.content {
  height: 300px;
  width: 1000px;
  background:  
  url(""https://lh4.ggpht.com/mJDgTDUOtIyHcrb69WM0cpaxFwCNW6f0VQ2ExA7dMKpMDrZ0A6ta64OCX3H-NMdRd20=w300-rw"") left/auto 100% no-repeat, 
  url(""https://lh4.ggpht.com/mJDgTDUOtIyHcrb69WM0cpaxFwCNW6f0VQ2ExA7dMKpMDrZ0A6ta64OCX3H-NMdRd20=w300-rw"") right/auto 100% no-repeat,
  grey;
}
```


```
<div class=""content"">
</div>
```






Or you can keep the `background-color` alone but you have to define it after the multiple backgrounds or it will get overridden by the background property:





```
/* This is correct */
.content {
  height: 300px;
  width: 1000px;
  background:  
  url(""https://lh4.ggpht.com/mJDgTDUOtIyHcrb69WM0cpaxFwCNW6f0VQ2ExA7dMKpMDrZ0A6ta64OCX3H-NMdRd20=w300-rw"") left/auto 100% no-repeat, 
  url(""https://lh4.ggpht.com/mJDgTDUOtIyHcrb69WM0cpaxFwCNW6f0VQ2ExA7dMKpMDrZ0A6ta64OCX3H-NMdRd20=w300-rw"") right/auto 100% no-repeat;
  background-color:grey;
}
/* This is not correct as background-color will have no effect*/
.content-alt {
  height: 300px;
  width: 1000px;
  background-color:grey;
  background:  
  url(""https://lh4.ggpht.com/mJDgTDUOtIyHcrb69WM0cpaxFwCNW6f0VQ2ExA7dMKpMDrZ0A6ta64OCX3H-NMdRd20=w300-rw"") left/auto 100% no-repeat, 
  url(""https://lh4.ggpht.com/mJDgTDUOtIyHcrb69WM0cpaxFwCNW6f0VQ2ExA7dMKpMDrZ0A6ta64OCX3H-NMdRd20=w300-rw"") right/auto 100% no-repeat;

}
```


```
<div class=""content""></div>
<div class=""content-alt""></div>
```






## For more details:


You can also consider another syntax for multiple backgrounds. Above I used the shorthand one, but you can separate properties. This can be useful if you will have many images that will share the same values (like the size and the repeat in your case). This will also remove the `background-color` restriction and you can for example easily apply a background-color with another class:





```
.content {
  height: 300px;
  width: 1000px;
  /* you can put back this on the top as there is no more background property used */
  background-color:grey;
  background-image:  
  url(""https://lh4.ggpht.com/mJDgTDUOtIyHcrb69WM0cpaxFwCNW6f0VQ2ExA7dMKpMDrZ0A6ta64OCX3H-NMdRd20=w300-rw""), 
  url(""https://lh4.ggpht.com/mJDgTDUOtIyHcrb69WM0cpaxFwCNW6f0VQ2ExA7dMKpMDrZ0A6ta64OCX3H-NMdRd20=w300-rw"");
  background-size:auto 100%; /* Same for both no need to define twice */
  background-position:left,right; /* we respect the same order as background-image*/
  background-repeat:no-repeat; /* Same for both no need to define twice */

}
```


```
<div class=""content"">
</div>
```






"
"Development of the Coq library. (Add LoadPath solution is not good enough.)
I am adding some theorems to the library
<https://github.com/coq-contribs/zfc>


But there is a not very good thing.
While I developing the code in the CoqIDE I have to add



```
Add LoadPath ""/home/user/0my/GITHUB/"".

```

and to rename all 



```
Require Import Axioms.

```

to 



```
Require Import ZFC.Axioms.

```

. All files of the library are in



```
/home/user/0my/GITHUB/ZFC

```

and the name of the last folder matters.


But when I want to run the ""make"" command I have to rename everything back.


The file ""Make"" contains the names of files and the prefix. Deleting the first line didn't solve the problem.


I don't think that it is the best practice of the developing with the CoqIDE, so what shall I do instead?


**Edited1:**


I have ""run\_coqide.sh"" which consist of



```
#!/usr/bin/env bash
COQPATH=/home/user/0my/GITHUB/ 
/home/user/opam-coq.8.8.1/4.02.3/bin/coqide

```

""From ZFC Require Import Sets. "" raises error ""Cannot find a physical path"".


**Edited2:**


I have find out that this is a working script:



```
#!/usr/bin/env bash
export COQPATH=/home/user/0my/GITHUB/ 
/home/user/opam-coq.8.8.1/4.02.3/bin/coqide

```

Is it normal run or a hack?


","Rename `Make` to `_CoqProject`, this is the name currently recognized by CoqIDE where it will look for project configuration (in particular the `-R . ZFC` option to make the Coq files in the directory visible to CoqIDE). 
It is also possible to change the name CoqIDE is looking for to `Make`, but `_CoqProject` seems to really be the new standard.


Note that the `-R . ZFC` option, that allows you to import libraries unqualified, corresponds to the command [`Add Rec LoadPath ""/.../ZFC"" as ZFC`](https://coq.inria.fr/distrib/current/refman/proof-engine/vernacular-commands.html#coq:cmd.add-rec-loadpath).


I would also suggest to actually switch the whole codebase to explicit qualification `ZFC.Axiom`, that you've been doing locally by hand, making it less error prone to work with different projects at the same time. I'm not sure why it was necessary to rename things back to run `make`, my understanding is that it shouldn't be necessary.


See also the Coq reference manual, [about the `coq_makefile` utility](https://coq.inria.fr/distrib/current/refman/practical-tools/utilities.html#building-a-coq-project-with-coq-makefile).


"
"Download snap ""lxd"" (10756) from channel ""stable"" (invalid credentials) when trying to install
Just upgraded from bionic to cosmic. Part of that process migrates from deb lxd to snap lxd. The upgraded succeeded except for the lxd package. I get this error when trying to complete the upgrade or when installing lxd snap using snap command.



```
snap install lxd
error: cannot perform the following tasks:
- Download snap ""lxd"" (10756) from channel ""stable"" (invalid credentials)

```

Why do I get this? What credentials? Where is the snapd error log?


Update as requested in attempted answers:


`grep snap /var/log/syslog` :



```
May 30 00:21:14 delays snapd[30230]: storehelpers.go:441: cannot refresh snap ""core"": snap has no updates available
May 30 00:21:14 delays snapd[30230]: autorefresh.go:389: auto-refresh: all snaps are up-to-date
May 30 10:21:14 delays snapd[30230]: storehelpers.go:441: cannot refresh snap ""core"": snap has no updates available
May 30 10:21:14 delays snapd[30230]: autorefresh.go:389: auto-refresh: all snaps are up-to-date
May 30 15:11:14 delays snapd[30230]: storehelpers.go:441: cannot refresh snap ""core"": snap has no updates available
May 30 15:11:14 delays snapd[30230]: autorefresh.go:389: auto-refresh: all snaps are up-to-date
May 30 15:37:32 delays snapd[30230]: api.go:1079: Installing snap ""lxd"" revision unset

```

","Could you submit the content of `/var/log/syslog` once you run `sudo snap install lxd`?


To do so, open a terminal and type:



```
$ grep snap /var/log/syslog

```

## Solution #1


Usually, when this error occurs, you can resolve it by removing your credentials from `~/.snap/auth.json` and by trying to re-authenticate.


**Edit:**


## Solution #2


First, try to stop `snapd`



```
$ sudo systemctl stop snapd.service

```

Take a back-up and remove the `/var/lib/snapd/state.json`



```
$ sudo mv /var/lib/snapd/state.json /var/lib/snapd/state.json.bck

```

Start again `snapd`



```
$ sudo systemctl start snapd.service

```

Login to Snap



```
$ snap login

```

Update all Snap packages



```
$ snap refresh 

```

Remove broken Snap



```
$ snap remove lxd

```

Try to install lxd again



```
$ snap install lxd

```

"
"Doctrine insert many data
I'm working on an import of 300000 row inside many csv.


First I take csv and import every row into a table inside my database.


After I want to parse all row and insert into the right table with some relation that data.


So I have tried this:



```
    $qb = $this->entityManager->createQueryBuilder();
    $flows = $qb->select('flow')
        ->from('AppBundle:FlowAndata', 'flow')
        ->getQuery()
        ->getResult();

    $countRows = 0;
    foreach ($flows as $row) {
         //some check 
         $entity = new TestTable();
         $entity->setCode($row->getCode());
         //many other fields
         $this->entityManager->persist($entity);
         $this->entityManager->flush();
    }

```

In this case all procedure took about 5 seconds for every row!


Now If I add setMaxResults like this:



```
    $qb = $this->entityManager->createQueryBuilder();
    $flows = $qb->select('flow')
        ->from('AppBundle:FlowAndata', 'flow')
        ->setMaxResults(100)
        ->getQuery()
        ->getResult();

```

It took less than 1 second!


So I have thought to get all rows and split it into a recursive function with setMaxResult like this:



```
    $qb = $this->entityManager->createQueryBuilder();
    $flows = $qb->select('flow')
        ->from('AppBundle:FlowAndata', 'flow')
        ->getQuery()
        ->getResult();

    $countFlows = count($flows);
    $numberOfQuery = $countFlows / 100;

    for ($i = 0; $i <= $numberOfQuery; $i++) {
         $this->entityManager->clear();
         $qb = $this->entityManager->createQueryBuilder();
         $flows = $qb->select('flow')
            ->from('AppBundle:FlowAndata', 'flow')
            ->setFirstResult($i * 100)
            ->setMaxResults(100)
            ->getQuery()
            ->getResult();

    }

```

In this way I create many query splitted into 100 rows.
Is a good practice or there is a better way to parse many rows and make an insert of It?


","The efficient way recommended by the [official documentation of Doctrine](https://www.doctrine-project.org/projects/doctrine-orm/en/latest/reference/batch-processing.html) is taking advantage of the transactional write-behind behavior of an `EntityManager`.


**Iterating Large Results for Data-Processing**



> 
> You can use the `iterate()` method just to iterate over a large result and no UPDATE or DELETE intention. The `IterableResult` instance returned from `$query->iterate()` implements the Iterator interface so you can process a large result without memory problems using the following approach. ([See example](https://www.doctrine-project.org/projects/doctrine-orm/en/latest/reference/batch-processing.html#iterating-large-results-for-data-processing))
> 
> 
> 


**Bulk Inserts**



> 
> Bulk inserts in Doctrine are best performed in batches, taking advantage of the transactional write-behind behavior of an `EntityManager`. [...] You may need to experiment with the batch size to find the size that works best for you. Larger batch sizes mean more prepared statement reuse internally but also mean more work during `flush`. ([See example](https://www.doctrine-project.org/projects/doctrine-orm/en/latest/reference/batch-processing.html#bulk-inserts))
> 
> 
> 


**Version mixing both techniques** (Inside Entity Repository):



```
$q = $this->_em->createQuery('SELECT f FROM AppBundle:FlowAndata f');
$iterableResult = $q->iterate();

$i = 0;
$batchSize = 100;

foreach ($iterableResult as $row) {
    // do stuff with the data in the row, $row[0] is always the object 
    /** @var AppBundle\Entity\FlowAndata $flow */
    $flow = $row[0];

    //some check 
    $entity = new TestTable();
    $entity->setCode($row->getCode());
    //many other fields

    $this->_em->persist($entity);

    $i++;
    if (($i % $batchSize) === 0) {
        $this->_em->flush();
        // Detaches all objects from Doctrine!
        $this->_em->clear(); 
    } else {
        // detach from Doctrine, so that it can be Garbage-Collected immediately
        $this->_em->detach($flow);
    }
}

$this->_em->flush(); //Persist objects that did not make up an entire batch
$this->_em->clear();

```

"
"What goes wrong when passing a std::sub\_match as argument to a std::thread?
I'm passing a `std::sub_match` as an argument to a `std::thread` (see my example code below). The thread function expects a const string reference. A sub\_match can be converted to a string. So everything compiles fine.


But sometimes the function receives the wrong string.
When I convert the sub\_match to a string before passing it to the thread it works as expected. What is the difference?


I think this is a race condition as the original sub\_match may not exist anymore when the thread executes. But I thought arguments to threads would be copied anyway.
How can I find out which arguments are safe to pass to a thread and which not?



```
#include <iostream>
#include <string>
#include <vector>
#include <thread>
#include <regex>
#include <unistd.h>

class test_t {
  public:
    test_t(void) {}
    ~test_t(void) {}

    void start(void){
     //-------------------------------------------------
     // Do some memory allocation.
     // The error seems to appear faster with that.
     std::vector<std::string> vec;
     for(unsigned int i = 0; i < 1000; ++i) {
        vec.push_back(""test_test_test"");
     }
     //-------------------------------------------------

     std::string event = ""operating"";
     std::smatch match;
     std::regex expr(""\\(operating\\)"",
         std::regex_constants::icase | 
         std::regex_constants::basic);

     if(std::regex_match(event, match, expr)) {
        std::cout << ""start thread"" << std::endl;
        m_thread = std::thread(&test_t::thread_func, this, match[1]);              //NOK
//        m_thread = std::thread(&test_t::thread_func, this, match[1].str());        // OK
//        m_thread = std::thread(&test_t::thread_func, this, (std::string)match[1]); // OK
        m_thread.detach();
        std::cout << ""thread started"" << std::endl;
     }
    }

  private:
    std::thread m_thread;

    void thread_func(const std::string& string) {
     if(string != ""operating"") {
        std::cout << ""ERROR: string: \"""" << string << ""\"""" << std::endl;
        exit(EXIT_FAILURE);
     } else {
        std::cout << ""string: \"""" << string << ""\"""" << std::endl;
     }
    }
};

int main(int argc, char** argv) {
  test_t test;
  while(1) {
    test.start();
    usleep(100);
  }
  return 0;
}

```

Compilation Message:



```
Compiled with: g++ --std=c++11 -pthread -o test main.cpp
g++ --version: g++ (SUSE Linux) 4.8.5

```

Expected output:



```
start thread
thread started
string: ""operating""
(repeat)

```

Actual output:



```
start thread
thread started
string: ""operating""
ERROR: string: ""test_test""

```

","`operator[]` for `std::smatch` returns `sub_match` which can be treated as pair of iterators to matched characters. 


After `regex_match` was called, you can use `operator[]` to access sub-matches as long as `event` exists. When `event` is deleted (you are not joining your thread, so `start` returns immediately and `event` is destroyed), sub-matches have dangling pointers and should not be accessed.




---



```
m_thread = std::thread(&test_t::thread_func, this, match[1]);

```

this doesn't work because when function goes out of scope, event is deleted and sub-match has dangling pointers.




---



```
m_thread = std::thread(&test_t::thread_func, this, match[1].str());

```

this works because `str()` returns copy of matched string.




---



```
m_thread = std::thread(&test_t::thread_func, this, (std::string)match[1]);

```

this also works because temporary string is created based on sub-match `match[1]`, and temp is passed into thread.


"
"get global ThreadId in 2 dimensions in OpenCl
How can i get the global threadId in 2 dimensions in OpenCL? 
I know that for 1 dimension, the formula is:



```
 int global_id = get_global_id(1) * get_global_size(0) + get_global_id(0);

```

But if i allocate like this:



```
size_t block_size[] = {2,2}
size_t grid_size[] = {35,20}

```

The above formula fails, giving indexes only from 0 to 35\*20.
The indexes should go from 0 to 35\*40\*2\*2.


Can you recommend any good documentation or writings that could give me the intuition to understand how all of this works? 
Thanks!


","If you're launching a 2D NDRange, then get\_global\_id(0) and get\_global\_id(1) will give you the Gx and Gy indices. You can also independently fetch the local ids using get\_local\_id(0/1).


There's no need to calculate it yourself.


Did you mean that you're launching a 2D thread block but want to map that thread to a position in a 1 dimensional buffer?


EDIT: After reading your comment, I thought an explanation is in order.


OpenCL launches as many kernels as get\_global\_size(0) \* get\_global\_size(1) (which is 35 \* 20), so you will have threads



```
(0 ,0) (0 ,1) ... (0,34)
(1 ,0) (1 ,1) ... (1,34)
.
.
.
(19,0) (19,1) ... (19,34)

```

Local worksize is simply a way of splitting up the total number of threads and distributing them across the compute units available. It is quite possible that only 2 \* 2 = 4 threads are running at any point of time.


The [clEnqueueNDRangeKernel](http://www.khronos.org/registry/cl/sdk/1.0/docs/man/xhtml/clEnqueueNDRangeKernel.html) documentation tells us that local\_work\_size can be null, in which case the implementation will determine the size to break up the total amount of work.


In no way does local work size increase the number of threads.


Perhaps this image explains it better than I can.


![OpenCL 2D NDRange](https://i.stack.imgur.com/5gZZr.png)


Note that the total number of kernel launches are still get\_global\_size(0) \* get\_global\_size(1).


If you want your 1D indices to go from 0..(35\*40\*2\*2 - 1) then launch the kernel so that get\_global\_size(0) \* get\_global\_size(1) is 35\*40\*2\*2 (perhaps 70 x 80 ?)


Hope this helps.


"
"How can I get rid of the ""Do you trust this printer"" message box and add my printer via GPO?
- Workstation: Windows 7 (x64) [Install target for printer]
- Server: Windows Server 2012 R2 (x64) [Active Directory, Print Server]


I have been bashing my head on the desk trying to get this printer to be installed via group policy! For some reason, I simply can NOT get this printer deployed with GPO. I have tried setting it up to deploy via `Computer Configuration->Policies->Windows Settings->Deployed Printers`, as well as `Computer Configuration->Preferences->Control Panel Settings->Printers` and `User Configuration->Preferences->Control Panel Settings->Printers`. I have also tried going through my Print Server Management console to add it via user and/or computer targeting. I have tried ALL KINDS of ways and nothing is working. I followed a bunch of tutorials and watched a bunch of videos just to make sure I was not missing something but it really is a simple task (in theory)... It just will not work.


In trying to debug the issue I found that if I went to `\\myserver\` and double clicked on the printer it would try to install the printer and then prompt me to install the drivers with a UAC type prompt.
[![enter image description here](https://i.stack.imgur.com/pWsX0.png)](https://i.stack.imgur.com/pWsX0.png)


I have tried everything I can think of to get that message box to stop popping up. I dug into it and found that if I was to edit a GPO called `Point and Print Restrictions` located at `Computer Configuration->Policies->Administrative Templates->Printers` and was as at `User Configuration->Policies->Administrative Templates->Control Panel->Printers` you could try setting the policy to `Disabled` or `Enabled` and choose `Do not show warning or elevation prompt` for the two Security Prompts listed at the bottom of the policy settings.


Well that was a bust too...


I did find that if I tried to manually install the printer by going to the unc and typing in my Administrator credentials it would download the drivers from the server and install the printer (as expected). If the user tried to remove the printer and was successful somehow as soon as they would log off and back on the GPO would do what I want and add the printer back. But it required me to add it manually the first time on EVERY PC.


After testing this and then removing the printer from the GPO, then logging off and on again. I could run the command `printui /s /t2` to bring up a GUI that would allow me to remove the installed drivers with ease to put the PC back to it's original state (asking for Administrator credentials). Also something else I learned was the printers were stored in the registry located at `HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Windows NT\CurrentVersion\Print\Connections`. When I was trying to remove a printer and it told me I couldn't, I just went to that registry key and removed the GUID key of the printer I was trying to delete. Then just restarted the Print Spooler service and boom it was gone. This was no help getting me get to where I wanted but it was helpful in removing the printer during debugging the issue.


I read somewhere that maybe the cause is some type of windows security update that changed something. It was released because of some article showing how you could pwn an entire network if you were able to pwn a single printer. Something about when users connected to the printer and downloaded the drivers it would install injected software and run on the machine, etc...


My main goal is to be able to deploy this printer to a set of users in this OU with the GPO I'm using. But everything I try requires an Administrator to be logged on to do it (at least the first time). Does anyone have any idea why my printer will not automatically add itself via the GPO and also how do I get that dang ""Do you trust this printer?"" message to go away?


","The ""fix"" is to download trusted, package-aware print drivers from the printer manufacturer; however, since not all manufacturers will produce these drivers, there is a work around that I found here: [Cannon Forum - Package-Aware Print Drivers](https://community.usa.canon.com/t5/Office-Printers/Package-Aware-Print-Drivers/td-p/180787#messageview_7) (Note: This does not work for unsigned drivers, but there are plenty of tutorials out there for self-signing a print driver.)


Here are the steps to get around the issue:


1. Install the required drivers on the print server
2. Note any drivers that have ""false"" listed under the ""Packaged"" column. All of these will have to be modified to deploy via Group Policy.
3. Edit the registry on your print server and go to the following locations:
	- For 64-bit drivers: HKLM\System\CurrentControlSet\Control\Print\Enviroments\**Windowsx64**\Drivers\Version-**X**\{Driver Name}
	- For 32-bit drivers: HKLM\System\CurrentControlSet\Control\Print\Enviroments\**Windows NT x86**\Drivers\Version-**X**\{Driver Name}
	- Where ""X"" is the print driver ""Type"", usually either ""3"" or ""4""
4. Edit the key named ""PrinterDriverAttributes"" by adding 1 to whatever value is currently set. (Example: if the current value is ""6"", change it to ""7"".) This will make the print server believe that these drivers are packaged.
5. Do this for every driver that is not listed as a ""Packaged"" driver.
6. Restart the print server.
7. Everything should now deploy through Group Policy (provided you have all the usual GPO settings configured correctly).


I am in the process of deploying this fix myself; however, since it requires a restart of the print server I cannot test it until this evening since our print server also runs a few networked applications.


An alternate solution to the registry edit is to edit the printer driver INF file and add the following:


For 32-bit drivers:



```
[PrinterPackageInstallation.x86]
PackageAware=TRUE

```

For 64-bit driver



```
[PrinterPackageInstallation.amd64]
PackageAware=TRUE

```

If you decide to edit the INF file, it'll be easier to remove the driver from the print server, edit the INF file from a clean download, and then install the edited driver.


Aside from this, double check the Group Policy settings for `Point and Print Restrictions` and `Package Point and print - Approved Servers`.


# Background Info


[Microsoft Security Bulletin MS16-087](https://technet.microsoft.com/library/security/MS16-087) detailed a security issue where a rogue print server could inject malicious code through a ""man in the middle"" style attack. [Security update KB3170455](https://support.microsoft.com/en-us/kb/3170455) was issued on July 12, 2016 to fix it, which then messed with the distribution of print drivers from the print server.


"
"Webbrowser Control - displaying text using ""write""
I'm using the IWebBrowser2 interface to render a page from an HTML string created at runtime. I have written a method (let's call it DisplayHtmlString) that takes an HTML string and renders it as shown in [this example](http://msdn.microsoft.com/en-us/library/aa752639%28v=vs.85%29.aspx). The method also calls Navigate2 with ""about:blank"" first, to ensure a document is present, and it also calls close after calling write.


The first time I call DisplayHtmlString, the page is always rendered correctly, i.e. the browser displays HTML according to the string I pass. The problem is that subsequent calls sometimes do not work correctly, but render a blank page instead. What could be causing this?


I have found out that when the blank page is shown, this is the result of navigating to about:blank. This was determined by navigating to a local file instead, which is then shown (whereas the HTML string should be shown instead, due to the subsequent write/close). So the call to Navigate2 works, while the calls to write and close sometimes don't.


I considered IE-internal security checks as a possible cause (cross-domain checking?), but my gut feeling is that this isn't what's happening here.


It seems more likely to me that it is some kind of synchronization problem, along the lines of ""IE hasn't finished rendering yet before the next call to DisplayHtmlString comes along"". My code originally didn't check the browser's ready state (because the example doesn't). I added an experimental waiting loop with a call to get\_readyState and observed that the state never got beyond ""loading"" before returning from the method - probably because rendering is asynchronous(?). I also notice that when successive calls to DisplayHtmlString work correctly, the program's main message loop has run (giving Windows a chance to process messages), which is not the case in the scenario where successive calls to DisplayHtmlString fail.


So I'm pretty sure I need to provide correct synchronization here, but how? I notice there's a method named onreadystatechange, but haven't experimented with that yet, due to the multitude of other things I tried while groping in the dark. Could that be the solution, and how does one use it correctly? Or, alternatively, shall I just process the message loop inside DisplayHtmlString until the ready state changes to ""complete""?


UPDATE: Added message loop processing to DisplayHtmlString. In the first call (which works), the ready state gets to ""interactive"", but no further (which doesn't seem to be a problem). In the subsequent call (when it fails), the ready state stays at ""loading"", even though the message loop is processed.


","You should handle `readystatechange` event on the `document` object. In JavaScript, it would look like this:



```
<body>
<body>Hi, this is going to be replaced!</body>
<script>
window.onload = function()
{
    document.open(""text/html"");
    document.onreadystatechange = function() { 
        if (document.readyState == ""complete"")
            alert(""Done!""); 
    }
    document.write(""<b>Hello again!</b>"");
    document.close();   
}
</script>
</body>

```

To get it done with C++ or C#, probably the easiest way would be to provide an implementation of `IDispatch` interface to `IHTMLDocument2::put_readystatechange`. Then `IDispatch::Invoke(DISPID_VALUE)` will be called back upon `readystatechange` event.


I might be able to help with a code sample if you specify the language in use.


**[EDITED]**
You can grab the complete example (C++/ATL/VS2012) from [here](http://goo.gl/XFuhUk). The code does it asynchronously by posting a custom message to the main window. It's too long to cite here, below are the relevant parts.


IDispatch implementation, for onreadystatechange Event Sink:



```
class CEventSink: 
  public CComObjectRoot,
  public IDispatch
{
private:
  HWND m_hwnd;
  UINT m_message;

public:
  CEventSink()
  {
    m_hwnd = NULL;
    m_message = NULL;
  }

  BEGIN_COM_MAP(CEventSink)
    COM_INTERFACE_ENTRY(IDispatch)
  END_COM_MAP()

  // Init
  void Init(HWND hwnd, UINT message)
  {
    m_hwnd = hwnd;
    m_message = message;
  }

  // IDispatch
  STDMETHODIMP GetTypeInfoCount(UINT* pctinfo) { 
    return E_NOTIMPL; }

  STDMETHODIMP GetTypeInfo(UINT itinfo, LCID lcid, ITypeInfo** pptinfo) {
    return E_NOTIMPL; }

  STDMETHODIMP GetIDsOfNames(REFIID riid, LPOLESTR* rgszNames, UINT cNames, LCID lcid, DISPID* rgdispid) { 
    return E_NOTIMPL; }

  STDMETHODIMP Invoke(
    DISPID dispidMember, REFIID riid, LCID lcid, WORD wFlags,
    DISPPARAMS* pDispParams, VARIANT* pvarResult,
    EXCEPINFO*  pExcepInfo, UINT* puArgErr)
  {
    if ( dispidMember != NULL )
      return DISP_E_MEMBERNOTFOUND;

    // Just post a message to notify the main window
    ::PostMessage(m_hwnd, m_message, 0, 0);
    return S_OK;
  }

};

```

Using it:



```
CComObject<CEventSink>* p = NULL;
hr = CComObject<CEventSink>::CreateInstance(&p);
if ( FAILED(hr) ) 
  return 0;
p->Init(m_hWnd, WM_DOCUMENTREADYSTATECHANGE);
m_eventSink = p; // does AddRef

// ...

m_htmlDocument2->put_onreadystatechange(CComVariant(m_eventSink));

```

For more details, get [the sources](http://goo.gl/XFuhUk) and look at `WebOcHost.cpp`. The error checks are very basic, for brevity. 


"
"How do I mask multi-output in Tensorflow 2 LSTM training?
I am training an LSTM model in Tensorflow 2 to predict two outputs, streamflow and water temperature.


- For some of the time steps there is a streamflow label ***and*** a temperature label,
- For some there is only a streamflow label ***or*** a temperature label,
- and for some there are ***neither***.


So the loss function needs to ignore the temperature and streamflow loss when they don't have a label. I've done quite a bit of reading in the TF docs, but I'm struggling to figure out how to best do this.


So far I've tried


- specifying `sample_weight_mode='temporal'` when compiling the model and then included a `sample_weight` numpy array when calling `fit`


When I do this, I get an error asking me to pass a 2D array. But that confuses me because there are 3 dimensions: `n_samples`, `sequence_length`, and `n_outputs`. 


Here's some code of what I am basically trying to do:



```
import tensorflow as tf
import numpy as np

# set up the model
simple_lstm_model = tf.keras.models.Sequential([
    tf.keras.layers.LSTM(8, return_sequences=True),
    tf.keras.layers.Dense(2)
])

simple_lstm_model.compile(optimizer='adam', loss='mae',
                          sample_weight_mode='temporal')

n_sample = 2
seq_len = 10
n_feat = 5
n_out = 2

# random in/out
x = np.random.randn(n_sample, seq_len, n_feat)
y_true = np.random.randn(n_sample, seq_len, n_out)

# set the initial mask as all ones (everything counts equally)
mask = np.ones([n_sample, seq_len, n_out])
# set the mask so that in the 0th sample, in the 3-8th time step
# the 1th variable is not counted in the loss function
mask[0, 3:8, 1] = 0

simple_lstm_model.fit(x, y_true, sample_weight=mask)

```

The error: 



```
ValueError: Found a sample_weight array with shape (2, 10, 2). In order to use timestep-wise sample weighting, you should
pass a 2D sample_weight array.

```

Any ideas? I must not understand what `sample_weights` do because to me it only makes sense if the `sample_weight` array has the same dimensions as the output. I could write a custom loss function and handle the masking manually, but it seems like there should be a more general or built in solution.


","## 1. `sample_weights`


Yes, you understand it incorrectly. In this case you have `2` samples, `10` timesteps with `5` features each. You could pass `2D` tensor like this so each timestep for each sample will contribute differently to total loss, all features are equally weighted (as is usually the case).


**That's not what you are after at all**. You would like to mask certain loss values after their calculation so those do not contribute.


## 2. Custom loss


One possible solution is to implement your own loss function which multiplies loss tensor by mask before taking `mean` or `sum`.


Basically, you pass `mask` and `tensor` concatenated together somehow and split it within function for use. This is sufficient:



```
def my_loss_function(y_true_mask, y_pred):
    # Recover y and mask
    y_true, mask = tf.split(y_true_mask, 2)
    # You could user reduce_sum or other combinations
    return tf.math.reduce_mean(tf.math.abs(y_true - y_pred) * mask)

```

Now your code (no weighting as it's not needed):



```
simple_lstm_model = tf.keras.models.Sequential(
    [tf.keras.layers.LSTM(8, return_sequences=True), tf.keras.layers.Dense(2)]
)

simple_lstm_model.compile(optimizer=""adam"", loss=my_loss_function)

n_sample = 2
seq_len = 10
n_feat = 5
n_out = 2

x = np.random.randn(n_sample, seq_len, n_feat)
y_true = np.random.randn(n_sample, seq_len, n_out)

mask = np.ones([n_sample, seq_len, n_out])
mask[0, 3:8, 1] = 0

# Stack y and mask together
y_true_mask = np.stack([y_true, mask])

simple_lstm_model.fit(x, y_true_mask)

```

And so it works. You could also stack the values in some other way, but I hope you get the feel of how one could do it.


## 3. Masking outputs


Please notice above introduces a few problems. If you have a lot of zeroes and take `mean` you might get a really small loss value and inhibit learning. On the other hand, if you go with `sum` it might explode.


"
"Android horizontal progress bar?
I'd like to implement a progress bar like this


![enter image description here](https://i.stack.imgur.com/gaX5Q.jpg)


Please help how can I do it or if there is a library


","I saw that a lot of people are looking on this post so I thought that I should edit it and share an example:


Note (this example is not complete, it's still in progress but I guess it's a starting point)


GitHub: [Github Example](https://github.com/ManolescuSebastian/AndroidCustomHorizontalProgressBar/tree/master)


![enter image description here](https://i.stack.imgur.com/5orDg.png)


The important part in this example is below:
Create in drawable a xml file custom\_progress\_bar\_horizontal.xml and add the content below.



```
<?xml version=""1.0"" encoding=""UTF-8""?>

<layer-list xmlns:android=""http://schemas.android.com/apk/res/android"">
    <item
        android:id=""@android:id/background"">
        <shape>
            <padding
                android:bottom=""3dip""
                android:top=""3dip""
                android:left=""3dip""
                android:right=""3dip""/>
            <corners
                android:radius=""5dip"" />
            <gradient
                android:startColor=""#2a2b2f""
                android:centerColor=""#2a2b2f""
                android:centerY=""0.50""
                android:endColor=""#2a2b2f""
                android:angle=""270"" />
        </shape>
    </item>
    <item
        android:id=""@android:id/progress"">
        <clip>
            <shape>
                <corners
                    android:radius=""5dip"" />
                <gradient
                    android:startColor=""#ff0e75af""
                    android:endColor=""#ff1997e1""
                    android:angle=""90"" />
            </shape>
        </clip>
    </item>
</layer-list>

```

Add the following code in styles.xml



```
<style name=""CustomProgressBar"" parent=""android:Widget.ProgressBar.Horizontal"">
    <item name=""android:indeterminateOnly"">false</item>
    <item name=""android:progressDrawable"">@drawable/custom_progress_bar_horizontal</item>
    <item name=""android:minHeight"">10dip</item>
    <item name=""android:maxHeight"">20dip</item>
</style>

```

After you have added the style in your activity layout add:



```
    <ProgressBar
        android:id=""@+id/customProgress""
        style=""@style/CustomProgressBar""
        android:layout_width=""match_parent""/>

```

The progress dispaly below in a frame it's a little tricky because you need to extend the ProgressBar class and make the changes there.


I will leave here some examples and notify when I will add them in my github project.


Great example if you want to disaply the progress in your way:
[NumberProgressBar](https://github.com/daimajia/NumberProgressBar)


Other progress bar examples:


[Custom progress bar 1](http://custom-android-dn.blogspot.ro/2013/01/how-to-create-custom-progress-bar-and.html)


[Custom progress bar 2](http://www.tiemenschut.com/how-to-customize-android-progress-bars/)


[Custom progress bar 3](http://www.learn-android-easily.com/2013/05/custom-progress-bar-in-android.html)


**UPDATE:**


Also check [DiscreteSeekBar](https://github.com/AnderWeb/discreteSeekBar) this is a great github project.


Gradle: compile 'org.adw.library:discrete-seekbar:1.0.1'


**Update #2**


Maybe this library will be also useful
<https://github.com/ManolescuSebastian/USeekbar>


Cheers,


"
"Rails 5 API + Real time notifications
I'm creating a Rails 5 API only and I need to create a real time web notifications.


Is that possible by using ActionCable? Does anyone has an example of Action Cable or any other solution?


Thanks in advance. 


","This is a web notification channel that allows you to trigger client-side web notifications when you broadcast to the right streams:


Create the server-side web notifications channel:



```
# app/channels/web_notifications_channel.rb
class WebNotificationsChannel < ApplicationCable::Channel
  def subscribed
    stream_for current_user
  end
end

```

Create the client-side web notifications channel subscription:



```
# app/assets/javascripts/cable/subscriptions/web_notifications.coffee
# Client-side which assumes you've already requested
# the right to send web notifications.
App.cable.subscriptions.create ""WebNotificationsChannel"",
  received: (data) ->
    new Notification data[""title""], body: data[""body""]

```

Broadcast content to a web notification channel instance from elsewhere in your application:



```
# Somewhere in your app this is called, perhaps from a NewCommentJob
WebNotificationsChannel.broadcast_to(
  current_user,
  title: 'New things!',
  body: 'All the news fit to print'
)

```

The `WebNotificationsChannel.broadcast_to` call places a message in the current subscription adapter's pubsub queue under a separate broadcasting name for each user. For a user with an ID of 1, the broadcasting name would be `web_notifications:1`.


"
"How can i add numbers near each line in index 0 in richTextBox?

```
for (int i = 0; i < richTextBox2.Lines.Length; i++)
{
    richTextBox2.Lines[i] = richTextBox2.Lines[i].Insert(0, i + "" "");
}

```

Before that tried with the Insert:



```
richTextBox2.Lines[i].Insert(0, i + "" "");

```

In both cases it's not adding any numbers.


For example if the lines are:



```
Hello world
Hi
Hello

```

Then i want it to be now:



```
1 Hello world
2 Hi
3 Hello

```

But the loop does nothing it's not adding any numbers.


","You can't modifiy the single line in that way. If you [look at MSDN](https://msdn.microsoft.com/en-us/library/system.windows.forms.textboxbase.lines(v=vs.110).aspx) you will find this remark 



> 
> By default, the collection of lines is a read-only copy of the lines
>  in the TextBox. To get a writable collection of lines, use code
>  similar to the following: textBox1.Lines = new string[] { ""abcd"" };
> 
> 
> 


So the correct way to reach you goal is 



```
string[]  lines = richTextBox2.Lines;

for (int i = 0; i < lines.Length; i++)
{
    lines[i] = (i+1) + "" "" + lines[i];
}
richTextBox2.Lines = lines;

```

"
"Pandas: Conditional column creating
I'm trying to **create column C**, based on the values in columns A and B given the following conditions:



```
if A < 5000: C = A * B
else: C = A

```

The following gives a syntax error:



```
df['C'] = df.apply(lambda x (x['A'] * x['B)'] if x['A'] < 5000 else x = x['A']),axis=1)

```

How far off am I? 


","Use vectorized [`numpy.where`](https://docs.scipy.org/doc/numpy/reference/generated/numpy.where.html):



```
df['C'] = np.where(df['A'] < 5000, df['A'] * df['B'], df['A'])

```

**Performance**:



```
np.random.seed(2019)

N = 1000
data = np.asarray([np.random.rand(N).tolist(), list(range(N))]).T
df = pd.DataFrame(data, columns=['A', 'B'])

In [56]: %timeit df['C'] = np.where(df['A'] < 5000, df['A'] * df['B'], df['A'])
536 µs ± 47.3 µs per loop (mean ± std. dev. of 7 runs, 1000 loops each)

In [57]: %timeit df['C'] = df.apply(lambda x: x.A * x.B if x.A > 0.5 else x.A, 1)
30.9 ms ± 597 µs per loop (mean ± std. dev. of 7 runs, 10 loops each)

```



---



```
N = 100000
data = np.asarray([np.random.rand(N).tolist(), list(range(N))]).T
df = pd.DataFrame(data, columns=['A', 'B'])

In [59]: %timeit df['C'] = np.where(df['A'] < 5000, df['A'] * df['B'], df['A'])
1.29 ms ± 23.9 µs per loop (mean ± std. dev. of 7 runs, 1000 loops each)

In [60]: %timeit df['C'] = df.apply(lambda x: x.A * x.B if x.A > 0.5 else x.A, 1)
3.32 s ± 374 ms per loop (mean ± std. dev. of 7 runs, 1 loop each)

```

"
"How to properly configure multiple accounts in mutt
There are many articles on this topic none of which work for me. The .muttrc has the following related to multiple accounts (only IMAP is used + of course SMTP):



```
source ""~/.mutt/account1""
folder-hook $folder 'source ~/.mutt/account1'
source ""~/.mutt/account2""
folder-hook $folder 'source ~/.mutt/account2'

macro index <f2> '<sync-mailbox><enter-command>source ~/.mutt/account1<enter><change-folder>!<enter>'
macro index <f3> '<sync-mailbox><enter-command>source ~/.mutt/account2<enter><change-folder>!<enter>'

# Mailboxes
bind    index   G  imap-fetch-mail
macro   index   gi ""<change-folder>=INBOX<enter>""       ""Go to Inbox""
macro   index   gs ""<change-folder>=$my_sent<enter>""    ""Go to Sent""
macro   index   gd ""<change-folder>=$my_drafts<enter>""  ""Go to Drafts""
macro   index   gt ""<change-folder>=$my_trash<enter>""   ""Go to Trash""

```

The account files are as (e.g. for account1, the other one would be the same but the names of the folders would be in different language):



```
set my_drafts             = ""Drafts""
set my_drafts_noquote     = ""Drafts""
set my_sent               = ""Sent<quote-char><space>Items""
set my_sent_noquote       = ""Sent Items""
set my_trash              = ""Deleted<quote-char><space>Items""
set my_trash_noquote      = ""Deleted Items""
set imap_user             = ""username""
set folder                = ""imaps://outlook.office365.com:993/""
set imap_authenticators   = ""login""
set mbox                  = ""+INBOX""
set spoolfile             = ""+INBOX""
set record                = +$my_sent_noquote
set postponed             = +$my_drafts_noquote
set trash                 = +$my_trash_noquote
set smtp_url              = ""smtp://username@smtp.office365.com:587""
set smtp_authenticators   = ""login""
set copy                  = ""yes""
account-hook $folder ""set imap_user=username""

```

So after starting mutt, i end up being in INBOX for `account2`. If I press `gs`, `gd` or `gt` all of them get me where I want to be (Sent folder, drafts or trash). Then I switch to account on by `F2` and get to INBOX folder correctly but `gs`, `gt`, `gd` no longer work (Note that account1 and account2 use different names for trash, drafts and sent). (What it shows is ""Sent items"" does not exist but the words ""sent items"" are written in language of account2 but with malformed characters).


How do i need to correct the configuration?


","I finally found the resolution, so if someone got in the same issue, the answer is to add to the account1 and account2 files mapping for the sortcuts, that is from `muttrc`, the lines



```
bind    index   G  imap-fetch-mail
macro   index   gi ""<change-folder>=INBOX<enter>""       ""Go to Inbox""
macro   index   gs ""<change-folder>=$my_sent<enter>""    ""Go to Sent""
macro   index   gd ""<change-folder>=$my_drafts<enter>""  ""Go to Drafts""
macro   index   gt ""<change-folder>=$my_trash<enter>""   ""Go to Trash""

```

also have to be at the end of sourced accout1 and account2 files. It seems that `mutt` fills in the value of the variables when reading the configuration for the first time rather than keeping the variable and evaluate the expression for shortcut each time the shortcut is run with the actual variable value.


"
"i want to get all dates exclude weekends between two dates in python
I want to get all the dates between two dates excluding weekends dates.
Below is the format of dates:



```
last_date = '2019-01-21'
curr_date = '2019-02-04'

```

","Using [date.weekday()](https://docs.python.org/2/library/datetime.html)



> 
> Return the day of the week as an integer, where Monday is 0 and Sunday
>  is 6.
> 
> 
> 



```
from datetime import timedelta, date

def daterange(date1, date2):
    for n in range(int ((date2 - date1).days)+1):
        yield date1 + timedelta(n)

start_dt = date(2019,1,21)
end_dt = date(2019,2,4)

weekdays = [5,6]
for dt in daterange(start_dt, end_dt):
    if dt.weekday() not in weekdays:                    # to print only the weekdates
        print(dt.strftime(""%Y-%m-%d""))

```

EDIT:


Using [date.isoweekday()](https://docs.python.org/2/library/datetime.html):



> 
> Return the day of the week as an integer, where Monday is 1 and Sunday
>  is 7.
> 
> 
> 



```
weekdays = [6,7]
for dt in daterange(start_dt, end_dt):
    if dt.isoweekday() not in weekdays:
        print(dt.strftime(""%Y-%m-%d""))

```

OUTPUT:



```
2019-01-21
2019-01-22
2019-01-23
2019-01-24
2019-01-25
2019-01-28
2019-01-29
2019-01-30
2019-01-31
2019-02-01
2019-02-04

```

"
"How laravel uses object as an array in Facade class
I've noted that for creating a facade class, laravel provides only name ""db""


**framework/src/Illuminate/Support/Facades/DB.php**



```
class DB extends Facade
{
    /**
     * Get the registered name of the component.
     *
     * @return string
     */
    protected static function getFacadeAccessor()
    {
        return 'db';
    }
}

```

I looked deeper and figured out that this method uses the provided name


**framework/src/Illuminate/Support/Facades/Facade.php**



```
protected static function resolveFacadeInstance($name)
{
    if (is_object($name)) {
        return $name;
    }
    if (isset(static::$resolvedInstance[$name])) {
        return static::$resolvedInstance[$name];
    }
    return static::$resolvedInstance[$name] = static::$app[$name];
}

```

I understand first and second If statements.


But I have problems with understanding this:



```
return static::$resolvedInstance[$name] = static::$app[$name]

```

As I understood that `$app` is a protected property of `Facade` class which contains an instance of `\Illuminate\Contracts\Foundation\Application` class.



```
/**
     * The application instance being facaded.
     *
     * @var \Illuminate\Contracts\Foundation\Application
     */
    protected static $app;

```

My two questions:


How is it possible to use an object as an array(`static::$app[$name]`) if Application class doesn't extends ArrayObject class?


How laravel understands which class to call with providing only a short name 'db'?


","Clicking through the Laravel source, I found this. As you can see, `ApplicationContract` (the `private static $app` from your question) is implemented by `Application`. This is in turn derived from `Container`, which implements the PHP core `ArrayAccess` interface. Carefully implementing this whole chain eventually makes `Applicatin` accessible like an array.


Turns out it boils down to good ole' object oriented programming :)



```
// Illuminate/Foundation/Application.php
class Application extends Container implements ApplicationContract, HttpKernelInterface
                          ^^^^^^^^^            ^-> the private static $app in your question.

// Illuminate/Container/Container.php
class Container implements ArrayAccess, ContainerContract
                           ^^^^^^^^^^^

// PHP core ArrayAccess documentation
/**
 * Interface to provide accessing objects as arrays.
 * @link http://php.net/manual/en/class.arrayaccess.php
 */
interface ArrayAccess {

```

"
"Node.js Express3 - Middleware to add render data to all render requests
Is it possible through express middleware or another method to add render data (the second option in `res.render`) to each call in the apps routes.


My app is using passport for authentication and I would like a middleware to always append the user information to each rendered template.


Currently each of my calls to `res.render` look similar to. I would like to remove the `user : req.user` and add that to a middleware function.



```
// Page 1
res.render(""somepage1"",{data : ""somepage1"", user : req.user});

// Page 2
res.render(""somepage2"",{data : ""somepage2"", user : req.user});

```

","Express does provide [`app.locals`](http://expressjs.com/api.html#app.locals) and [`res.locals`](http://expressjs.com/api.html#res.locals), which it [automatically merges](https://github.com/visionmedia/express/blob/3.1.0/lib/application.js#L467-474) with explicit locals passed in to `render` by a route handler.



```
app.use(function(req, res, next) {
    res.locals.user = req.user;
    next();
});

```

Alternatively, you could hot patch the `render` function. I'd avoid this since it's extra overhead, but if for some reason you needed to pass information not available before your route handler runs, it's an option:



```
app.use(function(req, res, next) {
    var render = res.render;
    res.render = function(view, locals, cb) {
        if (typeof locals == 'object') locals.user = req.user;
        render.call(res, view, locals, cb);
    };
    next();
});

```

"
"jQuery - appendTo previous / preceding element
I couldn't find this on here already, and am hoping somebody can help.


I need the link (.more) to be appended to the paragraph that precedes it - I'm sure this is simple enough, but can't quite get it right:



```
<p>Lorem ipsum</p>

<p>Lorem ipsum</p>

<p>Lorem ipsum</p>

<a href=""#"" class=""more"">Read more</a>

<p>Lorem ipsum</p>

```

There may be more paragraphs in the final HTML, the reason I need to do this is CMS-related. But based on the above, my aim would be:



```
<p>Lorem ipsum</p>

<p>Lorem ipsum</p>

<p>Lorem ipsum <a href=""#"" class=""more"">Read more</a></p>

<p>Lorem ipsum</p>

```

I've tried:



```
$("".more"").appendTo(""p"");

```

But obviously this will append it to all paragraphs... perhaps this is a job for prevAll, but I can't quite put my finger on it.


Many thanks in advance for your help!


","**HTML**



```
<p>Lorem ipsum</p>
<p>Lorem ipsum</p>
<p>Lorem ipsum</p>
<a href=""#"" class=""more"">Read more1</a>
<p>Lorem ipsum</p>
<a href=""#"" class=""more"">Read more2</a>

```

**JS**



```
$('p').each(function(e){
    if($(this).next('a').hasClass('more'))
       $(this).html($(this).html()+"" "").append($(this).next('a')); 
});​

```

[**Example1**](http://jsfiddle.net/heera/Me5Dm/7/) and [**Example2 with multiple a tags.**](http://jsfiddle.net/heera/Me5Dm/6/)


"
"wordpress htaccess how to remove all the php extension except index.php
I have a WordPress blog and I want to add some custom page in it and remove the `.php` extension and querystring from urls. 


All WordPress page urls are like `index.php?pagename=`, so if I tried add `RewriteRule ^(.*)$ $1.php` in the `.htaccess` file, the other pages return `500 internal server error`. How can I remove the php extension from everything except `index.php` and `login.php`?


Here is my existing `.htaccess` file:



```
# BEGIN WordPress
<IfModule mod_rewrite.c>
RewriteEngine On
RewriteBase /
RewriteCond %{REQUEST_FILENAME} !-f
RewriteCond %{REQUEST_FILENAME} !-d
RewriteRule . /index.php [L]
# RewriteRule ^(.*)$ $1.php  # add this line will broken the wordpress rewrite rule
</IfModule>
# END WordPress

```

","If you want add, for example, few custom php pages, you should add your own rewrites before standard Wordpress part, but after `RewriteBase /`. 


Here I have added a couple of rewrites able to handle custom pages stored inside in a private directory `custom`.



```
    # BEGIN WordPress
    <IfModule mod_rewrite.c>
    RewriteEngine On
    RewriteBase /

    # *** START CUSTOM PART ***
    RewriteRule ^page1$     /custom/page1.php    [L,QSA]
    RewriteRule ^page2$     /custom/page2.php    [L,QSA]
    # *** END CUSTOM PART ***

    RewriteRule ^index\.php$ - [L]
    RewriteCond %{REQUEST_FILENAME} !-f
    RewriteCond %{REQUEST_FILENAME} !-d
    RewriteRule . /index.php [L]
    </IfModule>
    # END WordPress

```

If your site is available on `http://www.example.com`, your custom pages will be accessible using `http://www.example.com/page1`, `http://www.example.com/page2` and so on...


Just few words regarding rewrite flags `[L,QSA]`:


- `L` - If rewrite match, stop the rewriting process immediately and don't apply any more rules.
- `QSA` - Appends any query string from the original request URL to custom pages. It means that calling `http://www.example.com/page1?foo=bar&c=1` will be rewrited internally into `/custom/page1.php?foo=bar&c=1`


Please note: 


1. if you already have a page named ""page1"" in your wordpress site, this page will be overridden by custom page added into rewrite rule.
2. These Apache httpd directives were standard and generated by Wordpress, configuring `Permalink Settings`. I have just added the custom part.
3. Any time you change `Permalink Settings`, your `.htaccess` will be overwritten, better if you change `.htaccess` permission making the file readonly. And, in any case, save a copy of your .htaccess configuration.


On the other hand if you only want your Wordpress having URLs more SEO friendly, i.e. URLs without `.php` pages and without the full query strings but only the post/page slug as unique identifier, you may access to `Wordpress Permalink Settings` and select `Post name` or the option you prefer:


![Wordpress permalink settings](https://i.stack.imgur.com/C0FWF.png) 


When you activate `Permalink Settings`, Apache httpd work is limited to the .htaccess part. This means that Apache redirects everything (request header and body) via .htaccess to the index.php page and Wordpress implements internally every rewrite.
In case you want more suggestions related to Wordpress configuration, I'll suggest to write your questions in <https://wordpress.stackexchange.com/>


"
"Rolling average over inverse diagonal in Python
I have a pandas pivot table that was previously shifted and now looks like this:



```
pivot
    A    B    C    D    E
0  5.3  5.1  3.5  4.2  4.5
1  5.3  4.1  3.5  4.2  NaN
2  4.3  4.1  3.5  NaN  NaN
3  4.3  4.1  NaN  NaN  NaN
4  4.3  NaN  NaN  NaN  NaN

```

I'm trying to calculate a rolling average with a variable window (in this case 3 and 4 periods) over the inverse diagonal iterating over every column and trying to store that value in a new dataframe, which would look like this:



```
expected_df with a 3 periods window
    A    B    C    D    E
0  4.3  4.1  3.5  4.2  4.5

expected_df with a 4 periods window
    A    B    C    D    E
0  4.5  4.3  3.5  4.2  4.5

```

So far, I tried to subset the original pivot table and create a different dataframe that only contains the specified window values for each column, to then calculate the average, like this:



```
subset
    A    B    C    D    E
0  4.3  4.1  3.5  4.2  4.5
1  4.3  4.1  3.5  4.2  NaN
2  4.3  4.1  3.5  NaN  NaN

```

For this, I tried to build the following for loop:



```
df2 = pd.DataFrame()
size = pivot.shape[0]
window = 3

for i in range(size): 
    df2[i] = pivot.iloc[size-window-i:size-i,i]

```

Which does not work even when this `pivot.iloc[size-window-i:size-i,i]` does return the values I need when I manually pass in the indexes, but in the for loop, it misses the first value of the second column and so on:



```
df2
    A    B    C    D    E
0  4.3  NaN  NaN  NaN  NaN
1  4.3  4.1  NaN  NaN  NaN
2  4.3  4.1  3.5  NaN  NaN

```

Does anyone have a good idea on how to calculate the moving average or on how to fix the for loop part? Thanks in advance for your comments.


","IIUC:


### `shift` everything back



```
shifted = pd.concat([df.iloc[:, i].shift(i) for i in range(df.shape[1])], axis=1)
shifted

     A    B    C    D    E
0  5.3  NaN  NaN  NaN  NaN
1  5.3  5.1  NaN  NaN  NaN
2  4.3  4.1  3.5  NaN  NaN
3  4.3  4.1  3.5  4.2  NaN
4  4.3  4.1  3.5  4.2  4.5

```

Then you can get your mean.



```
# Change this  to get the last n number of rows
shifted.iloc[-3:].mean()

A    4.3
B    4.1
C    3.5
D    4.2
E    4.5
dtype: float64

```

Or the rolling mean



```
#   Change this  to get the last n number of rows
shifted.rolling(3, min_periods=1).mean()

          A         B    C    D    E
0  5.300000       NaN  NaN  NaN  NaN
1  5.300000  5.100000  NaN  NaN  NaN
2  4.966667  4.600000  3.5  NaN  NaN
3  4.633333  4.433333  3.5  4.2  NaN
4  4.300000  4.100000  3.5  4.2  4.5

```



---


## Numpy strides


I'll use strides to construct a 3-D array and average over one of the axes. This is faster but confusing as all ... 


Also, I wouldn't use this. I just wanted to nail down how to grab diagonal elements via strides. This was more practice for me and I wanted to share.



```
from numpy.lib.stride_tricks import as_strided as strided

a = df.values

roll = 3
r_ = roll - 1  # one less than roll

h, w = a.shape
w_ = w - 1  # one less than width

b = np.empty((h + 2 * w_ + r_, w), dtype=a.dtype)
b.fill(np.nan)
b[w_ + r_:-w_] = a

s0, s1 = b.strides
a_ = np.nanmean(strided(b, (h + w_, roll, w), (s0, s0, s1 - s0))[w_:], axis=1)

pd.DataFrame(a_, df.index, df.columns)

          A         B    C    D    E
0  5.300000       NaN  NaN  NaN  NaN
1  5.300000  5.100000  NaN  NaN  NaN
2  4.966667  4.600000  3.5  NaN  NaN
3  4.633333  4.433333  3.5  4.2  NaN
4  4.300000  4.100000  3.5  4.2  4.5

```



---


### Numba


I feel better about this than I do using strides



```
import numpy as np
from numba import njit
import warnings

@njit
def dshift(a, roll):
  h, w = a.shape
  b = np.empty((h, roll, w), dtype=np.float64)
  b.fill(np.nan)

  for r in range(roll):
    for i in range(h):
      for j in range(w):
        k = i - j - r
        if k >= 0:
          b[i, r, j] = a[k, j]

  return b

with warnings.catch_warnings():
  warnings.simplefilter('ignore', category=RuntimeWarning)

  df_ = pd.DataFrame(np.nanmean(dshift(a, 3), axis=1, ), df.index, df.columns)

df_

          A         B    C    D    E
0  5.300000       NaN  NaN  NaN  NaN
1  5.300000  5.100000  NaN  NaN  NaN
2  4.966667  4.600000  3.5  NaN  NaN
3  4.633333  4.433333  3.5  4.2  NaN
4  4.300000  4.100000  3.5  4.2  4.5

```

"
"Create responsive Bootstrap navbar with two rows
I am trying to create a responsive Bootstrap 3 based navbar with two rows. However, I am having trouble with the structure of the HTML and the collapse functionality.


Below is a visual description of the desired result, and I was hoping someone could point me in the right direction in terms of HTML/CSS (with as much default Bootstrap functionality as possible).


**Essentially the menu is desired to do the following:**


- On tablet/desktop devices, the first row is the logo and a menu of
small secondary links (Link1-3). The second row is the main menu with
main links (LinkA-E).
- On mobile, the classic collapse design should
appear with the logo and hamburger icon. The expanded menu should
show the main links (LinkA-E) first, and then the secondary links
(Link1-3) last.


**Tablet/Desktop device:**



```
|----------------------------------------------------------|
|  LOGO/BRAND                         Link1  Link2  Link3  |
|----------------------------------------------------------|
|  LinkA  LinkB  LinkC  LindD  LinkE                       |
|----------------------------------------------------------|

```

**Mobile device (collapsed):**



```
|--------------------------------------|
|  LOGO/BRAND               HAMBURGER  |
|--------------------------------------|

```

**Mobile device (expanded):**



```
|--------------------------------------|
|  LOGO/BRAND               HAMBURGER  |
|--------------------------------------|
|  LinkA                               |
|  LinkB                               |
|  LinkC                               |
|  LinkD                               |
|  LinkE                               |
|--------------------------------------|
|  Link1                               |
|  Link2                               |
|  Link3                               |
|--------------------------------------|

```

","If i had an html/css of your project i use it to show how you should do that, but in this context **THIS HTML** came from [static navbar example of Bootstrap v3.3.7](http://getbootstrap.com/examples/navbar/)


actualy the idea is placing `<div id=""navbar"" class=""navbar-collapse collapse"">` in row below `<div class=""navbar-header"">` with floating and width it 100% of its space, then pulling up `<ul class=""nav navbar-nav navbar-right"">` with `margin-top: -50px;`


With that media query we would sure it behave right in moblie device.


**HTML**



```
<div class=""container-fluid"">
  <div class=""navbar-header"">
    <button type=""button"" class=""navbar-toggle collapsed"" data-toggle=""collapse"" data-target=""#navbar"" aria-expanded=""false"" aria-controls=""navbar"">
      <span class=""sr-only"">Toggle navigation</span>
      <span class=""icon-bar""></span>
      <span class=""icon-bar""></span>
      <span class=""icon-bar""></span>
    </button>
    <a class=""navbar-brand"" href=""#"">Logo/Brand</a>
  </div>
  <div id=""navbar"" class=""navbar-collapse collapse"">
    <ul class=""nav navbar-nav"">
      <li><a href=""#"">LinkA</a></li>
      <li><a href=""#"">LinkB</a></li>
      <li><a href=""#"">LinkC</a></li>
      <li><a href=""#"">LinkD</a></li>
      <li><a href=""#"">LinkE</a></li>
    </ul>
    <ul class=""nav navbar-nav navbar-right"">
      <li><a href=""#"">Link1</a></li>
      <li><a href=""#"">Link2</a></li>
      <li><a href=""#"">Link3</a></li>
    </ul>
  </div><!--/.nav-collapse -->
</div>

```

**CSS**



```
.navbar-collapse {
    float: left;
    width: 100%;
    clear: both;
}

@media (min-width: 768px) {
    .navbar-right {
        margin-top: -50px;
    }
}

```

**JUST** Let me know if it worked or not. 


"
"Convert string into two dimensional string array in Java
I like to convert string for example :



```
String data = ""1|apple,2|ball,3|cat"";

```

into a two dimensional array like this



```
{{1,apple},{2,ball},{3,cat}}

```

I have tried using the `split("""")` method but still no solution :(


Thanks..


Kai


","
```
    String data = ""1|apple,2|ball,3|cat"";
    String[] rows = data.split("","");

    String[][] matrix = new String[rows.length][]; 
    int r = 0;
    for (String row : rows) {
        matrix[r++] = row.split(""\\|"");
    }

    System.out.println(matrix[1][1]);
    // prints ""ball""

    System.out.println(Arrays.deepToString(matrix));
    // prints ""[[1, apple], [2, ball], [3, cat]]""

```

Pretty straightforward except that [`String.split`](http://java.sun.com/javase/6/docs/api/java/lang/String.html#split%28java.lang.String%29) takes regex, so metacharacter `|` needs escaping.


### See also


- [Regular expressions and escaping special characters](https://stackoverflow.com/questions/399078/regular-expressions-and-escaping-special-characters)
- [Java Arrays.equals() returns false for two dimensional arrays.](https://stackoverflow.com/questions/2721033/java-arrays-equals-returns-false-for-two-dimensional-arrays/)
	- Use `Arrays.deepToString` and `Arrays.deepEquals` for multidimensional arrays




---


### Alternative


If you know how many rows and columns there will be, you can pre-allocate a `String[][]` and use a `Scanner` as follows:



```
    Scanner sc = new Scanner(data).useDelimiter(""[,|]"");
    final int M = 3;
    final int N = 2;
    String[][] matrix = new String[M][N];
    for (int r = 0; r < M; r++) {
        for (int c = 0; c < N; c++) {
            matrix[r][c] = sc.next();
        }
    }
    System.out.println(Arrays.deepToString(matrix));
    // prints ""[[1, apple], [2, ball], [3, cat]]""

```

"
"A ThreadLocal Supplier?
If I need a ThreadLocal of a variable, is there a need to also use Supplier (also thread-safe)?


For example, isn't the Supplier unnecessary to accomplish thread-safety here?



```
private ThreadLocal<Supplier<MyClass>> myObject = new ThreadLocal<Supplier<MyClass>>();

```

Thanks.


","Your question doesn't show the typical way to use a Supplier with a ThreadLocal. If you want a ThreadLocal of MyClass, the old (pre-1.8) way to do that was typically:



```
ThreadLocal<MyClass> local = new ThreadLocal<MyClass>();

// later
if (local.get() == null) {
  local.put(new MyClass());
}
MyClass myClass = local.get();

```

The alternative was to delcare a subclass of `ThreadLocal` that overrode the `initialValue` method.


In 1.8, you can instead use a Supplier to handle that initialization:



```
ThreadLocal<MyClass> local = ThreadLocal.withInitial(() -> new MyClass());

```

Functionally, these two are basically identical, but the Supplier version is a lot less code to write.


"
"Cost of running colour laser printers
Has anyone got a good list of the costs of running colour laser printers, both for colour and black that take into account different usage volumes.


E.g if the dram needs replacing after 20,000 pages, but I am only do 10,000 in the next 3 years, I don’t care about the cost of a new dram. 


So a simple “cost per page” figure is each printer is not good enough to be useful in real life.


(A spread sheet / website that I could type my own usage values into would be great.)


","Some review sites make a point of carrying this information. trustedreviews.com does, [like this](http://www.trustedreviews.com/printers/review/2008/12/05/Lexmark-C543dn-Colour-Laser/p4) for instance, though you will need to find one local to your currency or still do some extra research on part prices yourself. Actually, even if you do find such a site with your local prices, it is always worth checking what the actual retail prices (and availability) are like as they can differ from the official RRP (which is what review sites are likely to quote) enough to make quite a difference to your costings.


You can usually get this data from the manufacturer's sites too.


I don't know of any site that offers an easily accessed database like you are looking for. Such a site might be difficult to run because as soon as it showed one manufacturer to be better than another they'd bit hit with a lawyers letter suggesting that using the product and pricing information that way is a breach of copyright or some such (while such use is legally fine in most jurisdictions, arguing the point against a major manufacturer's legal team would be too expensive, time consuming and stressful for most to deal with).


Calculating your expected pages-per-consumable figures can be quite complex too as it depends on your actual output. Most official per-page figures for toner assume ~5% coverage without giving an example of what 5% actually looks like - I used to be under the impression that this was approximately a full page of text in a common font at a common size, though some articles I've seen ([like this for example](http://www.stinkyinkshop.co.uk/blog/2009/12/10/what-does-5-look-like/)) suggest it is closer to half a page. If you are printing charts and such then the coverage will increase considerably, and for photos it will be even higher. So all you can really hope for without much study is a good view of comparative consumables costs, which will not be easy to compare against the fixed costs of buying the printers in the first place.


"
"iPad Icon names?
Do I have to put 2 icon images in my build for iPad? and what are the names of those 2 build, one is Icon.png and what is the other one?


Thanks for all your help


","For [*iPhone* apps](http://developer.apple.com/iphone/library/documentation/UserExperience/Conceptual/MobileHIG/IconsImages/IconsImages.html) you need to provide a 57x57 `Icon.png` and may optionally provide a 29x29 `Icon-Small.png`.


To submit to AppStore you need a 512x512 icon too.


There's no special requirement for iPad. A 57x57 `Icon.png` is sufficient. Although you can add a larger 72x72 icon and add it into the `CFBundleIconFiles` key. (See [How do I specify both icons for a universal iPhone/iPad app?](https://stackoverflow.com/questions/2480563/how-do-i-specify-both-icons-for-a-universal-iphone-ipad-app).)


"
"Static Class and Getter
I've got a static class, called Test, and a private List collection. Now I want to implement a getter to return my List to the main class. Is this the correct way?


Secondly is it okay to implement a static constructor? If not, how do I properly declare my List Collection?



```
static class Storage
{

   private static List<string> store;

   static Storage()
   {
       store = new List<string>();


   }

   //Is it okay to have a getter in my static class to return my List Collection
   public static List<string> getList
   {
       get
       {
           return store;
       }


   }

   public static void addString(string add)
   {
       store.Add(add);
   }

   public static void removeString(string remove)
   {
       store.Remove(remove);
   }

```

","If you return a reference to the list in the getter then any caller can add or remove items from the list without going through the add/remove methods on the static class. If you want to prevent anyone from modifying the list I would return a [read only collection](http://msdn.microsoft.com/en-us/library/e78dcd75.aspx) from the getter:



```
public static IEnumerable<string> getList
{
   get
   {
       return store.AsReadonly();
   }
}

```

**Edit**


As Marc pointed out you will need to watch out for multi threading issues. Imagine one thread enumerates the collection obtained via the getter while another thread modifies it by add or removing an item -> you will get an exception saying that you can't modify a collection while enumerating it. The only solution for that is to lock all access to the private collection and return a copy in the getter.


"
"How to find out size of a PhotoImage in Tkinter?
How do I access the width and height information on a `PhotoImage()` class object?
I tried `PhotoImage(...).winfo_width()` and `PhotoImage(...)[""Width""]`. Both of them didn't work.


","The PhotoImage objects have a `width` and `height` method:



```
import Tkinter as tk

image_data = '''
    R0lGODlhEAAQAMQZAMPDw+zs7L+/v8HBwcDAwLW1teLi4t7e3uDg4MLCwuHh4e7u7t/f38TExLa2
    tre3t7i4uL6+vu/v77q6uu3t7b29vby8vLm5ubu7u+3t7QAAAAAAAAAAAAAAAAAAAAAAACH5BAEA
    ABkALAAAAAAQABAAAAWNYCaOZFlWV6pWZlZhTQwAyYSdcGRZGGYNE8vo1RgYCD2BIkK43DKXRsQg
    oUQiFAkCI3iILgCLIEvJBiyQiOML6GElVcsFUllD25N3FQN51L81b2ULARN+dhcDFggSAT0BEgcQ
    FgUicgQVDHwQEwc+DxMjcgITfQ8Pk6AlfBEVrjuqJhMOtA4FBRctuiUhADs=
'''

root = tk.Tk()
image = tk.PhotoImage(data=image_data)
dimensions = ""image size: %dx%d"" % (image.width(), image.height())
label = tk.Label(root, compound=""top"", image=image, text=dimensions)
label.pack()
root.mainloop()

```

"
"Go http listener with data update every seconds
I'm trying to build a little website in Go containing a report based on data collected from a web service. It uses an API to query the service for data. The service can only be queried once every few seconds.


However, I have to query it a number of times to get the complete report data. Right now I just hammer the API to update my whole data structure each time the http handler (http.HandleFunc) is called. Of course, this is bad because it triggers lots of queries to the external API that are throttled. So, my report comes up very, very, very, slowly.


What I want to do instead is to have a function to updateReportData in a non-blocking way and store that data in some variable that the http.HandleFunc() can just ingest without calling the external API.


But, I'm very new to Go (and things like closures, semaphores, concurrency, etc) and so I'm not really sure how to build this. Should I be using channels? Should I use timers? How can I get the updateReportData to not block the http.HandleFunc, but still run on a fixed interval?


To sum up, I want to have a background routine update a data structure on a fixed interval and I want to be able to use http.HandleFunc to serve whatever data is in the data structure any time i make an http request to the program. I just have no idea how to start. Any advice would be appreciated.


","There are a few things you need to do:


1. Create a background service that polls for the data. This service can run as a goroutine that periodically checks for new data.
2. Create a channel that the background service uses to send new data to a centralized place to store the values. The background service should write data to this channel whenever it finds something new. Another option would be to protect the centralized data store with a mutex. Depending on the way the data is written and read, one option will be a better choice.
3. Create a HTTP handler that returns the current contents of the centralized data store.


Here is a simplified example showing how to use a goroutine and a `sync.RWMutext` to accomplish what you want:



```
package main

import (
    ""fmt""
    ""net/http""
    ""sync""
    ""time""
)

var (
    timeSumsMu sync.RWMutex
    timeSums   int64
)

func main() {
    // Start the goroutine that will sum the current time
    // once per second.
    go runDataLoop()
    // Create a handler that will read-lock the mutext and
    // write the summed time to the client
    http.HandleFunc(""/"", func(w http.ResponseWriter, r *http.Request) {
        timeSumsMu.RLock()
        defer timeSumsMu.RUnlock()
        fmt.Fprint(w, timeSums)
    })
    http.ListenAndServe("":8080"", nil)
}

func runDataLoop() {
    for {
        // Within an infinite loop, lock the mutex and
        // increment our value, then sleep for 1 second until
        // the next time we need to get a value.
        timeSumsMu.Lock()
        timeSums += time.Now().Unix()
        timeSumsMu.Unlock()
        time.Sleep(1 * time.Second)
    }
}

```

"
"Parsing a string in haskell
I've got some strings that I want to parse into a list of ""chunks"". My strings look like this



```
""some text [[anchor]] some more text, [[another anchor]]. An isolated [""

```

And I expect to get back something like this



```
[
   TextChunk ""some text "",
   Anchor ""anchor"",
   TextChunk "" some more text, ""
   Anchor ""another anchor"",
   TextChunk "". An isolated [""
]

```

I've managed to write a function and types that do what I need, but they seems overly ugly.
Is there a nicer way to do this?



```
data Token = TextChunk String | Anchor String deriving (Show)
data TokenizerMode = EatString | EatAnchor deriving (Show)

tokenize::[String] -> [Token]
tokenize xs =  
  let (_,_,tokens) = tokenize' (EatString, unlines xs, [TextChunk """"])
  in reverse tokens

tokenize' :: (TokenizerMode, String, [Token]) -> (TokenizerMode, String,[Token])
-- If we're starting an anchor, add a new anchor and switch modes
tokenize' (EatString, '[':'[':xs, tokens) = tokenize' (EatIdentifier, xs, (Identifier """"):tokens )
-- If we're ending an anchor ass a new text chunk and switch modes
tokenize' (EatAnchor, ']':']':xs, tokens) = tokenize' (EatString, xs, (TextChunk """"):tokens )
-- Otherwise if we've got stuff to consume append it
tokenize' (EatString, x:xs, (TextChunk t):tokens) = tokenize'( EatString, xs, (TextChunk (t++[x])):tokens)
tokenize' (EatAnchor, x:xs, (Identifier t):tokens) = tokenize'( EatAnchor, xs, (Identifier (t++[x])):tokens)
--If we've got nothing more to consume we're done.
tokenize' (EatString, [], tokens) = ( EatString, [], tokens)
--We'll only get here if we're given an invalid string
tokenize' xx = error (""Error parsing .. so far "" ++ (show xx))

```

","This should work, including lone brackets:



```
import Control.Applicative ((<$>), (<*), (*>))
import Text.Parsec

data Text = TextChunk String
          | Anchor String
          deriving Show

chunkChar = noneOf ""["" <|> try (char '[' <* notFollowedBy (char '[')) 
chunk     = TextChunk <$> many1 chunkChar
anchor    = Anchor <$> (string ""[["" *> many (noneOf ""]"") <* string ""]]"")
content   = many (chunk <|> anchor)

parseS :: String -> Either ParseError [Text]
parseS input = parse content """" input

```

Note the use of `try` to allow backtracking when the `chunkChar` parser matches two opening brackets. Without `try`, the first bracket would have been consumed at that point.


"
"MySQL insert query failing when data gets large
I have already followed the question:[Data Limit on MySQL DB Insert](https://stackoverflow.com/questions/7651854/data-limit-on-mysql-db-insert) I was unable to solve it with the limited info.
I am using WAMP.
I have numerous Rich Text editors and 4 images which are being sent over to another page by a POST request. After a certain threshold limit, the query is failing. Is there a way around?


EDIT: while displaying the query string it seems that I am able to retrieve every bit of data that was sent via POST. I am quite sure that it is DB related. 
Images are being stored as a BLOB.


EDIT #2: Error showing is ""MySQL server has gone away"".


","You may be violating the `max_allowed_packet` setting. See [here](http://dev.mysql.com/doc/refman/5.0/en/packet-too-large.html) for more data.


Quote,



> 
> If you are using the mysql client program, its default
>  max\_allowed\_packet variable is 16MB.
> 
> 
> 


If you are uploading uncompressed images, this value is fairly easy to reach.


Also, it would be great if you could name the specific database interface class that you use (PDO? mysql\_? mysqli\_?), as different classes handle errors differently. It could just as well not handle an oversized packet situation at all.


P.S.: You should really be checking your logs for the specific error you encounter. The first place to look for would be `/var/log/mysql/error.log` (could vary depending on your env)


**Update:**
`mysql_error()` returned `""MySQL server has gone away""`



> 
> From the manual pages for the error: ""You can also get these errors if
>  you send a query to the server that is incorrect or too large. If
>  mysqld receives a packet that is too large or out of order, it assumes
>  that something has gone wrong with the client and closes the
>  connection. If you need big queries (for example, if you are working
>  with big BLOB columns), you can increase the query limit by setting
>  the server's max\_allowed\_packet variable, which has a default value of
>  1MB. You may also need to increase the maximum packet size on the
>  client end...""
> 
> 
> 


(quote courtesy of @Colin Morelli)


"
"Backspace doesn't work in gvim 7.2 64-bit for Windows?
I just installed the [64-bit zipped version for Windows of gvim](http://vim-win3264.googlecode.com/files/vim72.zip) on my new machine at work, and [I changed some settings](https://stackoverflow.com/questions/5395947/how-to-get-gvim-to-play-nicely-with-visual-studio-2010) to make it compatible with Visual Studio 2010.


Now the backspace key doesn't work for some reason.


Here's my .vimrc file:



```
set nowrap
set ruler
set encoding=utf-8
set fileencoding=utf-8
set ff=dos

```

It's only a minor annoyance (the delete key works), but I was just wondering if anybody knew.


","Does it entirely not work, or does it just not backspace past where you went into insert mode? There's an option [`backspace`](http://vimdoc.sourceforge.net/htmldoc/options.html#%27backspace%27) which controls this:



```
Influences the working of <BS>, <Del>, CTRL-W and CTRL-U in Insert
mode.  This is a list of items, separated by commas.  Each item allows
a way to backspace over something:
value       effect
indent      allow backspacing over autoindent
eol         allow backspacing over line breaks (join lines)
start       allow backspacing over the start of insert; CTRL-W and CTRL-U
            stop once at the start of insert.

```

[owen\_water's suggestion](https://stackoverflow.com/questions/5419848/backspace-doesnt-work-in-gvim-7-2-64-bit-for-windows/5427461#5427461) is the same as Carpetsmoker's comment, enabling all three of these. That's generally what people want, since it's a bit surprising to have Vim refuse to let you backspace text that's right there. So take your pick of the two equivalent choices - I'd go with the more verbose and clear one:



```
set backspace=2
set backspace=indent,eol,start

```

In some cases, bad terminals can mess with backspace and delete, but I doubt that's your problem in Windows. There is some ability to fix this; see [`:help fixdel`](http://vimdoc.sourceforge.net/htmldoc/options.html#:fixdel).


"
"What's wrong with this initialization of unique\_ptr?
Can somebody tell me, what is wrong with the following initialization of unique\_ptr?



```
int main()
{
  unique_ptr<int> py(nullptr);
  py = new int;
  ....
}

```

g++ -O2 xxx.cc -lm -o xxx -std=c++11 says:



```
error: no match for ‘operator=’ (operand types are    ‘std::unique_ptr<int>’ and ‘int*’)
   py = new int;
      ^

```

Doing



```
unique_ptr<int> px(new int);

```

works just fine.


","The initialization is fine in both pieces of code, `unique_ptr` has [constructors](http://en.cppreference.com/w/cpp/memory/unique_ptr/unique_ptr) for both `nullptr` and naked pointers.


What is failing in the first snippet is the assignment, that is because `unique_ptr` does not have an [`operator=`](http://en.cppreference.com/w/cpp/memory/unique_ptr/operator%3D) overload that accepts a naked pointer as its right hand side. It does accept another `unique_ptr` though, so you could do this:



```
py = unique_ptr<int>{new int};
py = std::make_unique<int>(); // Since c++14

```

Or you could look at [`reset`](http://en.cppreference.com/w/cpp/memory/unique_ptr/reset) that also accepts a naked pointer and has more or less the same meaning:



```
py.reset(new int);

```

"
"Unable to mock AWSSecretsManager in Spring boot using Mockito
I am new to Java and I have created sample class for which I need to write unit test case



```
import com.amazonaws.services.secretsmanager.AWSSecretsManager;
import com.amazonaws.services.secretsmanager.AWSSecretsManagerClientBuilder;
import com.amazonaws.services.secretsmanager.model.GetSecretValueRequest;
import com.amazonaws.services.secretsmanager.model.GetSecretValueResult;
import com.amazonaws.services.secretsmanager.model.InvalidParameterException;
import com.amazonaws.services.secretsmanager.model.ResourceNotFoundException;
import org.springframework.stereotype.Component;

@Component
public class MySecretService {

    private AWSSecretsManager client;

    public MySecretService() {
        this.client = AWSSecretsManagerClientBuilder.standard().withRegion(""US-west-2"").build();
    }

    public String GetMyKey(String secretId) {
        GetSecretValueRequest getSecretValueRequest = new GetSecretValueRequest();
        getSecretValueRequest.setSecretId(secretId);
        try {
            GetSecretValueResult getSecretValueResult = client.getSecretValue(getSecretValueRequest);

            return getSecretValueResult.getSecretString();
        } catch (ResourceNotFoundException resourceNotFoundException) {
            throw resourceNotFoundException;
        } catch (InvalidParameterException invalidParameterException) {
            throw invalidParameterException;
        }
        catch (Exception ex)
        {
            throw  ex;
        }
    }
}

```

I tried to write unit test case as below



```
import com.amazonaws.services.secretsmanager.AWSSecretsManager;
import com.amazonaws.services.secretsmanager.AWSSecretsManagerClientBuilder;
import com.amazonaws.services.secretsmanager.model.GetSecretValueRequest;
import com.amazonaws.services.secretsmanager.model.GetSecretValueResult;
import com.amazonaws.services.secretsmanager.model.ResourceNotFoundException;
import org.junit.jupiter.api.BeforeEach;
import org.junit.jupiter.api.Test;
import org.junit.runner.RunWith;
import org.mockito.Mock;
import org.mockito.junit.MockitoJUnitRunner;

import static org.junit.jupiter.api.Assertions.*;
import static org.mockito.Mockito.when;

@RunWith(MockitoJUnitRunner.class)
class MySecretServiceTest {

    @Mock
    private AWSSecretsManager client;

    private MySecretService mySecretService;

    @BeforeEach
    void setUp() {
        this.mySecretService = new MySecretService();
    }

    @Test
    void getMyKey() {
        GetSecretValueRequest getSecretValueRequest = new GetSecretValueRequest();
        getSecretValueRequest.setSecretId(""test"");

        GetSecretValueResult getSecretValueResult = new GetSecretValueResult();
        getSecretValueResult.setSecretString(""xyz"");
        when(this.client.getSecretValue(getSecretValueRequest))
                .thenReturn( getSecretValueResult);

        assertEquals(""xyz"", mySecretService.GetMyKey(""test""));
    }
}

```

However my test is getting failed where it says **AWSSecretsManager** is null with below error.


*java.lang.NullPointerException
at com.example.demo.MySecretServiceTest.getMyKey(MySecretServiceTest.java:40)*


I tried different solution but won't help me to execute the test cases successfully.


","I think that way how did you structure your code, you make it really hard to inject `AWSSecretsManager` client. This is because it is not exposed as a dependency in the `MySecretService` service.


What I suggest doing is the following:


1. Create a configuration class and put an instance for `AWSSecretsManager` in the IoC container:



```
@Configuration
public class AWSConfig {
    @Bean
    public AWSSecretsManager awsSecretsManager() {
        return AWSSecretsManagerClientBuilder.standard()
                .withRegion(""us-east-1"")
                .build();
    }
}

```

2. Inject the bean of `AWSSecretsManager` inside your service:



```
@Service
public class MySecretService {
    private final AWSSecretsManager awsSecretsManager;

    public SecretsService(AWSSecretsManager awsSecretsManager) {
        this.awsSecretsManager = awsSecretsManager;
    }

    public String getMyKey(String secretId) {
        GetSecretValueRequest getSecretValueRequest = new GetSecretValueRequest().withSecretId(secretId);
        try {
            GetSecretValueResult getSecretValueResult = awsSecretsManager.getSecretValue(getSecretValueRequest);
            return getSecretValueResult.getSecretString();
        } catch (ResourceNotFoundException | InvalidParameterException resourceNotFoundException) {
            throw resourceNotFoundException;
        }
    }
}

```

3. Now, in your test I see that you are mixing JUnit 4 and Junit 5 dependencies, which is not a good idea. The latest Spring Boot starter for testing comes with JUnit 5, I suggest sticking to it. Knowing this, we can do the following:



```
import com.amazonaws.services.secretsmanager.AWSSecretsManager;
import com.amazonaws.services.secretsmanager.model.GetSecretValueRequest;
import com.amazonaws.services.secretsmanager.model.GetSecretValueResult;
import org.junit.jupiter.api.Test;
import org.junit.jupiter.api.extension.ExtendWith;
import org.mockito.InjectMocks;
import org.mockito.Mock;
import org.mockito.junit.jupiter.MockitoExtension;

import static org.junit.jupiter.api.Assertions.assertEquals;
import static org.mockito.Mockito.when;

@ExtendWith(MockitoExtension.class)
class MySecretServiceTest {

    @Mock
    private AWSSecretsManager client;

    @InjectMocks
    private SecretsService mySecretService;

    @Test
    void getMyKey() {
        GetSecretValueRequest getSecretValueRequest = new GetSecretValueRequest();
        getSecretValueRequest.setSecretId(""test"");

        GetSecretValueResult getSecretValueResult = new GetSecretValueResult();
        getSecretValueResult.setSecretString(""xyz"");
        when(client.getSecretValue(getSecretValueRequest))
                .thenReturn(getSecretValueResult);

        assertEquals(""xyz"", mySecretService.GetMyKey(""test""));
    }
}

```

"
"html5 input type datetime value not showing on ASP.NET webform
*Edit* 
I have an Asp.Net FormView, with an itemtemplate:



```
 <asp:TemplateField HeaderText=""DateStart"" SortExpression=""DateStart"">
                    <EditItemTemplate>
                        <asp:TextBox ID=""TextBoxEditStageDetailsDateStart"" type=""datetime-local"" runat=""server""
                            Text='<%# Bind(""DateStart"") %>' CssClass=""TextBoxDateTime""></asp:TextBox>
                    </EditItemTemplate>
                    <ItemTemplate>
                        <asp:Label ID=""Label1"" runat=""server"" Text='<%# Bind(""DateStart"") %>'></asp:Label>
                    </ItemTemplate>
                    <ControlStyle Width=""160px"" />
                </asp:TemplateField>

```

When the Form is in View Mode, the date is correctly showed (of course when the date is set):


![enter image description here](https://i.stack.imgur.com/HrudP.jpg)


However, when the Form change in Edit Mode, the input with type=""datetime-local"" is showed **but** is ***empty***, requiring the user to choose a *new* date. This is the **Popup** showed:


![enter image description here](https://i.stack.imgur.com/oBRfY.jpg)


However, as I can see inspecting the generated html source, the **value** is correctly set...



```
<input name=""ctl00$MainContent$StageDetailsView$TextBoxEditStageDetailsDateStart"" value=""28/05/2013 10:00:00"" id=""MainContent_StageDetailsView_TextBoxEditStageDetailsDateStart"" class=""TextBoxDateTime"" type=""datetime-local"" style=""width:160px;"">

```

So two problems:
first is that in Edit mode, I am not able to see what the actual value is;
second if I only want to change the day or hour I have to re-enter the whole date...


My question is: Is this the Html5 planned behaviour or I am missing something? 


","According to the [W3 specification](http://www.w3.org/TR/html-markup/input.datetime-local.html#input.datetime-local.attrs.value):



> 
> **value** = local date and time 
> 
> 
> A string representing a local date and time. The following parts, **in
>  exactly the following order**: 
> 
> 
> - A date.
> - The literal string ""T"".
> - A time.
> 
> 
> Example: 
> 
> 
> 1985-04-12T23:20:50.52
> 
> 
> 1996-12-19T16:39:57
> 
> 
> 


So, in your case:



```
value=""2013-05-28T10:00:00""

```

will work.


**[UPDATE]**


You may use the following `aspx` markup:



```
<asp:TemplateField HeaderText=""DateStart"" SortExpression=""DateStart"">
    <EditItemTemplate>
        <asp:TextBox ID=""TextBoxEditStageDetailsDateStart"" type=""datetime-local"" runat=""server""
            Text='<%# GetFormattedDate(Eval(""DateStart"")) %>' CssClass=""TextBoxDateTime""></asp:TextBox>
    </EditItemTemplate>
    <ItemTemplate>
        <asp:Label ID=""Label1"" runat=""server"" Text='<%# Bind(""DateStart"") %>'></asp:Label>
    </ItemTemplate>
    <ControlStyle Width=""160px"" />
</asp:TemplateField>

```

In your code behind, define the following method:



```
protected string GetFormattedDate(object dateTimeObject)
{
    DateTime dateTime;
    if (DateTime.TryParse(dateTimeObject.ToString(), out dateTime))
    {
        return String.Format(""{0}-{1}-{2}T{3}:{4}:{5}"",
            dateTime.Year,
            dateTime.Month.ToString().PadLeft(2, '0'),
            dateTime.Day.ToString().PadLeft(2, '0'),
            dateTime.Hour.ToString().PadLeft(2, '0'),
            dateTime.Minute.ToString().PadLeft(2, '0'),
            dateTime.Second.ToString().PadLeft(2, '0')
            );
    }
    return null;
}

```

"
"How can I use one function from main application and bootloader? (embedded)
First of all I need to say I develop application for embedded device based on cortex m4.


I have functions that are common for bootloader and main application. For now I compile source files 2 times once for bootloader and app. But I am running short on space for dual bank dfu and I would like to have those functions only once in ROM. Any idea how can I achieve this? 


**EDIT:**


Using functions pointers maybe danger in some cases, check on my problems - [Using pointer functions - 2 seperate applications on 1 device](https://stackoverflow.com/questions/46520290/using-pointer-functions-2-seperate-applications-on-1-device)


","This is only a partial answer and assumes you can jump into your bootloader from your main code using the same address space. Then, a common technique is to provide your ""bootloader API"" as a table of function pointers.


Say for example you have the following functions in your bootloader:



```
static int do_something(void)
{
    return 42;
}

static int do_something_else(int arg)
{
    return arg+5;
}

```

Then you would declare your API in a header like this:



```
struct bootloaderApi
{
    int (*do_something)(void);
    int (*do_something_else)(int arg);
};

```

In the implementation of your bootloader, you define this table in its own section:



```
// this is GCC syntax, use whatever your compiler provides to specify the section
struct bootloaderApi api __attribute__((section(""API""))) = {
    do_something,
    do_something_else
};

```

Then when building the bootloader, make sure your section is placed at a suitable fixed address. When e.g. using the GNU linker, you could have something like this in your linker script:



```
SECTIONS {
  // standard sections, e.g.:
  .text : { *(.text) }
  .data : { *(.data) } 
  .bss :  { *(.bss)  *(COMMON) }

  // your API table:
  .API 0x10000 : { *(.API) }
}

```

This now assumes your API table will be placed at `0x10000`. Then you could do the following to access the API from your main code:



```
struct bootloaderApi *api = (struct bootloaderApi *)0x10000;

api->do_something();

```

All of this is just a sketch to give you an idea how to do this in a sensible way. It will heavily depend on your target platform and the toolchain you're using.


"
"What is C equivalent of python's map()
Here is the situation:


Input consist of space-delimited integers/float/doubles type of data.The datatype is going to be the same. The length for the input is decided during runtime. The input appears as follows (single test case per line)



```
1 5 7 9
2 32 5 5 6
2 5 7 43 73 364
1 2 

```

In python using the map function, returns a list of the input with the required data type. I need to implement the same using C. Here's what I have tried so far:


1. Use `fgets` and parse the input, but too many chances of failure and/or slow
2. Use functions like `strtol()`, `strok()` etc, but they only work when the size of the input is known.


Any suggestions on how to go about it?


PS: The maximum number of inputs is known, the number of inputs is always below this arbitrarily large value, but their number can vary.


","Your question contains *a lot* of talk about things that have nothing to do with `map()`. You should try to avoid this, and be more focused. If there are several things you wonder about, separate them out into their own questions, don't just rant away about the entire problem.


In C there's no way of writing a general ""apply this function"" code, without first inventing a way to describe all the data involved, then it might be possible.


If you data is integer, you need something like this:



```
int * map_int(const int *input, size_t input_len, int (*func)(int x));

```

The above should return a dynamically allocated array of the same size as the input, where each element is the result of running `func()` on the corresponding input element. That's probably as close to a generic `map()` as you need to come, here.


This could be trivially implemented, like so:



```
int * map_int(const int *input, size_t input_len, int (*func)(int x))
{
  int *out = malloc(input_len * sizeof *out);
  if(out != NULL)
  {
    for(size_t i = 0; i < input_len; ++i)
      out[i] = func(input[i]); 
  }
  return out;
}

```

Note that the above will assume `func` to be valid; a more reasonable general-purpose function should probably treat a `NULL` function pointer as the identity map. Also note that it does heap-allocation and moves ownership to the caller. A more reasonable implementation should perhaps support caller-owned buffers, too.


"
"Stopping a method after a specified time period
I'm running parallel operations on a list of PCs I'm getting from the Active Directory. I was using this method to check PC states, such as whether the computer was online, or if a specific directory existed. However due to the nature of these, sometimes slow, operations, I wanted to include a timeout so my application could continue to operate.



```
public static T MethodTimeout<T>(Func<T> f, int timeout, out bool completed)
{
    T result = default(T);
    var thread = new Thread(() => result = F());
    thread.Start();
    Completed = thread.Join(Timeout);
    if (!Completed) thread.Abort();
    return result;
}

```

This works for the most part, but the processing usage seemed to spike a little, and in a few cases I ran into Out of Memory exceptions. So, I since changed the method to use tasks in the hope that the `ThreadPool` would eliminate the aforementioned problems:



```
public static T MethodTimeout<T>(Func<T> f, int timeout, out bool completed)
{
    T result = default(T);

    var timedTask = Task.Factory.StartNew(() => result = F());
    Completed = timedTask.Wait(Timeout);

    return result;
}

```

However, I have a feeling I'm just filling up the `ThreadPool` with processes that are hanging waiting for these potentially long tasks to finish. Since I'm passing the task's function as a parameter I don't see a way to use a cancellation token. However, my experience is very limited with these classes and I might be missing some superior techniques.


Here's how I've been using the method above:



```
bool isReachable; // Check if the directory exists (1 second)
MethodTimeout(() => Directory.Exists(startDirectory), 1000, out isReachable);

```

A quick note, I'm only running the above check after I've already confirmed the PC is online via a WMI call (also using `MethodTimeout`). I'm aware through my early testing that checking the directory before that ends up being inefficient.


I'm also open to replacing this approach with something better.


","I may be the harbinger of bad news here, but this scenario is way more difficult to deal with than most people think. It seems that you are already picking up on this. Using the cooperative cancellation mechanisms is all good, but you have to be able to actually cancel the time consuming operation to be able to use them effectively. The problem is that `Directory.Exists` can not be cancelled.


The problem with your first solution is that you are aborting a thread. This is not a good idea because it halts the thread at unpredictable points. This may lead to the corruption of data structures for anything that was executing on the call stack when the abort got injected. In this particular case I would not be surprised if the `Thread.Abort` call actually hung. The reason is that aborts are normally delayed while the execution is in unmanaged code. It is likely `Directory.Exists` defers to an unmanaged module. If that is the case then the abort will not work anyway.


The problem with the second solution is that you are leaving the task orphaned. That `Directory.Exists` call would still be executing on a thread pool thread somewhere. This is because you have not actually cancelled it.


To be honest I am not really sure what to do about this. The lack of a cancellable `Directory.Exists` method is very problematic. My first thought is to attempt a write or read from the directory you want to test as kind of a proxy for checking its existence. The `FileStream` class does have cancellable operations. In fact, many of the methods even accept a `CancellationToken` as a parameter. Or you could close the `FileStream` and that will cancel any pending operations as well. Another option might be to use the Win32 API functions to do the IO. Then you could call [CancelSynchronousIo](http://msdn.microsoft.com/en-us/library/windows/desktop/aa363794%28v=vs.85%29.aspx) if necessary.


I know this is not much of an answer because all I have really done is tell you what you cannot do, but did not offer a definitive solution. The point is that the best solution starts with a cancellable operation. It is unfortunate that some of the BCL classes do not provide these even when they should. 


"
"How do I access the content of folders inside a jar file?
I need to take a look at the names of files inside a specific package. Currently, I'm doing the following:



```
ClassLoader loader = Thread.currentThread().getContextClassLoader();
URL packageUrl = loader.getResource(getPackage().replace('.', '/'));
File packageDir = new File(packageUrl.getFile());

// Work with the files inside the directory

```

This actually works just fine in Eclipse, since it doesn't package my application by default. When I run it from a jar file, however, I get a `NullPointerException` on `packageUrl`.


So, how do I obtain a list of the contents of a package that's inside a jar? My research suggested using `getResourceAsStream`, but I'm not quite sure how to explore a directory with an `InputStream`, if that's even possible.


","You could get path to the jar file and open it with ZipInputStream list all the files inside that jar. 


To know the path of the running jar, try using: 



```
InputStream in = MyClass
                .class
                .getProtectionDomain()
                .getCodeSource()
                .getLocation()
                .openStream();

```

See also: [How do I list the files inside a JAR file?](https://stackoverflow.com/questions/1429172/list-files-inside-a-jar)


**update**


I've compiled an ran your solution and works perfect:



```
C:\java\injar>dir
 El volumen de la unidad C no tiene etiqueta.
 El número de serie del volumen es: 22A8-203B

 Directorio de C:\java\injar

21/02/2011  06:23 p.m.    <DIR>          .
21/02/2011  06:23 p.m.    <DIR>          ..
21/02/2011  06:23 p.m.             1,752 i18n.jar
21/02/2011  06:23 p.m.    <DIR>          src
21/02/2011  06:21 p.m.    <DIR>          x

C:\java\injar>jar -tf i18n.jar
META-INF/
META-INF/MANIFEST.MF
I18n.class
x/
x/y/
x/y/z/
x/y/z/hola.txt

C:\java\injar>type src\I18n.java
import java.util.*;
import java.net.*;
import java.util.jar.*;
class I18n {
    public static void main( String ... args ) {
        getLocaleListFromJar();
    }
    private static List<Locale> getLocaleListFromJar() {
        List<Locale> locales = new ArrayList<Locale>();
        try {
            URL packageUrl = I18n.class.getProtectionDomain().getCodeSource().getLocation();
            JarInputStream jar = new JarInputStream(packageUrl.openStream());
            while (true) {
                JarEntry entry = jar.getNextJarEntry();
                if (entry == null) {
                    System.out.println( ""entry was null "");
                    break;
                }
                String name = entry.getName();
                System.out.println( ""found : "" +name );
                /*if (resourceBundlePattern.matcher(name).matches()) {
                    addLocaleFromResourceBundle(name, locales);
                }*/
           }
        } catch (Exception e) {
            System.err.println(e);
            return null;
            //return getLocaleListFromFile(); // File based implementation in case resources are not in jar
        }
        return locales;
    }
}


C:\java\injar>java -jar i18n.jar
found : I18n.class
found : x/
found : x/y/
found : x/y/z/
found : x/y/z/hola.txt
entry was null

```

"
"Why does Optional.map make this assignment work?

```
Optional<ArrayList<String>> option = Optional.of(new ArrayList<>());

Optional<ArrayList<?>> doesntWork = option;

Optional<ArrayList<?>> works = option.map(list -> list);

```

The first attempted assignment does not compile, but the second one with the `map` does. It feels like the `map` shouldn't actually accomplish anything, but for some reason it turns my `Optional<ArrayList<String>>` into an `Optional<ArrayList<?>>`. Is there some sort of implicit cast going on?


","If you look into the code of `map` and follow all the method calls, you'll see that `option.map(list -> list)` ends up returning `new Optional<>(option.get())`. So you can replace your last assignment with:



```
Optional<ArrayList<?>> works = new Optional<>(option.get());

```

This creates a new `Optional<ArrayList<?>>` and initializes its `value` instance variable (whose type is `ArrayList<?>`) with the `ArrayList<String>` returned by `map.get()`. This is a valid assignment.



> 
> Is there some sort of implicit cast going on?
> 
> 
> 


No, `map` returns a new `Optional` instance. It doesn't cast the original instance on which it was called.


Here's the chain of method calls:



```
option.map(list -> list)

```

returns (since `option` is not empty)



```
Optional.ofNullable(mapper.apply(value))

```

which in your case is the same as



```
Optional.ofNullable(value)

```

which returns (since the value is not null):



```
Optional.of(value)

```

which returns



```
new Optional<>(value)

```

"
"How to call a controller's method from a view?
I'm developing a small application in Ruby-On-Rails. I want to call a controller's method from a view. This method will only perform some inserts in the database tables. What's the correct way to do this? I've tried something like this but apparently the method code is not executed:



```
<%= link_to 'Join', method: join_event %>

```

","The method option in a link\_to method call is actually the HTTP method, not the name of the action. It's useful for when passing the HTTP Delete option, since the RESTful routing uses the DELETE method to hit the destroy action.


What you need to do here, is setup a route for your action. Assuming it's called `join_event`, add the following to your routes.rb:



```
match '/join_event' => 'controllername#join_event', :as => 'join_event'

```

Be sure to change controllername to the name of the controller you are using. Then update your view as follows: 



```
<%= link_to 'Join', join_event_path %>

```

The \_path method is generated based on the as value in the routes file.


To organize your code, you might want to encapsulate the inserts into a static model method. So if you have a model called `MyModel` with a name column, you could do



```
class MyModel
  # ...
  def self.insert_examples
    MyModel.create(:name => ""test"")
    MyModel.create(:name => ""test2"")
    MyModel.create(:name => ""test3"")
  end
end

```

Then just execute it in your action via:



```
MyModel.insert_examples

```

"
"Does anybody know of a proc to turn a row into an INSERT statement?
Does anybody know of a proc or script which will generate any row into an insert statement into the same table?


Basically, I'd like to call something like



```
exec RowToInsertStatement 'dbo.user', 45;

```

And the following code would be output



```
insert into dbo.MyTable( FirstName, LastName, Position)
values( 'John', 'MacIntyre', 'Software Consultant');

```

I realize I could 



```
insert into dbo.MyTable
select * from dbo.MyTable where id=45;

```

But this obviously won't work, because the ID column will complain (I hope it complains) and there's no way to just override that one column without listing all columns, and in some tables there could be hundreds.


So, does anybody know of a proc that will write this simple insert for me?


**EDIT 3:04**: The purpose of this is so I can make a copy of the row, so after the INSERT is generated, I can modify it into something like



```
insert into dbo.MyTable( FirstName, LastName, Position)
values( 'Dave', 'Smith', 'Software Consultant');

```

.. no obviously this contrived example is so simple it doesn't make sense, but if you have a table with 60 columns, and all you need is to change 3 or 4 values, then it starts to be a hassle.


Does that make sense?


","**Update**


I believe the following dynamic query is what you want:



```
declare @tableName varchar(100), @id int, @columns varchar(max), @pk varchar(20)
set @tableName = 'MyTable'
set @pk = 'id'
set @id = 45

set @columns = stuff((select ',['+c.name+']' [text()] from sys.tables t
join sys.columns c on t.object_id = c.object_id
where t.name = @tableName and c.name <> @pk for xml path('')),1,1,'')

print 'insert into [' + @tableName + '] (' + @columns + ')
select ' + @columns + '
from [' + @tableName + '] 
where ' + @pk + ' = ' + cast(@id as varchar)

```

**Update 2**


The actual thing that you wanted:



```
declare @tableName varchar(100), @id int, @columns nvarchar(max), @pk nvarchar(20), @columnValues nvarchar(max)
set @tableName = 'MyTable'
set @pk = 'id'
set @id = 45

set @columns = stuff((select ',['+c.name+']' [text()] from sys.tables t
join sys.columns c on t.object_id = c.object_id
where t.name = @tableName and c.name <> @pk for xml path('')),1,1,'')

set @columnValues = 'set @actualColumnValues = (select' +
stuff((select ','','''''' + cast(['+c.name+'] as varchar(max)) + '''''''' [text()]' [text()] 
from sys.tables t
join sys.columns c on t.object_id = c.object_id
where t.name = @tableName and c.name <> @pk for xml path('')),1,1,'')
+ 'from [' + @tableName + ']
where ' + @pk + ' = ' + cast(@id as varchar) 
+ 'for xml path(''''))'

--select @columnValues
declare @actualColumnValues nvarchar(max), @columnValuesParams nvarchar(500)
SET @columnValuesParams = N'@actualColumnValues nvarchar(max) OUTPUT';
EXECUTE sp_executesql @columnValues, @columnValuesParams, @actualColumnValues OUTPUT;
--SELECT stuff(@actualColumnValues, 1,1, '')

declare @statement nvarchar(max)
set @statement =
'insert into [' + @tableName + '] (' + @columns + ')
select ' + stuff(@actualColumnValues,1,1,'')

print @statement

```

What it does is this:
It generates the `insert` statement and then it queries the actual data from the table and generates the `select` statement with that data. May not work correctly for some really complex datatypes but for varchars, datetimes and ints should work like a charm.


"
"How can I exclude SVN files from harvesting with heat (WiX)?
I hate to practically duplicate existing questions, but the supplied answers haven't worked:


Here's what my .wxs looks like:



```
<?xml version=""1.0"" encoding=""utf-8""?>
<Wix xmlns=""http://schemas.microsoft.com/wix/2006/wi"">
<Fragment>
<DirectoryRef Id=""SDKCONTENTDIR"">
<Directory Id=""dirE2EC21E8B765C611E918FB22F30721D1"" Name="".svn"" />
<Directory Id=""dir7DC42F44E7FE9E20277B180A353D0263"" Name=""bin"" />
</DirectoryRef>
</Fragment>
<Fragment>
<ComponentGroup Id=""sdkContent"">
<Component Id=""cmp5E86312F0CA2C53B8173AECD6A428747"" Directory=""dirE2EC21E8B765C611E918FB22F30721D1"" Guid=""{E87F312D-9DA2-4A68-B6C5-BCE2FF90720C}"">
<File Id=""filB766A28A7577EB4311FD03CD707BC211"" KeyPath=""yes"" Source=""$(var.publishContentDir)\.svn\all-wcprops"" />
</Component>
<Component Id=""cmp6EF52B3E331F226299060D45F533DC07"" Directory=""dirE2EC21E8B765C611E918FB22F30721D1"" Guid=""{5EA6AB2D-20C3-4B07-8E0A-7C28135BE922}"">
<File Id=""fil83205196F05211A66F9D25A7A5496FBA"" KeyPath=""yes"" Source=""$(var.publishContentDir)\.svn\entries"" />
</Component>

```

...


I'm using this .xsl code to exclude:



```
<xsl:key name=""svn-search"" match=""wix:Component[ancestor::wix:Directory/@Name = '.svn']"" use=""@Id"" />
<xsl:template match=""wix:Directory[@Name='.svn']"" />
<xsl:template match=""wix:Component[key('svn-search', @Id)]"" />

```

But I'm getting many ""Error 48 Unresolved reference to symbol"" errors as its not removing all child elements.


Ideas?


","Here is what I got working:



```
<?xml version=""1.0"" ?>
<xsl:stylesheet version=""1.0""
    xmlns:xsl=""http://www.w3.org/1999/XSL/Transform""
    xmlns:wix=""http://schemas.microsoft.com/wix/2006/wi"">

    <!-- Copy all attributes and elements to the output. -->
    <xsl:template match=""@*|*"">
        <xsl:copy>
            <xsl:apply-templates select=""@*"" />
            <xsl:apply-templates select=""*"" />
        </xsl:copy>
    </xsl:template>

    <xsl:output method=""xml"" indent=""yes"" />

    <!-- Create searches for the directories to remove. -->
    <xsl:key name=""svn-search"" match=""wix:Directory[@Name = '.svn']"" use=""@Id"" />
    <xsl:key name=""tmp-search"" match=""wix:Directory[@Name = 'tmp']"" use=""@Id"" />
    <xsl:key name=""prop-base-search"" match=""wix:Directory[@Name = 'prop-base']"" use=""@Id"" />
    <xsl:key name=""text-base-search"" match=""wix:Directory[@Name = 'text-base']"" use=""@Id"" />
    <xsl:key name=""props-search"" match=""wix:Directory[@Name = 'props']"" use=""@Id"" />

    <!-- Remove directories. -->
    <xsl:template match=""wix:Directory[@Name='.svn']"" />
    <xsl:template match=""wix:Directory[@Name='props']"" />
    <xsl:template match=""wix:Directory[@Name='tmp']"" />
    <xsl:template match=""wix:Directory[@Name='prop-base']"" />
    <xsl:template match=""wix:Directory[@Name='text-base']"" />

    <!-- Remove Components referencing those directories. -->
    <xsl:template match=""wix:Component[key('svn-search', @Directory)]"" />
    <xsl:template match=""wix:Component[key('props-search', @Directory)]"" />
    <xsl:template match=""wix:Component[key('tmp-search', @Directory)]"" />
    <xsl:template match=""wix:Component[key('prop-base-search', @Directory)]"" />
    <xsl:template match=""wix:Component[key('text-base-search', @Directory)]"" />

    <!-- Remove DirectoryRefs (and their parent Fragments) referencing those directories. -->
    <xsl:template match=""wix:Fragment[wix:DirectoryRef[key('svn-search', @Id)]]"" />
    <xsl:template match=""wix:Fragment[wix:DirectoryRef[key('props-search', @Id)]]"" />
    <xsl:template match=""wix:Fragment[wix:DirectoryRef[key('tmp-search', @Id)]]"" />
    <xsl:template match=""wix:Fragment[wix:DirectoryRef[key('prop-base-search', @Id)]]"" />
    <xsl:template match=""wix:Fragment[wix:DirectoryRef[key('text-base-search', @Id)]]"" />
</xsl:stylesheet>

```

"
"Pandas dataframe manipulating with long string column to 2 columns dynamically
I have dataframe:





| id | greetingsname |
| --- | --- |
| 1 | hello+peter|goodbye+john |
| 2 | hi+cheff|gutentag+rudolf|goodafternoon+Alex |



and i want





| id | greeting | name |
| --- | --- | --- |
| 1 | hello | peter |
| 1 | goodbye | john |
| 2 | hi | cheff |
| 2 | gutentag | rudolf |
| 2 | goodafternoon | Alex |



I dont know, how dynamically split column greetingsname, to get what I want, because column greetingsname has different string lengths. But the delimeter distribution remains the same `greeting DELIMETER(+) name DELIMETER(|) greeting DELIMETER(+) name`


And in this sense it can have different lengths (several names and greetings and in another column a different number of names and greetings)


Thx


","You can use explode for this:



```
df['greetingsname']=df['greetingsname'].apply(lambda x: x.split('|'))

res=df.explode('greetingsname')

print(res)

   id       greetingsname
0   1         hello+peter
0   1        goodbye+john
1   2            hi+cheff
1   2     gutentag+rudolf
1   2  goodafternoon+Alex

res[['greeting', 'name']] = res['greetingsname'].str.split('+', expand=True)

del res['greetingsname']

res.reset_index(drop=True, inplace=True)

print(res)

   id       greeting    name
0   1          hello   peter
1   1        goodbye    john
2   2             hi   cheff
3   2       gutentag  rudolf
4   2  goodafternoon    Alex

```

"
"Easy way to show a floating point integer without decimal positions in Haskell?
Let's suppose we have 2 numbers, 2.0 and 2.1. There is an easy way to show or convert them respectively to ""2"" and ""2.1""?


Sure I can do, for example, showFFloat (Just 0) 2 """", or showFFloat Nothing 2 """", but I use different arguments for each parameter, so the solution don't work for both numbers.


I can, for example, strip "".0"", if present, from the end of the string after I showed 2.0 and 2.1, and that works, but it isn't exactly simple. And many other solutions solve the problem, but I can't think about a simple one.


","Printing floating-point numbers is [surprisingly hard task](http://dl.acm.org/citation.cfm?id=93559&dl=ACM&coll=DL&CFID=610197641&CFTOKEN=10140114). The only implementation in `base` package I'm aware about is [floatToDigits](https://hackage.haskell.org/package/base-4.8.2.0/docs/Numeric.html#v:floatToDigits), though it returns digits instead of strings:



```
Prelude Numeric> floatToDigits 10 2.0
([2],1)
Prelude Numeric> floatToDigits 10 2.1
([2,1],1)

```

(The second component of the result is exponent.)


"
"What is the signal for when a widget loses focus?
In a dialog, when the `tab` key is pressed, the focus changes to another widget. In Qt, is there any signal for when a widget loses its focus? Can I use it to check if the input is valid or not? If not, can I set the focus back and ask the user to re-input?


","There's no signal but if you want to know when your widget has lost focus, override and reimplement `void QWidget::focusOutEvent(QFocusEvent* event)` in your widget. It will be called whenever your widget has lost focus. To give focus to a widget, use `QWidget::setFocus(Qt::FocusReason)`.


To validate input in a `QLineEdit` or `QComboBox` you can subclass `QValidator` and implement your own validator, or use one of the existing subclasses, `QIntValidator`, `QDoubleValidator`, or `QRegExpValidator`. Set the validator with `QLineEdit::setValidator(const QValidator*)` and `QComboBox::setValidator(const QValidator*)` respectively.


If you want to validate the contents of a modal dialog box, one way would be to override `QDialog::exec()` with an implementation like this:



```
int MyDialog::exec() {
  while (true) {
    if (QDialog::exec() == QDialog::Rejected) {
      return QDialog::Rejected;
    }
    if (validate()) {
      return QDialog::Accepted;
    }
  }
}

bool MyDialog::validate() {
  if (lineEdit->text().isEmpty()) {
    QMessageBox::critical(this, ""Invalid value"", ""The specified value is not valid"");
    lineEdit->setFocus();
    lineEdit->selectAll();
    return false;
  }
  return true;
}

```

It will not allow the user to close the dialog with the OK button or any other button with the Accepted role unless the contents of the dialog is successfully validated. In this example I assume the dialog has a `QLineEdit` named `lineEdit` and the `validate` function will make sure that its content is not empty. If it is, it will set the focus to the `QLineEdit` and show the dialog again.


"
"How to use receive POST data in asp.net web api?
I've googled a whole day but still can't find the answer. I need to `POST` data via `jQuery.post` to `Web API MVC-4` but unable to. This is my routing:



```
routes.MapRoute(
    name: ""Default"",
    url: ""{controller}/{action}/{id}"",
    defaults: new { controller = ""Home"", action = ""Index"", id = UrlParameter.Optional }
);

```

and this is my `Controller` (the `GET` works!):



```
    public string Get(int id)
    {
        return ""value"";
    }

    public void Post([FromBody]string data)
    {
        //body...
    }

```

This is the `jQuery.post:`



```
$.post('api/mycontroller', { key1: 'val1' });

```

Any idea ?


# Edit:


@Darin: I tried this:



```
public class UnitDetails{
    public string id { get; set; }
}

```

and:



```
public void Post(UnitDetails id) {
    //body...
}

```

and:



```
$.post('api/mycontroller', {id:'string1'});

```

But still I miss something.. it doesn't stop in `Post(...){...}`. Again - `Get(...){...}` does work.. ?


","This is by design and the only way to make this work with a primitive type such as a string is the following:



```
$.post('/api/mycontroller', '=' + encodeURIComponent('val1'));

```

So the body of the POST request must contain the following:



```
=val1

```

instead of:



```
data=val1

```

This has been discussed in [this thread](http://aspnetwebstack.codeplex.com/discussions/359687).


As an alternative you could define a view model:



```
public class MyViewModel
{
    public string Data { get; set; }
}

```

and then have your controller action take this view model as parameter:



```
public void Post(MyViewModel model)
{
    //body...
}

```

Contrary to primitive types, complex types use formatters instead of model binding. Here's [`an article`](http://blogs.msdn.com/b/jmstall/archive/2012/04/16/how-webapi-does-parameter-binding.aspx) which covers how does the Web API does parameter binding.


"
"C# Merging 2 dictionaries
I'm developing an app in C# targeting .NET 3.5. In it, I have 2 similar dictionaries that contain validation criteria for a specific set of elements in my app. Both dictionaries have identical signatures. The first dictionary has the default settings and the 2nd dictionary contains some user defined settings. 



```
var default_settings = new Dictionary<string, MyElementSettings>();
var custom_settings = new Dictionary<string, MyElementSettings>();

```

I would like to combine the 2 dictionaries into one that contains the elements of both dictionaries. 


The problem that I am running into is it is possible for both dictionaries to have the some of the same key values. The basic rule I want is to have a combination of both dictionary and if there are any keys in the custom\_settings that already exist in the default\_settings, the custom\_settings value will overwrite the default\_settings value. The best solution i have is just a foreach loop, check if the key exists in the other dictionary, and if not, add it.



```
foreach (var item in custom_settings)
{
    if (default_settings.ContainsKey(item.Key))
        default_settings[item.Key] = item.Value;
    else
        default_settings.Add(item.Key, item.Value);
}

```

I've done some basic LINQ queries, but I'm still working on learning the more advanced stuff. I've seen a few queries that will merge 2 dictionaries, but most involve grouping any element with duplicate keys, or only return a collection with just the duplicate keys/ Is there a LINQ query or expression that will mimic the behavior of the foreach loop I am using?


","Two points:


1. LINQ isn't great for executing side effects. In this case, you're attempting to mutate an existing collection rather than execute a query, so I would shy away from a pure LINQ solution.
2. The [setter on the generic dictionary's indexer](http://msdn.microsoft.com/en-us/library/9tee9ht2.aspx) already has the effect of adding the key-value pair if the key doesn't exist, or overwriting the value if it does.



> 
> When you set the property value, if
>  the key is in the Dictionary, the value associated with
>  that key is replaced by the assigned
>  value. If the key is not in the
>  Dictionary, the key and
>  value are added to the dictionary.
> 
> 
> 


So your `foreach` loop is essentially equivalent to:



```
foreach (var item in custom_settings)
{
   default_settings[item.Key] = item.Value;
}

```

Now that's pretty terse already, so I don't think LINQ is going to help you all that much.


"
"Searching for a value in an unknown NSMutableArray depth
Ok, i kind of asked the wrong question so I've edited the original question.


I'm storing Arrays within Arrays, as well as NSDictionaries. It's a utility kind of application and there is no set structure, the user can enter nested information as much as they require.


Ideally I need a method to scroll through the entire contents of my array given a set parameter (a type of class, maybe a dictionary key). Here's an example..



```
NSMutableArray *array = [[NSMutableArray alloc]init];

NSMutableDictionary *enteredItem = [[NSMutableDictionary alloc]init];

[enteredItem setObject:@""i'm a title""       forKey:@""title""];
[enteredItem setObject:@""i'm an id""         forKey:@""id""];
[enteredItem setObject:@""i'm a description"" forKey:@""description""];
[enteredItem setObject:@""i'm a timestamp""   forKey:@""timestamp""];
[enteredItem setObject:array                forKey:@""items""];


[array addObject:enteredItem];
[array addObject:anotherDictionary];
[array addObject:moreDictionaries];

```

**So in the example above, I would need to find the dictionary (and return it) that contains @""i'm an id"".**


Hopefully my question is clear. Thanks for any help you can offer.


","Recursive approach is correct, but I'm not sure the code samples were very helpful if you don't already know recursion. Here's a working solution:


Add these methods:



```
- (id)findObjectWithKey:(id)key inArray:(NSArray *)array
{
    for (id object in array)
    {
        if ([object isKindOfClass:[NSArray class]])
        {
            return [self findObjectWithKey:key inArray:object];
        }
        else if ([object isKindOfClass:[NSDictionary class]])
        {
            return [self findObjectWithKey:key inDictionary:object];
        }
    }
    return nil;
}

- (id)findObjectWithKey:(id)key inDictionary:(NSDictionary *)dict
{
    for (id subKey in dict)
    {
        id object = [dict objectForKey:subKey];
        if ([subKey isEqual:key])
        {
            return object;
        }
        else if ([object isKindOfClass:[NSArray class]])
        {
            return [self findObjectWithKey:key inArray:object];
        }
        else if ([object isKindOfClass:[NSDictionary class]])
        {
            return [self findObjectWithKey:key inDictionary:object];
        }
    }
    return nil;
}

```

Then to find your object, just say:



```
id object = [self findObjectForKey:@""title"" inArray:array];

```

To modify the methods to find a specific object and return the dictionary key, do this instead:



```
- (id)findKeyWithObject:(id)key inArray:(NSArray *)array
{
    for (id object in array)
    {
        if ([object isKindOfClass:[NSArray class]])
        {
            return [self findKeyWithObject:key inArray:object];
        }
        else if ([object isKindOfClass:[NSDictionary class]])
        {
            return [self findKeyWithObject:key inDictionary:object];
        }
    }
    return nil;
}

- (id)findKeyWithObject:(id)object inDictionary:(NSDictionary *)dict
{
    for (id key in dict)
    {
        id subObject = [dict objectForKey:key];
        if ([subObject isEqual:object])
        {
            return key;
        }
        else if ([subObject isKindOfClass:[NSArray class]])
        {
            return [self findKeyWithObject:object inArray:object];
        }
        else if ([subObject isKindOfClass:[NSDictionary class]])
        {
            return [self findKeyWithObject:object inDictionary:object];
        }
    }
    return nil;
}

```

Then to find your key, just say:



```
id key = [self findKeyWithObject:object inArray:array];

```

"
"How come `e.what()` prints ""bad allocation""?
The `new` expression in the `try` block throws a `bad_alloc`exception in my computer.


Note that the catch clause receives an exception object by value, not by reference. How come `e.what()` prints `""bad allocation""` ? I thought it would be sliced.



```
#include <iostream>

int main()
{
    try
    {
        int* p = new int[0x1F000000];
    }
    catch(std::exception e)
    {
        std::cout << e.what() << std::endl;
    }
}

```

","Visual Studio (Dinkumware?) uses an implementation of `std::exception` that contains internal storage† for the message. (Complete with a non-standard constructor that accepts a string.)


Because of this, no virtual dispatch is actually needed to get the error message, it survives any slicing.


A more orthodox implementation would indeed print a generic exception message, because the derived object was sliced off. (Effectively, MS has made `std::exception` and `std::runtime_error` equivalent. There's nothing wrong with this, since the return value of `std::exception::what` is implementation-defined, but it explains your results.)




---


†Internal storage here is used loosely. It doesn't have an internal *buffer*, but it has a `const char*` and a `bool`. The `const char*` points to the message (the return value of `what()`), and the `bool` is a flag determining if the buffer should be deleted.


It's like this:



```
class msvc_exception // for exposition
{
public:
    msvc_exception(const char* msg) :
    mMsg(msg),
    mDoDelete(false)
    {}

    msvc_exception(const std::string& msg) :
    mMsg(copy_string(msg)),
    mDoDelete(true)
    {}

    virtual ~msvc_exception()
    {
        if (mDoDelete)
            delete [] mMsg;
    }

    virtual const char* what() const throw()
    {
        return mMsg ? mMsg : ""unknown"";
    }

private:
    const char* copy_string(const std::string& str)
    {
        const char* result = new char[str.size() + 1];

        std::copy(str.begin(), str.end(), result);
        result[str.size()] = 0; // null-terminate

        return result;
    }
};

```

You see now that `bad_alloc` works like this:



```
    class msvc_bad_alloc : // for exposition
        public msvc_exception
    {
    public:
        msvc_bad_alloc() :
        msvc_exception(""bad_alloc"") // note: a static string, no dynamic storage
        {}
    };

```

Slicing doesn't affect the message because the message ""exists"" in the base class.


Other compilers, like GCC and LLVM, implement it a bit more straight-forwardly:



```
class orthodox_exception
{
public:
    orthodox_exception(){}
    virtual ~orthodox_exception() {}

    virtual const char* what() const throw()
    {
        return ""orthodox_exception"";
    }
};

class orthodox_bad_alloc :
    public orthodox_exception
{
public:
    const char* what() const throw()
    {
        return ""orthodox_bad_alloc"";
    }
};

```

Here, slicing would affect your outcome. (That said, after all this: always catch by reference.)


"
"What is difference between method process.exit(1) and process.exit(0) in node.js?
In node.js applications i saw usage of both these methods `process.exit(1)` and `process.exit(0)`. Can anybody give me the exact answer ?


","
> 
> Node normally exits with a 0 status code when no more async operations
>  are pending. There are other exit codes which are described below:
> 
> 
> **1 - Uncaught Fatal Exception**: There was an uncaught exception, and it was not handled by a domain or an uncaughtException event handler.
> 
> 
> **2 - Unused**: Reserved by Bash for built in misuse.
> 
> 
> **3 - Internal JavaScript Parse Error**: The JavaScript source code internal in Node's bootstrapping process caused a parse error. This is
>  extremely rare, and generally can only happen during the development
>  of Node itself.
> 
> 
> **4 - Internal JavaScript Evaluation Failure**: The JavaScript source code internal in Node's bootstrapping process failed to return a
>  function value when evaluated. This is extremely rare, and generally
>  can only happen during the development of Node itself.
> 
> 
> **5 - Fatal Error**: There was a fatal unrecoverable error in V8. Typically, a message will be printed to stderr with the prefix FATAL
>  ERROR.
> 
> 
> **6 - Non-function Internal Exception Handler**: There was an uncaught exception, but the internal fatal exception handler function was
>  somehow set to a non-function, and could not be called.
> 
> 
> **7 - Internal Exception Handler Run-Time Failure**: There was an uncaught exception, and the internal fatal exception handler function
>  itself threw an error while attempting to handle it.
> 
> 
> **8 - Unused**
> 
> 
> **9 - Invalid Argument**: Either an unknown option was specified, or an option requiring a value was provided without a value.
> 
> 
> **10 - Internal JavaScript Run-Time Failure**: The JavaScript source code internal in Node's bootstrapping process threw an error when the
>  bootstrapping function was called. This is extremely rare, and
>  generally can only happen during the development of Node itself.
> 
> 
> **11 - Invalid Debug Argument**: The `--debug` and/or `--debug-brk` options were set, but an invalid port number was chosen
> 
> 
> **>128 - Signal Exits**: If Node receives a fatal signal such as `SIGKILL` or `SIGHUP`, then its exit code will be 128 plus the value
>  of the signal code. This is a standard Unix practice, since exit codes
>  are defined to be 7-bit integers, and signal exits set the high-order
>  bit, and then contain the value of the signal code.
> 
> 
> 


**Source:** <https://www.tutorialspoint.com/nodejs/nodejs_process.htm>


"
"Return ALL available IP addresses for an Android device?
I have the need to list all available IP addresses for any particular Android Device.


I have found some sample code but this only results in returning one IP address, which happens to be an IPv6 address. I need to get all available IPs for any particular device. I do the same on an iOS version of this app and it returns 3 IPv6 addresses, one `192.` address and one `10.` address. I am trying to replicate the same on Android. I pass all values to an Array and display them in a list.


The code I have in place is:



```
public String getLocalIpAddress()
{
    try {
        for (Enumeration<NetworkInterface> en = NetworkInterface.getNetworkInterfaces(); en.hasMoreElements();) {
            NetworkInterface intf = en.nextElement();
            for (Enumeration<InetAddress> enumIpAddr = intf.getInetAddresses(); enumIpAddr.hasMoreElements();) {
                InetAddress inetAddress = enumIpAddr.nextElement();
                if (!inetAddress.isLoopbackAddress()) {
                    IPAddresses.setText(inetAddress.getHostAddress().toString());
                    return inetAddress.getHostAddress().toString();
                }
            }
         }
    } catch (SocketException ex) {
        String LOG_TAG = null;
        Log.e(LOG_TAG, ex.toString());
    }

    return null;
}

```

","It looks to me like your code is just returning the first match - is that not the problem? I would have expected you to build up list of addresses, and return that instead of just a single string. Something like this:



```
public String[] getLocalIpAddress()
{          
    ArrayList<String> addresses = new ArrayList<String>();
    try {
        for (Enumeration<NetworkInterface> en = NetworkInterface.getNetworkInterfaces(); en.hasMoreElements();) {
            NetworkInterface intf = en.nextElement();
            for (Enumeration<InetAddress> enumIpAddr = intf.getInetAddresses(); enumIpAddr.hasMoreElements();) {
                InetAddress inetAddress = enumIpAddr.nextElement();
                if (!inetAddress.isLoopbackAddress()) {
                    IPAddresses.setText(inetAddress.getHostAddress().toString());
                    addresses.add(inetAddress.getHostAddress().toString());
                }
             }
         }
     } catch (SocketException ex) {
         String LOG_TAG = null;
         Log.e(LOG_TAG, ex.toString());
     }
     return addresses.toArray(new String[0]);
}

```

I'm not sure what the `IPAddresses.setText` call is doing, so I've left it in, but I expect that would also need to be adjusted in some way to handle the fact that you could have multiple addresses being matched.


"
"How secure is JWT
I know this question is not anything new, possibly already been discuss all over the internet.


I'm new to it but after some study, I agree that it is safe as anonymous could sniff the Token but unable to append anything on it. I'm planning to store JWT in HTML5Storage, and decode the payload for some in-sensitive information: DisplayName, email\_address, and role\_info and etc. 


Here is my question, could anonymous sniff my JWT Token and act on-behalf of me?
If that is possible, how can I avoid that?


","In short, JWT by itself is not safe it's just clear text. JWT in basic terms is a standard protocol for how information is defined otherwise known as claims passed between to parties. In combination with JWS (signature) and JWE (encyption) will make it secure. The over arching topic for this is JOSE - Javascript Object Signing and Encryption. Aside from reading the RFC themselves which you should refer to as well there are ton of information online without stating the obvious nor insulting your search abilities. Check out <http://jose.readthedocs.io/en/latest/> (includes reference links to RFCs)


So to answer your question if signed and encrypted with industry secure standards in flight (over the wire) and at rest (ie: database) yes it's secure.


In terms of spoofing as you need to pre-cautions in areas of session jacking and/or token jacking and prevention against things like rainbow tables. 


JOSE really is just a standard no matter what standard, what you are really asking is about best practices for security measures, you should refer to OWASP as well <https://www.owasp.org/> 


Does that help clarify your question?


"
"Simple linear regression fit manually via matrix equations does not match lm() output
I am trying to fit a linear model using matrices to my data set even though I can use OLS and do it without matrices as a simple tutorial for myself to better understand both `R` and matrix notation. 


This is the model I am trying to fit:


$$\bf Y=X\boldsymbol\beta+\varepsilon$$


where $\bf Y$ is a $1\times n$ matrix, $\bf X$ is a $n\times k$ matrix (where $k$ is the number of $\beta$'s, which in this case is 2), $\boldsymbol \beta$ is a $k\times 1$ matrix and lastly our error term is $n\times 1$. I understand this portion. 


When I simply use the `lm()` command to fit my data, I get the following from the `summary()` command:



```
Call:
lm(formula = y ~ x)

Residuals:
    Min      1Q  Median      3Q     Max 
-4.0503 -1.4390  0.4921  1.0589  3.9446 

Coefficients:
            Estimate Std. Error t value Pr(>|t|)    
(Intercept)   2.9849     0.8219   3.632  0.00191 ** 
x             0.5612     0.1084   5.178 6.32e-05 ***

```

So the `summary()` is telling me that the $\beta$ matrix is a $2\times 1$ matrix, with the first number (which is $\beta\_0$) as 2.0949 and the second number (which is $\beta\_1)$ as 0.1084. My question is this:


We know that the matrix $\beta$ is actually:


$$\boldsymbol\beta=(\bf{X}^T\bf{X})^{-1}(\bf{X}^T\bf{Y})$$


and when I try to simply carry out this calculate by hand using R using `b=(t(x)*x)^-1*t(x)*y`, I get a $1\times 20$ vector (where $20$ of course is $n$, the number of observations). Why am I not getting a $2\times 1$ matrix like I should be getting?


","You've made two mistakes in your R code for `b`.


1. `solve` is used for matrix inversion. Raising `X` to the $-1$ power inverts each element of `X`, which can occasionally be useful, but is not what we want here.
2. R uses the operator `%*%` for matrix multiplication. Otherwise, it does element-wise multiplication and requires your arrays to be conformable according to R's handling of vectors. Again, occasionally useful.


This is all explained in more detail in the R documentation, or the nigh-uncountable intro to R materials online, such as [this one](http://www.statmethods.net/stats/regression.html).


`b<-solve(t(X)%*%X)%*%t(X)%*%y` is the literal representation of the normal equations, while `b<-solve(crossprod(X), crossprod(X,y))` is faster and more idiomatic, and matches the output of `lm()`. But you definitely don't want to explicitly compute the normal equations directly for numerical reasons (and R won't warn you until you lose all significant figures). The `lm` method uses QR decomposition by default, which is more numerically stable. If you apply the correct ""hand cranked"" computation and it still differs from the R output, it's plausibly because of loss of numerical precision due to the explicit inversion of $X^TX$.


Don't forget to add a column of $1$ for the intercept!


"
"Django Channel Custom Authentication Middleware \_\_call\_\_() missing 2 required positional arguments: 'receive' and 'send'
I am writing a custom authentication middleware for django channels



```
class TokenAuthMiddleware:
    def __init__(self, inner):
        # Store the ASGI application we were passed
        self.inner = inner

    def __call__(self, scope):

        return TokenAuthMiddlewareInstance(scope, self)


class TokenAuthMiddlewareInstance:

    def __init__(self, scope, middleware):
        self.middleware = middleware
        self.scope = dict(scope)
        self.inner = self.middleware.inner

    async def __call__(self, receive, send):
        ## my logic to get validate user and store the user in user data
        ...
        ...
        ...
        self.scope['user'] = user_data
        inner = self.inner(self.scope)
        return await inner(receive, send)

```

but on trying to connect to web socket from front end I get the following error



```
TypeError: __call__() missing 2 required positional arguments: 'receive' and 'send'

```

","For your reference: <https://channels.readthedocs.io/en/stable/releases/3.0.0.html>


change from in routing.py



```
websocket_urlpatterns = [
    re_path(r'ws/chat/(?P<room_name>\w+)/$', consumers.ChatConsumer),
]

```

to
Consumers now have an as\_asgi() class method you need to call when setting up your routing:



```
websocket_urlpatterns = [
    re_path(r'ws/chat/(?P<room_name>\w+)/$', consumers.ChatConsumer.as_asgi()),
]

```

then
if you need custom authentication <https://channels.readthedocs.io/en/stable/topics/authentication.html#custom-authentication>



```
from channels.auth import AuthMiddlewareStack
from channels.db import database_sync_to_async
from django.contrib.auth import get_user_model

User = get_user_model()

@database_sync_to_async
def get_user(user_id):
    try:
        return User.objects.get(id=user_id)
    except User.DoesNotExist:
        return AnonymousUser()
        
class QueryAuthMiddleware:
    """"""
    Custom middleware (insecure) that takes user IDs from the query string.
    """"""
        
    def __init__(self, app):
        # Store the ASGI application we were passed
        self.app = app
        
    async def __call__(self, scope, receive, send):
        # Look up user from query string (you should also do things like
        # checking if it is a valid user ID, or if scope[""user""] is already
        # populated).
        scope['user'] = await get_user(int(scope[""query_string""]))
        
        return await self.app(scope, receive, send)
TokenAuthMiddlewareStack = lambda inner: QueryAuthMiddleware(AuthMiddlewareStack(inner))   

```


> 
> use requirements.txt as following list, and also download package in this order
> 
> 
> 



```
Django==3.0.8
djangorestframework==3.11.0
websocket-client==0.57.0
redis==3.5.3
asgiref==3.2.10
channels-redis==2.4.2
channels==3.0.1

```

"
"Symbol stripping of Objective-C code still leaves method names, etc. in binary
I am trying to strip our iOS application binary of debug symbols to make it more difficult for hackers to modify the binary. I have tried both xcode's symbol stripping (enabling strip linked product and deployment postprocessing) and using ""strip -S -x"". Both do reduce the number of symbols, but running the binary through ""strings"" still returns loads of hits.


How do I remove them?


","Objective-C is a dynamic language. Method calls are resolved at runtime based on selectors (effectively the method name as a string). This is different from a language like C++ that binds method calls at compilation/link time.


Removing the method names (selectors) from the binary would make the application unusable. 


Applications written in Objective-C are pretty much open books when it comes to their internals. Just look at tools like [`otool`](https://developer.apple.com/library/mac/#documentation/Darwin/Reference/ManPages/man1/otool.1.html), [`class-dump`](http://www.codethecode.com/projects/class-dump/), [F-Script](http://www.fscript.org) or [DTrace](http://www.dtrace.org) to see how much is accessible and modifiable in a running Objective-C application.


However, there are linker flags (`P_LNOATTACH`) which stop DTrace connecting to the running application. You can also call `ptrace` with the `PT_DENY_ATTACH` flag. iTunes is an example of an app that does this as Apple doesn't want you poking around inside their DRM.


There appears to be a previous post on Objective-C code obfuscation. See [iPhone/iPad App Code Obfuscation - Is it Possible? Worth it?](https://stackoverflow.com/questions/5556849/iphone-ipad-app-code-obfuscation-is-it-possible-worth-it) for more details


"
"Why don't my udev rules work inside of a running docker container?
I have udev rules written to create SYMLINKS when a device is connected. The rules are working fine on the host machine, but when I start a container with these same rules installed in /etc/udev/rules.d, they don't work inside my container.


I am trying to detect when an external drive is plugged in and create a corresponding symlink. /dev/sdX is created on boot of the container iff the drive was present at time of `docker run`, but it won't appear after the run command, nor will it disappear upon the drive being removed.


Here's an example rule that works just fine on the host machine:


`KERNEL==""sd?"", SYMLINK+=""test_%k""`


","I figured it out. What I've seen done on the internet is to mount the host's `/dev` inside the container:


`docker run -v=/dev:/dev` *(Note: not safe)*


But this is very dangerous and pretty much destroys the host computer by mucking with permissions (e.g. psuedo-terminals can't be spawned).


However, if **I set up a udev rule on the host machine to create devices in a unique subdirectory, like `/dev/foo/sdX`**, I can then just share `dev/foo` with my container:


**`docker run -v=/dev/foo:/dev/foo`**


Now, when I insert a drive that matches my udev rule, the host machine creates a symlink in `/dev/foo/sdX`, which is now suddenly visible to my container. When the drive is removed, `/dev/foo/sdX` also disappears.




---


The one missing feature that would be nice is the ability to trigger a script inside the container when the device is created. A udev rule can do that on the host machine, but no udev rules seem to be tripped inside the container. So manual polling it is, for now.


"
"algorithm for ai to chase player in straight line?
I am making a game in libgdx(java) where squares chase the main player. I can't figure out how to make the squares move in a straight line directly to the player at a constant speed. this means i want the squares to travel the same distance every second. I make a variable and set it equal to `chasePattern1`. `saidvariable[0]` is added to the enemies `x` value, and `saidvariable[1]` is added to the enemies `y` value.



```
private float[] chasePattern1(float enemy_x,float enemy_y,float speed){
    float[] returnvalue={0,0};

    if(enemy_x>mainsquare.getX()){
        returnvalue[0]=-1*speed;
    }
    if(enemy_x<mainsquare.getX()){
        returnvalue[0]=speed;///does float work?no
    }
    if(enemy_y>mousey){
        returnvalue[1]=-1*speed;
    }
    if(enemy_y<mousey){
        returnvalue[1]=speed;
    }
    return returnvalue;
}

```

","if `d` is the distance between the player and a square, `v` is the velocity that you want the squares to move at, `(x1,y1)` is the position of the player and `(x0,y0)` is the position of the square, then



```
vx = (x1 - x0) * v / d
vy = (y1 - y0) * v / d

```

where `vx` and `vy` are the values that you need to add to the square's coordinates every second.


### Edit:


You can view it like this: If you want to travel from your position to the destination `d` ""points"" with a speed of `v`, it would take you `t=d/v` moves.


Now, the distance you have to travel horizontally is `x1-x0` and the distance vertically is `y1-y0`, if you want to travel both distances in `t` moves, you need to divide each distance in `t` pieces and that's the speed for each component (`x` and `y`). When you divide `x1-x0` for `t`, you get that value, same for `y`


"
"Where is Import Table in ELF file?
I found "".dynsym"" in String Table, got index. 
Then I found section with sh\_name = index && sh\_type = SHT\_DYNSYM.
So I got sh\_offset = 464 and sh\_size = 64. 
But you can see in the attached picture, that on the offset 464 there are only zeros. 


I suppose that Import Table starts on offset 528. Question is: how calculate it %)



> 
> ![enter image description here](https://i.stack.imgur.com/HAfR2.png)
> 
> 
> 


","
> 
> But you can see in the attached picture, that on the offset 464 there are only zeros.
> 
> 
> 


Wrong: `01`, `20`, `29`, `12` etc. are *not* ""only zeros"" last time I checked.



> 
> I suppose that Import Table starts on offset 528
> 
> 
> 


No, it does not. For some reason you are expecting to find a Microsoft `PE`-style import table in an ELF file. It's not there.


An *equivalent* of an import table in ELF is contained in *two* tables. One contains `Elf{32,64}_Sym` fixed-size records:



```
typedef struct
{
  Elf32_Word    st_name;                /* Symbol name (string tbl index) */
  Elf32_Addr    st_value;               /* Symbol value */
  Elf32_Word    st_size;                /* Symbol size */
  unsigned char st_info;                /* Symbol type and binding */
  unsigned char st_other;               /* Symbol visibility */
  Elf32_Section st_shndx;               /* Section index */
} Elf32_Sym;

```

and is contained in the `.dynsym` section.


The other table is contained in `.dynstr` section (which, in your file starts at offset 528), and has just the (variable-size) strings separated by `NUL` character.


The `.st_name` in the first table refers to offset in `.dynstr`.


"
"Kusto: How can I get the value from the column that doesn't participate in a SUMMARIZE?
Having the below table and Kusto query, how can I obtain a result that has the Purchase column?



```
let ProductsTable = datatable(Supplier: string, Fruit: string, Price: int, Purchase: datetime)
[
    'Contoso', 'Grapes', 220, datetime(2018-10-01 01:00),
    'Fabrikam', 'Lemons', 31, datetime(2018-10-01 02:00),
    'Contoso', 'Lemons', 29, datetime(2018-10-02 03:00),
    'Contoso', 'Grapes', 210, datetime(2018-10-02 04:00),
    'Fabrikam', 'Lemons', 30, datetime(2018-10-03 05:00),
    'Contoso', 'Bananas', 12, datetime(2018-10-03 06:00),
    'Contoso', 'Bananas', 12, datetime(2018-10-04 07:00),
    'Contoso', 'Lemons', 29, datetime(2018-10-04 08:00),
    'Contoso', 'Grapes', 200, datetime(2018-10-05 09:00),
];
ProductsTable
    | summarize Price = min(Price) by Supplier, Fruit
    | order by Supplier asc, Fruit asc, Price asc

```

Result



```
Contoso Bananas 12
Contoso Grapes      200
Contoso Lemons      29
Fabrikam    Lemons      30

```

DESIRED Result



```
Contoso Bananas 12  2018-10-03 06:00
Contoso Grapes      200 2018-10-05 09:00
Contoso Lemons      29  2018-10-02 03:00
Fabrikam    Lemons      30  2018-10-03 05:00

```

I know that there could be multiple results, for example for ***Contoso-Bananas-12*** we can have any of the following


- 2018-10-03 0**6**:00
- 2018-10-04 0**7**:00


","try using `arg_min()`: <https://learn.microsoft.com/en-us/azure/kusto/query/arg-min-aggfunction>



```
let ProductsTable = datatable(Supplier: string, Fruit: string, Price: int, Purchase: datetime)
[
    'Contoso', 'Grapes', 220, datetime(2018-10-01 01:00),
    'Fabrikam', 'Lemons', 31, datetime(2018-10-01 02:00),
    'Contoso', 'Lemons', 29, datetime(2018-10-02 03:00),
    'Contoso', 'Grapes', 210, datetime(2018-10-02 04:00),
    'Fabrikam', 'Lemons', 30, datetime(2018-10-03 05:00),
    'Contoso', 'Bananas', 12, datetime(2018-10-03 06:00),
    'Contoso', 'Bananas', 12, datetime(2018-10-04 07:00),
    'Contoso', 'Lemons', 29, datetime(2018-10-04 08:00),
    'Contoso', 'Grapes', 200, datetime(2018-10-05 09:00),
];
ProductsTable
| summarize Price = arg_min(Price, *) by Supplier, Fruit
| order by Supplier asc, Fruit asc, Price asc

```

"
"Slurm: Why use srun inside sbatch?
In a sbatch script, you can directly launch programs or scripts (for example an executable file `myapp`) but in many tutorials people use `srun myapp` instead.


Despite reading some documentation on the topic, **I do not understand the difference and when to use each of those syntaxes**.


I hope this question is precise enough (1st question on SO), thanks in advance for your answers.


","The `srun` command is used to create job 'steps'. 


First, it will bring **better reporting** of the resource usage ; the [sstat](https://slurm.schedmd.com/sstat.html) command will provide real-time resource usage for processes that are started with `srun`, and each step (each call to srun) will be reported individually in the accounting.


Second, it can be used to **setup many instances of a serial program** (program that only use one CPU) into a single job, and micro-schedule those programs inside the job allocation.


Finally, for parallel jobs, `srun` will also play the important role **of starting the parallel program** and setup the parallel environment. It will start as many instances of the program as were requested with the `--ntasks` option on the CPUs that were allocated for the job. In the case of a [MPI](https://en.wikipedia.org/wiki/Message_Passing_Interface) program, it will also handle the communication between the MPI library and Slurm.


"
"Segmentation fault when jumping to goto over VLA array
The following example demonstrates the issue:



```
#include <cstdio>

int main()
{
        unsigned int remaining=1;

        goto loop;

        while(remaining) {
                unsigned char tmp[remaining];
                printf(""&tmp: %p\n"",tmp);
loop:
                remaining = 512;//or something else;
        }
}

```

Initially, the initialization of ""remaining"" variable was a bit long and I used `goto` to initialize it on one line. However, now this example gives segmentation fault on the `printf` line.


It looks like the array is not initialized properly.


Even gdb cannot print the tmp array's address:



```
Program received signal SIGSEGV, Segmentation fault.
0x00000000004005b8 in main () at test.cpp:11
11          printf(""&tmp: %p\n"",tmp);
(gdb) p tmp
$1 = 0xfffffffffffffe00 <error: Cannot access memory at address 0xfffffffffffffe00>

```

My gcc version:



```
gcc (Ubuntu 4.8.2-19ubuntu1) 4.8.2

```

compiling with:



```
g++ -o testc test.cpp

```

If I remove the goto, or replace the variadic array with a fixed array, the segmentation fault is gone.
What is happening actually?


Is this a gcc bug? If the combination of `goto` and variadic arrays is not allowed, there should be a warning?


","Variable length arrays(*VLA*) are a C99 feature that gcc supports as [an extension in C++](https://gcc.gnu.org/onlinedocs/gcc/Variable-Length.html) and in C99 a jump across a VLA declaration is undefined behavior, from the draft C99 standard section `6.8.6.1` *The goto statement*:



> 
> A goto statement shall not jump from outside the scope of an identifier having
>  a variably modified type to inside the scope of that identifier.
> 
> 
> 


`clang` and `gcc 4.9` actually makes this an error and says:



```
error: goto into protected scope
    goto loop;
    ^

note: jump bypasses initialization of variable length array
            unsigned char tmp[remaining];
                          ^

```

See the [gcc bug report: Jumps into VLA or VM scope not rejected for C++](https://gcc.gnu.org/bugzilla/show_bug.cgi?id=21113).


The rules for jumping past a declaration of an automatic variable in C++ is covered in section `6.7` *[stmt.dcl]* which says:



> 
> It is possible to transfer into a block, but not in a way that
>  bypasses declarations with initialization. A program that jumps87 from
>  a point where a variable with automatic storage duration is not in
>  scope to a point where it is in scope is ill-formed unless the
>  variable has scalar type, class type with a trivial default
>  constructor and a trivial destructor, a cv-qualified version of one of
>  these types, or an array of one of the preceding types and is declared
>  without an initializer (8.5). [ Example:
> 
> 
> 
> ```
> void f() {
>     // ...
>     goto lx; // ill-formed: jump into scope of a
> ly:
>     X a = 1;
>     // ...
> lx:
>     goto ly; // OK, jump implies destructor
>              // call for a followed by construction
>             // again immediately following label ly
> }
> 
> ```
> 
> —end example ]
> 
> 
> 


"
"How to divide an array in to two different arrays in Perl
I have an Array of Hash, and I need to divide it into two different arrays. The array of Hash looks like this:



```
 old_array = [
   {
      'a'=>1,
      'b'=>2,
      'c'=>3,
      'd'=>4
   },
   {  
      'e'=>5,
      'f'=>6,
      'g'=>7,
      'h'=>8
   },
  {  
      'i'=>9,
      'j'=>10,
      'k'=>11,
      'l'=>12
   },
  {  
      'm'=>13,
      'n'=>14,
      'o'=>15,
      'p'=>16
   }

];

```

I need to divide it into two different arrays in the manner:



```
   new_array_1 = [{
      'a'=>1,
      'b'=>2,
      'c'=>3,
      'd'=>4
   },
   {  
      'e'=>5,
      'f'=>6,
      'g'=>7,
      'h'=>8
   }
 ];


 new_array_2 = [{  
      'i'=>9,
      'j'=>10,
      'k'=>11,
      'l'=>12
   },
  {  
      'm'=>13,
      'n'=>14,
      'o'=>15,
      'p'=>16
   }
 ];

```

I have now tried the below code as mentioned in one of the answers:



```
     my ($new_array_1, $new_array_2) =
    map { [ $_ ] }
          @$old_array;

```

This is splitting the array but removing the middle hashes and, I am getting output as:



```
   $VAR1 = [
      {
        'c' => 3,
        'a' => 1,
        'b' => 2,
        'd' => 4
      }
    ];

$VAR1 = [
      {
        'g' => 7,
        'f' => 6,
        'e' => 5,
        'h' => 8
      }
    ];

```

What shall I do to get the desired output? The hashes may increase to 10 or 20 or 15, etc.


","*The question has changed significantly, where now the arrayref has not two elements but more, and it need be split so that its halves go into two arrays. Elements need not be made arrayrefs.*


Now one has to go to indices. (Well, or to alternate between new arrays while moving/copying element by element.) One way



```
my $midind = ($#$old_arrayref-1)/2;  # or: (@$oldarrayref/2)-1

my $new_arrayref_1 = [ @$old_arrayref[0..$midind] ];
my $new_arrayref_2 = [ @$old_arrayref[$midind+1 .. $#$old_array] ];

```

This preserves the original array as it was. (The syntax `$#$arrayref` is for the last index in the arrayref `$arrayref`.) If the old array doesn't matter, can do instead



```
my $new_arrayref_1 = [ splice @$old_arrayref, 0, $midind+1 ];
my $new_arrayref_2 = [ @$old_arrayref ];

```

**NOTE** If elements of the original array themselves contain complex data structures (so, references) then they need be ""deep-copied,"" for example using `Storable::dclone`.


I still recommend working through resources linked in the original post (below), to learn how to work with arrays in this language.




---


*The original answer. The question before it was changed had the initial array(ref) with strictly two elements, and once th ey are split into two arrays they (seemingly) needed to be wrapped into arrayrefs. This should still be useful for the list of basic resources (if the reader needs that).*


There is a range of operations supported on arrays in Perl. As for basics, we can remove and return an element from front or back, with [shift](https://perldoc.perl.org/functions/shift) or `pop`, and can add with `unshift` or `push`.


Then there is more, including `splice` you tried, and then there are libraries for more complex operations. But the basics readily solve your quest. One way



```
my $old_arrayref = [ { a=>1, b=>2 }, { c=>3, d=>4 } ];

my @new_array_1 = shift @$old_array;
my @new_array_2 = shift @$old_array;

```

(An [array](https://perldoc.perl.org/perlintro#Arrays) can be simply assigned a list of elements, including just one. But if there is anything in the array then that would be all gone after such an assignment so don't forget that.)


Here I take the source (`old_array`) to be an array-reference, following the question. Now that `$old_array` is empty, if it had only those two elements, which are now in `@new_array_N`'s. †


Another, even more elemental way, is to [index into the array](https://perldoc.perl.org/perldata#Subscripts) to access individual elements and copy them to new arrays. This way `$old_array` would stay as it is.


Perhaps a more reasonable arrangement is to have references for these two new arrays as elements of an array, which may be what your code attempts to do. For that you only need †



```
my @new_array = map { [ $_ ] } @$old_array;

```

The [map](https://perldoc.perl.org/functions/map) applies the code in the block to each element of its input (`@$old_array`), returning a list, assigned to `@new_array`. That `[]` in the body constructs an [anonymous array (reference)](https://perldoc.perl.org/perlreftut#Make-Rule-2), which is a scalar and as such can be placed on an array. The only element placed in this array-reference is the currently processed element of `@$old_array`, available in the (ubiquitous) [`$_` variable](https://perldoc.perl.org/perlvar#$_).


See `perlintro` and `perldata` linked above, and links in `shift`. For many additional operations on lists and arrays, provided in libraries, see [List::Util](https://perldoc.perl.org/List::Util) and [List::MoreUtils](https://metacpan.org/pod/List::MoreUtils) for starters.


For references see [perlreftut](https://perldoc.perl.org/perlreftut) and for arrays with references see [perllol](https://perldoc.perl.org/perllol) (lists-of-lists) and [perldsc](https://perldoc.perl.org/perldsc) (data structures cookbook).


A complete program†



```
use warnings;
use strict;
use feature 'say';
use Data::Dumper;

my @old_array = ( { 'a'=>1, 'b'=>2 }, { 'c'=>3, 'd'=>4 } );  

my @new_array = map { [ $_ ] } @old_array;

say Dumper \@new_array;

```

This prints (just condensed, for readability in this post)



```

$VAR1 = [
          [ { 'a' => 1, 'b' => 2 } ],
          [ { 'c' => 3, 'd' => 4 } ]
        ];

```



---


† If the new array(s) need be array-references as well



```
my $new_arrayref_1 = [ shift @$old_array ];
my $new_arrayref_2 = [ shift @$old_array ];

```

In the second example, and in the short ""complete program""



```
my $new_arrayref = [ map { [ $_ ] } @$old_array ];

```

"
"Encoding ExistentialQuantification with RankNTypes
I've read in a few places claims that equivalent functionality to `ExistentialQuantification` can be had using `RankNTypes`. Could someone provide an example of why this is or is not possible?


","Normally, all type variables in Haskell are implicitly universally quantified at the outermost scope of the type. `RankNTypes` allows a universal quantifier `forall` to appear nested, e.g. the type `forall a b. (a -> a) -> b -> b` is very different from `forall b. (forall a. a -> a) -> b -> b`.


There is a sense in which types on the left side of a function arrow are logically ""negated"", in roughly the same sense that `(->)` is logical implication. Logically, the universal and existential quantifiers are related by a De Morgan duality: (∃x. P(x)) is equivalent to ¬(∀x. ¬P(x)), or in other words ""there exists an x such that P(x)"" corresponds to ""it is not the case that, for all x, P(x) is false"".


So a forall on the left of a function arrow is ""negated"" and behaves like an existential. If you put the whole thing to the left of another function arrow it's double-negated and behaves as a universal quantifier again, modulo some fiddly details.


The same idea of negation applies to values as well, so to encode the type `exists x. x` we want: 


1. `forall x.` in contravariant (negated) position
2. a value of that type `x` in *covariant* (positive) position.


Since the value must be inside the scope of the quantifier, our only choice is double-negation--a CPS transform, basically. To avoid restricting things otherwise, we'll then universally quantify over the type on the right of the function arrows. So `exists x. x` is translated to `forall r. (forall x. x -> r) -> r`. Compare the placement of the types and quantifiers here to the requirements above to verify that it meets the requirement.


In more operational terms this just means that given a function with the above type, because we give it a function with a universally quantified argument type, it can apply that function to any type `x` it likes, and since it has no other way of getting a value of type `r` we know it will apply that function to *something*. So `x` will refer to *some* type, but we don't know what--which is basically the essence of existential quantification.


In more practical, day to day terms, any universally quantified type variable can be regarded as existential if you're looking at it from the ""other side"" of a function type. Because the unification performed as part of type inference transcends quantifier scope, you can sometimes end up in a situation where GHC would have to unify a type variable in an outer scope with a quantified type from a nested scope, which is how you get compiler errors about escaping types and skolems and whatnot, the latter (I assume) being related to [Skolem normal form](http://en.wikipedia.org/wiki/Skolem_normal_form).


The way this relates to data types using existentials is that while you can declare a type like this:



```
data Exists = forall a. Exists a

```

That represents an existential type, to get at the ""existential type"" you need to unwrap it by pattern matching:



```
unexist :: Exists -> r
unexist (Exists x) = foo x

```

But if you consider what the type of `foo` would have to be in this definition, you end up with something like `forall a. a -> r`, which is equivalent to the CPS-style encoding above. There's a close relationship between CPS transforms and Church encoding of data types, so the CPS form can also be seen as a reified version of the pattern match.


Finally, relating things back to logic--since that's where the term ""existential quantifier"" comes from--note that, if you think of being left of an arrow as negation and kinda squint to ignore the `forall r. ...` CPS cruft, these encodings of existential types are exactly the same thing as the De Morgan dualized form ¬(∀x. ¬P(x)) that was the starting point. So these really all are just different ways of looking at the same concept.


"
"How to write multiple conditions of if-statement in Robot Framework
I have trouble writing `if` conditions in Robot Framework.


I want to execute



```
Run Keyword If '${color}' == 'Red' OR '${color}' == 'Blue'  OR '${color}' == 'Pink'    Check the quantity

```

I can use this ""`Run keyword If`"" keyword with one condition, but for more than one conditions, I got this error:



> 
> FAIL: Keyword name cannot be empty.
> 
> 
> 


And also I would like to use these keywords:



```
Run Keyword If '${color} == 'Blue' AND '${Size} == 'Small' AND '${Design}' != '${Simple}'      Check the quantity

```

And



```
Run Keyword Unless '${color}' == 'Black' OR '${Size}' == 'Small' OR '${Design}' == 'Simple'

```

But I just end up getting errors.


","You should use small caps ""or"" and ""and"" instead of OR and AND.


And beware also the spaces/tabs between keywords and arguments (you need at least two spaces).


Here is a code sample with your three keywords working fine:


Here is the file `ts.txt`:



```
  *** test cases ***
  mytest
    ${color} =  set variable  Red
    Run Keyword If  '${color}' == 'Red'  log to console  \nexecuted with single condition
    Run Keyword If  '${color}' == 'Red' or '${color}' == 'Blue' or '${color}' == 'Pink'  log to console  \nexecuted with multiple or

    ${color} =  set variable  Blue
    ${Size} =  set variable  Small
    ${Simple} =  set variable  Simple
    ${Design} =  set variable  Simple
    Run Keyword If  '${color}' == 'Blue' and '${Size}' == 'Small' and '${Design}' != '${Simple}'  log to console  \nexecuted with multiple and

    ${Size} =  set variable  XL
    ${Design} =  set variable  Complicated
    Run Keyword Unless  '${color}' == 'Black' or '${Size}' == 'Small' or '${Design}' == 'Simple'  log to console  \nexecuted with unless and multiple or

```

and here is what I get when I execute it:



```
$ pybot ts.txt
==============================================================================
Ts
==============================================================================
mytest                                                                .
executed with single condition
executed with multiple or
executed with unless and multiple or
mytest                                                                | PASS |
------------------------------------------------------------------------------

```

"
"Install Debian kerner 5.14
I just got a laptop with a wireless adapter (MEDIATEK 7961) supported only by kernel 5.14+


I installed the latest kernel (5.15.0-2) from the unstable Debian repository but I can't get it working. I would like to try an older version, like 5.14.


I got 5.15.0-2 simply pointing to unstable Debian repo and upgraded the system



```
deb http://deb.debian.org/debian unstable main contrib non-free
deb-src http://deb.debian.org/debian unstable main contrib non-free

```

Is there a way to specify which kernel version I want to install, or any other way to install it?


","Unfortunately pointing `apt` to `unstable` and upgrading the system, unless you did an `apt full-upgrade`, has probably left you with a mix of `unstable` and whatever you had before running the upgrade. There are no guarantees at all that such a [FrankenDebian](https://wiki.debian.org/DontBreakDebian) mix will actually work.


If you only recently got the laptop, reinstalling might be the easiest and quickest way back to a stable configuration.


The easiest way to get a newer kernel to a stable version of Debian is to use the [backports repository](https://backports.debian.org/) designated for that release.


For example, [the backports repository for Debian 11 (`bullseye`)](https://backports.debian.org/changes/bullseye-backports.html) currently has kernel version 5.14.9-2~bpo11+1 available. (The `~bpo11` part of the version number indicates it's from the backports repository for Debian 11.)


To get a 5.14 kernel for Debian 11 from the backports repository, you should:


1. add `deb http://deb.debian.org/debian bullseye-backports main` to your `/etc/apt/sources.list` (or create a file `/etc/apt/sources.list.d/backports.list` for it, and add it there)
2. run `apt update` to download the package index from the newly-added backports repository
3. run `apt install linux-image-amd64/bullseye-backports` to install the kernel package from the backports repository (currently version 5.14.9+2~bpo11+1)


The distribution's normal packages will be used by default by `apt`. Whenever you want to install a newer version from the backports repository, you will have to explicitly specify that you want the version from the `bullseye-backports`. If the backports package you already have installed gets updated in the backports repository, `apt` should handle updates for it just as usual. But it won't automatically add more packages from the backports repository without your explicit request.


"
"ERC721. How to add additional information related to property using erc721 standard
I'm not having so much experience with ERC721 token standard, currently I'm working on a real-estate DAPP. I have a question. if I want to add external information related to a specific property like Location, Price, property number etc. every time if a new property register, what will be the best way to do that..?? but I don't want this with solidity Struct, is it possible to extend the `ERC721 Metadata Interface Contract`?? or any other solution??


I have tried almost everything but I think I'm missing something.


","If you want to store the data on-chain, a `mapping (uint256 => Property)`, where the `uint256` is the token ID and `Property` is ""struct(location, price, ...)"", containing the data is probably the cheapest option gas-wise.


But since your question states you don't want to use struct, you can chose to store the data on-chain with a series of mappings:



```
mapping (uint256 => string) tokenIdToLocation;
mapping (uint256 => uint64) tokenIdToUsdPrice;
// etc.

```

You can also decide to store the data off-chain, and link to this storage from your contract. In that case, you'd implement the `tokenURI()` function of the `ERC721Metadata` interface (both defined in the [ERC-721 standard](https://github.com/ethereum/EIPs/blob/master/EIPS/eip-721.md)). The `tokenUri()` would return a (string) URL of an off-chain resource where you can display the (off-chain) data.


"
"Is it allowed to use decltype in an initializer for the variable that is decltyped?
Triggered by [this question](https://stackoverflow.com/questions/60830000/is-this-an-incorrect-warning), I was wondering if, this is allowed:



```
template <typename T>
T foo(){return T{};}

struct bar {};

int main()
{

    bar a = foo<decltype(a)>();
}

```

Compilers I tried took it without complaints, but I am not sure if this is really legal or if I am missing any pitfalls (and it looks weird to use the type of `a` during its declaration). 


For background: in the linked question OP wants to avoid `auto` and spelling out the type (here it is `bar`, `SomeComplexTypeAndNotAuto` in that question) twice at the same time, hence they use a (unused) parameter to deduce `T`. I didn't like misusing a parameter merely to deduce the type so my first idea was `decltype`.


","It's kosher alright.



> 
> **[basic.scope.pdecl]**
> 
> 
> [1](http://eel.is/c++draft/basic.scope#pdecl-1) The point of declaration for a name is immediately after its
>  complete declarator ([dcl.decl]) and before its initializer (if any),
>  except as noted below. [ Example:
> 
> 
> 
> ```
> unsigned char x = 12;
> { unsigned char x = x; }
> 
> ```
> 
> Here, the initialization of the second x has undefined behavior,
>  because the initializer accesses the second x outside its lifetime
>  ([basic.life]). — end example ]
> 
> 
> 


So you can use `a` in its own initializer, since its declared at the point. The only question is how. `decltype` can be applied to an id-expression naming any variable in scope. And since the expression of `decltype` [is an unevaluated operand](http://eel.is/c++draft/dcl.type#decltype-1.sentence-2), there is no UB. Only the type of the variable is examined, not its indeterminate value.


No accounting for taste though.


"
"Firebase Crashlytics Bitmap OutOfMemoryError which activity
I see a user of mine got a crash. How do I know when this crash happened? Which activity did it happen in? Which line of code?


I don't use Bitmap factory anywhere, but I do display some pictures in ImageView.


Here is the crash as shown to me in the Firebase console:



```
Fatal Exception: java.lang.OutOfMemoryError
   at android.graphics.Bitmap.nativeCreateFromParcel(Bitmap.java)
   at android.graphics.Bitmap.access$000(Bitmap.java:31)
   at android.graphics.Bitmap$1.createFromParcel(Bitmap.java:1308)
   at android.graphics.Bitmap$1.createFromParcel(Bitmap.java:1300)
   at android.app.IUiAutomationConnection$Stub$Proxy.takeScreenshot(IUiAutomationConnection.java:224)
   at android.app.UiAutomation.takeScreenshot(UiAutomation.java:599)
   at com.google.android.apps.mtaas.crawler.platform.common.Utils.takeScreenshot(Utils.java:42)
   at com.google.android.apps.mtaas.crawler.platform.hybrid.ViewScreenStateBuilder.traverseAndBuild(ViewScreenStateBuilder.java:41)
   at com.google.android.apps.mtaas.crawler.platform.hybrid.HybridStateExtractor.tryExtractingScreenState(HybridStateExtractor.java:40)
   at com.google.android.apps.mtaas.crawler.platform.hybrid.HybridStateExtractor.getStableScreen(HybridStateExtractor.java:8)
   at com.google.android.apps.mtaas.crawler.controller.remote.RemotePlatform.handlePerformScrape(RemotePlatform.java:40)
   at com.google.android.apps.mtaas.crawler.controller.remote.RemotePlatform.access$400(RemotePlatform.java:81)
   at com.google.android.apps.mtaas.crawler.controller.remote.RemotePlatform$ControllerMessageHandler.handleMessage(RemotePlatform.java:10)
   at android.os.Handler.dispatchMessage(Handler.java:99)
   at android.os.Looper.loop(Looper.java:176)
   at android.os.HandlerThread.run(HandlerThread.java:61)

```

I never got this crash myself? I can see which device it happened on, (Galaxy S3), so I tried it in an emulator, and everything worked.


Thanks


Here is screen shot from email
[![enter image description here](https://i.stack.imgur.com/lSApy.png)](https://i.stack.imgur.com/lSApy.png)


","I don't think this is coming from a user. The origin of your stack trace is `com.google.android.apps.mtaas.crawler`. This is most likely coming from the [Goole Play Pre-launch Test Report](https://developer.android.com/distribute/best-practices/launch/pre-launch-crash-reports) when you submitted your \*.apk to Google Play. This happens automatically. And as one of the the other comments stated, it looks like it's running out of memory when the crawler is taking a screenshot. 


If you log into Google Play, you should see an option to view the Prelaunch test report (from the left menu go to `Release Management -> Pre-launch report`). In fact, you should have gotten a warning right before you published the app I believe. This is probably a bug in that service and not a user. So you probably don't have to worry about it. Though it might be a good idea to file a bug Google Play.


"
"TypeError: Can't convert 'bytes' object to str implicitly while working with sockets
So I'm trying to convert code from Python 2.7 to Python 3, and it seems as though something has changed. I'm trying to receive binary data over a socket and now it doesn't work. Here's my code.


EDIT: I have added my send code. Also, I don't really like the way it works right now, it's overcomplicated. If you can it would be nice to have a better way of sending/receiving data.



```
def recv(self):
    # Receive the length of the incoming message (unpack the binary data)
    dataLength = socket.ntohl(struct.unpack(""I"", self._recv(4))[0])

    # Receive the actual data
    return self._recv(dataLength)

def _recv(self, length):
    try:
        data = ''
        recvLen = 0
        while recvLen < length:
            newData = self.sock.recv(length-recvLen)

            if newData == '':
                self.isConnected = False
                raise exceptions.NetworkError(errors.CLOSE_CONNECTION, errno=errors.ERR_CLOSED_CONNECTION)

            data = data + newData # TypeError here
            recvLen += len(newData)

        return data
    except socket.error as se:
        raise exceptions.NetworkError(str(se))

def send(self, data):
    if type(data) is not str:
        raise TypeError()

    dataLength = len(data)

    # Send the length of the message (int converted to network byte order and packed as binary data)
    self._send(struct.pack(""I"", socket.htonl(dataLength)), 4)

    # Send the actual data
    self._send(data, dataLength)

def _send(self, data, length):
    sentLen = 0
    while sentLen < length:
        try:
            amountSent = self.sock.send(data[sentLen:])
        except Exception:
            self.isConnected = False
            raise exceptions.NetworkError(errors.UNEXPECTED_CLOSE_CONNECTION)

        if amountSent == 0:
            self.isConnected = False
            raise exceptions.NetworkError(errors.UNEXPECTED_CLOSE_CONNECTION)

        sentLen += amountSent

```

","Python 3 sends data as bytes so you have to decode to string 



```
 data = data + newData.decode('utf-8')

 # or 

 data = data + newData.decode('ascii')

```

if you need bytes data then use 



```
 data = b''

```

and keep without `.decode()`



```
 data = data + newData

```



---


**EDIT:** for new code in question.


When you send you have to convert/encode string to bytes and after that get its length. Native chars has length 1 as unicode but they can use 2 bytes (or more).


When you receive you have to work with bytes `b''` and at the end convert/decode bytes to string again.


See comments `# <--` in code 



```
def send(self, data):
    if not isinstance(data, str): # <-- prefered method
    #if type(data) is not str:
        raise TypeError()

    data = data.encode('utf-8') # <-- convert to bytes

    # get size of bytes
    dataLength = len(data)

    # Send the length of the message (int converted to network byte order and packed as binary data)
    self._send(struct.pack(""I"", socket.htonl(dataLength)), 4)

    # Send the actual data
    self._send(data, dataLength)


def recv(self):
    # Receive the length of the incoming message (unpack the binary data)
    dataLength = socket.ntohl(struct.unpack(""I"", self._recv(4))[0])

    # Receive the actual data
    return self._recv(dataLength).decode('utf-8') # <-- convert to string again

def _recv(self, length):
    try:
        data = b'' # <-- use bytes
        recvLen = 0
        while recvLen < length:
            newData = self.sock.recv(length-recvLen)

            #if newData == b'': # <-- use bytes
            if not newData:    # <-- or 
                self.isConnected = False
                raise exceptions.NetworkError(errors.CLOSE_CONNECTION, errno=errors.ERR_CLOSED_CONNECTION)

            data = data + newData # TypeError here
            recvLen += len(newData)

        return data
    except socket.error as se:
        raise exceptions.NetworkError(str(se))

```

"
"custom exception behavior in symfony2
Im trying to figure out how to make a custom exception behavior. When i throw a exception using



```
 throw new \Exception('Error occurred with your request please try again');

```

I automatically get status 500 and the message as internal server error 


However i would instead like my response to include my exception message instead of just internal server error so for it to display something like so:



```
 {
   ""error"":{
      ""code"":500, 
      ""message"":""Error occurred with your request please try again""
   }
 }

```

and on top of that possibly do some extra things such as email myself the error. However I only want this to happen when i throw a \Exception as opposed to using something like



```
    throw new HttpException

```

Any help or ideas on how to accomplish this. 


I should also mention that I am not using Twig or any templates for this. This is strictly a API type response


","Take a look at <http://symfony.com/doc/current/cookbook/controller/error_pages.html> There is enough information to get you started.


In short, you should create app/Resources/TwigBundle/views/Exception/exception.json.twig and there you have access to the exception.message and error\_code.


here's solution for you:



```
{% spaceless %}
{
  ""error"":{
    ""code"": {{ error_code }}, 
    ""message"":{{ exception.message }}
  }
}
{% endspaceless %}

```

Another solution is to use Exception Listener:



```
namespace Your\Namespace;

use Symfony\Component\HttpFoundation\JsonResponse;
use Symfony\Component\HttpKernel\Event\GetResponseForExceptionEvent;

class JsonExceptionListener
{
    public function onKernelException(GetResponseForExceptionEvent $event)
    {
        $exception = $event->getException();
        $data = array(
            'error' => array(
                'code' => $exception->getCode(),
                'message' => $exception->getMessage()
            )
        );
        $response = new JsonResponse($data, $exception->getCode());
        $event->setResponse($response);
    }
}

```

update your services config:



```
json_exception_listener:
    class: Your\Namespace\JsonExceptionListener
    tags:
        - { name: kernel.event_listener, event: kernel.exception, method: onKernelException, priority: 200 }

```

cheers


"
"jQuery add a class after a delay then delay and a another class
How do I add a class after a delay then delay and a another class?


I can only seem to add one.



```
$(document).ready(function() {
    $("".websites-container"").delay(5000).queue(function() {
            $(this).addClass(""active"");
    });

    $("".websites-container"").delay(8000).queue(function() {
            $(this).addClass(""gone"")
    });
});

```

","You need to call [`.dequeue()`](https://api.jquery.com/dequeue/) in the `.queue()` callback, otherwise the following items in the queue never get executed.





```
$(document).ready(function() {
    $("".websites-container"").delay(1000).queue(function() {
            // I'm impatient. I shortened the times.
            $(this).addClass(""active"").dequeue();
    }).delay(2000).queue(function() {
            $(this).addClass(""gone"");
    });
});
```


```
.active {font-weight: bold}
.gone {color:red}
```


```
<script src=""https://ajax.googleapis.com/ajax/libs/jquery/1.11.1/jquery.min.js""></script>
<div class=""websites-container"">abc</div>
```






(This was almost solved by [@adeneo](https://stackoverflow.com/questions/30697921/jquery-add-a-class-after-a-delay-then-delay-and-a-another-class/30698162#comment49454493_30697921), with the exception of placing dequeue in the callback.)


"
"Programmatically convert Word (docx) to PDF
Ok before you think ""Not another question like this"" please read this first.


I have an application (web application in ASP.NET MVC 3) the generates Word files in DocX using the [DocX library](http://docx.codeplex.com/).


The application takes a template and fills it in with all the data from a database.
Now I want to create a PDF version of that created docx-file.


I know [apose.word](http://www.aspose.com/categories/.net-components/aspose.words-for-.net/default.aspx) is an option, but not for me since I have little budget. Other libs where I have to spend some money on are also out of the question.


I don't have a sharepoint server so [Word Automation Services](http://msdn.microsoft.com/en-us/library/ee558830.aspx) isn't an option either.


So I have 2 options (that I know) left and they both use [iTextSharp](http://itextpdf.com/). Don't know which is better.


1. I could use the generated XML from the docx file and transform it to a version that is usable by iTextSharp.
2. I could create the PDF like I create the docx with a template.


Anybody has any idea on how much work it is, which of those 2 has better performance and if it is even possible to do.


I know that the second option has the downside that when I change a template I have to change it for both versions.


If you have a better solution (free that is), you are welcome to share it.


","Another option, even if it needs some work: install OpenOffice on server and, using UNO libraries (including them as assemblies in your app), you can open docx document and save it in PDF directly.  

In a few minutes I post an example...


**PARTIAL EXAMPLE:**  

This is a class I created a long time ago and used to convert files to pdf 



```
using unoidl.com.sun.star.lang;
using unoidl.com.sun.star.uno;
using unoidl.com.sun.star.container;
using unoidl.com.sun.star.frame;
using unoidl.com.sun.star.beans;
using unoidl.com.sun.star.view;
using System.Collections.Generic;
using System.IO;

namespace QOpenOffice
{
    public enum AppType
    {
        Writer,
        Calc,
        Impress,
        Draw,
        Math
    }

    public enum ExportFilter{
        Word97,
        WriterPDF,
        CalcPDF,
        DrawPDF,
        ImpressPDF,
        MathPDF
    }

    class OpenOffice
    {
        private XComponentContext context;
        private XMultiServiceFactory service;
        private XComponentLoader component;
        private XComponent doc;

        private List<string> filters = new List<string>();

        #region Constructors
        public OpenOffice()
        {
            /// This will start a new instance of OpenOffice.org if it is not running, 
            /// or it will obtain an existing instance if it is already open.
            context = uno.util.Bootstrap.bootstrap();

            /// The next step is to create a new OpenOffice.org service manager
            service = (XMultiServiceFactory)context.getServiceManager();

            /// Create a new Desktop instance using our service manager
            component = (XComponentLoader)service.createInstance(""com.sun.star.frame.Desktop"");

            // Getting filters
            XNameContainer filters = (XNameContainer)service.createInstance(""com.sun.star.document.FilterFactory"");
            foreach (string filter in filters.getElementNames())
                this.filters.Add(filter);
        }

        ~OpenOffice()
        {
            if (doc != null)
                doc.dispose();
            doc = null;
        }
        #endregion

        #region Private methods
        private string FilterToString(ExportFilter filter)
        {
            switch (filter)
            {
                case ExportFilter.Word97: return ""MS Word 97"";
                case ExportFilter.WriterPDF: return ""writer_pdf_Export"";
                case ExportFilter.CalcPDF: return ""calc_pdf_Export"";
                case ExportFilter.DrawPDF: return ""draw_pdf_Export"";
                case ExportFilter.ImpressPDF: return ""impress_pdf_Export"";
                case ExportFilter.MathPDF: return ""math_pdf_Export"";
            }
            return """";
        }
        #endregion

        #region Public methods
        public bool Load(string filename, bool hidden)
        {
            return Load(filename, hidden, """", """");
        }
        public bool Load(string filename, bool hidden, int filter_index, string filter_options)
        {
            return Load(filename, hidden, filters[filter_index], filter_options);
        }
        public bool Load(string filename, bool hidden, string filter_name, string filter_options)
        {
            List<PropertyValue> pv = new List<PropertyValue>();
            pv.Add(new PropertyValue(""Hidden"", 0, new uno.Any(hidden), PropertyState.DIRECT_VALUE));
            if (filter_name != """")
            {
                pv.Add(new PropertyValue(""FilterName"", 0, new uno.Any(filter_name), PropertyState.DIRECT_VALUE));
                pv.Add(new PropertyValue(""FilterOptions"", 0, new uno.Any(filter_options), PropertyState.DIRECT_VALUE));
            }

            try
            {
                doc = component.loadComponentFromURL(
                    ""file:///"" + filename.Replace('\\', '/'), ""_blank"",
                    0, pv.ToArray());
                return true;
            }
            catch
            {
                doc = null;
                return false;
            }
        }
        public bool Print()
        {
            return Print(1, """");
        }
        public bool Print(int copies, string pages)
        {
            List<PropertyValue> pv = new List<PropertyValue>();
            pv.Add(new PropertyValue(""CopyCount"", 0, new uno.Any(copies), PropertyState.DIRECT_VALUE));
            if (pages != """")
                pv.Add(new PropertyValue(""Pages"", 0, new uno.Any(pages), PropertyState.DIRECT_VALUE));
            //if (doc is XPrintable)
            try
            {
                ((XPrintable)doc).print(pv.ToArray());
                return true;
            }
            catch { return false; }
        }
        public bool Save(string filename, ExportFilter filter)
        {
            return Save(filename, FilterToString(filter));
        }
        public bool Save(string filename, string filter)
        {
            List<PropertyValue> pv = new List<PropertyValue>();
            pv.Add(new PropertyValue(""FilterName"", 0, new uno.Any(filter), PropertyState.DIRECT_VALUE));
            pv.Add(new PropertyValue(""Overwrite"", 0, new uno.Any(true), PropertyState.DIRECT_VALUE));
            try
            {
                filename = filename.Replace(""\\"", ""/"");
                ((XStorable)doc).storeToURL(""file:///"" + filename, pv.ToArray());
                return true;
            }
            catch { return false; }
        }
        public bool ExportToPdf(string filename)
        {
            filename = Path.ChangeExtension(filename, "".pdf"");
            bool ret = Save(filename, ""writer_pdf_Export"");
            if (!ret) ret = Save(filename, ""impress_pdf_Export"");
            if (!ret) ret = Save(filename, ""calc_pdf_Export"");
            if (!ret) ret = Save(filename, ""draw_pdf_Export"");
            if (!ret) ret = Save(filename, ""impress_pdf_Export"");
            if (!ret) ret = Save(filename, ""math_pdf_Export"");
            return ret;
        }
        public void Close()
        {
            doc.dispose();
            doc = null;
        }

        public bool New(AppType app, bool hidden)
        {
            try
            {
                string sapp = ""private:factory/"";
                switch (app)
                {
                    case AppType.Writer:
                        sapp += ""swriter"";
                        break;
                    case AppType.Calc:
                        sapp += ""scalc"";
                        break;
                    case AppType.Impress:
                        sapp += ""simpress"";
                        break;
                    case AppType.Draw:
                        sapp += ""sdraw"";
                        break;
                    case AppType.Math:
                        sapp += ""smath"";
                        break;
                }
                PropertyValue pv = new PropertyValue(""Hidden"", 0, new uno.Any(hidden), PropertyState.DIRECT_VALUE);
                doc = component.loadComponentFromURL(sapp, ""_blank"", 0, new PropertyValue[1] { pv });
                return true;
            }
            catch
            {
                doc = null;
                return false;
            }
        }
        #endregion


        #region Properties
        public List<string> Filters
        {
            get { return filters; }
        }
        #endregion
    }
}

```

"
"How to convert string to decimal with 3 decimal places?

```
string num = 23.6;

```

I want to know how can I convert it into decimal with 3 decimal places
like



```
decimal nn = 23.600 

```

Is there any method?


","I try my best..


First of all your `string num = 23.6;` won't *even* compile. You need to use double quotes with your strings like `string num = ""23.6"";`


If you wanna get this as a `decimal`, you need to parse it first with a `IFormatProvider` that have `.` as a [`NumberDecimalSeparator`](https://msdn.microsoft.com/en-us/library/system.globalization.numberformatinfo.numberdecimalseparator.aspx) like [`InvariantCulture`](https://msdn.microsoft.com/en-us/library/system.globalization.cultureinfo.invariantculture%28v=vs.110%29.aspx)(if your `CurrentCulture` uses `.` already, you *don't* have to pass second paramter);



```
decimal nn = decimal.Parse(num, CultureInfo.InvariantCulture);

```

Now we have a `23.6` as a `decimal` value. But as a value, `23.6`, `23.60`, `23.600` and `23.60000000000` are totally same, right? No matter which one you parse it to decimal, you will get the same value as a `23.6M` in debugger. Looks like these are *not* true. See Jon Skeet comments on this answer and his ""Keeping zeroes"" section on [Decimal floating point in .NET](http://csharpindepth.com/Articles/General/Decimal.aspx) article.


Now what? Yes, we need to get it's *textual* representation as `23.600`. Since we need only decimal separator in a textual representation, [The `""F""` Format Specifier](https://msdn.microsoft.com/en-us/library/dwhawy9k%28v=vs.110%29.aspx#FFormatString) will fits out needs.



```
string str = nn.ToString(""F3"", CultureInfo.InvariantCulture); // 23.600

```

"
"Combine dimensions of lists of lists into separate vectors
Here's an example of my data.



```
ll <- list(
  ll1 = list(Mi = 1:4,
       Mj = 10:13,
       dn = ""l1""),
  ll2 = list(Mi = 5:8,
             Mj = 14:17,
             dn = ""l2""))
> str(ll)
List of 2
 $ ll1:List of 3
  ..$ Mi: int [1:4] 1 2 3 4
  ..$ Mj: int [1:4] 10 11 12 13
  ..$ dn: chr ""l1""
 $ ll2:List of 3
  ..$ Mi: int [1:4] 5 6 7 8
  ..$ Mj: int [1:4] 14 15 16 17
  ..$ dn: chr ""l2""

```

I'm trying to combine each Mi, each Mj, and each dn together. So the final result will be 3 vectors: 1 combined Mi, 1 combined Mj, and 1 combined dn. I.e., the final result for Mi would be something like



```
> c(ll$ll1$Mi,ll$ll2$Mi)
[1] 1 2 3 4 5 6 7 8

```

Similar final results for Mj and dn. I'm wondering how to do this simply, preferably using something from tidyverse. The number of lists will vary and won't be 2 generally, which is why I don't think c is a great solution. I am guessing the map functions can work here, but I have not yet found a solution.


","With `pmap` from `purrr`. `pmap` takes a list of inputs (in this case a list of two lists) and apply the `c` function to `ll[[1]][[1]]`, `ll[[2]][[1]]`, ... then `ll[[1]][[2]]`, `ll[[2]][[2]]`, ... and so on:



```
library(purrr)
pmap(ll, c)

```

**Output:**



```
$Mi
ll11 ll12 ll13 ll14 ll21 ll22 ll23 ll24 
   1    2    3    4    5    6    7    8 

$Mj
ll11 ll12 ll13 ll14 ll21 ll22 ll23 ll24 
  10   11   12   13   14   15   16   17 

$dn
 ll1  ll2 
""l1"" ""l2""

```

If we don't want to keep the vector names, we can use the `use.names` argument in `c`:



```
pmap(ll, c, use.names = FALSE)

```

**Output:**



```
$Mi
[1] 1 2 3 4 5 6 7 8

$Mj
[1] 10 11 12 13 14 15 16 17

$dn
[1] ""l1"" ""l2""

```

"
"C++ char variable 2 words

```
#include <iostream>
using namespace std;

void main(){
  char name[20];
  cin>>name;         // when I input ""This is"" 
  cout<<name<<endl;  // output was ""This"" 
}

```

How to do that when I input ""This is"" , output too will be ""This is"" , not only ""This"" ?


","You could use 



```
cin.get( name, 20 );

```

or



```
cin.getline(name, 20);

```

Depending on whether you want the new line character in your string.


EDIT:


If you want to further simplify your code you could use a string instead of a char array, you wouldn't have to worry about the user exceeding your buffer size then ie.



```
string name;
getline(cin, name);
cout << name << endl;

```

EDIT 2:


As David Heffernan also pointed out in a comment I should explain why it's not working for you as expected. The extraction operator (`operator >>`) ends when a whitespace character -- a space in your case -- is reached. The operator also terminates when a null-character or the end of file is reached. 


"
"Typeclass constraints on data declarations
Apparently it's a bad idea to put a typeclass constraint on a data declaration [[src]](https://stackoverflow.com/questions/2354707/in-haskell-is-there-num-a-infinity-a?rq=1#comment2333396_2354747), [[src]](http://www.haskell.org/haskellwiki/Data_declaration_with_constraint).


I haven't personally come across a desire to constrain the types within data types I've created, but it's not obvious to me why the language designers ""decided it was a bad idea to allow"". Why is that?


","
> 
> I haven't personally come across a desire to constrain the types within data types I've created, but it's not obvious to me why the language designers ""decided it was a bad idea to allow"". Why is that?
> 
> 
> 


Because it was misleading and worked completely backwards from what would actually be useful.


In particular, it *didn't actually constrain the types* within the data type in the way you're probably expecting. What it *did* do was put a class constraint on *the data constructor itself*, which meant that you needed to satisfy the instance when constructing a value... but that was all.


So, for instance, you couldn't simply define a binary search tree with an `Ord` constraint and then know that any tree has sortable elements; the lookup and insert functions would still need an `Ord` constraint themselves. All you'd prevent would be constructing an empty tree that ""contains"" values of some non-ordered type. As far as pattern matching was concerned, there was no constraint on the contained type at all.


On the other hand, the people working on Haskell didn't think that the sensible version (that people tended to assume data type contexts provided) was a bad idea at all! In fact, class constraints on a data type [declared with GADT syntax](http://www.haskell.org/ghc/docs/latest/html/users_guide/data-type-extensions.html#gadt-style) (generalized algebraic data types, enabled in GHC with the `GADTs` language pragma) do work in the obvious way--you need a constraint to *construct* the value, and the instance in question also gets stored in the GADT, so that you don't need a constraint to work with values, and pattern matching on the GADT constructor lets you use the instance it captured.


"
"Unexplained syntax error with logical OR ( || )
I'm working with NetBeans for Mac, and I'm running CakePHP (though I don't think the framework has anything to do with it) in a shared hosting in Linux. This is not a big issue, but it is frustrating.


I wonder why I can't simply do this:



```
if($this->Session->read('User.value1') || $this->Session->read('User.value2')){
  ...
}

```

The error message I get is:



```
Error: syntax error, unexpected '$this' (T_VARIABLE)

```

Why is there a syntax error? I can't see it.


I can do this with no problems:



```
if($this->Session->read('value1')){
  ...
}

```

I can also do this with no problems (no whitespace around `||`):



```
if($this->Session->read('User.value1')||$this->Session->read('User.value2')){
  ...
}

```

But if I put spaces around the `||` operator, it stops working. Or rather — and this is the most confusing part — *sometimes* it stops working when I put spaces around the `||` operator, and sometimes it doesn't.


I thought this might be a bug in Netbeans 7.4, but when I ignored the warning from NetBeans and tried to run the code anyway, PHP gave me the same error.


What is happening here?


","
> 
> I'm working With NetBeans for MAC
> 
> 
> 


## When is a space not a space?


When it's a [non-breaking space](https://superuser.com/questions/167886/disabling-altspace-non-breaking-space-on-mac-os-x-10-6)!


The intention is:



```
"" || ""
207C7C20 (hex)

```

But what is actually in the source file is almost certainly:



```
"" || ""
207C7CA0 (hex)

```

(*on stack overflow it won't be* but I bet it is in the source file).


With a mac the problem is (using my own keyboard layout, but I am assuming it's similar in your case):



```
""|"" = alt + 1
"" "" = alt + space (accidental)

```

So typing away, with the sequence "" || "" it's *very* easy for the alt key to still be depressed when the space bar is pressed and: voilà you get unexpected ""wat"" syntax errors which at face value make no sense - until you realize what the problem is.


Example:



```
-> cat foo.php 
<?php

$foo = ""x"";
if (true || $foo) {
}
-> php -l foo.php 

Parse error: syntax error, unexpected '$foo' (T_VARIABLE) in foo.php on line 4

Errors parsing foo.php

```

"
"how to read contact name from contact number?
i am trying to show contact number as well as contact names in a list view. 
i have done...:



```
public String DisplayName(String number) {
    Uri uri = Uri.withAppendedPath(
            ContactsContract.PhoneLookup.CONTENT_FILTER_URI,
            Uri.encode(number));
    String name = ""?"";

    ContentResolver contentResolver = getContentResolver();
    Cursor contactLookup = contentResolver.query(uri, new String[] {
            BaseColumns._ID, ContactsContract.PhoneLookup.DISPLAY_NAME },
            null, null, null);

    try {
        if (contactLookup != null && contactLookup.getCount() > 0) {
            contactLookup.moveToNext();
            name = contactLookup.getString(contactLookup
                    .getColumnIndex(ContactsContract.Data.DISPLAY_NAME));
            // String contactId =
            // contactLookup.getString(contactLookup.getColumnIndex(BaseColumns._ID));
        }
    } finally {
        if (contactLookup != null) {
            contactLookup.close();
        }
    }

    return name;
}

```

but my app keep on force closing , i dont know why?? i have include the permission also...
""""


any guess? why


","use this 2 methods:



```
/*
 * Returns contact's id
 */
private String getContactId(String phoneNumber, Context context) {
    ContentResolver mResolver = context.getContentResolver();

    Uri uri = Uri.withAppendedPath(PhoneLookup.CONTENT_FILTER_URI,
            Uri.encode(phoneNumber));

    Cursor cursor = mResolver.query(uri, new String[] {
            PhoneLookup.DISPLAY_NAME, PhoneLookup._ID }, null, null, null);

    String contactId = """";

    if (cursor.moveToFirst()) {
        do {
            contactId = cursor.getString(cursor
                    .getColumnIndex(PhoneLookup._ID));
        } while (cursor.moveToNext());
    }

    cursor.close();
    cursor = null;
    return contactId;
}

/*
 * Returns contact's name
 */
private String getContactName(String contactId, Context context) {

    String[] projection = new String[] { Contacts.DISPLAY_NAME };
    Cursor cursor = mResolver.query(Contacts.CONTENT_URI, projection,
            Contacts._ID + ""=?"", new String[] { contactId }, null);
    String name = """";
    if (cursor.moveToFirst()) {
        name = cursor.getString(0);
    }

    cursor.close();
    cursor = null;
    return name;
}

```

"
"when for\_each empty, ignore dynamic resource
I have the following in my main.tf:



```
data ""aws_iam_policy_document"" ""task_role_policy"" {
  dynamic ""statement"" {
    for_each = var.policy_statements
    content {
      actions   = statement.value.actions
      resources = statement.value.resources
      effect    = ""Allow""
    }
  }
}

```

When var.policy\_statements is empty list or nothing I get the following error when running `terraform apply`:



```
Error: Error creating IAM policy dev-chatbot-engine-policy: MalformedPolicyDocument: Syntax errors in policy.
    status code: 400, request id: a181b065-b659-4261-87d5-9aae8c4454aa

  on .terraform/modules/service/main.tf line 68, in resource ""aws_iam_policy"" ""task_role"":
  68: resource ""aws_iam_policy"" ""task_role"" {

```

","It looks like this policy is still being reference in the `aws_iam_policy.task_role` resource when `var.policy_statements` is empty.


This would cause `aws_iam_policy.task_role` to be created with an empty `Statement` (which causes that malformed-policy error you are seeing).


I would recommend setting a `count` flag on the policy itself so that it doesn't even attempt to create it when the statements are empty, e.g.



```
resource ""aws_iam_policy"" ""task_role"" {
  count = length(var.policy_statements) == 0 ? 0 : 1

  // Your other args here...
}

```

This may have cascading effects to other resources (such as those that consume `aws_iam_policy.task_role`). You'll need to handle those effects by providing defaults that don't break or adding a `count` there as well.


"
"Is there a problem doing another API call inside interceptor in Android?
I have an issue where I am trying to refresh an access token inside my interceptor. I have read it is better to use an Authenticator for this issue, but the way my API works is that I get a 403 error for an unauthorized access token (not my design but gotta work with it). 


So I am trying to call the API endpoint to get a new access token inside my interceptor but it seems like the second API call creates an OKHTTP Response without closing it.



```
override fun intercept(chain: Interceptor.Chain?): Response {

        var currentRequest = chain!!.request()
        val currentRequestResponse = chain.proceed( currentRequest )

        if ( currentRequestResponse.code() >= HTTP_ERROR_CODE && isNotLoginRequest( chain ) ) {

            try {

                // we check the response code, if there is 403 code we need to relog the user
                // before executing the request
                // try to reconnect the user with synchronous request


                val userManager = UserManager.getInstance()

                val clientId = getQueryParameter(UserService.CLIENT_ID_QUERY_KEY, currentRequest)
                val redirectUri = getQueryParameter(UserService.REDIRECT_URI_QUERY_KEY, currentRequest)
                val clientSecret = getQueryParameter(UserService.CLIENT_SECRET_QUERY_KEY, currentRequest)

                // we get a new access token with a synchronous request
                val accessToken = userManager.refreshToken(clientId, clientSecret, redirectUri)

                currentRequest = rebuildRequestWithNewToken( accessToken, currentRequest )

                return chain.proceed( currentRequest )

            } catch (e: Throwable) {

                e.printStackTrace()
            }
        }

        return currentRequestResponse
    }

```

I believe the problem comes from the userManager.refreshToken which is the API call to get the new access token.


Here is the error I receive:


java.lang.IllegalStateException: cannot make a new request because the previous response is still open: please call response.close()


","You can make the refresh token call inside the network interceptor like this



```
if (response.code() == 403) {
            try {
                val body = RequestBody.create(contentType, """")

                val authRequest = request.newBuilder()
                        .setDefaultHeaders()
                        .addHeader(""Authorization"", ""Your token"")
                        .url(refreshtokenurl)
                        .post(body)
                        .build()
                val tokenRefreshResponse = chain.proceed(authRequest)
                if (tokenRefreshResponse.code() == 401 || tokenRefreshResponse.code() == 403) {
                   //error in refreshing token
                } else if (tokenRefreshResponse.code() == 200) {
                   //save the new token and refresh token in preference and continue with the earlier request
                    currentRequestResponse = chain.proceed(currentRequest.newBuilder()
                                .setDefaultHeaders()
                                .addHeader(your header)
                                .build())
                }
            } catch (e: Exception) {

            }

        }


```

"
"How to save a plot in R in a subdirectory of the working directory
Is it possible to save a plot in R into a subdirectory of the current working directory? I tried the following, but that doesn't work. I'm not sure how to concatenate the working directory to the file name I want. 



```
  wd <- getwd()

  png(filename=wd+""/img/name.png"")

  counts <- table(dnom$Variant, dnom$Time)
  barplot(counts, main=""Distribution of Variant and words of time"",
    xlab=""Temporal nouns"", col=c(""paleturquoise3"", ""palegreen3""),
    legend = rownames(counts))

```

Also, what's the default save directory for the image export functions?


When running David's suggestion below the error returned is:



```
Error in png(filename = paste0(wd, ""/img/name.png"")) : 
  unable to start png() device
In addition: Warning messages:
1: In png(filename = paste0(wd, ""/img/name.png"")) :
  unable to open file 'D:/Dropbox/Corpuslinguïstiek project/antconc resultaten/img/name.png' for writing
2: In png(filename = paste0(wd, ""/img/name.png"")) : opening device failed

```

","Try this:



```
File <- ""./img/name.png""
if (file.exists(File)) stop(File, "" already exists"")
dir.create(dirname(File), showWarnings = FALSE)

png(File)

# ... whatever ...

dev.off()

```

Omit the `if` statement if its ok to overwrite the file.


If `img` exists then `dir.create` could be optionally omitted. (If you try to create a directory that is already there it won't cause a problem.)


**Notes**


1) Another possility is to put `img` in the home directory. We could use `png(""~/img/name.png"")` to save the file to the `img` directory in the home directory. If unsure which directory is the home directory try `path.expand(""~"")`.


2) Also note the `savePlot` command which is given after (rather than before) the plotting command.


"
"Why does boost not have a make\_scoped()?
Boost's [`make_shared()`](http://www.boost.org/doc/libs/1_50_0/libs/smart_ptr/make_shared.html) function promises to be exception-safe while attempting to create a `shared_ptr`.


**Why is there no `make_scoped()` equivalent? Is there a common best practice?**


Here's a code example from the [`boost::scoped_ptr` documentation](http://www.boost.org/doc/libs/1_50_0/libs/smart_ptr/scoped_ptr.htm) that seems unsafe to me:



```
    boost::scoped_ptr<Shoe> x(new Shoe);

```

This line of code will do these three things **in order**:


- Allocate heap memory for `Shoe`
- Call the constructor for `Shoe`
- Call the constructor for `boost::scoped_ptr<Shoe>`


If the constructor for `Shoe` throws an exception, memory will be leaked. *(see R. Martinho Fernandes answer)* The `scoped_ptr` won't handle the deallocation because it hasn't been constructed yet.


Is this an oversight? Or is there a solution that I've failed to notice?


","If the constructor fails, no memory is leaked. That's part of the semantics of `new`, no smart pointers involved:



```
struct Foo { Foo() { throw 23; } };
new Foo(); // no memory leaked

```

The added exception safety provided by `make_shared` comes from when you're initializing two `shared_ptr`s in an expression and *the two initializations are not sequenced*, as is the case in function call arguments:



```
struct Bar {
    Bar(bool fail) {
        if(fail) throw 17;
    }
}
f(shared_ptr<Bar>(new Bar(true)), shared_ptr<Bar>(new Bar(false)));

```

Since there is no sequencing between the evaluations of `new Bar(true)`, `shared_ptr<Bar>(new Bar(true))`, `new Bar(false)` and `shared_ptr<Bar>(new Bar(false))`, the following could happen:


1. `new Bar(false)` is evaluated and succeeds: memory is allocated;
2. `new Bar(true)` is evaluated and fails: it doesn't leak memory resulting from this evaluation;


No `shared_ptr` was constructed at this time, and so the memory allocated in #1 is now leaked.


"
"How to split monolithic node.js javascript
I have a growing *node.js* server application, and I'd like to get it split into several files. Here is a working code snippet demonstrating what monolithic *server.js* roughly looks like:



```
var express = require('express');
var app = express();
// other initialization code etc

//************** start of part to be moved foo.js

var fooTestData = {data: ""data"", id:1};

app.get(""/foo/ajax"", function(req, res) {
    res.json(fooTestData);
});

// more REST stuff and other foo-specific code

//************** end of part to be moved

// more stuff which remains in server.js

// http://localhost:8888/foo/ajax
app.listen(8888);

```

An ideal answer has two pieces of code with identical functionality: What *server.js* looks like after moving the indicated part, and what *foo.js* looks like with copied code and any needed extra code. 


","foo.js



```
var fooTestData = {data: ""data"", id:1};

exports.setApp = function (app) {

    app.get(""/foo/ajax"", function(req, res) {
        res.json(fooTestData);
    });

    // more REST stuff and other foo-specific code
};

```

new sever.js



```
var express = require('express');
var app = express();

require('./foo.js').setApp(app);

// other initialization code etc


// more stuff which remains in server.js

// http://localhost:8888/foo/ajax
app.listen(8888);

```

It's a simple example. If you need some things more complex you may want to take a look to other express program on github. It could be a good source of inspiration. 


"
"What does href expression  do?
I have seen the following href used in webpages from time to time. However, I don't understand what this is trying to do or the technique. Can someone elaborate please?



```
<a href=""javascript:;""></a>

```

","An `<a>` element is invalid HTML unless it has either an `href` or `name` attribute.


If you want it to render correctly as a link (ie underlined, hand pointer, etc), then it will only do so if it has a `href` attribute.


Code like this is therefore sometimes used as a way of making a link, but without having to provide an actual URL in the `href` attribute. The developer obviously wanted the link itself not to do anything, and this was the easiest way he knew.


He probably has some javascript event code elsewhere which is triggered when the link is clicked, and that will be what he wants to actually happen, but he wants it to look like a normal `<a>` tag link.


Some developers use `href='#'` for the same purpose, but this causes the browser to jump to the top of the page, which may not be wanted. And he couldn't simply leave the href blank, because `href=''` is a link back to the current page (ie it causes a page refresh).


There are ways around these things. Using an empty bit of Javascript code in the `href` is one of them, and although it isn't the best solution, it does work.


"
"How to insert in multiple tables with a single API call in Supabase
[![the simplified structure of my  tables - 2 main tables, one relation table.](https://i.stack.imgur.com/4DspE.png)](https://i.stack.imgur.com/4DspE.png)


This is the simplified structure of my tables - 2 main tables, one relation table.  

What's the best way to handle an insert API for this?
If I just have a Client and Supabase:



```
- First API call to insert book and get ID
- Second API call to insert genre and get ID
- Third API call to insert book-genre relation

```

This is what I can think of, but 3 API calls seems wrong.  

**Is there a way where I can do insert into these 3 tables with a single API call from my client**, like a single postgres function that I can call?  

Please share a general example with the API, thanks!


","Is there any reason you need to do this with a single call? I'm assuming from your structure that you're not going to create a new `genre` for every book you create, so most of the time, you're just inserting a `book` record and a `book_gen_rel` record. In the real world, you're probably going to have books that fall into multiple genres, so eventually you're going to be changing your function to handle the insert of a single book along with multiple genres in a single call.


That being said, there are two ways too approach this. You can make multiple API calls from the client (and there's really no problem doing this -- it's quite common). Second, you could do it all in a single call if you create a PostgreSQL function and call it with `.rpc()`.


Example using just client calls to insert a record in each table:



```
    const { data: genre_data, error: genre_error } = await supabase
      .from('genre')
      .insert([
        { name: 'Technology' }
      ]);
    const genre_id = genre_data[0].id;  
    const { data: book_data, error: book_error } = await supabase
      .from('book')
      .insert([
        { name: 'The Joys of PostgreSQL' }
      ]);
    const book_id = book_data[0].id;
    const { data: book_genre_rel_data, error: book_genre_rel_error } = await supabase
      .from('book_genre_rel_data')
      .insert([
        { book_id, genre_id }
      ]);

```

Here's a single SQL statement to insert into the 3 tables at once:



```
WITH genre AS (
  insert into genre (name) values ('horror') returning id
),
book AS (
  insert into book (name) values ('my scary book') returning id
)
insert into book_genre_rel (genre_id, book_id) 
  select genre.id, book.id from genre, book

```

Now here's a PostgreSQL function to do everything in a single function call:



```
CREATE OR REPLACE FUNCTION public.insert_book_and_genre(book_name text, genre_name text)
    RETURNS void language SQL AS
$$
  WITH genre AS (
    insert into genre (name) values (genre_name) returning id
  ),
  book AS (
    insert into book (name) values (book_name) returning id
  )
  insert into book_genre_rel (genre_id, book_id) 
    select genre.id, book.id from genre, book
$$

```

Here's an example to test it:



```
select insert_book_and_genre('how to win friends by writing good sql', 'self-help')

```

Now, if you've created that function (inside the Supabase Query Editor), then you can call it from the client like this:



```
const { data, error } = await supabase
  .rpc('insert_book_and_genre', {book_name: 'how I became a millionaire at age 3', genre_name: 'lifestyle'})

```

Again, I don't recommend this approach, at least not for the `genre` part. You should insert your genres first (they probably won't change) and simplify this to just insert a `book` and a `book_genre_rel` record.


"
"python: find all keys with a value
I have a dictionary like below:



```
mydict = {'a' : 'apple',
     'b' : 'bobb',
    'c' : {
         'd' : 'dog'
       },
      'e' : 'dog'
     }

```

Suppose I want to find all the keys with value '`dog`'In case of nesting, the keys should be separated by a dot.


So the output should be a list `['e', 'c.d']`


Now if I write below code in python 3, it only outputs '`e`'. 



```
print(list(mydict.keys())[list(mydict.values()).index('dog')])

```

How can I get the nested keys?


","You can use a recursion function like following:



```
def find_key(mydict, pre=tuple()):
    for key, value in mydict.items():
        if isinstance(value, dict):
            yield from find_key(value, pre=pre+(key,))
        elif value == 'dog':
            if pre:
                yield '.'.join(pre + (key,))
            else:
                yield key

```

test: 



```
In [23]: list(find_key(mydict))
Out[23]: ['e', 'c.d']
In [26]: mydict = {'a' : 'apple',
     'b' : 'bobb',
    'c' : {
         'd' : 'dog'
       },
      'e' : 'dog',
     'k':{'f':{'c':{'x':'dog'}}}}

In [27]: 

In [27]: list(find_key(mydict))
Out[27]: ['k.f.c.x', 'e', 'c.d']

```

"
"How and where the TCP connection has been created in httpwebrequest, and how is it related to servicepoint?
I am trying to find out when the TCP connection has been established while using HttpWebRequest, how these connections have been pooled and reused using ServicePoint.


I have looked at the system.dll, and tried to browse through the code using ILSpy and Reflector, somehow didn't see any references to sockets, establishing tcp connection etc.


Below I have pasted the decompiled code - can any please give me tips or redirect me so that I can understand:


1. When the TCP connection has been created?
2. How these connections are kept alive, pooled and reused using ServicePoint?


**Code snippet from HttpWebRequest of System.dll:**



```
public override Stream GetRequestStream()
    {
        TransportContext context;
        return this.GetRequestStream(out context);
    }

    public Stream GetRequestStream(out TransportContext context)
    {
        if (Logging.On)
        {
            Logging.Enter(Logging.Web, this, ""GetRequestStream"", """");
        }
        context = null;
        this.CheckProtocol(true);
        if ((this._WriteAResult == null) || !this._WriteAResult.InternalPeekCompleted)
        {
            lock (this)
            {
                if (this._WriteAResult != null)
                {
                    throw new InvalidOperationException(SR.GetString(""net_repcall""));
                }
                if (this.SetRequestSubmitted())
                {
                    throw new InvalidOperationException(SR.GetString(""net_reqsubmitted""));
                }
                if (this._ReadAResult != null)
                {
                    throw ((Exception) this._ReadAResult.Result);
                }
                this._WriteAResult = new LazyAsyncResult(this, null, null);
                this.Async = false;
            }
            this.CurrentMethod = this._OriginVerb;
            while (this.m_Retry && !this._WriteAResult.InternalPeekCompleted)
            {
                this._OldSubmitWriteStream = null;
                this._SubmitWriteStream = null;
                this.BeginSubmitRequest();
            }
            while (this.Aborted && !this._WriteAResult.InternalPeekCompleted)
            {
                if (!(this._CoreResponse is Exception))
                {
                    Thread.SpinWait(1);
                }
                else
                {
                    this.CheckWriteSideResponseProcessing();
                }
            }
        }
        ConnectStream connectStream = this._WriteAResult.InternalWaitForCompletion() as ConnectStream;
        this._WriteAResult.EndCalled = true;
        if (connectStream == null)
        {
            if (Logging.On)
            {
                Logging.Exception(Logging.Web, this, ""EndGetRequestStream"", this._WriteAResult.Result as Exception);
            }
            throw ((Exception) this._WriteAResult.Result);
        }
        context = new ConnectStreamContext(connectStream);
        if (Logging.On)
        {
            Logging.Exit(Logging.Web, this, ""GetRequestStream"", connectStream);
        }
        return connectStream;
    }

```

","K, after browsing through code some time I think I kind of understood the abstractions. Basically servicepoint, servicepoint manager, how the tcp connection has been created, connections have been pooled, queued etc. always confused me. Below information kind of helped me - hopefully this is useful for others who are curious or tried to understand these details:


**ServicePoint**: High level abstraction of 'connection' to a particular host (destination Host Ip:port)
(that's why for ex, function static ServicePoint **FindServicePoint**(string host, int port) is defined in servicePointManger.


**ServicePointManager**: as the name indicates its the global (static) class which manages service points.


**Connection (internal class)**: Basically this is the one I think that represents TCP connection. it basically derives from System.Net.PoolStream (internal class - it has the definitions of the sockets its using) which derives from stream. 


**ConnectionGroup (internal class)**: Each HttpWebRequest is associated with a connection group. 
(basically based on connectionLimit it creates at most **connectionLimit** (can be configured globally through ServicePointManager, and also per httpwebrequest using its servicePoint property) number of connection objects per httpwebrequest) 


If the connection limit is reached, its simply queued and passed to the wire (most likely - but still didn't get the code which does this).


And if you are connecting to service on the local machine, the servicepoint.connectionlimit no longer equal to servicepointmanager.defaultconnectionlimit. it defaults to; int.Maxvalue (**2147483647** or **7FFFFFFF**) ( you may refer to: <http://blogs.microsoft.co.il/idof/2011/06/20/servicepointmanagerdefaultconnectionlimit-2-depends/> )


**Update:**


Looks like following two links are useful too:


[System.Net.ServicePointManager.DefaultConnectionLimit and .MaxServicePointIdleTime](https://stackoverflow.com/questions/3363183/system-net-servicepointmanager-defaultconnectionlimit-and-maxservicepointidleti)


<http://blogs.msdn.com/b/jpsanders/archive/2009/05/20/understanding-maxservicepointidletime-and-defaultconnectionlimit.aspx>


Best Regards!


"
"Replace Duplicates in Clojure Vector
I am trying to replace duplicates in a vector with empty strings. However, the only functions I can find are to remove duplicates, not replace them. How can I take 


`[""Oct 2016"" ""Oct 2016"" ""Nov 2016"" ""Nov 2016"" ""Nov 2016"" ""Nov 2016""]`


and output:


`[""Oct 2016"" """" ""Nov 2016"" """" """" """"]`


Everything I can find will return `[""Oct 2016"" ""Nov 2016""]` I'm currently achieving the desired output by doing a nested `doseq`, but it seems inefficient. Is there a better way to achieve this?
Thanks!


","Here is a strategy for a solution.


1. `loop` over the items of the vector.
2. Maintain a `set` of visited items. It can be used to check for uniqueness.
3. For each item: if the set contains the current item then insert `""""` into the result vector.
4. If the current item is unique then insert it into the result vector and also the set.
5. Return the result vector when all items are visited.
6. Optionally: Use a [transient](https://clojure.org/reference/transients) result vector for better performance.


Code:



```
(defn duplicate->empty [xs]
  (loop [xs     (seq xs)
         result []
         found #{}]
        (if-let [[x & xs] (seq xs)]
          (if (contains? found x)
            (recur xs (conj result """") found)
            (recur xs (conj result x) (conj found x)))
          result)))

```

Calling it:



```
(duplicate->empty [""Oct 2016"" ""Oct 2016"" ""Nov 2016"" ""Nov 2016"" ""Nov 2016"" ""Nov 2016""])
=> [""Oct 2016"" """" ""Nov 2016"" """" """" """"]

```

"
"How to find a text inside stored procedures across multiple databases
I have 40 + databases, and I want to find procedures in all databases that use the text `sp_reset_data`. [This query helped me a lot](https://stackoverflow.com/questions/674623/how-to-find-a-text-inside-sql-server-procedures-triggers):



```
DECLARE @Search varchar(255)
SET @Search='sp_reset_data'

SELECT DISTINCT
    o.name AS Object_Name,o.type_desc
    FROM sys.sql_modules        m 
        INNER JOIN sys.objects  o ON m.object_id=o.object_id
    WHERE m.definition Like '%'+@Search+'%'
    ORDER BY 2,1

```

But, this get the procedures only for the current database. Is there a way to improve this kind of query to look in every server's DB without manually changing the current DB?


","
```
DECLARE @sql NVARCHAR(MAX);
SET @sql = N'';

SELECT @sql = @sql + 'SELECT db = ''' + name + ''', o.name, o.type_desc
  FROM ' + QUOTENAME(name) + '.sys.sql_modules AS m
  INNER JOIN ' + QUOTENAME(name) + '.sys.objects AS o
  ON m.[object_id] = o.[object_id]
  WHERE m.definition LIKE N''%'' + @Search + ''%'' 
  ORDER BY o.type_desc, o.name;'
FROM sys.databases
WHERE database_id > 4 AND state = 0; -- online

EXEC sp_executesql @sql, N'@Search NVARCHAR(255)', N'sp_reset_data';

```

Strictly speaking, if you only want procedures, then it is a little simpler (the above will also include functions, triggers, even views):



```
DECLARE @sql NVARCHAR(MAX);
SET @sql = N'';

SELECT @sql = @sql + 'SELECT db = ''' + name + ''', o.name
  FROM ' + QUOTENAME(name) + '.sys.sql_modules AS m
  INNER JOIN ' + QUOTENAME(name) + '.sys.procedures AS o
  ON m.[object_id] = o.[object_id]
  WHERE m.definition LIKE N''%'' + @Search + ''%'' 
  ORDER BY o.name;'
FROM sys.databases
WHERE database_id > 4 AND state = 0; -- online

EXEC sp_executesql @sql, N'@Search NVARCHAR(255)', N'sp_reset_data';

```

"
"Redirect to a specific page after successful action - Redux saga
I have a user profile page where the user can update the data. Once the data is submitted, it should be redirected to dashboard page. 


MyProfile.js



```
  handleProfileUpdate = () => {
    let user = this.state;
    this.props.updateUserProfile(user);
  }

```

Action is defined as follows



```
export const updateUserProfile = (obj) => ({
  type: types.USER_PROFILE_UPDATE_REQUEST,
  payload: obj
});

```

Saga file is defined as follows.



```
function* updateUserProfile(action) {
  try {
    let data = action.payload;
    yield call(userProvider.updateUserProfile, data);

    // dispatch a success action to the store
    yield put({ type: types.USER_PROFILE_UPDATE_SUCCESS, data });
    yield put(push('/dashboard'));
    yield put(constants.successMsg(userProfileUpdateSuccess));

  } catch (error) {
    // dispatch a failure action to the store with the error
    yield put(constants.DEFAULT_ERROR_MSG);
  }
}

export function* watchUserProfileUpdateRequest() {
  yield takeLatest(types.USER_PROFILE_UPDATE_REQUEST, updateUserProfile);
}


export default function* userSaga() {
  yield all([
    watchUserProfileRequest(),
    watchUserProfileUpdateRequest()
  ]);
}

```

But the code **yield put(push('/dashboard'))** is not redirecting on to the page.


Any idea on how to fix this?


","I founded a solution myself


history.js



```
import createHistory from 'history/createBrowserHistory';
const history = createHistory();
export default history;

```

Import history file into the saga



```
import history from '../../history';

```

Modify the watcher saga as follows.



```
function* updateUserProfile(action) {
  try {
    let data = action.payload;
    yield call(userProvider.updateUserProfile, data);

    // dispatch a success action to the store
    yield put({ type: types.USER_PROFILE_UPDATE_SUCCESS, data });
    yield call(forwardTo, '/dashboard');
    yield put(constants.successMsg(userProfileUpdateSuccess));

  } catch (error) {
    // dispatch a failure action to the store with the error
    yield put(constants.DEFAULT_ERROR_MSG);
  }
}

```

Now, you need to define the forwardTo function as follows.



```
function forwardTo(location) {
  history.push(location);
}

```

"
"JavaScript gets rid of trailing slash in ![]() tag
I have the following JS code:



```
validateConfigName.html('<img src=""/rta/images/check-24.png"" />');

```

But when it executes in Firefox I see this as the generated code:



```
<img src=""/rta/images/check-24.png"">

```

Why?


","In HTML the `<img>` tag is supposed to be `<img>`, in XHTML it'll be `<img />`...so depending on what DOCTYPE your page is using, this will vary.


[From the HTML 4.0 Spec for `<img>`](http://www.w3.org/TR/REC-html40/struct/objects.html#h-13.2):



> 
> Start tag: **required**, End tag: **forbidden**
> 
> 
> 


[In XHTML](http://www.w3.org/TR/xhtml1/) elements must be closed: 



> 
> Well-formedness is a new concept introduced by [XML]. Essentially this means that all elements must either **have closing tags** or be written in a special form (as described below), and that all the elements must nest properly.
> 
> 
> 


"
"getUserMedia - Cannot access camera from two browsers at the same time
I've been experimenting with webRTC for my next project to create a video chat and testing has been difficult. I have this simple code to access the camera: 



```
        navigator.getUserMedia = navigator.getUserMedia ||
                                 navigator.webkitGetUserMedia ||
                                 navigator.mozGetUserMedia ||
                                 navigator.msGetUserMedia;

        var video = document.querySelector('#av-chat video');

        if (navigator.getUserMedia) {
          navigator.getUserMedia({audio: true, video: true}, function(stream) {
            video.src = window.URL.createObjectURL(stream);
          }, errorCallback);
        }

```

which works fine on chrome and mozilla but when I try to do it together, it doesn't work. What I mean by together is like opening the same file with this code in two browsers. It seems that when one browser has access to the camera, it blocks it for anyone else.


I have not seen this issue discussed on the internet so I was wondering, is it just me? If not, is there a solution?


","This is a limitation on Windows. On Mac OS you can use the same camera in multiple applications at the same time but not in Windows unfortunately. You have a few options here:


1. Buy a Mac
2. Get a second USB camera to use in your second application while testing.
3. Install some [software for splitting your camera](http://www.addictivetips.com/windows-tips/how-to-split-a-webcam-between-multiple-applications-easily/) or [creating a virtual camera](https://manycam.com/).
4. Use fake devices in either Chrome or Firefox (or both) for testing (like Philipp suggested). It's actually quite easy to turn on fake devices in Firefox and it's free so that's probably the best option if you don't already have another USB camera sitting around.
	1. Just go to `about:config` in your URL bar
	2. Accept the warning message
	3. Add a new Boolean value. You do this by right clicking and choosing New->Boolean.
	4. Then enter `media.navigator.streams.fake`
	5. Choose true as the value.
	[![enter image description here](https://i.stack.imgur.com/d03ZY.png)](https://i.stack.imgur.com/d03ZY.png)
	Now when you use your camera you will get a fake video and audio sound.


"
"List all files in a directory given a regular expression / a set of extensions (Matlab)
I have a regular expression defining the filenames of interest. What is the best way to list all files in a directory that match this condition?


My attempt at this is:



```
f = dir(DIR);
f = {f([f.isdir] == 0).name};
result = f(~cellfun(@isempty, regexpi(f, '.*(avi|mp4)')));

```

However, I wonder if there is a faster and/or cleaner solution to this? 


Is is possible to simplify it if instead of a regular expression I have only a list of possible file extensions?


","Fundamentally your approach is what I would go for. However, your lines of code can be **simplified** to (directories are lost in the regex and empty cells in the final concatenation):



```
f = dir('C:\directory');
f = regexpi({f.name},'.*txt|.*pdf','match');
f = [f{:}];

```

Also, note that the function `dir()` accepts **wildcards** (`*`) but not multiple extensions:



```
dir('C:\directory\*.avi')

```

This means you can retrieve immediately only those files that match an extension, however you have to **loop** for the number of extensions:



```
d   = 'C:\users\oleg\desktop';
ext = {'*.txt','*.pdf'};
f   = [];
for e = 1:numel(ext)
    f = [f; dir(fullfile(d,ext{e}))];
end

```

## Alternative (not recommended)



```
ext = {'*.txt','*.pdf'};
str = ['!dir ' sprintf('%s ',ext{:}) '/B'];
textscan(evalc(str),'%s','Delimiter','')

```

where `str` is `!dir *.txt *.pdf /B` and `evalc()` captures the evaluation of the string and `textscan()` parses it.


"
"Force method call on null variable to throw a NullPointerException

> 
> **Possible Duplicate:**  
> 
> [How come invoking a (static) method on a null reference doesn’t throw NullPointerException?](https://stackoverflow.com/questions/3293353/how-come-invoking-a-static-method-on-a-null-reference-doesnt-throw-nullpointe)  
> 
> [Static fields on a null reference in Java](https://stackoverflow.com/questions/11579953/static-fields-on-a-null-reference-in-java) 
> 
> 
> 


I tried the code from [this old video](http://www.metacafe.com/watch/370425/impossible_java_program/):



```
class Impossible {
    public static void main(String[] args) {
        Thread t = null;
        System.out.println(t.currentThread().getName());
    }
}

```

Output: `main`


Well, just what the heck is that?! Does `java.lang.Thread` breach any the `NullPointerException` rule?


But what I'm most interested in it: **How can I make that variable behave to throw a NullPointerException?**


","
> 
> Does java.lang.Thread breach any the NullPointerException rule?
> 
> 
> 


**No**, reason for a `NPE` been thrown is not related to a `class`. It is related to an `instance` of that class, on which invocation is done. Also, it depends upon which type of method, or field you are accessing.


What is happening here is, `currentThread()` is a static method of Thread class. Which is boudn to a class, rather than an instance. So, even if you invoke it on a reference of Thread class, it is actually invoked on class name.


So, 



```
Thread t = null
t.currentThread();

```

is actually invoked as: -



```
Thread.currentThread();

```

So, when accessing a static member through an object reference expression, only the declared type of the reference matters. This means that:


- It doesn't matter if the reference is actually pointing to null, since no
instance is required.
- If the reference is not null, it doesn't matter what the type
of the object the reference is pointing to, there is no dynamic dispatch.




---



> 
> How can I make that variable behave to throw a NullPointerException?
> 
> 
> 


Well, the current print statement will never throw a `NPE`. The first part is already explained above. Now, let's move ahead.



```
Thread.currentThread();

```

The above invocation will never return `null`. It always returns the current thread instance. And in Java, you are always inside one or the other thread. Even when inside `public static void main` method, you are executing the `Main Thread`. So, `currentThread` can't be `null`.


And hence, further invocation: -



```
Thread.currentThread().getName();

```

will work fine, and return the name of current thread.


"
"Create a link on the ios filesystem with Objective-C
I'm trying to create a comic reader app which supports both online reading & off-line reading (by downloading). 


I've found [MWPhotoBrowser](https://github.com/mwaterfall/MWPhotoBrowser) as my image viewer, which supports [SDWebImage](https://github.com/rs/SDWebImage) as the image cache. 


My problem is that if some user read little part of a comic on-line, and then they decides to download it for offline reading. Since the already read part is cached by SDWebImage, I don't wanna download them again from the web server. But since user asks to download them locally, I don't wanna keep them on the image cache neither as this will make the downloaded images out of our hand. 


Copy the image from the image cache to the place I wanna put is a feasible solution, but it takes storage space. So, what I'm trying to do is cut the image from the image cache to the right place, and then make some soft-link in the image cache. In this way, there should be only one image copy in the storage file system.


","You don't want a symbolic (soft) link. A symlink doesn't prevent the original file (in the cache) from being deleted. If the original file is deleted, you'll get an error when you try to open it through the symlink.


You want a hard link. You can create a hard link using `-[NSFileManager linkItemAtURL:toURL:error:]` or `-[NSFileManager linkItemAtPath:toPath:error:]`. Take a look at the [*NSFileManager Class Reference*](http://developer.apple.com/library/ios/#documentation/cocoa/Reference/Foundation/Classes/NSFileManager_Class/Reference/Reference.html).


If you need to learn more about hard links and symbolic links, you can quickly find a lot of information by searching for “hard link” or “symbolic link” in your favorite search engine.


### NOTE FOR FUTURE SEARCHERS


You can create a symbolic link using `-[NSFileManager createSymbolicLinkAtURL:withDestinationURL:error:]` or `-[NSFileManager createSymbolicLinkAtPath:withDestinationPath:error:]`.


"
"Is there a Julia equivalent of Python's %timeit
In Python, a convenient way to evaluate the performance of code is [`timeit`](https://docs.python.org/2/library/timeit.html). In Julia we have `@time`, but this has the disadvantage of only running a piece of code once, which means you need to execute it several times to get a good picture of code performance. Is there a better way to time code in Julia, more akin to Python's `timeit`?


","The package `BenchmarkTools` has `@benchmark` and `@btime` which has a statistical way of determining the number of runs. 



```
julia> A = rand(100,100);
julia> B = rand(100,100);

julia> using BenchmarkTools

julia> @benchmark A*B
BenchmarkTools.Trial:
  memory estimate:  78.20 KiB
  allocs estimate:  2
  --------------
  minimum time:     48.302 μs (0.00% GC)
  median time:      72.015 μs (0.00% GC)
  mean time:        74.314 μs (6.52% GC)
  maximum time:     3.232 ms (95.17% GC)
  --------------
  samples:          10000
  evals/sample:     1

julia> @btime A*B;
  49.180 μs (2 allocations: 78.20 KiB)

```

"
"Deploy to Azure Web App from Multiple Repositories
Looking for a way to deploy to an azure web app from 3 separate github repositories. One of the repositories is the main website, the other two are common libraries used by other projects. We want the deployment to pull down from all three repositories into an artifact directory, run msbuild and copy over the artifacts to wwwroot.


I am assuming i will need a custom .deployment file along with a kudu batch file. But I can't find any examples on how to pull down from multiple github repositories.


edit: spelling


","According to your description, you could create your own .deployment and deploy.cmd file to firstly clone the second and third github repositories common libraries to the web app folder then use msbuild command to build it.


Notice: You couldn't clone the second and third github repositories to artifact directory. Because, you have 3 separate github repositories, each repository has its own .git file. We couldn't combine this three. So I suggest you could firstly clone the second and third repository to new folder and use msbuild to build it.


More details, you could refer to below steps:


Firstly, I suggest you could download the deployment script from the KUDU console. 


Notice:After you have already deployed from the git, this file could be downloaded.We use this file as a example to add codes. If you doesn't have you could follow my deployment.cmd. Remember changing the project name.


As below:


Open Kudu.


[![enter image description here](https://i.stack.imgur.com/9ttsP.png)](https://i.stack.imgur.com/9ttsP.png)


Download the deploy script.


[![enter image description here](https://i.stack.imgur.com/MGSEX.png)](https://i.stack.imgur.com/MGSEX.png)


Then you could change find it contains the .deployment and deploy.cmd file.


Add below codes to the deploy.cmd file and add this two file into your git folder to push.



```
  git clone https://github.com/{yourgithubname}/BrandoGitTestLibrary.git D:\home\site\{foldername}

  echo  second project
  :: 1. Start restore second project
  dotnet restore ""D:\home\site\{foldername}\BrandoGitTestLibrary\BrandoGitTestLibrary.csproj""

  :: 2. Build to DEPLOYMENT_TEMP wait copy
  dotnet build  ""D:\home\site\{foldername}\BrandoGitTestLibrary\BrandoGitTestLibrary.csproj"" --output ""%DEPLOYMENT_TEMP%""

```

This code is used to build your common library to tempfolder wait for copying.


After the main website published, it will copy the tempfolder to the wwwrot.


The total deploy.cmd file.


Notice: This example just build two repositories. If you want to build and publish 3 repositories, you just need add the codes to clone again.



```
    git clone https://github.com/BrandoTest/BrandoGitTestLibrary.git D:\home\site\sencondproject

    echo  second project
    :: 3. Start restore second project
    dotnet restore ""D:\home\site\sencondproject\BrandoGitTestLibrary\BrandoGitTestLibrary.csproj""

    :: 4. Build and publish
    dotnet build  ""D:\home\site\sencondproject\BrandoGitTestLibrary\BrandoGitTestLibrary.csproj"" --output ""%DEPLOYMENT_TEMP%""




@if ""%SCM_TRACE_LEVEL%"" NEQ ""4"" @echo off

:: ----------------------
:: KUDU Deployment Script
:: Version: 1.0.15
:: ----------------------

:: Prerequisites
:: -------------

:: Verify node.js installed
where node 2>nul >nul
IF %ERRORLEVEL% NEQ 0 (
  echo Missing node.js executable, please install node.js, if already installed make sure it can be reached from current environment.
  goto error
)

:: Setup
:: -----

setlocal enabledelayedexpansion

SET ARTIFACTS=%~dp0%..\artifacts

IF NOT DEFINED DEPLOYMENT_SOURCE (
  SET DEPLOYMENT_SOURCE=%~dp0%.
)

IF NOT DEFINED DEPLOYMENT_TARGET (
  SET DEPLOYMENT_TARGET=%ARTIFACTS%\wwwroot
)

IF NOT DEFINED NEXT_MANIFEST_PATH (
  SET NEXT_MANIFEST_PATH=%ARTIFACTS%\manifest

  IF NOT DEFINED PREVIOUS_MANIFEST_PATH (
    SET PREVIOUS_MANIFEST_PATH=%ARTIFACTS%\manifest
  )
)

IF NOT DEFINED KUDU_SYNC_CMD (
  :: Install kudu sync
  echo Installing Kudu Sync
  call npm install kudusync -g --silent
  IF !ERRORLEVEL! NEQ 0 goto error

  :: Locally just running ""kuduSync"" would also work
  SET KUDU_SYNC_CMD=%appdata%\npm\kuduSync.cmd
)
IF NOT DEFINED DEPLOYMENT_TEMP (
  SET DEPLOYMENT_TEMP=%temp%\___deployTemp%random%
  SET CLEAN_LOCAL_DEPLOYMENT_TEMP=true
)

IF DEFINED CLEAN_LOCAL_DEPLOYMENT_TEMP (
  IF EXIST ""%DEPLOYMENT_TEMP%"" rd /s /q ""%DEPLOYMENT_TEMP%""
  mkdir ""%DEPLOYMENT_TEMP%""
)

IF DEFINED MSBUILD_PATH goto MsbuildPathDefined
SET MSBUILD_PATH=%ProgramFiles(x86)%\MSBuild\14.0\Bin\MSBuild.exe
:MsbuildPathDefined
::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::
:: Deployment
:: ----------

echo Handling ASP.NET Core Web Application deployment.

:: 1. Restore nuget packages
call :ExecuteCmd dotnet restore ""%DEPLOYMENT_SOURCE%\TestForCore2.csproj""
IF !ERRORLEVEL! NEQ 0 goto error

:: 2. Build and publish
call :ExecuteCmd dotnet publish ""%DEPLOYMENT_SOURCE%\TestForCore2.csproj"" --output ""%DEPLOYMENT_TEMP%"" --configuration Release
IF !ERRORLEVEL! NEQ 0 goto error

:: 5. KuduSync
call :ExecuteCmd ""%KUDU_SYNC_CMD%"" -v 50 -f ""%DEPLOYMENT_TEMP%"" -t ""%DEPLOYMENT_TARGET%"" -n ""%NEXT_MANIFEST_PATH%"" -p ""%PREVIOUS_MANIFEST_PATH%"" -i "".git;.hg;.deployment;deploy.cmd""
IF !ERRORLEVEL! NEQ 0 goto error

::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::::
goto end

:: Execute command routine that will echo out when error
:ExecuteCmd
setlocal
set _CMD_=%*
call %_CMD_%
if ""%ERRORLEVEL%"" NEQ ""0"" echo Failed exitCode=%ERRORLEVEL%, command=%_CMD_%
exit /b %ERRORLEVEL%

:error
endlocal
echo An error has occurred during web site deployment.
call :exitSetErrorLevel
call :exitFromFunction 2>nul

:exitSetErrorLevel
exit /b 1

:exitFromFunction
()

:end
endlocal
echo Finished successfully.

```

Result:


[![enter image description here](https://i.stack.imgur.com/KcOG7.png)](https://i.stack.imgur.com/KcOG7.png)


"
"Why is appending on a preallocated slice quicker than indexing?
I wrote the following Go benchmarks (Go1.9.2, Linux amd64) to test indexing vs appending:



```
func BenchmarkIndex(b *testing.B) {
    for i := 0; i < b.N; i++ {
        existing := make([]int64, 1000, 1000)
        init := make([]int64, 1000) // len 1000, cap 1000
        for index, element := range existing {
            init[index] = element
        }
    }
}

func BenchmarkAppend(b *testing.B) {
    for i := 0; i < b.N; i++ {
        existing := make([]int64, 1000, 1000)
        init := make([]int64, 0, 1000) // len 0, capacity 1000
        for _, element := range existing {
            init = append(init, element)
        }
    }
}

```

And got the following results:



```
goos: linux
goarch: amd64
BenchmarkIndex-4         1000000          2183 ns/op
BenchmarkAppend-4        1000000          1933 ns/op
PASS

```

I'm a little confused on why `append` seems to perform better than in general than indexing into the slice. Does it have anything to do with omitting the `index` inside the range? If not, what is append doing under the hood to make it quicker than just indexing?


For reference, [here is append's source](https://github.com/golang/go/blob/master/src/cmd/compile/internal/gc/ssa.go#L2167). Thanks! 


","On computer one:



```
$ go test same_test.go -run=! -bench=. -benchmem -count=3
goos: linux
goarch: amd64
BenchmarkIndex-4     1000000    1290 ns/op    0 B/op    0 allocs/op
BenchmarkIndex-4     1000000    1290 ns/op    0 B/op    0 allocs/op
BenchmarkIndex-4     1000000    1288 ns/op    0 B/op    0 allocs/op
BenchmarkAppend-4    1000000    1299 ns/op    0 B/op    0 allocs/op
BenchmarkAppend-4    1000000    1302 ns/op    0 B/op    0 allocs/op
BenchmarkAppend-4    1000000    1301 ns/op    0 B/op    0 allocs/op
PASS
ok      command-line-arguments  7.868s
$ go version
go version devel +38c725b148 Sun Nov 5 17:30:11 2017 +0000 linux/amd64
$ 

```

The difference between `BenchmarkAppend` and `BenchmarkIndex` is:



```
((1299+1302+1301) - (1290+1290+1288)) * 100 / (1290+1290+1288) =  +0.879%

```

They are the same.


On computer two:



```
$ go test same_test.go -run=! -bench=. -benchmem -count=3
goos: linux
goarch: amd64
BenchmarkIndex-8     2000000    706 ns/op    0 B/op    0 allocs/op
BenchmarkIndex-8     2000000    713 ns/op    0 B/op    0 allocs/op
BenchmarkIndex-8     2000000    710 ns/op    0 B/op    0 allocs/op
BenchmarkAppend-8    2000000    711 ns/op    0 B/op    0 allocs/op
BenchmarkAppend-8    2000000    718 ns/op    0 B/op    0 allocs/op
BenchmarkAppend-8    2000000    718 ns/op    0 B/op    0 allocs/op
PASS
ok      command-line-arguments  12.945s
$ go version
go version devel +38c725b148 Sun Nov 5 17:30:11 2017 +0000 linux/amd64
$ 

```

The difference between `BenchmarkAppend` and `BenchmarkIndex` is:



```
((711+718+718) - (706+713+710)) * 100 / (706+713+710) =  +0.845%

```

They are the same.




---


Let's measure one thing at a time, indexing versus appending over a preallocated slice, which avoids confounding issues like memory management.



```
package main

import ""testing""

func BenchmarkIndexing(b *testing.B) {
    existing := make([]int64, 1000, 1000)
    init := make([]int64, 1000) // len 1000, cap 1000
    b.ReportAllocs()
    b.ResetTimer()
    for i := 0; i < b.N; i++ {
        init = init[:cap(init)]
        for index, element := range existing {
            init[index] = element
        }
    }
}

func BenchmarkAppending(b *testing.B) {
    existing := make([]int64, 1000, 1000)
    init := make([]int64, 0, 1000) // len 0, capacity 1000
    b.ReportAllocs()
    b.ResetTimer()
    for i := 0; i < b.N; i++ {
        init = init[:0]
        for _, element := range existing {
            init = append(init, element)
        }
    }
}

```

Output:



```
$ go test same_test.go -run=! -bench=ing -benchmem -count=3
goos: linux
goarch: amd64
BenchmarkIndexing-4     2000000    736 ns/op    0 B/op    0 allocs/op
BenchmarkIndexing-4     2000000    735 ns/op    0 B/op    0 allocs/op
BenchmarkIndexing-4     2000000    735 ns/op    0 B/op    0 allocs/op
BenchmarkAppending-4    2000000    733 ns/op    0 B/op    0 allocs/op
BenchmarkAppending-4    2000000    733 ns/op    0 B/op    0 allocs/op
BenchmarkAppending-4    2000000    733 ns/op    0 B/op    0 allocs/op
PASS
ok      command-line-arguments  13.303s
$ go version
go version devel +38c725b148 Sun Nov 5 17:30:11 2017 +0000 linux/amd64
$

```

The difference between `BenchmarkAppend` and `BenchmarkIndex` is:



```
((733+733+733) - (736+735+735)) * 100 / (736+735+735) =  −0.317%

```

They are the same.


"
"Mongoose: how to set a schema-field to be the ID?
Given the following schema: 



```
var UserSchema = new Schema({
   , email   :  { type: String }
   , passwordHash   :  { type: String }
   , roles  :  { type: [String] }
});

```

I'd like `email` to be the key. 
How can I define this? 


I could do: 



```
var UserSchema = new Schema({
       , _id:  { type: String }
       , passwordHash   :  { type: String }
       , roles  :  { type: [String] }
    });

```

so MongoDB would recognize it as the id-field, and adapt my code to refer to `_id` instead of `email` but that doesn't feel clean to me. 


Anyone?


","Since you're using Mongoose, one option is to use the email string as the `_id` field and then add a [virtual field](http://mongoosejs.com/docs/guide.html#virtuals) named `email` that returns the `_id` to clean up the code that uses the email. 



```
var userSchema = new Schema({
    _id: {type: String},
    passwordHash: {type: String},
    roles: {type: [String]}
});

userSchema.virtual('email').get(function() {
    return this._id;
});

var User = mongoose.model('User', userSchema);

User.findOne(function(err, doc) {
    console.log(doc.email);
});

```

Note that a virtual field is not included by default when converting a Mongoose doc to a plain JS object or JSON string. To include it you have to set the `virtuals: true` option in the [`toObject()`](http://mongoosejs.com/docs/api.html#document_Document-toObject) or [`toJSON()`](http://mongoosejs.com/docs/api.html#document_Document-toJSON) call:



```
var obj = doc.toObject({ virtuals: true });
var json = doc.toJSON({ virtuals: true });

```

"
"Popup a shape in Excel Desired Cell
I have been trying to popup a shape whenever double click on a cell and I have wrote a code but its not working for me.


All I want is that Whenever I double click on a cell between `Range(""A1""A100000"")` then shape will be popup otherwise it will be hidden.


My code:



```
Private Sub Worksheet_BeforeDoubleClick(ByVal Target As Range, Cancel As Boolean)
     If Intersect(Target, Range(""A1:A5000"")) Is Nothing Then
         ActiveSheet.Shapes(""Group 4"").Visible = True
     End If
End Sub

```

","
> 
> I have been trying to popup a shape whenever DoubleClick on a cell and i have wrote a code but its not working for me.
> 
> 
> 


You have to use `If Not Intersect` to check if user clicked on `A1:A5000`.


Is this what you are trying?



```
Option Explicit

Private Sub Worksheet_BeforeDoubleClick(ByVal Target As Range, Cancel As Boolean)
    If Not Intersect(Target, Range(""A1:A5000"")) Is Nothing Then
        Shapes(""Group 4"").Visible = True
    Else
        Shapes(""Group 4"").Visible = False
    End If
End Sub

```

**EDIT**



> 
> Thank you very much its working great but its not moving cell to cell. I mean when i click on cell A100 it remains still where it was before. – Mento 4 mins ago
> 
> 
> 


That is because you are not moving the shape in the code. Is this what you are trying?



```
Option Explicit

Private Sub Worksheet_BeforeDoubleClick(ByVal Target As Range, Cancel As Boolean)
    If Not Intersect(Target, Range(""A1:A5000"")) Is Nothing Then
        With Shapes(""Group 4"")
            .Top = Range(""B"" & Target.Row).Top
            .Left = Range(""B"" & Target.Row).Left
            .Visible = True
        End With
    Else
        Shapes(""Group 4"").Visible = False
    End If
End Sub

```

**Note**: If you do not want the cursor blinking when you double click in `A1:A5000`, then use `Cancel = True` in the first part of the `If`


"
"Dynamic basedir for Ant in Eclipse
The title might not be the best so let me explain what I am trying to do.


I have made an Ant buildfile that will help me with Maven goals. It's so much easier having to scope to Ant view and choose what I want to do. The biggest benefit with this is that I can use it in almost any project. Whether I deploy to tomcat or jboss or sakai(:deploy) or I handle mutiple instaces of the same server or skip tests... worst case I just change some path properties.


The drawback right now is that I have to keep a copy of this buildfile in every project.


What I am trying to do is have only 1 buildfile, in the workspace, and dinamicaly optain, from eclipse, the current project that I am working on. Be it module or parent I can refine that later.
So basically change the basedir for ant tasks based on the selected project in eclipse.


I have tried Ant Runtime Properties but for some reason properties like 



> 
> ${project\_path}
> 
> 
> 


fail to give me what their description say. I get this:



> 
> Variable references empty selection: ${project\_path}
> 
> 
> 


I hope it is clear what I am trying to do. So my question, I know it's possible, I'm just missing something and I hope some of you can help me with this.


I believe the trouble is in the Ant Runtime properties. I'm not 100% sure how I should use those.


Thank you!


","**EDIT** after comment and further investigation..  

Make the projectdir available as basedir property for ant like that :  

`Window > Preferences > Ant > Runtime > Properties`  

and create a property named basedir with value either :  



`${project_loc}`  

or  

`${workspace_loc}/${project_name}`  



and 



```
<project basedir=""${basedir}"">
 <echo>$${basedir} = ${basedir}</echo>
</project>

```

will work as expected, means echoing eclipse/yourworkspace/projectdir


Maybe there are other ways, i.e. via [Ant Addon Ant4Eclipse](http://www.ant4eclipse.org/), which aims to make Eclipse settings available for ant, never used it.


"
"How to link HTML5 form action to Controller ActionResult method in ASP.NET MVC 4
I have a basic form for which I want to handle buttons inside the form by calling the `ActionResult` method in the View's associated `Controller` class. Here is the following HTML5 code for the form:



```
<h2>Welcome</h2>

<div>

    <h3>Login</h3>

    <form method=""post"" action= <!-- what goes here --> >
        Username: <input type=""text"" name=""username"" /> <br />
        Password: <input type=""text"" name=""password"" /> <br />
        <input type=""submit"" value=""Login"">
        <input type=""submit"" value=""Create Account""/>
    </form>

</div>

<!-- more code ... -->

```

The corresponding `Controller` code is the following:



```
[HttpPost]
public ActionResult MyAction(string input, FormCollection collection)
{
    switch (input)
    {
        case ""Login"":
            // do some stuff...
            break;
        case ""Create Account""
            // do some other stuff...
            break;
    }

    return View();
}

```

","you make the use of the HTML Helper and have



```
    @using(Html.BeginForm())
    {
        Username: <input type=""text"" name=""username"" /> <br />
        Password: <input type=""text"" name=""password"" /> <br />
        <input type=""submit"" value=""Login"">
        <input type=""submit"" value=""Create Account""/>
    }

```

or use the Url helper



```
<form method=""post"" action=""@Url.Action(""MyAction"", ""MyController"")"" >

```

`Html.BeginForm` has several (13) overrides where you can specify more information, for example, a normal use when uploading files is using:



```
@using(Html.BeginForm(""myaction"", ""mycontroller"", FormMethod.Post, new {enctype = ""multipart/form-data""}))
{
    < ... >
}

```



---


If you don't specify any arguments, the `Html.BeginForm()` will create a `POST` form that **points to your current controller and current action**. As an example, let's say you have a controller called `Posts` and an action called `Delete`



```
public ActionResult Delete(int id)
{
   var model = db.GetPostById(id);
   return View(model);
}

[HttpPost]
public ActionResult Delete(int id)
{
    var model = db.GetPostById(id);
    if(model != null) 
        db.DeletePost(id);

    return RedirectToView(""Index"");
}

```

and your html page would be something like:



```
<h2>Are you sure you want to delete?</h2>
<p>The Post named <strong>@Model.Title</strong> will be deleted.</p>

@using(Html.BeginForm())
{
    <input type=""submit"" class=""btn btn-danger"" value=""Delete Post""/>
    <text>or</text>
    @Url.ActionLink(""go to list"", ""Index"")
}

```

"
"How to insert jsonb[] data into column using pg-promise
Given a table with a column of type `jsonb[]`, how do I insert a json array into the column? 


Using the provided formatters `:array`, `:json` won't work in this instance - unless I am missing the correct combination or something.



```
const links = [
    {
        title: 'IMDB',
        url: 'https://www.imdb.com/title/tt0076759'
    },
    {
        title: 'Rotten Tomatoes',
        url: 'https://www.rottentomatoes.com/m/star_wars'
    }
];

 const result = await db.none(`INSERT INTO tests (links) VALUES ($1:json)`, [links]);

```

","You do not need the library's `:json` filter in this case, as you need an array of JSON objects, and not a JSON with an array of JSON objects.


The former is formatted correctly by default, which then only needs `::json[]` type casting:



```
    await db.none(`INSERT INTO tests(links) VALUES($1::json[])`, [links]);

```

**Other Notes**


- Use [pg-monitor](https://github.com/vitaly-t/pg-monitor) or event [query](https://vitaly-t.github.io/pg-promise/global.html#event:query) to output queries being executed, for easier diagnostics.
- Method [none](https://vitaly-t.github.io/pg-promise/Database.html#none) can only resolve with `null`, no point storing the result in a variable.
- Library `pg-promise` does not have any `:array` filter, see [supported filters](https://github.com/vitaly-t/pg-promise#formatting-filters).


"
"Google reCAPTCHA ERROR for site owner: Invalid site key
I cant get reCAPTCHA v3 working on my website. I have put the right code in from the Google reCAPTCHA v3 website but it shows ""ERROR for site owner: Invalid site key"" in the bottom right.


I'm only 7 and if I have missed something, please take it easy



```
 <!DOCTYPE html>
<html lang=""en"">
<head>
    <meta charset=""UTF-8"">
    <title>Recaptcha test</title>
    <script src=""https://www.google.com/recaptcha/api.js""></script>
    <script>
      function onSubmit(token) {
          document.getElementById(""demo-form"").submit();
      }
    </script>
    <script src=""https://www.google.com/recaptcha/api.js?render=reCAPTCHA_site_key""></script>


<body>
<button class=""g-recaptcha""
        data-sitekey=""reCAPTCHA_site_key""
        data-callback='onSubmit'
        data-action='submit'>Submit</button>
<script>
    function onClick(e) {
        e.preventDefault();
        grecaptcha.ready(function() {
            grecaptcha.execute('reCAPTCHA_site_key', {action: 'submit'}).then(function(token) {
                // Add your logic to submit to your backend server here.
            });
        });
    }
</script>


</head>
</body>
</html>

```

Then, I tried reCAPTCHA v2 Checkbox, same things happened. Although, ""ERROR for site owner: Invalid site key"" was now in the top left:



```
    <!DOCTYPE html>?
<html lang=""en"">
<head>
    <meta charset=""UTF-8"">
    <title>Recaptcha test</title>
    <script src=""https://www.google.com/recaptcha/api.js"" async defer></script>

    <script src=""https://www.google.com/recaptcha/api.js?onload=onloadCallback&render=explicit""
            async defer>
    </script>



<body>
<form action=""?"" method=""POST"">
    <div class=""g-recaptcha"" data-sitekey=""your_site_key""></div>
    <br/>
    <input type=""submit"" value=""Submit"">
</form>
<script type=""text/javascript"">
    var onloadCallback = function() {
        alert(""grecaptcha is ready!"");
    };
</script>



</head>
</body>
</html>

```

","I have tried your code of reCAPTCHA v3 which you have mentioned. 


- reCAPTCHA\_site\_key is at three places please check if you have replaced it with correct site key under heading: `""COPY SITE KEY""`.
- Reason that it's showing ""ERROR for site owner: Invalid site key"" is you might be trying file system to access the key. It doesn't work on file systems ie, this will not work: `file:///E:/Personal%20Data/stackoverflow/recapcha.html`
- Host your site on local server like IIS/Node etc, and add the same domain in domain configurations of recaptcha. 


`for example if your website is: https://www.testingrecapcha.com/ then add testingrecapcha.com in domains of rechapcha configurations. (to reach admin console of recapcha, use this: https://www.google.com/recaptcha/admin and select your relevant site and click on settings icon).`


"
"Parallel preapproved paypal payment?
I have parallel payments working on my test server very nicely, i.e. I can send money to more than 1 person at a time in parallel based on the example parallel.php file I downloaded from the developer paypal site.


I also downloaded the preapproval\_flow.php file, but in the code, I can't see a variable/array which allows me to set parallel payment receivers, or even a single receiver for that matter.


Is it not possible to preapprove parallel payments with paypal api's?


If it is possible, how do I do it?


With straight forward parallel payments, I can use the following to set the parallel receivers of a payment:



```
$receiverEmailArray = array(
    'email0',
    'email1',
    'email2',
    'email3',
    'email4',
    'email5',
    'email6',
    'email7',
    'email8',
    'email9'
);

```

According to [cms.paypal.com](https://cms.paypal.com/us/cgi-bin/?cmd=_render-content&content_ID=developer/e_howto_api_APPreapproval) and [x.com](https://www.x.com/developers/paypal/documentation-tools/api/preapproval-api-operation), the preappoval\_flow has an option called `feesPayer`, for which the documentation says:



> 
> SENDER – Sender pays all fees (for personal, implicit
>  simple/**parallel payments**; do not use for chained or unilateral
>  payments)
> 
> 
> 


The bold bit, tells me that it's possible to do pre-approved parallel payments, but I can't figure out how to sent the receivers...


Also, on [x.com](https://www.x.com/developers/paypal/products/adaptive-payments), it says:



> 
> Developers and merchants can combine;
> 
> 
> - Parallel and preapproved payments.
> 
> 
> 


","The first step in pre-approvals, is on obtaining one from the buyer.


This pre-approval is not tied to any specific seller, but rather to the API caller, and it allows you to make a Pay API call in the future without the buyer having to confirm it.


The pre-approval flow is a separate API call from the Payment flow. In it, you specify the total amount you want to pre-approve, the start and end dates that the pre-approval is valid for, and a few other parameters. (See chapter 8 of the Adaptive Payments dev guide for the full spec: <https://cms.paypal.com/cms_content/US/en_US/files/developer/PP_AdaptivePayments.pdf> )


In this API call, you redirect the buyer to the PayPal site, where he will approve **future** payments ***up to*** the amount you specified. (e.g. if you specified a total amount of $1000, the buyer can make 20 purchases of $50, or 1 big purchase of $1000)


Once the buyer approves this, you have a permission to charge him (In the form of a pre-approval token `PA-xxxxxxxxxxxxxx`, but at this point no money has changed hands.


Now that you have the token, you can proceed to make the Parallel Payment Pay API call, just like you are doing now, **BUT** you will specify one additional parameter:
`preapprovalKey=<The PA key obtained previously>`


(And you also need to make sure that `actionType` is set to **PAY** )


As soon as you make this API call, because PayPal sees that you have already obtained permission from the buyer (in the form of the approved PA-key), it will immediately move the money to the receivers specified. There is no need for the buyer to be redirected to approve this payment, since he has already *pre* approved it. (See the logic here?)


So to recap:


1. Make call to Pre-Approval API, to obtain **permission** to charge from the buyer
2. Make call to Pay API (that includes the PA key from step 1) to execute on the permission given


"
"reset a pointer for a for-loop in php
given the following code



```
<?php
$a = array(1,2,3,4,5,6);
$c=0;
foreach($a as $v){
    if($v==5&&$c==0){
        $c=1;
        reset($a);
    }
    var_dump($v);
}

```

How do I reset the pointer so it will print 1,2,3,4,5,1,2,3,4,5,6 ?


I know that in this case I can simply



```
<?php
$a = array(1,2,3,4,5,6);
$c=0;
for($i=0;$i<count($a);++$i){
    $v = $a[$i];
    if($v==5&&$c==0){
        $c=1;
        $i=-1;  //because of the loop ++$i
    }
    var_dump($v);
}

```

But I have a much more complex piece of code and the solution is not as simple as rewrite the loop (not numeric keys).


Any PHP guru out there that can help me here ?


","## NOTE: Answer uses function deprecated as of PHP 7.2


As @TheodoreR.Smith pointed out in the comments. list() = each() will not work for any PHP version >= 7.2. See:


[How can I update code that uses the deprecated each() function?](https://stackoverflow.com/questions/46492621/how-can-i-update-code-that-uses-the-deprecated-each-function)


## Original Answer for version < 7.2


As documented:



> 
> Note: When foreach first starts executing, the internal array pointer is automatically reset to the first element of the array. This means that you do not need to call reset() before a foreach loop.
> 
> 
> As foreach relies on the internal array pointer, **changing it within the loop may lead to unexpected behavior**.
> 
> 
> 


<http://php.net/manual/en/control-structures.foreach.php>


Not sure what ""unexpected behavior"" is, as I've never tried this...but maybe safer to manually use each()...also more clear in your code.



```
reset($a);
while(list($key, $val) = each($a)) {
    if($val==5&&$c==0){
        $c=1;
        reset($a);
    }
    var_dump($val);
}

```

**Foreach Issues Explained**


I thought I had seen before that you couldn't rely on the internal pointer in the foreach...but couldn't find it in the documentation this time...only ""unexpected results"". However, thanks to a commenter, I found the text that used to be in the foreach:



> 
> Unless the array is referenced, foreach **operates on a copy of the specified array** and not the array itself. foreach has some side effects on the array pointer. Don't rely on the array pointer during or after the foreach without resetting it.""
> 
> 
> 


<http://php.net/manual/en/control-structures.foreach.php#114759>


"
"Parallel simulation setting in Dymola
I am building a large system in Dymola, I plan to exploit the parallel simulation potential in Dymola, but I just find two general settings in the help document.


My questions are:


1. Is there any other setting I could use in Dymola so the code would run faster?
2. In my model code, is there any method that allows the model to run in a parallel mode? For example, I know MATLAB code style is using vectors, so does that mean in Modelica I should use vectors too?


[![enter image description here](https://i.stack.imgur.com/pOb7Z.jpg)](https://i.stack.imgur.com/pOb7Z.jpg)


[![enter image description here](https://i.stack.imgur.com/dwETk.jpg)](https://i.stack.imgur.com/dwETk.jpg)


","The `Advanced.ParallelizeCode`-flag does generate code that runs in parallel (hopefully, check the log), and vectorization as in Matlab does not help here (it may even cause problems).


What you can do instead is:


- As described in User manual 1B section Handling of external functions in parallel code add `annotation(__Dymola_ThreadSafe=true);` to called functions, after checking that they are thread-safe.
- Try to ensure that there is some form of dynamics separating the different parts of the model.
- If it is not possible to naturally treat the different parts in parallel you might use the decouple operator (look in the `dymola\Modelica\Libraries\DecoupleBlocks.mo` ); but be careful to verify the results first.


The flag `Advaned.ParallelSimulations` does not seem relevant for *this case*, as it runs entire simulations with different parameter values in parallel (much more coarse-grained). If you only run one simulation at a time that gives no benefit. 


**However, if you could change to use this instead it is in general a better idea** - and you should then (in general) not enable the other parallelization variant.


"
"Service will not inject into controller
I've made a handful of Angular applications in the past, but have never really made use of modules. I'm creating an Ionic app right now (which uses Angular) and for some reason I cannot get my services to inject into the controllers.


**controllers.js**



```
angular.module('myapp.controllers', [])
.controller('StreamCtrl', ['$scope', 'StreamService',
function($scope, StreamService) {
    $scope.playlists = [1,2,3,4];
}]);

```

**services.js**



```
angular.module('myapp.services', [])
.factory('StreamService', ['$scope', function($scope) {
    var self = {
        'fetching' : false,
        'streamItems' : []
    };
    return self;
]);

```

**app.js**



```
angular.module('myapp', ['ionic',
                        'myapp.services',
                        'myapp.controllers',
                        'ngCordova',
                        'LocalStorageModule'])
// And lots of other code that isn't needed for this question.

```

So when I try and load the `StreamCtrl` I end up getting the following error:



```
Error: [$injector:unpr] Unknown provider: $scopeProvider <- $scope <-StreamService
http://errors.angularjs.org/1.3.6/$injector/unpr?p0=<ion-nav-view name=""menuContent"" class=""view-container"" nav-view-transition=""ios"">copeProvider%20%3C-%20%24scope%20%3C-%20StreamService
at REGEX_STRING_REGEXP (http://localhost:8100/lib/ionic/js/ionic.bundle.js:7888:12)
at http://localhost:8100/lib/ionic/js/ionic.bundle.js:11806:19
at Object.getService [as get] (http://localhost:8100/lib/ionic/js/ionic.bundle.js:11953:39)
at http://localhost:8100/lib/ionic/js/ionic.bundle.js:11811:45
at getService (http://localhost:8100/lib/ionic/js/ionic.bundle.js:11953:39)
at Object.invoke (http://localhost:8100/lib/ionic/js/ionic.bundle.js:11985:13)
at Object.enforcedReturnValue [as $get] (http://localhost:8100/lib/ionic/js/ionic.bundle.js:11847:37)
at Object.invoke (http://localhost:8100/lib/ionic/js/ionic.bundle.js:11994:17)
at http://localhost:8100/lib/ionic/js/ionic.bundle.js:11812:37
at getService (http://localhost:8100/lib/ionic/js/ionic.bundle.js:11953:39)

```

","See this working [fiddle](http://jsfiddle.net/HB7LU/11116/) ,


You are treating service as if it was a controller, never inject a `$scope` into service, its not there



```
var myApp = angular.module('myApp',['myapp.services','myapp.controllers']);

angular.module('myapp.controllers', [])
.controller('StreamCtrl', ['$scope', 'StreamService',
function($scope, StreamService) {
    $scope.playlists = [1,2,3,4];
}]);

angular.module('myapp.services', []).factory('StreamService',  function() {
    var self = {
        'fetching' : false,
        'streamItems' : []
    };
    return self;}
);

```

"
"How can I stop .exe's being ran from removeable media, such as USB drives?
I have an issue with users running .exe's from removable storage such as Memory Sticks and SD cards. 


I am trying to set up blocking of exe's being ran from all removable storage to combat this, however under the Group Policy settings under ""User Configuration > Windows Settings > Security Settings > Software Restriction Policies > Additional Rules"" you can create a ""path"" variable. I would like to use a system variable for *all* removable storage but do not know if there is one that works. Do I have to go through and create a blocklist for all potential drive letters that could be assigned to removeable media (E:\ through to about L:\) or is there one that actually works? I've found `%~dp0` which, apparently, only works for for loops, if statements and batch parameters.


If there are other, better, or more reliable measures please let me know.


Oh, I've looked at AppLocker but our DC is running Server 2008 and I believe that AppLocker is part of Windows 7 and 2008 R2. As mentioned in the comments to the answer below I don't think Server 2008 has the ""Deny Execute Access"" setting, so are there any other options?


","You can stop execution of software on removable devices via a GPO. The setting is under Computer > Administrative Templates > System > Removable Storage Access > Removable Disks: Deny Execute Access


![enter image description here](https://i.stack.imgur.com/dAokD.png)


**Edit:**


Do you have access to a Server 2008 R2 System? If so, you could create the policy setting, back it up to disk and transfer it to your Server 2008 System and import the policy back. This won't give you the ability to modify the policy, but you should be able to see the settings as `Extra Registry Settings` and it should work fine on your Windows 7 clients.


If it helps, I've just created a backup of such a policy, and have put it up on [Dropbox](http://dl.dropbox.com/u/3789379/PolicyBackup.zip) for you to download. The usual use at your own risk, etc. applies.


"
"Angular Input property change is not detected when I push new item in array
I have a list of an array(dataSource) on which I am performing Add and Delete operations. I have passed an array(dataSource) in the child component and added \*ngFor loop in the child. Array I have passed as getter setter for detect @Input change.


This is my **AppComponent** :



```
<div>
    <vertital-tabs
          [config]=""config""
          [dataSource]=""dataSource""    <-- Array changes I want to detect in Child
          [(selectedItem)]=""selectedItem"">
    </vertital-tabs>

    <div style=""background-color: yellow;margin: 10px"">
        {{ selectedItem | json }}
     </div>
</div>

```

In **ChildComponent** (vertical-tabs) :



```
get dataSource(): any {
    return this._dataSource;
}

@Input() set dataSource(value: any) {
    this._dataSource = value;
    // Not called this on Add New button even if Input Datasource is changed.
    // I want this to be called on Add New Item button click....
    debugger;
}

```

**Problem** is that when I am adding a new item on the array, it is not calling the setter @Input change method. When I am deleting an item it is working fine and call @Input change.


**Note:** I have lot of properties as an Input in real scenario so I don't want to use ngOnChanges()


Here is an example I have created: <https://stackblitz.com/edit/angular-vertical-tabs-component-split-yzynef>


","Angular only checks whether the *reference* has changed - if it has not changed since the last check the setter will not be called.


In `app.component.ts`:



```
let nextItem = this.dataSource.length;
this.dataSource.push({
  id: nextItem.toString(),
  text: 'item ' + nextItem.toString(),
  icon: 'settings',
});

```

Here you add something to your array instead of creating a new array and assigning it to `dataSource`.


In `vertical-tabs.ts`:



```
onDelete(item, index) {
  this.dataSource = this.dataSource.filter((x) => x !== item);
}

```

Here you create a new array and assign it to `dataSource`.


That's why deletion works as you expect, but adding doesn't. Copying the array and assigning it to `dataSource` should solve your problem:



```
let nextItem = this.dataSource.length;
this.dataSource = [
  ...dataSource,
  {
    id: nextItem.toString(),
    text: 'item ' + nextItem.toString(),
    icon: 'settings',
  }
];

```

"
"Exporting query results as JSON via Google BigQuery API
I've got jobs/queries that return a few hundred thousand rows. I'd like to get the results of the query and write them as json in a storage bucket. 


Is there any straightforward way of doing this? Right now the only method I can think of is:


- set allowLargeResults to true
- set a randomly named destination table to hold the query output
- create a 2nd job to extract the data in the ""temporary"" destination table to a file in a storage bucket
- delete the random ""temporary"" table.


This just seems a bit messy and roundabout. I'm going to be wrapping all this in a service hooked up to a UI that would have lots of users hitting it and would rather not be in the business of managing all these temporary tables. 


","1) As you mention the steps are good. You need to use Google Cloud Storage for your export job. Exporting data from BigQuery is explained [here](https://cloud.google.com/bigquery/exporting-data-from-bigquery), check also the variants for different path syntax.


Then you can download the files from GCS to your local storage.


[Gsutil](https://cloud.google.com/storage/docs/gsutil?hl=en) tool can help you further to download the file from GCS to local machine.


With this approach you first need to export to GCS, then to transfer to local machine. If you have a message queue system (like [Beanstalkd](https://github.com/kr/beanstalkd/wiki/client-libraries)) in place to drive all these it's easy to do a chain of operation: submit jobs, monitor state of the job, when done initiate export to GCS, then delete the temp table.


Please also know that you can [update a table](https://cloud.google.com/bigquery/docs/reference/v2/tables/update) via the API and set the [expirationTime](https://cloud.google.com/bigquery/docs/reference/v2/tables#resource) property, with this aproach you don't need to delete it.


2) If you use the BQ Cli tool, then you can set output format to JSON, and you can redirect to a file. This way you can achieve some export locally, but it has certain other limits.


this exports the first 1000 line as JSON



```
bq --format=prettyjson query --n=1000 ""SELECT * from publicdata:samples.shakespeare"" > export.json

```

"
"Windows XP Home issuing about 20 ARP requests a second
I have been asked to remotely fix a family member's computer that is running ""slow"". I've had them run any number of S&D and AVG scans, even in safe mode, but they find nothing of interest. In using Remote Assistance to poke around, I installed Wireshark and discovered that the box is issuing up to 20 ARP requests per second, mostly for its local subnet (directly connected to WebSTAR 2000 USB cable modem) as seen in the graphic below. 


The IP address at the time was 70.119.184.xx/255.255.240.0. The default gateway was 70.119.176.1 and the DHCP server was 10.212.0.1.


Is it possible this is legitimate traffic, or is this a symptom of a worm like WootBot trying to spread?


![WireShark capture](https://i.stack.imgur.com/hEfod.png)


EDIT: I think the machine is infected with Trojan.Opachki or something very like it. 


EDIT: The ARP traffic is actually sourced elsewhere so it is not an issue with the machine in question. There was signs of an Opachki infection but I think that has been cleared up. I'll put him down for a router next Christmas.


","Not to be a party pooper, but the size of the subnet is irrelevant to the number of ARP broadcasts per second. Just because a subnet is large enough for x number of hosts doesn't mean that a host will ARP for x number of ip addresses in that subnet. A host sends an ARP packet when it needs to send a packet to another host. The only time a host will ARP for x number of ip addresses in it's subnet (or a large number of ip addresses in it's subnet) is if it's scanning the ip address range for it's subnet (using an IP scanning program), it's infected with malware, or has a faulty NIC or NIC driver. At no other time does a host normally send a large number of ARP packets like what you're seeing. Additionally a host will not send an ARP packet for an ip address that's not on it's local subnet as it knows that the ip address is not local and that it needs to send it to it's default gateway, and therefore will not send an ARP packet for that ip address.


You have 5 hosts sending ARP packets on the network:


10.212.0.1 - This seems normal. This is the default gateway and there's only one ARP packet in your screenshot. The default gateway will send ARP packets in to the network when it needs to pass traffic to an internal host and that hosts' MAC address is not in it's ARP cache (like every other network device does).


24.170.135.1 - I don't understand this one. This is a non-local ip address. Where is it coming from? Do you have multiple networks bridged together? Do any of the computers have multiple NIC's connected to multiple networks or multiple connections, such as a VPN connection, etc.


24.233.137.1 - Again, this is a non-local ip adress.


70.119.248.1 - This is probably normal, although the ip addresses it's ARP'ing for seem a little out of place. They're in the same subnet but far separated from what I would consider a normal ip addressing scheme.


70.119.176.1 - This is the one that worries me as it's the one sending the bulk of the ARP packets. I suspect that this one is either performing a subnet scan for all ip adresses in the subnet, it's infected with malware, or it has a bad NIC or NIC driver.


ARP floods (which is what I believe you're dealing with) are not a normal condition in a network. ARP broadcasts exceeding about 3 -5 % of all network traffic is a very good indication that something is wrong.


***EDIT***


After re-reading your question with your recent edits, I have a different opinion of what's going on: if 70.119.176.1 is the default gateway for the network, and it's the one sending the bulk of the ARP requests for addresses in the subnet, then I'm thinking that someone external to you is performing an ip address\port scan against your network and your firewall is not blocking it. For every ip address being probed, your default gateway is sending an ARP request to try and find a host on the ip address being probed. Does your firewall, router, or modem have a log that you can look at?


I still don't understand where the 24.x.x.x addresses are coming from.


"
"Android BLE BluetoothGatt.writeDescriptor() return sometimes false
I'm trying to write BLE Android app. I found that sometimes when I call 
[BluetoothGatt.writeDescriptor()](https://developer.android.com/reference/android/bluetooth/BluetoothGatt.html#writeDescriptor(android.bluetooth.BluetoothGattDescriptor)) it returns false.


I have not found in documentation any note of limitation to this function. But ppl on stack overflow says that I need to wait for [BluetoothGattCallback.onDescriptorWrite()](https://developer.android.com/reference/android/bluetooth/BluetoothGattCallback.html#onDescriptorWrite(android.bluetooth.BluetoothGatt,%20android.bluetooth.BluetoothGattDescriptor,%20int)) before I try to write another descriptor. 


Here is one reply saying that [BLE is busy](https://stackoverflow.com/a/36169096/1093607) with writeDescriptor() and can not do other write. 


Here is [another thread](https://stackoverflow.com/a/24178815/1093607) saying that you can not call twice writeCharacteristic().


My questions are


- is it really true?
- is there really missing some internal android API buffer for serializing BLE requests and every developer has to do it on it's own?
- Is it true for different functions? For example when I call `writeDescriptor()` I understand I can not call second time `writeDescriptor()` before I receive `onDescriptorWrite()`. But do I have to wait for `onDescriptorWrite()` when I want to call `writeCharacteristic()`?
- Also if there is inter-function dependency then what else function have this limitation (namely: `readCharacteristic()`, `readDescriptor()`, `requestMtu()`...)?
- And additionally is there interdependency between BluetoothGattServer and BluetoothGatt. So for example when I call `BluetoothGattServer.notifyCharacteristicChanged()` shall I wait for[`BluetoothGattServerCallback.onNotificationSent`](https://developer.android.com/reference/android/bluetooth/BluetoothGattServerCallback.html#onNotificationSent(android.bluetooth.BluetoothDevice,%20int)) before I can call `BluetoothGatt.writeDescriptor()` or `BluetoothGatt.writeCharacteristic()`? (BTW praise for google documentation `onNotificationSent()` is by luck documented properly. Doc says:



> 
> When multiple notifications are to be sent, an application must wait
>  for this callback to be received before sending additional
>  notifications.
> 
> 
> 


- Lastly having all this questions - I feel that Android BLE API is under-documented. Or am I wrong and there is documented somewhere what are allowed methods calling sequences? If yes can you please point me to such documentation? If not is there some channel we can open issue with google and ask them to add to documentation something? I mean it may not be much text - some function like `onNotificationSent()` is arleady properly documented. They just need to copy this sentence to other functions.


","The documentation lacks information. However you can read the source code to find out the rules, which (currently) are the following:


For each `BluetoothGatt` object, you can only have one outstanding request at a time, including `requestMtu`, `readCharacteristic`, `writeCharacteristic`, `readDescriptor`, `writeDescriptor` and `executeReliableWrite`. So if you issue a read request you need to wait for the read response before you issue a write request. While they implemented the code that returns false if there is an ongoing operation in [BluetoothGatt.java](https://android.googlesource.com/platform/frameworks/base/+/master/core/java/android/bluetooth/BluetoothGatt.java), they forgot to do this for `requestMtu`, so if you have multiple requests at a time where `requestMtu` is one of them, you will get random errors sooner or later (in the latest versions at the time of this post).


So yes, every developer has to manually serialize the requests. Note that the Bluetooth stack actually has a queue of requests, but it is limited to only one request per client (i.e. BluetoothGatt object). So if two apps on the same phone talk to the same device simultaneously you will never get ""busy"" errors. The only exception is if you use Write Without Response for which the current data flow implementation is quite buggy (see <https://issuetracker.google.com/issues/37121017> which Google seems to have ignored).


You can send notifications at the same time as you write a characteristic, since the server and client roles are separated.


Regarding updating the documentation, you can always try to file an issue at <https://issuetracker.google.com> (but I get the feeling nobody reads that), or, since Android is open source, send a pull request to <https://android-review.googlesource.com/> which updates the Javadoc from which the documentation is generated.


"
"pdo::beginTransaction does it lock table?
Can I use this function instead of 'LOCK TABLES' query?


Example:



```
pdo::beginTransaction(); 

    SELECT id AS last_id FROM t WHERE...
    INSERT INTO t (id,...) VALUES (last_id+1,....)

pdo::commit();

```

","The answer to this is: it depends. The most important resource to help you understand exactly what it does, what it doesn't do and how it works is [here](http://www.php.net/manual/en/pdo.transactions.php).


For a kick off, it depends whether the underlying driver (MySQL, MSSQL, etc etc) supports transaction functionality at all. If the driver does not support transactions, `pdo::beginTransaction();` will fail and return `FALSE`, and all your queries will be executed immediately. This is not to say that a [`LOCK TABLES`](http://dev.mysql.com/doc/refman/5.0/en/lock-tables.html) query would fail - it depends whether the underlying database engine supports it.


In fact, in MySQL at least, `pdo::beginTransaction()` follows the same rules as a [`START TRANSACTION`](http://dev.mysql.com/doc/refman/5.5/en/commit.html) statement. It does not lock the table immediately, it simply ensures that that the queries that are part of the transaction follow the rules of [ACID](http://en.wikipedia.org/wiki/ACID).


I won't go into full details of exactly which combinations will work and what they will do because they are covered at length in the documents I linked to, and there is no sense in me writing it all out again.


"
"Trying to identify this symbol ⎋ used for a key in a tutorial
I was watching a Linux tutorial where the keystrokes where displayed.
I cannot identify the key matching this:


![symbol](https://i.stack.imgur.com/VQgut.png)


What key is it?


Edit: It is from [this tutorial](https://medium.com/@schtoeffel/you-don-t-need-more-than-one-cursor-in-vim-2c44117d51db).


","It's 'Broken Circle with Northwest Arrow' Unicode U+238B UTF-8 E2 BE BB


# ⎋


It's often used as the symbol for the  `Esc`  key - <https://en.wikipedia.org/wiki/Esc_key#Symbol> and even has its own ISO Standard - <https://commons.wikimedia.org/wiki/File:ISOIEC-9995-7-029--ISO-7000-2029--Symbol-for-Escape.svg>


It seems symbols like this are mainly used by Apple. They are all valid unicode. It enables menu shortcuts to be displayed easily, e.g.


[![enter image description here](https://i.stack.imgur.com/23HJF.png)](https://i.stack.imgur.com/23HJF.png)


After many comments on the topic, I doubt Apple were instrumental in the adoption of this symbol, they merely started using it once it was. Menus used to use the older


# ␛


glyph back in the day.


"
"How do I override icons in Material UI DataGrid?
We need to customize most of the icons used in the [DataGrid](https://material-ui.com/api/data-grid/data-grid/) from Material UI. Reading the very limited docs, we cannot figure out how to do that.


The [documentation](https://material-ui.com/api/data-grid/data-grid/#icons-slots) talks about *slots* (but with no explanation on how to use those). There is a `icons` property, but we're not sure if `IconsOptions` is the same as slots. Regardless, we tried the following:



```
      <div style={{ height: 400, width: ""100%"" }}>
        <DataGrid
          rows={rows}
          columns={columns}
          pageSize={5}
          disableSelectionOnClick
          icons={{
            ColumnMenuIcon: AcUnit
          }}
          />
      </div>

```

and the following:



```
      <div style={{ height: 400, width: ""100%"" }}>
        <DataGrid
          rows={rows}
          columns={columns}
          pageSize={5}
          disableSelectionOnClick
          ColumnMenuIcon={AcUnit}
          />
      </div>


```

To no avail. We tried using `<AcUnit/>` instead, but no change. Tried reading the code but that did not really help either...


What is the proper way of doing this? If that is a standard way of doing things in either React or Material UI, where can we find some documentation?


Here is a [Code Sandbox](https://codesandbox.io/s/material-datagrid-icons-override-ukzyk?file=/demo.js).


","You can find this documented here: <https://material-ui.com/components/data-grid/components/#icons> with the list of available icons to customize in the [documentation you noted](https://material-ui.com/api/data-grid/data-grid/#icons-slots). The example shows using the `components` prop to customize the icons.


The following syntax works:



```
        <DataGrid
          rows={rows}
          columns={columns}
          pageSize={5}
          disableSelectionOnClick
          components={{
            ColumnMenuIcon: AcUnit
          }}
        />

```

[![Edit Material DataGrid Icons Override](https://codesandbox.io/static/img/play-codesandbox.svg)](https://codesandbox.io/s/material-datagrid-icons-override-n2vzk?fontsize=14&hidenavigation=1&theme=dark)


"
"If cell contains 1 or more keywords, change value of a different cell
I have a column with some string description in it.


for example:
```

Bob davids
mowing the lawn
tipping cows
```



In addition I will have on a different sheet or column a list of keywords
For example work keyword list 1:



```

davids
work
```

play keyword list:
```

mowing
cows
```



So as the main column is populated with these text strings, I would like them checked automatically each keyword list to see if those words exist, and when it finds a match, place the title of the list (work/play) in the cell next to it.


I know this is possible in VBA and can even do it ""on the fly"" in `SelectionChange` function. But is this possible to do without VBA such as a conditional formatting?


","This is pretty easy to do with just formulas, as long as you don't care too much about possibly improperly finding parts of words. Ignore that caveat for a second though. First, here is a formula that will tell you if any of several strings are found anywhere within a source string:



```
=OR(NOT(ISERROR(FIND(<array of strings to look for>,<string to look in>))))

```

This needs to be entered as an array formula for it to work. You do that by entering it with Ctrl-Shift-Enter. To understand how it works, consider what Excel does in evaluating a real example:



```
=OR(NOT(ISERROR(FIND({""a"",""b"",""c""},""q b q""))))

```

'FIND' finds the position of one string within another. When called with an array for the first argument, it will return an array of positions (or #VALUE! if the search string isn't found). You can trace the evaluation by entering that formula and then using the F9 key on expressions within it:



```
=OR(NOT(ISERROR({#VALUE!,3,#VALUE!})))
=OR(NOT({TRUE,FALSE,TRUE}))
=OR({FALSE,TRUE,FALSE})
=TRUE

```

So, for your example, say you had your strings you want searched in $B$6:$B$8, your work strings in $D$2:$D$3, and your play strings in $E$2:$E$3. You could put the formula



```
=OR(NOT(ISERROR(FIND(D$2:D$3,$B6))))

```

in cell D6, enter it as an array formula, and then drag it through the range D6:E8 to find which strings in B had work or play words in them. Then you could use those results to drive further formulas or conditional formatting.


However, as mentioned above, you'll note that any substring within the string being searched will get found, so 



```
=OR(NOT(ISERROR(FIND({""a"",""b"",""c""},""bad""))))

```

will evaluate to TRUE. (And if your fun list includes ""id"", the ""id"" in ""davids"" will match.)


As is often the case with Excel, if you're doing something you understand with a limited data set, you might not care about this. But it can defeat an attempt to use this kind of formula as part of a general ""application"" that has users who don't know fancy array tricks or exactly what 'FIND' does. (You can sort of get around that by putting a space after your search words, etc., but that is just more mysterious voodoo waiting to be broken if you hand it someone else.) For a quick and dirty scan, though, it's fine.


"
"SQL Server Stored Procedure capture return value in T-SQL
I have a SQL Server stored procedure; I need to capture the return value from the stored procedure. Is this the correct way of doing this?



```
  declare valback varchar(30)
  set valback = exec storeproc1 

```

In this case, `storeproc1` is my stored procedure.


","To start, use proper T-SQL syntax:



```
declare @valback int;
exec @valback = storeproc1;

```

The only return type allowed for a stored procedure is `int`. Stored procedures return status via the [`return`](http://msdn.microsoft.com/en-us/library/ms174998.aspx) statement. 


I somehow have a feeling that you really want something else, namely:
to have an [OUTPUT parameter](http://msdn.microsoft.com/en-us/library/aa258259%28v=sql.80%29.aspx) in the procedure:



```
declare @valback varchar(30);
exec storedproc1 @valback OUTPUT;

```

or capture the procedure result set via `INSERT ... EXEC`. See [How to Share Data Between Stored Procedures](http://www.sommarskog.se/share_data.html).


"
"Android piracy prevention with server requests
I'm working on an application that requires a server to do most of the heavy lifting. I want to prevent pirated clients from sending requests to that server. Is there a way to send some identifier with the requests so that my server can ask the Android Market if someone with that id actually bought the app? How would I go about doing that?


Note that just protecting the app with LVL won't work, because people could fairly easily write an application to interface with the server and still provide the same functionality as the paid app.


","**First: There is no 100% security for anything you run on a device that is not under your control (like Android devices in your case).**


You could make ""abuse"" harder by several measures:


- issue a random session key (cryptographically secure) after a successfull login with a time-limit so a new login needs to happen after a certain time has passed by
- issue a random interaction key (cryptographically secure) for every communication step which gets invalidated right after one usage
- when a successfull login happens terminate any other session associated with the same credentials that might be active before that login
- ""throttle"" usage, i.e. limit how many calls per minute/hour or similar are allowed (might be impossible depending on the specific application)


IF you really really want to make it very hard you can issue a device-specific client-certificate (when the client buys your app) and use cert-based client authentication (defined in SSL standard) - you can invalidate the cert associated with the device if you see abuse without harm for the legitimate users of other devices...


"
"Regex that finds consecutive words with first letter capitalized
I am looking for a regex that can identify in a sentence that consecutive words in a sentence start with capital letters.


If we take the text below as an example:



> 
> The A-Z Group is a long-established
> market leader in the provision of
> information for the global air cargo
> community, and also for the defence
> and security sectors through BDEC
> Limited, publishers of the British
> Defence Equipment Catalogue and
> British Defence Industry Directory.
> 
> 
> 


I want to be able to retrieve the following:



> 
> The A-Z Group
> 
> 
> BDEC Limited Defence Equipment
> 
> 
> Catalogue British Defence
> 
> 
> IndustryDefence Industry
> 
> 
> 


Is this even possible with a regex?
If so, can anyone suggest one?


","(Update: I misunderstood your question at first.)


A simple case is



```
/([A-Z][\w-]*(\s+[A-Z][\w-]*)+)/

```

It may need to be modified if there are special cases of different language construct.


`ruby-1.9.2-p0 > %Q{The A-Z Group is a long-established market leader in the provision of information for the global air cargo community, and also for the defence and security sectors through BDEC Limited, publishers of the British Defence Equipment Catalogue and British Defence Industry Directory.}.scan(/([A-Z][\w-]*(\s+[A-Z][\w-]*)+)/).map{|i| i.first}`


`=> [""The A-Z Group"", ""BDEC Limited"", ""British Defence Equipment Catalogue"", ""British Defence Industry Directory""]`


"
"Why does Java not allow foreach on iterators (only on iterables)?

> 
> **Possible Duplicate:**  
> 
> [Why is Java's Iterator not an Iterable?](https://stackoverflow.com/questions/839178/why-is-javas-iterator-not-an-iterable)
> 
> 
> [Idiomatic way to use for-each loop given an iterator?](https://stackoverflow.com/questions/3883131/idiomatic-way-to-use-for-each-loop-given-an-iterator)
> 
> 
> [Can we use for-each loop for iterating the objects of Iterator type?](https://stackoverflow.com/questions/2162917/can-we-use-for-each-loop-for-iterating-the-objects-of-iterator-type)
> 
> 
> 


The foreach loop are as far as I know syntax sugar added in Java 5. So



```
Iterable<O> iterable;
for(O o : iterable) {
    // Do something
}

```

will essentially produce the same bytecode as



```
Iterable<O> iterable;
for(Iterator<O> iter = iterable.iterator(); iter.hasNext(); /* NOOP */) {
    O o = iter.next();
    // Do something
}

```

However, if I do not have an iterable in the first place, but only an iterator (say, because a class offers two different iterators), I cannot use the syntax sugar foreach loop. Obviously I can still do the plain old style iteration. However, I'd actually like to do:



```
Iterator<O> iter;
for(O o : iter /* Iterator<O>, not Iterable<O>! */) {
     // Do something
}

```

And of course I can do a fake `Iterable`:



```
class Adapter<O> implements Iterable<O> {
    Iterator<O> iter;

    public Adapter(Iterator<O> iter) {
        this.iter = iter;
    }

    @Override
    public Iterator<O> iterator() {
        return iter;
    }
}

```

(Which in fact is an ugly abuse of the Iterable API, as it can only be iterated once!)


If it were designed around `Iterator` instead of iterable, one could do a number of interesting things:



```
for(O o : iterable.iterator()) {} // Iterate over Iterable and Collections

for(O o : list.backwardsIterator()) {} // Or backwards

Iterator<O> iter;
for(O o : iter) {
    if (o.something()) { iter.remove(); }
    if (o.something()) { break; }
}
for(O : iter) { } // Do something with the remaining elements only.

```

Does anyone know *why* the language was designed this way? To avoid ambiguity if a class would implement both `Iterator` and `Iterable`? To avoid programmer errors that assume that ""for(O o : iter)"" will process all elements twice (and forget to get a fresh iterator)? Or is there some other reason for this?


Or is there some language trick I just do not know?


","So I have a somewhat reasonable explanation now:


Short version: **Because the syntax also applies to *arrays*, which don't have iterators.**


If the syntax were designed around `Iterator` as I proposed, it would be inconsistent with arrays. Let me give three variants:


**A)** as chosen by the Java developers:



```
Object[] array;
for(Object o : array) { }
Iterable<Object> list;
for(Object o : list) { }
Iterator<Object> iter;
while(iter.hasNext()) { Object o = iter.next(); }

```

The behaves the same way and is *highly consistent* across arrays and collections.
Iterators however have to use the classic iteration style (which at least is not likely to cause errors).


**B)** allow arrays and `Iterators`:



```
Object[] array;
for(Object o : array) { }
Iterable<Object> list;
for(Object o : list.iterator()) { }
Iterator<Object> iter;
for(Object o : iter) { }

```

Now arrays and collections are inconsistent; but arrays and ArrayList are very closely related and should behave the same way. Now if at any point, the language is *extended* to make e.g. arrays implement `Iterable`, it becomes inconsistent.


**C)** allow all three:



```
Object[] array;
for(Object o : array) { }
Iterable<Object> list;
for(Object o : list) { }
Iterator<Object> iter;
for(Object o : iter) { }

```

Now if we end up in unclear situations when either someone implements *both* `Iterable` and `Iterator` (is the for loop supposed to get a new iterator or iterate over the current - happens easily in tree-like structures!?!). A simple tie-braker ala ""Iterable beats Iterator"" unfortunately won't do: it suddenly introduces runtime vs. compile time difference and generics issues.


Now suddenly, we need to pay attention to whether we want to iterate over collections/iterables or arrays, at which point we have gained very little benefits at the cost of a big confusion.


The way ""for each"" is in Java (A) is very consistent, it causes very little programming errors, and it allows for the possible future change of turning arrays into regular objects.


There is a variant **D)** that would probably also work okay:
for-each for Iterators only. Preferrably by adding a `.iterator()` method to primitive arrays:



```
Object[] array;
for(Object o : array.iterator()) { }
Iterable<Object> list;
for(Object o : list.iterator()) { }
Iterator<Object> iter;
for(Object o : iter) { }

```

But this requires changes to the runtime environment, not just the compiler, and breaks backwards compatibility. Plus, the mentioned confusion is still present that



```
Iterator<Object> iter;
for(Object o : iter) { }
for(Object o : iter) { }

```

Only iterates over the data once.


"
"Coarse-grained vs fine-grained
What is the difference between coarse-grained and fine-grained?


I have searched these terms on Google, but I couldn't find what they mean.


","From [Wikipedia (granularity)](http://en.wikipedia.org/wiki/Granularity):



> 
> Granularity is the extent to which a
>  system is broken down into small
>  parts, either the system itself or its
>  description or observation. It is the
>  extent to which a larger entity is
>  subdivided. For example, a yard broken
>  into inches has finer granularity than
>  a yard broken into feet.
> 
> 
> Coarse-grained systems consist of
>  fewer, larger components than
>  fine-grained systems; a coarse-grained
>  description of a system regards large
>  subcomponents while a fine-grained
>  description regards smaller components
>  of which the larger ones are composed.
> 
> 
> 


"
"What is the difference between C# , .NET and CLI?
What is the difference between C# , .NET and CLI?


","C# is one of the .NET languages - VB.NET, F# and [many more](http://en.wikipedia.org/wiki/List_of_CLI_languages) exist.


[.NET](http://en.wikipedia.org/wiki/.NET_Framework) is a platform (libraries, compiler and runtime) on which code written by any of these languages runs.


[CLI](http://en.wikipedia.org/wiki/Common_Language_Infrastructure) is the Common Language Infrastructure - it is a specification that describes the executable code and runtime that make the .NET platform. Apart from the Microsoft one, there are other implementations - [mono](http://en.wikipedia.org/wiki/Mono_(software)) and [Portable.NET](http://en.wikipedia.org/wiki/Portable.NET) the most known.


If you want some more acronyms:


- There is also the [CLR](http://en.wikipedia.org/wiki/Common_Language_Runtime) - the common language runtime, which is the Microsoft implementation of the CLI, mentioned above.
- In addition, on top of the CLR, Microsoft have created the [DLR](http://en.wikipedia.org/wiki/Dynamic_Language_Runtime) - the dynamic language runtime is a set of services for several dynamic languages such as python and ruby.


"
"Why does %TEMP% resolve to a non-deterministic path of the form %TEMP\?
I have a Windows Server 2008 32 bit and when I'm typing %temp% it takes me to ..AppData\Local\Temp\1\ or ..AppData\Local\Temp\2 instead of ..AppData\Local\Temp


I've already checked the Environment Variables and they are ok.


I'm asking that because I have an application that's supposed to copy files to %temp%\my application and instead it copies the files to %temp%\1\my application


Does someone have a solution for that?


","Explanation, discussion, workarounds, etc, all under the Old New Thing article [Why does the name of my TEMP directory keep changing](http://blogs.msdn.com/b/oldnewthing/archive/2011/01/25/10119675.aspx) - it ensures multiple terminal services sessions for the same user don't (by default) share a temp directory. 



> 
> The problem lay in the [Administrative Templates\Windows Components\Terminal Services\Temporary folders](http://technet.microsoft.com/en-us/library/cc759190%28WS.10%29.aspx) group policy. If you don't select **Do not use temporary folders per session**, then these TEMP subdirectories are created. There is also [a knowledge base article describing the registry keys behind these group policies](http://support.microsoft.com/kb/243215).
> 
> 
> 


Personally, I find the first comment in reply to be the most useful:



> 
> Why does it matter if the temp location jumps around anyway, it's temporary!
> 
> 
> 


## See also


- [Use separate temporary folders for each session](http://technet.microsoft.com/en-us/library/cc759190%28WS.10%29.aspx)


"
"How to sort date with yyyy-mm-dd format. By using datatables [js]?
I need to sort date with yyyy-mm-dd format. 


[From this document](http://datatables.net/blog/2014-12-18). Here is my code.



```
$( function () {
    $.fn.dataTable.moment('YYYY-MM-DD');

    $('#tbTest').DataTable();
});

```

But I still got the wrong sort (See picture below)


![enter image description here](https://i.stack.imgur.com/D60YC.png)


How can I fix this?


","If you check the unshift function in the datatable plugin you see that moment has the strict parameter defined true. 



```
    // Add type detection
    types.detect.unshift( function ( d ) {
    return moment( d, format, locale, true ).isValid() ?
        'moment-'+format :
        null;
    } );

```

From the [moment docs](http://momentjs.com/docs/#/parsing/):



> 
> Moment's parser is very forgiving, and this can lead to undesired behavior. As of version 2.3.0, you may specify a boolean for the last argument to make Moment use strict parsing. Strict parsing requires that the format and input match exactly.
> 
> 
> 


So instead of:



```
$.fn.dataTable.moment('YYYY-MM-DD');

```

which would require a date like 2015-04-01, you have to use:



```
$.fn.dataTable.moment('YYYY-M-D');

```

"
"pandas data transformation long-wide-long
how to get data from this form (long representation of data):



```
import pandas as pd
df = pd.DataFrame({
    'c0': ['A','A','B'],
    'c1': ['b','c','d'],
    'c2': [1, 3,4]})

print(df)

```

Out:



```
   c0 c1  c2
0  A  b   1
2  A  c   3
3  B  d   4

```

to this form :



```
   c0 c1  c2
0  A  b   1
2  A  c   3
3  A  d   NaN
4  B  b   NaN
5  B  c   NaN
6  B  d   4

```

Is long to wide to long transformation the only approach to doing this? 


","***method 1***  

`unstack` and `stack`



```
df.set_index(['c0', 'c1']).unstack().stack(dropna=False).reset_index()

```

[![enter image description here](https://i.stack.imgur.com/1dLr7.png)](https://i.stack.imgur.com/1dLr7.png)


***method 2***  

`reindex` with product



```
df.set_index(['c0', 'c1']).reindex(
    pd.MultiIndex.from_product([df.c0.unique(), df.c1.unique()], names=['c0', 'c1'])
).reset_index()

```

[![enter image description here](https://i.stack.imgur.com/1dLr7.png)](https://i.stack.imgur.com/1dLr7.png)


"
"Grammar production class implementation in C#
Grammar by definition contains productions, example of very simple grammar:



```
E -> E + E
E -> n

```

I want to implement Grammar class in c#, but I'm not sure how to store productions, for example how to make difference between terminal and non-terminal symbol.
i was thinking about:



```
struct Production
{
   String Left;       // for example E
   String Right;      // for example +
}

```

Left will always be non-terminal symbol (it's about context-free grammars)
But right side of production can contain terminal & non-terminal symbols


So now I'm thinkig about 2 ways of implementation:


1. Non-terminal symbols will be written using brackets, for example:


E+E will be represented as string ""[E]+[E]""
2. Create additional data structure NonTerminal


struct NonTerminal
{
 String Symbol;
}


and E+E will be represented as array/list:



```
[new NonTerminal(""E""), ""+"", new NonTerminal(""E"")]

```

but think that there are better ideas, it would be helpfull to hear some response


","I'd use 



```
 Dictionary<NonTerminalSymbol,Set<List<Symbol>>> 

```

enabling lookup by Nonterminal of the *set* of production rule right-hand-sides (themselves represented as lists of Terminal/Nonterminal Symbols) associated with the Nonterminal. (OP's question shows that the Nonterminal E might be associated with *two* rules, but we only need the right-hand sides if we have the left hand side).


This representation works only for a vanilla BNF grammar definitions, in which there is no syntactic sugar for common grammar-defining idioms. Such idioms typically include *choice*, *Kleene star/plus*, ... and when they are avialable in defining the grammar you get an so-called Extended BNF or EBNF. If we write EBNF only allowing *choice* denoted by **|**, the Expression grammar in flat form hinted at by OP as an example is:



```
         E = S ;
         S = P | S + P | S - P ; 
         P = T | P * T | P / T ;
         T = T ** M | ( E ) | Number | ID ;

```

and my first suggestion *can* represent this, because the alternation is only used to show different rule right-hand-sides. However, it won't represent this:



```
         E = S ;
         S = P A* ;
         A = + P | - P ;
         P = T M+ ; -- to be different
         M = * T | / T ;
         T = T ** M | ( E ) | Number | ID | ID ( E  ( # | C) * ) ; -- function call with skipped parameters
         C = , E ;

```

The key problem that this additional notation introduces is the ability to compose the WBNF operators repeatedly on sub-syntax definitions, and that's the whole point of EBNF. 


To represent EBNF, you have to store productions essentially as *trees* that represent the, well, expression structure of the EBNF (in fact, this is essentially the same problem as representing any expression grammar).


To represent the EBNF (expression) tree, you need to define the tree structure of the EBNF. 
You need tree nodes for:


- symbols (terminal or not)
- Alternation (having a list of alternatives)
- Kleene \*
- Kleene +
- ""Optional"" ?
- others that you decide your EBNF has as operators (e.g., comma'd lists, a way to say that one has a list of grammar elements seperated by a chosen ""comma"" character, or ended by a chosen ""semicolon"" character, ...)


The easiest way to do that is to first write an EBNF grammar for the EBNF itself:



```
EBNF = RULE+ ;
RULE = LHS ""="" TERM* "";"" ;
TERM = STRING | SYMBOL | TERM ""*"" 
       | TERM ""+"" | ';' STRING TERM | "","" TERM STRING 
      ""("" TERM* "")"" ;

```

Note that I've added comma'd and semicolon'ed list to the EBNF (extended, remember?)


Now we can simply inspect the EBNF to decide what is needed.
What you now need is a set of records (OK, classes for C#'er) to represent each of these rules.
So:


- a class for EBNF that contains a set of rules
- a class for a RULE having an LHS symbol and a LIST
- an abstract base class for TERM with several concrete variants, one for each alternative of TERM (a so-called ""discriminated union"" typically implemented by inheritance and instance\_of checks in an OO language).


Note that some of the concrete variants can refer to other class types in the representation, which is how you get a tree. For instance:



```
   KleeneStar inherits_from TERM {
        T: TERM:
   }

```

Details left to the reader for encoding the rest.


This raises an *unstated* problem for the OP: how do you *use* this grammmar representation to drive parsing of strings?


The simple answer is *get a parser generator*, which means you need to figure out what EBNF *it* uses. (In this case, it might simply be easier to store your EBNF as text and hand it to that parser generator, which kind of makes this whole discussion moot).


If you can't get one (?), or want to build one of your own, well, now you have the representation you need to climb over to build it. One other alternative is to build a recursive descent parser driven by this representation to do your parsing. The approach to do that is too large to contain in the margin of this answer, but is straightforward for those with experience with recursion.


EDIT 10/22: OP clarifies that he insists on parsing *all context free grammars* and ""especially NL"". For all context free grammars, he will need very a stong parsing engine (Earley, GLR, full backtracking, ...). For Natural Language, he will need parsers much stronger than those; people have been trying to build such parsers for decades with only some, but definitely not easy, success. Either of these two requirements seems to make the discussion of representing the grammar rather pointless; if he does represent a straight context free grammar, it won't parse natural language (proven by those guys trying for decades), and if he wants a more powerful NL parser, he'll need to simply use what the bleeding edge types have produced. Count me a pessimist on his probable success, unless he decides to become a real expert in the area of NL parsing.


"
"Elastica Client with cluster
I have an Elasticsearch cluster with Compose.io but I can't connect with Elastica Client. This is my configuration :



```
    $elasticaClient = new \Elastica\Client(array(
        'servers' => array(
            array('host' => 'https://myusername:mypass@aws-us-east-1-portal2.dblayer.com', 'port' => 10050),
            array('host' => 'https://myusername:mypass@aws-us-east-1-portal3.dblayer.com', 'port' => 10062)
        )
    ));
    $elasticaIndex = $elasticaClient->getIndex('test');

```

I got this error :


Couldn't resolve host
500 Internal Server Error


How to correctly connect to the database ?


","Parameter `host` must be specified without protocol. 


If you want to use https, you should set `transport` parameter to `Https` (instead of `Http` which used by default).



```
$elasticaClient = new \Elastica\Client([
    'connections' => [
        ['transport' => 'Https', 'host' => 'myusername:mypass@aws-us-east-1-portal2.dblayer.com', 'port' => 10050],
        ['transport' => 'Https', 'host' => 'myusername:mypass@aws-us-east-1-portal3.dblayer.com', 'port' => 10062],
    ],
]);
$elasticaIndex = $elasticaClient->getIndex('test');

```

"
"Comparing lists in Lisp
I could figure out some way to do this myself but I have a feeling there's a simpler, perhaps built-in way to do this. I want to see if any two lists share an element. These are the two lists I'm dealing with at the moment:



```
((0 1 2) (3 4 5) (6 7 8) (0 3 6) (1 3 7) (2 4 8) (0 4 8) (2 4 6))

((0 1 7) (0 1 6) (0 1 3) (0 3 7) (0 3 6) (0 6 7) (1 3 7) (1 3 6) (1 6 7) (3 6 7)) 

```

Since both lists contain `(1 3 7)`, I'd like a comparison of the lists to return `T`.


Is there a better way to do this than just setting up a couple `DOLIST`s?


","How about [INTERSECTION](http://www.ai.mit.edu/projects/iiip/doc/CommonLISP/HyperSpec/Body/fun_intersect_intersection.html)?



```
(defvar a '((0 1 2) (3 4 5) (6 7 8) (0 3 6) (1 3 7) (2 4 8) (0 4 8) (2 4 6)))
=> A
(defvar b '((0 1 7) (0 1 6) (0 1 3) (0 3 7) (0 3 6) (0 6 7) (1 3 7) (1 3 6) (1 6 7) (3 6 7)))
=> B
(intersection a b :test 'equal)
=> ((1 3 7) (0 3 6))

```

"
"After injecting html by jquery, the event handlers doesn't work with/without delegate
I have a list of `<div>` s with same html but different values inside html. The hierarchy is the following ;



```
<div id=""element"">
    <div class=""likecomm"">
        <a class=""commenticon"" href=""#"">...some value according to the returning value...</a>
    </div>
</div>

<div id=""element"">
    <div class=""likecomm"">
        <a class=""commenticon"" href=""#"">...some value according to the returning value...</a>
    </div>
</div>

```

In an event, I inject html to the top of the list which is another `<div id=""element> ... </div>`


I have the following event handler for comment icon click;



```
$('.commenticon').click(function(){
    $(this).closest('.actionframe').next('nav').slideToggle(300);
    return false;
});

```

It works correct until I insert the new `<div id=""element> ... </div>`. The comment icon click event handler doesn't match. I searched about that problem, all were saying that `.delegate()` should be used. 


My problem is I don't know where to use the delegate function. Before, I get the results from ajax in order to inject the `<div id=""element> ... </div>` , I used `.delegate()` function in ajax call that injects html.



```
$.ajax({

    success:function(data) {

        var html=... // the necessary div html binded with data
        // prepend function call()

        $('.likecomm').delegate("".commenticon"",""click"" , function() {                      
                       $(this).closest('.actionframe').next('nav').slideToggle(300);
                       return false;
        }); 

    }        
});

```

Currently, it doesn't work. So, any ideas how to make this thing work ?


","You need to bind the event handler to a common ancestor of the elements on which it should be triggered. For example, if your `#element` gets appended inside a `div` with an `id` of `parent`:



```
$(""#parent"").delegate("".commenticon"", ""click"", function() {
    //Do stuff
});

```

This would be for an HTML structure like so:



```
<div id=""parent"">
    <div class=""element"">

    </div>
    <div class=""element"">

    </div>
</div>

```

The reason this works is that DOM events bubble up the tree from the point at which they originate. The `delegate` method captures the event at an ancestor element and checks whether or not it originated at an element matching the selector.


Also note that it's invalid to have duplicate `id` values in the same document. Here I've changed your `element` ID values to class names instead.


Finally, if you are using jQuery 1.7+ you should use the [`on`](http://api.jquery.com/on/) method instead. It will have the same effect, but notice the reversal of the first 2 arguments:



```
$(""#parent"").on(""click"", "".commenticon"", function() {
    //Do stuff
};

```

"
"How to get a data structure from inside a string (Python)
I have a program which outputs a list in a .txt file on a server and I want to save that list in a new variable.  

I'm doing: 



```
with open (""file.txt"", ""r"") as myfile:
    data = myfile.read().replace('\n', '')
var = data

```

but obviously `print(var)` returns a string `""[list, of, stuff]""`


Yeah I can bypass the txt-save procedure but what in case I could not?


","you can use AST



```
import ast
var=""[list, of, stuff]""
var=ast.literal_eval(var)
print var[0]
#output list

```

From docs



> 
> *ast.literal\_eval(node\_or\_string) Safely evaluate an expression node or a string containing a Python expression. The string or node
>  provided may only consist of the following Python literal structures:
>  strings, bytes, numbers, tuples, lists, dicts, sets, booleans, and
>  None.*
> 
> 
> 


If you're original data was serialized using [JSON](http://json.org/) then you could also use the [json](https://docs.python.org/2/library/json.html) module from the Python standard library.


Even if it was just `print(some_list)` you could still use the [json](https://docs.python.org/2/library/json.html) library; i.e:



```
>>> from json import dumps, loads
>>> xs = [1, 2, 3]
>>> loads(dumps(xs)) == loads(repr(xs))
True

```

"
"EF Code First, exposing a linked table/object from a many to many relationship?
I am trying to expose the linked class that EF code first automatically creates in many to many relationships as a separate object because **that linked object needs to be referenced in other classes**, however I seem to have a problem with getting the data that exists in the database.


I've got the following 3 objects:



```
 public class Role : Entity
{
    public virtual ICollection<User> Users { get; set; }
    public string Name { get; set; }
    public string Description { get; set; }
    public virtual ICollection<Permission> Permissions { get; set; }
}

public class User: Entity
{
    public string FirstName { get; set; }
    public string MiddleName { get; set; }
    public string LastName { get; set; }

    public virtual ICollection<Role> Roles { get; set; }
 }

 public class UserRole : Entity
{
    public User User { get; set; }
    public Role Role { get; set; }
}

```

This creates the following tables:


![enter image description here](https://i.stack.imgur.com/FCPLS.png)


Now I can see the problem is that it is creating a `RoleUsers` table, when it shouldn't and should just use my `UserRoles` table. How can I force the linked table to be `UserRoles` so that I can have that linked object exposed in EF so I can use it in other objects?


Also, how would I be traversing the objects then when writing queries? Would I still be able to use User.Roles.Any(y => y.Name == ""blah""), or would I now have to query by going User.UserRoles.Any(y => y.Role.Name == ""blah"")? I suppose it isn't possible to maintain the transparent linking while simultaneously exposing the linked object?


Edit: One of the classes that references UserRole as a navigation property looks like this:



```
 public class UserRoleEntity : Entity
{
    public UserRole UserRole { get; set; }
    public Guid EntityId { get; set; }
    public EntityType EntityType { get; set; }
}

```

I don't want to individually store the User and Role object in this class because then they are not tightly coupled and data could be orphaned.


","You don't need to explicitly create `UserRole` entity in your code. since you have the navigational property of `Roles` in `Users` and `Users` in `Roles`, EF will automatically create a relationship between them.


EDIT : If you want to create the link table and use that as a property of another class, you can do it like this. 



```
public class User 
{
    public int UserID { set; get; }
    public string FirstName { get; set; }     
    public string LastName { get; set; }
    public virtual ICollection<UserRole> UserRoles { get; set; }      
}
public class Role 
{      
    public int RoleID { set;get;}
    public string Name { get; set; }
    public string Description { get; set; }
    public virtual ICollection<UserRole> UserRoles { get; set; }
}
public class UserRole 
{
    public int UserRoleID { set; get; }
    public virtual User User { get; set; }
    public virtual Role Role { get; set; }
}
public class AnotherEntity
{
    public int ID { set; get; }
    public int UserRoleID { set; get; }     
}

```

and you will have your tables created like this


![enter image description here](https://i.stack.imgur.com/EBd53.png)


And you are correct, You have to access it like this



```
 StringBuilder stRoleNames = new StringBuilder();
 var user1 = dbContext.Users.Where(x => x.UserID == 34).SingleOrDefault();
 var userRoles = user1.UserRoles;
 foreach (var userRole in userRoles)
 {
     stRoleNames.Append(userRole.Role.Name);
 } 

```

"
"Terraform - Place variable inside EOF tag
I have a terraform file that I'm reusing to create several AWS Eventbridge (as triggers for some lambdas).
On a different part of the file I'm able to use `For Each` method to create several eventbridge and naming them accordingly. My problem is that I'm not being able to do the same thing inside EOF tag (that need to be different on each Eventbridge) since it takes everything as a string.
I would need to replace the ARN in `""prefix"": ""arn:aws:medialive:us-west-2:11111111111:channel:3434343""` with a variable. How could I do that?


This is the EOF part of the terraform code:



```
      event_pattern = <<EOF
{
  ""source"": [""aws.medialive""],
  ""detail-type"": [""AWS API Call via CloudTrail""],
  ""detail"": {
    ""eventSource"": [""medialive.amazonaws.com""],
    ""eventName"": [""StopChannel""],
    ""responseElements"": {
      ""arn"": [{
        ""prefix"": ""arn:aws:medialive:us-west-2:11111111111:channel:3434343""
      }]
    }
  }
}
EOF
}

```

","It's called a [Heredoc String](https://www.terraform.io/language/expressions/strings), not an EOF tag. `""EOF""` just happens to be the string you are using to tag the beginning and ending of a multi-line string. You could use anything there that doesn't occurr in your actual multiline string. You could be replacing `""EOF""` with `""MYMULTILINESTRING""`.


To place the value of a variable in a Heredoc String in Terraform, you do the exact same thing you would do with other strings in Terraform: You use [String Interpolation](https://www.terraform.io/language/expressions/strings).



```
      event_pattern = <<EOF
{
  ""source"": [""aws.medialive""],
  ""detail-type"": [""AWS API Call via CloudTrail""],
  ""detail"": {
    ""eventSource"": [""medialive.amazonaws.com""],
    ""eventName"": [""StopChannel""],
    ""responseElements"": {
      ""arn"": [{
        ""prefix"": ""${var.my_arn_variable}""
      }]
    }
  }
}
EOF
}

```

"
"Given a regular expression, how would I generate all strings that match it?
I'm using a simple language of only `()`, `|`, spaces, and alpha characters.  

Given a regular expression like the following:



```
(hello|goodbye) (world(s|)|)

```

How would I go about generating the following data? 



```
hello worlds
hello world
hello 
goodbye worlds
goodbye world
goodbye

```

I'm not quite sure if I need to build a tree first, or if it can be done recursively. I'm stuck on what data structures to utilize, and how to generate the strings as I go. Will I have to keep a bunch of markers, and index back into partially built strings to concatenate more data on? I don't know how best to approach this problem. Would I need to read the whole expression first, and re-order it a certain way?


The function signature is going to look the following way:



```
std::vector<std::string> Generate(std::string const&){
   //...
}

```

What do you suggest I do?


**EDIT:**  

Let me clarify that the results should always be finite here. In my particular example, there are only 6 strings that would ever be true for the expression. I'm not sure if my terminology is correct here, but what I'm looking for, is a perfect match of the expression- not any string that contains a substring which matches.


","Somewhat following [Kieveli](https://stackoverflow.com/users/15852/kieveli)'s advice, I have come up with a working solution. Although not previously mentioned, it was important for me to also get a count of how many results could potentially be generated. I was using a python script called ""[exrex](https://github.com/asciimoo/exrex)"" which I had found on github. Embarrassingly, I did not realize that it had the capability to also count. Nonetheless, I implemented it the best I could in C++ using my simplified regular expression language. If interested in my solution, please read on.


From an object oriented stand point, I wrote a scanner to take the regular expression(string), and convert it into a list of tokens(vector of strings). The list of tokens was then sent to a parser which generated an n-ary tree. All of this was packed inside an ""expression generator"" class that could take an expression and hold the parse tree, as well as the generated count.  

![object overview](https://i.stack.imgur.com/GxmAe.png)  

The scanner was important because it tokenized the empty string case which you can see in my question appearing as ""|)"". Scanning also created a pattern of [word] [operation] [word] [operation] ... [word].  

For example, scanning: `""(hello|goodbye) (world(s|)|)""`  

will create: `[][(][hello][|][goodbye][)][ ][(][world][(][s][|][][)][][|][][)][]`


The parse tree was a vector of nodes. Nodes contain a vector of vector of nodes.
![parse structure](https://i.stack.imgur.com/Kpb36.jpg)  

The orange cells represent the ""or""s, and the other boxes that draw the connections, represent the ""and""s. Below is my code.


**Node header**



```
#pragma once
#include <string>
#include <vector>

class Function_Expression_Node{

public:
    Function_Expression_Node(std::string const& value_in = """", bool const& more_in = false);

    std::string value;
    bool more;
    std::vector<std::vector<Function_Expression_Node>> children;

};

```

**Node source**



```
#include ""function_expression_node.hpp""

    Function_Expression_Node::Function_Expression_Node(std::string const& value_in, bool const& more_in)
    : value(value_in)
    , more(more_in)
    {}

```

**Scanner header**



```
#pragma once
#include <vector>
#include <string>

class Function_Expression_Scanner{

    public: Function_Expression_Scanner() = delete;
    public: static std::vector<std::string> Scan(std::string const& expression);

};

```

**Scanner source**



```
#include ""function_expression_scanner.hpp""

std::vector<std::string> Function_Expression_Scanner::Scan(std::string const& expression){

    std::vector<std::string> tokens;
    std::string temp;

    for (auto const& it: expression){

        if (it == '('){
            tokens.push_back(temp);
            tokens.push_back(""("");
            temp.clear();
        }

        else if (it == '|'){
            tokens.push_back(temp);
            tokens.push_back(""|"");
            temp.clear();
        }

        else if (it == ')'){
            tokens.push_back(temp);
            tokens.push_back("")"");
            temp.clear();
        }

        else if (isalpha(it) || it == ' '){
            temp+=it;
        }

    }

    tokens.push_back(temp);

    return tokens;
    }

```

**Parser header**



```
#pragma once
#include <string>
#include <vector>
#include ""function_expression_node.hpp""

class Function_Expression_Parser{

    Function_Expression_Parser() = delete;

//get parse tree
public: static std::vector<std::vector<Function_Expression_Node>> Parse(std::vector<std::string> const& tokens, unsigned int & amount);
    private: static std::vector<std::vector<Function_Expression_Node>> Build_Parse_Tree(std::vector<std::string>::const_iterator & it, std::vector<std::string>::const_iterator const& end, unsigned int & amount);
        private: static Function_Expression_Node Recursive_Build(std::vector<std::string>::const_iterator & it, int & total); //<- recursive

    //utility
    private: static bool Is_Word(std::string const& it);
};

```

**Parser source**



```
#include ""function_expression_parser.hpp""

bool Function_Expression_Parser::Is_Word(std::string const& it){
        return (it != ""("" && it != ""|"" && it != "")"");
    }
Function_Expression_Node Function_Expression_Parser::Recursive_Build(std::vector<std::string>::const_iterator & it, int & total){

    Function_Expression_Node sub_root("""",true); //<- contains the full root
    std::vector<Function_Expression_Node> root;

    const auto begin = it;

    //calculate the amount
    std::vector<std::vector<int>> multiplies;
    std::vector<int> adds;
    int sub_amount = 1;

    while(*it != "")""){

        //when we see a ""WORD"", add it.
        if(Is_Word(*it)){
            root.push_back(Function_Expression_Node(*it));
        }

        //when we see a ""("", build the subtree,
        else if (*it == ""(""){
            ++it;
            root.push_back(Recursive_Build(it,sub_amount));

            //adds.push_back(sub_amount);
            //sub_amount = 1;
        }

        //else we see an ""OR"" and we do the split
        else{
            sub_root.children.push_back(root);
            root.clear();

            //store the sub amount
            adds.push_back(sub_amount);
            sub_amount = 1;
        }

        ++it;
    }

    //add the last bit, if there is any
    if (!root.empty()){
        sub_root.children.push_back(root);

        //store the sub amount
        adds.push_back(sub_amount);
    }
    if (!adds.empty()){
        multiplies.push_back(adds);
    }


    //calculate sub total
    int or_count = 0;
    for (auto const& it: multiplies){
        for (auto const& it2: it){
            or_count+=it2;
        }

        if (or_count > 0){
            total*=or_count;
        }
        or_count = 0;
    }

    /*
    std::cout << ""---SUB FUNCTION---\n"";
    for (auto it: multiplies){for (auto it2: it){std::cout << ""{"" << it2 << ""} "";}std::cout << ""\n"";}std::cout << ""--\n"";
    std::cout << total << std::endl << '\n';
    */

    return sub_root;
}
std::vector<std::vector<Function_Expression_Node>> Function_Expression_Parser::Build_Parse_Tree(std::vector<std::string>::const_iterator & it, std::vector<std::string>::const_iterator const& end, unsigned int & amount){

    std::vector<std::vector<Function_Expression_Node>> full_root;
    std::vector<Function_Expression_Node> root;

    const auto begin = it;

    //calculate the amount
    std::vector<int> adds;
    int sub_amount = 1;
    int total = 0;

    while (it != end){

        //when we see a ""WORD"", add it.
        if(Is_Word(*it)){
            root.push_back(Function_Expression_Node(*it));
        }

        //when we see a ""("", build the subtree,
        else if (*it == ""(""){
            ++it;
            root.push_back(Recursive_Build(it,sub_amount));

        }

        //else we see an ""OR"" and we do the split
        else{
            full_root.push_back(root);
            root.clear();

            //store the sub amount
            adds.push_back(sub_amount);
            sub_amount = 1;
        }

        ++it;
    }

    //add the last bit, if there is any
    if (!root.empty()){
        full_root.push_back(root);

        //store the sub amount
        adds.push_back(sub_amount);
        sub_amount = 1;
    }

    //calculate sub total
    for (auto const& it: adds){ total+=it; }

    /*
    std::cout << ""---ROOT FUNCTION---\n"";
    for (auto it: adds){std::cout << ""["" << it << ""] "";}std::cout << '\n';
    std::cout << total << std::endl << '\n';
    */
    amount = total;

    return full_root;
}
std::vector<std::vector<Function_Expression_Node>> Function_Expression_Parser::Parse(std::vector<std::string> const& tokens, unsigned int & amount){

    auto it = tokens.cbegin();
    auto end = tokens.cend();
    auto parse_tree = Build_Parse_Tree(it,end,amount);
    return parse_tree;
}

```

**Generator header**



```
#pragma once
#include ""function_expression_node.hpp""

class Function_Expression_Generator{

    //constructors
    public: Function_Expression_Generator(std::string const& expression);
    public: Function_Expression_Generator();

    //transformer
    void Set_New_Expression(std::string const& expression);

    //observers
    public: unsigned int Get_Count();
    //public: unsigned int Get_One_Word_Name_Count();
    public: std::vector<std::string> Get_Generations();
        private: std::vector<std::string> Generate(std::vector<std::vector<Function_Expression_Node>> const& parse_tree);
            private: std::vector<std::string> Sub_Generate(std::vector<Function_Expression_Node> const& nodes);

private:
    std::vector<std::vector<Function_Expression_Node>> m_parse_tree;
    unsigned int amount;

};

```

**Generator source**



```
#include ""function_expression_generator.hpp""

#include ""function_expression_scanner.hpp""
#include ""function_expression_parser.hpp""

//constructors
Function_Expression_Generator::Function_Expression_Generator(std::string const& expression){
    auto tokens = Function_Expression_Scanner::Scan(expression);
    m_parse_tree = Function_Expression_Parser::Parse(tokens,amount);
}
Function_Expression_Generator::Function_Expression_Generator(){}

//transformer
void Function_Expression_Generator::Set_New_Expression(std::string const& expression){
    auto tokens = Function_Expression_Scanner::Scan(expression);
    m_parse_tree = Function_Expression_Parser::Parse(tokens,amount);
}

//observers
unsigned int Function_Expression_Generator::Get_Count(){
    return amount;
}
std::vector<std::string> Function_Expression_Generator::Get_Generations(){
    return Generate(m_parse_tree);
}
std::vector<std::string> Function_Expression_Generator::Generate(std::vector<std::vector<Function_Expression_Node>> const& parse_tree){
    std::vector<std::string> results;
    std::vector<std::string> more;

    for (auto it: parse_tree){
        more = Sub_Generate(it);
        results.insert(results.end(), more.begin(), more.end());
    }

    return results;
}
std::vector<std::string> Function_Expression_Generator::Sub_Generate(std::vector<Function_Expression_Node> const& nodes){
    std::vector<std::string> results;
    std::vector<std::string> more;
    std::vector<std::string> new_results;

    results.push_back("""");
    for (auto it: nodes){
        if (!it.more){
            for (auto & result: results){
                result+=it.value;
            }
        }
        else{
            more = Generate(it.children);
            for (auto m: more){
                for (auto r: results){
                    new_results.push_back(r+m);
                }
            }
            more.clear();
            results = new_results;
            new_results.clear();
        }
    }

    return results;
}

```

In conclusion, I recommend using [exrex](https://github.com/asciimoo/exrex), or any other programs mentioned in this thread, if you need to generate matches for regular expressions.


"
"Why do you have to call URLConnection#getInputStream to be able to write out to URLConnection#getOutputStream?
I'm trying to write out to [`URLConnection#getOutputStream`](http://download.oracle.com/javase/1.5.0/docs/api/java/net/URLConnection.html#getOutputStream%28%29), however, no data is actually sent until I call [`URLConnection#getInputStream`](http://download.oracle.com/javase/1.5.0/docs/api/java/net/URLConnection.html#getInputStream%28%29). Even if I set [`URLConnnection#doInput`](http://download.oracle.com/javase/1.5.0/docs/api/java/net/URLConnection.html#doInput) to false, it still will not send. Does anyone know why this is? There's nothing in the API documentation that describes this.


Java API Documentation on URLConnection: <http://download.oracle.com/javase/6/docs/api/java/net/URLConnection.html>


Java's Tutorial on Reading from and Writing to a URLConnection: <http://download.oracle.com/javase/tutorial/networking/urls/readingWriting.html>



```
import java.io.IOException;
import java.io.OutputStreamWriter;
import java.net.URL;
import java.net.URLConnection;

public class UrlConnectionTest {

    private static final String TEST_URL = ""http://localhost:3000/test/hitme"";

    public static void main(String[] args) throws IOException  {

        URLConnection urlCon = null;
        URL url = null;
        OutputStreamWriter osw = null;

        try {
            url = new URL(TEST_URL);
            urlCon = url.openConnection();
            urlCon.setDoOutput(true);
            urlCon.setRequestProperty(""Content-Type"", ""text/plain"");            

            ////////////////////////////////////////
            // SETTING THIS TO FALSE DOES NOTHING //
            ////////////////////////////////////////
            // urlCon.setDoInput(false);

            osw = new OutputStreamWriter(urlCon.getOutputStream());
            osw.write(""HELLO WORLD"");
            osw.flush();

            /////////////////////////////////////////////////
            // MUST CALL THIS OTHERWISE WILL NOT WRITE OUT //
            /////////////////////////////////////////////////
            urlCon.getInputStream();

            /////////////////////////////////////////////////////////////////////////////////////////////////////////
            // If getInputStream is called while doInput=false, the following exception is thrown:                 //
            // java.net.ProtocolException: Cannot read from URLConnection if doInput=false (call setDoInput(true)) //
            /////////////////////////////////////////////////////////////////////////////////////////////////////////

        } catch (Exception e) {
            e.printStackTrace();                
        } finally {
            if (osw != null) {
                osw.close();
            }
        }

    }

}

```

","The API for URLConnection and HttpURLConnection are (for better or worse) designed for the user to follow a very specific sequence of events:


1. Set Request Properties
2. (Optional) getOutputStream(), write to the stream, close the stream
3. getInputStream(), read from the stream, close the stream


If your request is a POST or PUT, you need the optional step #2.


To the best of my knowledge, the OutputStream is not like a socket, it is not directly connected to an InputStream on the server. Instead, after you close or flush the stream, AND call getInputStream(), your output is built into a Request and sent. The semantics are based on the assumption that you will want to read the response. Every example that I've seen shows this order of events. I would certainly agree with you and others that this API is counterintuitive when compared to the normal stream I/O API.


The [tutorial](http://download.oracle.com/javase/tutorial/networking/urls/readingWriting.html) you link to states that ""URLConnection is an HTTP-centric class"". I interpret that to mean that the methods are designed around a Request-Response model, and make the assumption that is how they will be used.


For what it's worth, I found this [bug report](http://bugs.sun.com/bugdatabase/view_bug.do?bug_id=4192018) that explains the intended operation of the class better than the javadoc documentation. The evaluation of the report states ""The only way to send out the request is by calling getInputStream.""


"
"Angular 2 Shared Data Service is not working
I have built a shared data service that's designed to hold the users login details which can then be used to display the username on the header, but I cant get it to work.


Here's my (abbreviated) code:



```
// Shared Service
@Injectable()
export class SharedDataService {

    // Observable string source
    private dataSource = new Subject<any>();

    // Observable string stream
    data$ = this.dataSource.asObservable();

    // Service message commands
    insertData(data: Object) {
        this.dataSource.next(data)
    }
}

```

...



```
// Login component
import { SharedDataService } from 'shared-data.service';
@Component({
    providers: [SharedDataService]
})
export class loginComponent {
    constructor(private sharedData: SharedDataService) {}

    onLoginSubmit() {
        // Login stuff
        this.authService.login(loginInfo).subscribe(data => {
             this.sharedData.insertData({'name':'TEST'});
        }
    }
}

```

...



```
// Header component
import { SharedDataService } from 'shared-data.service';
@Component({
    providers: [SharedDataService]
})
export class headerComponent implements OnInit {
    greeting: string;

    constructor(private sharedData: SharedDataService) {}

    ngOnInit() {
        this.sharedData.data$.subscribe(data => {
            console.log('onInit',data)
            this.greeting = data.name
        });
    }
}

```

I can add a console log in the service insertData() method which shoes the model being updated, but the OnInit method doesn't reflect the change.


The code I've written is very much inspired by this [plunkr](https://plnkr.co/edit/neM6EdYYUkGkRpF0fKGS?p=preview) which does work, so I am at a loss as to what's wrong.


Before posting here I tried a few other attempts. [This one](https://plnkr.co/edit/EjxffJLwTAusMI6BtEUw?p=preview) and [this one](https://plnkr.co/edit/neM6EdYYUkGkRpF0fKGS?p=preview) again both work on the demo, but not in my app.


I'm using Angular 2.4.8.


Looking through different tutorials and forum posts all show similar examples of how to get a shared service working, so I guess I am doing something wrong. I'm fairly new to building with Angular 2 coming from an AngularJS background and this is the first thing that has me truly stuck.


Thanks


","This seems to be a recurring problem in understanding Angular's dependency injection.


The basic issue is in how you are configuring the providers of your service. 


**The short version:**
Always configure your providers at the NgModule level **UNLESS** you want a separate instance for a specific component. Only then do you add it to the providers array of the component that you want the separate instance of.


**The long version:**
Angular's new dependency injection system allows for you to have multiple instances of services if you so which (which is in contrast to AngularJS i.e. Angular 1 which ONLY allowed singletons). If you configure the provider for your service at the NgModule level, you'll get a singleton of your service that is shared by all components/services etc. But, if you configure a component to also have a provider, then that component (and all its subcomponents) will get a different instance of the service that they can all share. This option allows for some powerful options if you so require. 


That's the basic model. It, is of course, not quite so simple, but that basic rule of configuring your providers at the NgModule level by default unless you explicitly want a different instance for a specific component will carry you far. 


And when you want to dive deeper, [check out the official Angular docs](https://angular.io/docs/ts/latest/guide/hierarchical-dependency-injection.html)


Also note that lazy loading complicates this basic rule as well, so again, check the docs.


**EDIT:**


So for your specific situation,



```
@Component({
    providers: [SharedDataService]  <--- remove this line from both of your components, and add that line to your NgModule configuration instead
})

```

"
"Python set() and \_\_hash\_\_ confusion
One of the classes I defined is used in a `set()` to filter out equal objects. But it doesn't work as I expect it, so I obviously understand something wrong.



```
class Foo(object):

    def __hash__(self):
        return 7

x = set()
x.add(Foo())
assert len(x) == 1
x.add(Foo())
assert len(x) == 1    # AssertionError

```

I expect the set to consist of only one element, but it has two. Why is that?


","Hash collisions are know to occur in sets (hash maps), no hash algorithm is good enough to have a unique hash for every item, or else it would take a long time to calculate. When a collision does occur, python falls back to checking the equality of the values with `__eq__` to make sure they are not the same.



```
class Foo(object):

    def __hash__(self):
        return 7
    def __eq__(self, other):
        return True

>>> x = set()
>>> x.add(Foo())
>>> assert len(x) == 1
>>> x.add(Foo())
>>> assert len(x) == 1
>>> 

```

This is a reason why you see frightening runtimes over [here](http://wiki.python.org/moin/TimeComplexity) but note that you can expect O(1) amortized membership checks in sets, even though they have O(N) worst case (everything is a hash collision). The worst case is very very very unlikely to occur due to Python's smart implementation.


"
"Setting number of columns programmatically in TableLayout
I have an XML layout which contains a TableLayout with an unknown number of TableRows...
The number of Rows will be established durin runtime, what I do know though is that I want two columns...
So I have a couple of questions regarding this :


- is there a way to set the whole TableLayout to have 2 columns ?

- is there a way programmatically to give an id to the (during runtime) created TableRows which will be placed within the TableLayout, so I can reference them later on from other parts of the software ?


","You can build your table rows via XML parts and LayoutInflater. Say you had this as your table\_cell.xml:



```
<TextView android:id=""@+id/text""
    android:text=""woot"" />

```

And this as your table\_row.xml (unless you're doing something fancy with your TableRow, you may not need to put it in it's own XML file, and instead just create it programmatically. The result will be the same):



```
<TableRow />

```

Assuming your TableLayout reference was called ""table"", you could do something like this:



```
TableLayout table = (TableLayout)findViewById(R.id.table);
LayoutInflater inflater = getLayoutInflater();

for(int i = 0; i < 5; i++) {
    TableRow row = (TableRow)inflater.inflate(R.layout.table_row, table, false);
    View v = (View)inflater.inflate(R.layout.table_cell, row, false);
    row.addView(v);
    v = (View)inflater.inflate(R.layout.table_cell, row, false);
    row.addView(v);

    // you can store your reference to `row` here for later use
    table.addView(row);
}

```

With this technique, you can still set up your layout in XML, making it easier to read/organize/edit, and you still have programmatic control over how many columns and rows are in the table. You can also store references to each table row for later use.


"
"What is the difference between areNotificationsEnabled and areNotificationsPaused
What is the difference between the [areNotificationsEnabled](https://developer.android.com/reference/android/app/NotificationManager#areNotificationsEnabled()) and [areNotificationsPaused](https://developer.android.com/reference/android/app/NotificationManager#areNotificationsPaused()) methods in the [NotificationManager](https://developer.android.com/reference/android/app/NotificationManager) class?


","**areNotificationsEnabled**


The first method, i.e areNotificationsEnabled will be able to tell if the notifications we completely blocked by the user for your app or the package name you want
The user can block the notification in this way


[![Notification Block](https://i.stack.imgur.com/PIVjT.png)](https://i.stack.imgur.com/PIVjT.png)


**areNotificationsPaused**


The second method will be able to tell if the app was snoozed or paused in other terms, that is, a temporary stoppage of notifications from a certain app
The user can snooze notifications from the notification panel by sliding left or right partially and then set the time for which they want the notification to be snoozed


[![Here you go ](https://i.stack.imgur.com/xgt75.jpg)](https://i.stack.imgur.com/xgt75.jpg)


[![Notification Pause](https://i.stack.imgur.com/0HtjG.png)](https://i.stack.imgur.com/0HtjG.png)


"
"Spring Boot Redirect to requested URL after login
I have a Spring Boot UI application. I am trying to redirect users to the originally requested URL after login.


When a user requests <http://www.example.com/myapp/user/22>, the application aptly redirects to <http://www.example.com/myapp/login>. Once the user logs in, the application redirects to <http://www.example.com/myapp/dashboard>. I would like the application to redirect to <http://www.example.com/myapp/user/22>.


I have gone through several links and feel I have a proper configuration, yet, redirection is not working as expected.


My Security Config is 



```
public class SecurityConfig extends WebSecurityConfigurerAdapter {
.....
....

    @Autowired
    private MyAuthenticationSuccessHandler authenticationSuccessHandler;

    @Override
    protected void configure(HttpSecurity http) throws Exception {
        http.
        authorizeRequests()
                .antMatchers(""/user/**"").authenticated()
                .and().csrf().disable().formLogin()
                .successHandler(authenticationSuccessHandler)
......

```

and My Success Handler is



```
    @Component
    public class MyAuthenticationSuccessHandler extends SimpleUrlAuthenticationSuccessHandler {
    ...
public MyAuthenticationSuccessHandler() {
        super();
        this.setDefaultTargetUrl(""/myapp/dashboard"");
        this.setUseReferer(true);
    }

        @Override
        public void onAuthenticationSuccess(HttpServletRequest request, HttpServletResponse response,
                Authentication authentication) throws IOException, ServletException {
            //Do something ..........
            ........
            .........
            super.onAuthenticationSuccess(request, response, authentication);
}

```

I tried using SavedRequestAwareAuthenticationSuccessHandler too.


I notice that my success handler is invoked, but the target URL is always /user/login and my login controller is invoked..



```
@RequestMapping(""/login"")
public ModelAndView login(@ModelAttribute() {
    if(!userIdentified) {
        //go to login page
    } else {
        new ModelAndView(""redirect:/myapp/dashboard"");
    }
}

```

and the user is redirected to ""dashboard"".


What else am I missing?


","Use ""Referer"" from session attribute to get the latest request URL. On my app, i use this one



```
public class CustomAuthenticationSuccessHandler extends SimpleUrlAuthenticationSuccessHandler {

    public static final String REDIRECT_URL_SESSION_ATTRIBUTE_NAME = ""REDIRECT_URL"";

    @Override
    public void onAuthenticationSuccess(HttpServletRequest request, HttpServletResponse response,
            Authentication authentication) throws IOException, ServletException {

        Object redirectURLObject = request.getSession().getAttribute(REDIRECT_URL_SESSION_ATTRIBUTE_NAME);

        if(redirectURLObject != null)
            setDefaultTargetUrl(redirectURLObject.toString());
        else{
            setDefaultTargetUrl(""/"");
        }

        request.getSession().removeAttribute(REDIRECT_URL_SESSION_ATTRIBUTE_NAME);
        super.onAuthenticationSuccess(request, response, authentication);
    }

}

```

Edit :


Sorry i forgot to show the login controller



```
@RequestMapping(method = RequestMethod.GET, value = {""/login""})
    String login(Model model, Principal principal, HttpServletRequest request) throws Exception{
        String referer = request.getHeader(""Referer""); //Get previous URL before call '/login'

        //save referer URL to session, for later use on CustomAuthenticationSuccesshandler
        request.getSession().setAttribute(CustomAuthenticationSuccessHandler.REDIRECT_URL_SESSION_ATTRIBUTE_NAME, referer); 


        return principal == null ?  ""login"" : ""redirect:/""; 
    }

```

"
"setting values for ntree and mtry for random forest regression model
I'm using R package `randomForest` to do a regression on some biological data. My training data size is `38772 X 201`.


I just wondered---what would be a good value for the number of trees `ntree` and the number of variable per level `mtry`? Is there an approximate formula to find such parameter values?


Each row in my input data is a 200 character representing the amino acid sequence, and I want to build a regression model to use such sequence in order to predict the distances between the proteins. 


","The default for mtry is quite sensible so there is not really a need to muck with it. There is a function `tuneRF` for optimizing this parameter. However, be aware that it may cause bias. 


There is no optimization for the number of bootstrap replicates. I often start with `ntree=501` and then plot the random forest object. This will show you the error convergence based on the OOB error. You want enough trees to stabilize the error but not so many that you over correlate the ensemble, which leads to overfit. 


Here is the caveat: variable interactions stabilize at a slower rate than error so, if you have a large number of independent variables you need more replicates. I would keep the ntree an odd number so ties can be broken. 


For the dimensions of you problem I would start `ntree=1501`. I would also recommended looking onto one of the published variable selection approaches to reduce the number of your independent variables. 


"
"Why is this an example of a noninformative prior?
From *[Bayesian Data Analysis 3rd Edition](https://amzn.to/37QhtUh)* [Gelman et. al], they give this as an example when introducing non-informative priors:



> 
> ""We return to the problem of estimating the mean θ of a normal model
> with known variance $σ^2$ , with a $N(μ\_0 , τ\_0^2 )$ prior distribution on
> $θ$. If the prior precision, $1/τ\_0^2$, is small relative to the data
> precision, $n/σ^2$, then the posterior distribution is approximately as
> if $τ\_0^2 = ∞$:
> 
> 
> $$p(θ|y) ≈ N(θ|y, σ^2 /n)$$
> 
> 
> **Putting this another way, the posterior distribution is approximately
> that which would result from assuming $p(θ)$ is proportional to a
> constant for $θ ∈ (−∞, ∞)$. Such a distribution is not strictly
> possible, since the integral of the assumed $p(θ)$ is infinity, which
> violates the assumption that probabilities sum to 1.** In general, we
> call a prior density $p(θ)$ proper if it does not depend on data and
> integrates to 1. (If $p(θ)$ integrates to any positive finite value, it
> is called an unnormalized density and can be renormalized—multiplied
> by a constant—to integrate to 1.) The prior distribution is improper
> in this example, but the posterior distribution is proper, given at
> least one data point.""
> 
> 
> 


In particular, I don't really understand the bold part. It doesn't look to me like the marginal $p(θ)$ is proportional to a constant.
Is my understanding correct that the integral is infinity because if the variance of the posterior is approximately $∞$, then $θ$ is equally likely to be found anywhere, and the pdf is uniform on $(-∞, ∞)$.
Also, why is the posterior distribution proper given at least one data point?


I don't understand the intuition here


","I cannot really help with the ""intuition"" but here are some items of explanation:



> 
> It doesn't look to me like the marginal $p(θ)$ is proportional to a constant.
> 
> 
> 


If the posterior $p(\theta|y)$ is approximately $\mathcal N(θ|y, σ^2 /n)$, it means it is approximately equal to the likelihood since
$$f(y|\theta)=\frac{\exp(-n\{\theta-y\}^2/2\sigma^2)}{\sqrt{2\pi\sigma^2/n}}=f(\theta|y)$$by symmetry of the Normal density. Thus, $p(\theta|y)\approx f(y|\theta)\propto f(y|\theta)\times c$ where $c$ is an arbitrary constant and by Bayes theorem
$$p(\theta)f(y|\theta)\propto p(\theta|y) \propto f(y|\theta)\times c$$
where all the [proportionality symbols](https://stats.stackexchange.com/q/375140/7224) are in terms of functions of $\theta$ (as $y$ is observed, hence fixed).



> 
> Is my understanding correct that the integral is infinity because if
> the variance of the posterior is approximately $∞$, then $θ$ is equally
> likely to be found anywhere, and the pdf is uniform on $(−∞,∞)$.
> 
> 
> 


A positive function with an infinite mass, $$\int p(\theta)\,\text{d}\theta=\infty$$ cannot be interpreted as a probability density. Hence the probability of $\theta$ to be anywhere does not exist if $p(\theta)$ is improper. (It is *improperly* called a prior since it is not a probability density.) It is also improper to call a constant prior *uniform* because this is not a Uniform probability distribution. Note that (i) the integral of the function is infinity because it is constant, *not because* the variance is infinite (as there exist plenty of true probability distributions with an infinite variance). And (ii) the variance of the prior, not the posterior, is infinite, although it is again improper to call it a variance since it is not a true probability distribution.



> 
> why is the posterior distribution proper given at least one data point?
> 
> 
> 


It happens that with this specific improper prior, $p(\theta)=1$, the integral
$$\int f(y\_1|\theta)\,\text{d}\theta=1$$
is finite. The proof is direct since $f(y\_1|\theta)=f(\theta|y\_1)$ in this specific case. There exist infinitely other cases that require an arbitrary number $m$ of observations for the posterior to exist. And yet infinitely other cases that never produce posterior distributions, no matter the number of observations. Note also that the very name of [noninformative priors](https://stats.stackexchange.com/a/20535/7224) is itself debatable.


"
"How to develop line-of-business tablet apps as a .NET developer?
My company has a lot of experience in .NET development, and one of our products in an ERP system. Recently, a customer asked us if we could provide a tablet interface to that system, i.e., a software that allows the customer to view product information and create orders on a tablet.


Of course, we are not thrilled about the idea of investing a lot of time and money into learning Objective-C, purchasing Mac development workstations, paying fees to Apple, etc. just for this one project (we *might* be able to sell the app to a few additional customers afterwards, but the market is very small, since it would only be useful for existing customers of our ERP system).


So, what should we do? As far as I can see, we have the following options:


- Write a ""**plain old Windows application**"" (WPF) and run it on a Windows 7 tablet, such as the Samsung Slate or Acer Iconia.


Drawbacks: Heavy, expensive devices with short running time (as compared to ""real"" tablets).
- Wait for **Windows 8 ARM-based tablets** and write a Metro (WinRT) app.


Drawbacks: Wait for at least one year; it's unclear whether Windows 8 ARM will support installation of custom B2B apps without going through the app store.
- Use **[mono for Android](http://xamarin.com/)** and write a .NET app for Android.


Drawbacks: Yet another UI library (different from WPF and Silverlight); some providers disallow sideloading of apps.


So far, options 1 and 3 seem to be the most realistic. Did I miss any obvious drawbacks or advantages? Is there another option that I haven't considered yet? Have you been in a similar situation and (successfully) chosen one particular option?


","[JQuery Mobile](http://jquerymobile.com/) + [Phone Gap Build](http://build.phonegap.com). 


This is basically saying ""use HTML5 and JavaScript to build your app"", as has been said before, but with an important twist.


Nitobi's [Phone Gap Build](http://build.phonegap.com) service (now owned by Adobe) allows developers to convert HTML5/JavaScript apps to ""native"" apps (really hybrid apps) that can be deployed locally to a device. It's my understanding that basically what's happening under the hood is packaging a small native binary that calls the native browser and loads up your site from a file:// URL. 


You don't need to target any specific JavaScript framework - the same HTML & JavaScript that would work really well in a mobile web app will work fine. 


Offline support isn't hard, either. With local browser storage well supported with many mobile devices you can build truly powerful offline apps this way. Its best practice to package your external dependencies locally as opposed to using a CDN, so that your app works well offline. 


Frameworks like [KnockoutJS](http://knockoutjs.com/) and [BackboneJS](http://documentcloud.github.com/backbone/) are very helpful at allowing you to build well engineered JavaScript apps, and they work just fine with Phone Gap's build service.


When the device is online, you can easily have it hit ASP.NET/MVC, WebAPI or WCF service back ends to refresh data.


The resulting apps are really quite good, and can be distributed in the Apple and Android markets. There are already lots of apps in those marketplaces built with Phone Gap Build and other similar products, and 99% of people (including most devs) can't tell the difference.


Obviously you aren't going to try to build Angry Birds this way (though, with Canvas I suppose you might try), it works wonderfully well with the kinds of apps you are talking about.


Don't take my word for it. PhoneGap has been doing the rounds on the PodCast circuit, having recently been on [Hanselminutes](http://www.hanselminutes.com/304/making-your-first-phonegap-application-with-peter-mourfield), [DotNetRocks](http://www.dotnetrocks.com/default.aspx?showNum=557), and the [Tablet Show](http://thetabletshow.com/?ShowNum=17). Also, I wrote about it in a recent [blog post](http://paymentnetworks.wordpress.com/2012/02/14/4-indications-that-html5-mobile-development-is-getting-big/). 


"
"How multiple consumer group consumers work across partition on the same topic in Kafka?
I was reading [this SO answer](https://stackoverflow.com/questions/23136500/how-kafka-broadcast-to-many-consumer-groups) and many such blogs.


What I know:


Multiple consumers can run on a single partition when running multiple consumers with multiple consumer group id and only one consumer from a consumer group can consume at a given time from a partition.


My question is related to **multiple consumers** from **multiple consumer groups** consuming from the **same topic**:


1. What happens in the case of multiple consumers(different groups) consuming a single topic(eventually the same partition)?
2. Do they get the same data?
3. How offset is managed? Is it separate for each consumer?
4. (Might be opinion based) How do you or generally recommended way is to handle overlapping data across two consumers of a separate group operating on a single partition?


Edit:
""overlapping data"": means two consumers of separate consumer groups operating on the same partition getting the same data.


","1. Yes they get the same data. Kafka only stores one copy of the data in the topic partitions' commit log. If consumers are not in the same group then they can each get the same data using fetch requests from the clients' consumer library. The assignment of which partitions each group member will get is managed by the lead consumer of each group. The entire process in detailed steps is documented here <https://community.hortonworks.com/articles/72378/understanding-kafka-consumer-partition-assignment.html>
2. Offsets are ""managed"" by the consumers, but ""stored"" in a special \_\_consumer\_offsets topic on the Kafka brokers.
3. Offsets are stored for each `(consumer group, topic, partition)` tuple. This combination is also used as the key when publishing offsets to the \_\_consumer\_offsets topic so that log compaction can delete old unneeded offset commit messages and so that all offsets for the same `(consumer group, topic, partition)` tuple are stored in the same partition of the \_\_consumer\_offsets topic (which defaults to 50 partitions)


"
"Python mint hashcash token
This is a Python program to mint a hashcash token, but my code is a lot slower than using a library. What is slowing my program down? It takes over 10 seconds to mint a 20-bit stamp, but using a library, it takes less than a second.


This is my program:



```
from os import urandom
from hashlib import sha1
import time

def mint(name, bits):
    t = time.strftime(""%y%m%d"")
    out = "" ""
    count = 0
    while out[:bits] != bits*""0"":
        RandomData = urandom(8).encode(""base64"").replace(""\n"", """")
        data = ""1:%d:%s:%s::%s:%s""%(bits, t, name, RandomData, format(count, 'x'))
        out = (''.join(format(ord(x), '08b') for x in sha1(data).digest()))
        count += 1
    return data

print mint(""email_address"", 20)

```

[This](http://gnosis.cx/download/gnosis/util/hashcash.py) is the library I am using for comparison.


","The first thing to do is to profile your code. Hashcash is really designed to make you perform 2^20 hashes: if you do more work than those hashes, something's wrong!


## Profiling


`python -m cProfile -s cumtime original.py`


Let's look at all the places where the algorithms spends more than 0.001 seconds (I've asked for 15 bits since 20 bits is too slow on my computer) :



```
   Ordered by: internal time

   ncalls  tottime  percall  cumtime  percall filename:lineno(function)
  1414287    4.087    0.000    7.665    0.000 original.py:12(<genexpr>)
  1414287    2.433    0.000    2.433    0.000 {format}
   134695    1.737    0.000    9.402    0.000 {method 'join' of 'str' objects}
  1346940    1.295    0.000    1.295    0.000 {ord}
        1    0.886    0.886   12.709   12.709 original.py:5(mint)
    67347    0.536    0.000    0.969    0.000 base64.py:310(encodestring)
    67347    0.432    0.000    0.432    0.000 {posix.urandom}
    67347    0.258    0.000    1.292    0.000 base64_codec.py:13(base64_encode)
    67347    0.231    0.000    1.525    0.000 {method 'encode' of 'str' objects}
    67348    0.141    0.000    0.141    0.000 {_hashlib.openssl_sha1}
    67347    0.131    0.000    0.131    0.000 {method 'digest' of '_hashlib.HASH' objects}
   134694    0.131    0.000    0.131    0.000 {len}
    67347    0.126    0.000    0.126    0.000 {binascii.b2a_base64}
    67348    0.110    0.000    0.110    0.000 {range}
    67347    0.109    0.000    0.109    0.000 {method 'replace' of 'str' objects}
    67347    0.066    0.000    0.066    0.000 {method 'append' of 'list' objects}
        1    0.001    0.001    0.001    0.001 hashlib.py:55(<module>)
        1    0.001    0.001    0.001    0.001 base64.py:3(<module>)

```

This specific run took 12.7 seconds (original.py:5). It spent 4 seconds in the `for x in ...` generator, 2.4 seconds in `format(ord(x), '08b')`, 1.7 seconds in `''.join(...)`, and 1.3 seconds in `ord(x)`.


That's a total of... 9.5 seconds out of 12.7 seconds spent in the `out = (''.join(format(ord(x), '08b') for x in sha1(data).digest()))` line. This would have been impossible to figure out without profiling (as you can see in kyle k's comment). It's the number one rule of optimizations: since you'll spent 80% of your time in 20% of the code, only optimize what you know to be slow!


## Removing the 80%


So, how can we replace this line? The idea is to use `hexdigest` instead of `digest`. The issue with `hexdigest` is that we only see hexadecimal digits, not bits. Since 20 bits is 5 hexadecimal numbers, that's not an issue here. If you needed 21 bits, then you would ask for 6 hexadecimal numbers, which is more hashes than necessary, but still worth the speed up.


Here's the new code:



```
from os import urandom
from hashlib import sha1
import time
from math import ceil

def mint(name, bits):
    t = time.strftime(""%y%m%d"")
    out = "" ""
    count = 0
    digits = int(ceil(bits/4.0))

    while out[:digits] != digits*""0"":
        RandomData = urandom(8).encode(""base64"").replace(""\n"", """")
        RandomData = 'AHZlx9LYMyo='
        data = ""1:%d:%s:%s::%s:%s""%(bits, t, name, RandomData, format(count, 'x'))
        out = sha1(data).hexdigest()
        count += 1
    return data

print mint(""email_address"", 15)

```

And the new profiling:



```
   ncalls  tottime  percall  cumtime  percall filename:lineno(function)
        1    0.750    0.750    3.212    3.212 improved.py:6(mint)
    67347    0.529    0.000    0.959    0.000 base64.py:310(encodestring)
    67347    0.427    0.000    0.427    0.000 {posix.urandom}
    67347    0.251    0.000    1.274    0.000 base64_codec.py:13(base64_encode)
    67347    0.227    0.000    1.503    0.000 {method 'encode' of 'str' objects}
    67347    0.155    0.000    0.155    0.000 {format}
    67348    0.138    0.000    0.138    0.000 {_hashlib.openssl_sha1}
    67347    0.134    0.000    0.134    0.000 {method 'hexdigest' of '_hashlib.HASH' objects}
   134694    0.131    0.000    0.131    0.000 {len}
    67347    0.122    0.000    0.122    0.000 {binascii.b2a_base64}
    67348    0.109    0.000    0.109    0.000 {range}
    67347    0.105    0.000    0.105    0.000 {method 'replace' of 'str' objects}
    67348    0.066    0.000    0.066    0.000 {method 'join' of 'str' objects}
    67347    0.066    0.000    0.066    0.000 {method 'append' of 'list' objects}
        1    0.001    0.001    0.001    0.001 hashlib.py:55(<module>)
        1    0.001    0.001    3.213    3.213 improved.py:1(<module>)
        1    0.001    0.001    0.001    0.001 base64.py:3(<module>)

```

Yay! Only 3.2 seconds in `mint()`.


## Removing another 80%


But hashing is still not the most expensive part. It's now the computing of RandomData (0.5 + 0.2 seconds in base64 encoding, 0.4 seconds in urandom, and so on). @cHao is right, you only need to compute that RandomData once and find the count that provides the correct sha1 hash. Let's move it out of the while loop:



```
from os import urandom
from hashlib import sha1
import time
from math import ceil

def mint(name, bits):
    t = time.strftime(""%y%m%d"")
    out = "" ""
    count = 0
    digits = int(ceil(bits/4.0))
    RandomData = urandom(8).encode(""base64"").replace(""\n"", """") 
    RandomData = 'AHZlx9LYMyo='

    while out[:digits] != digits*""0"":
        data = ""1:%d:%s:%s::%s:%s""%(bits, t, name, RandomData, format(count, 'x'))
        out = sha1(data).hexdigest()
        count += 1
    return data

print mint(""email_address"", 15)

```

The new profile is:



```
   ncalls  tottime  percall  cumtime  percall filename:lineno(function)
        1    0.414    0.414    0.795    0.795 improved.py:6(mint)
    67347    0.134    0.000    0.134    0.000 {format}
    67348    0.124    0.000    0.124    0.000 {_hashlib.openssl_sha1}
    67347    0.121    0.000    0.121    0.000 {method 'hexdigest' of '_hashlib.HASH' objects}
        1    0.001    0.001    0.001    0.001 hashlib.py:55(<module>)
        1    0.001    0.001    0.796    0.796 improved.py:1(<module>)
        1    0.001    0.001    0.001    0.001 base64.py:3(<module>)

```

We only took 0.8 seconds (down from 12.7). `format` still takes time, but only one third of the time. I don't know what to do about it, so I'll leave it as it is, but you got your 10x speedup with two simple changes.


"
"Bool support Oracle SQL
It's a constant frustration of mine that Oracle PL/SQL supports the `bool` data-type, while Oracle SQL does not. It's a big pain in the proverbial when you want to process a PL/SQL boolean return-value back into your everyday SQL (example below).


Even the ask-Tom website is blasé about this misfit, reporting that you should code boolean columns as fixed-values `'Y'/'N' CHAR` columns, which is a such a bad cop-out answer on so many different levels that I don't know where to start criticising it. In fact, the only redeeming quality of this response is the fact that (as far as I've recently discovered), many other database-engines don't support the boolean data-type either.


Anyhow - the question...


I have a work-around for the following problem (albeit messy and verbose), so I'm asking this question out of curiosity rather than necessity. But one of the few things that surprises me any more is the ingenuity of clever programmers, so here's hoping that one of you can come up with a solution to the following.


In the following sample, the function `stock_pkg.is_in_stock()` (which is an inherent part of my application) returns a BOOL value, rendering the SQL invalid (remember, SQL doesn't support BOOL):



```
SELECT part_no, stock_pkg.is_in_stock(part_no) in_stock
FROM   parts_table

```

What I need is to find a way of using the above function-call to generate a valid string (varchar) output of the format:



```
PART_NO IN_STOCK
------- ------------
AA      YES
BB      NO
CC      NO

```

(You may substitute 'yes/no' for 'true/false', 'green/red', 'tory/labour' or even numeric 1/0 for all I care - just so long as the output falls into one of two distinct categories.)


Unfortunately, I don't have the privilege to rewrite the original function to return a different data-type. And besides, there are thousands of functions like this dotted around the larger application, making it impractical to rewrite them all.


So in this sense, the solution must be a 'generic' one (i.e. not specific to this function call). For example, it is not sufficient to rewrite the function as `stock_pkg.is_in_stock_chr()`, because that would mean having to re-write all the other similar functions in my application too.


I've already tried:



```
SELECT part_no,
       CASE WHEN stock_pkg.is_in_stock(part_no) THEN 'y' ELSE 'n' END in_stock
FROM   parts_table

```

and even my own wrapper function:



```
SELECT part_no,
       my_bool_to_str(stock_pkg.is_in_stock(part_no)) in_stock
FROM   parts_table

```

But even wrapping booleans inside other functional constructs doesn't seem to be allowed by Oracle SQL (at least not in Oracle 10g).


There's also the option of writing a sub-select inside the `in_stock` column, but that could get excessively complicated in extreme examples too, and would also be case-specific.


As I say, I hope there's an ingenious solution out there somewhere (or at least a very simple one which I happen to have overlooked).


Thanks for your time.


","You can write your own wrapper like this:



```
CREATE OR REPLACE FUNCTION my_bool_to_str(f varchar2) RETURN VARCHAR2 IS

  b varchar2(2);

BEGIN

  EXECUTE IMMEDIATE 'declare bl boolean; begin bl := ' || f ||
                    '; if bl then :1 := ''y''; else :1 := ''n''; end if; end;'
    using out b;

  return b;

END;

```

Then you can call it like this:



```
SELECT part_no,
       my_bool_to_str('stock_pkg.is_in_stock('|| part_no|| ')') in_stock
FROM   parts_table

```

The difference from your wrapper is that it gets a varchar as input and not a boolean which the SQL engine doesn't recognize


"
"word-wrap:break-word is not working with iText 2.1.7
I am using **Java** & **Itext2.1.7** for generating **pdf** and I have written the following **HTML**. But it is not breaking words when they are too long.



```
String td =""<td colspan=\""2\"" align=\""left\"" style=\""table-layout: fixed;word-wrap:break-word;white-space: normal;\""><b>""
                    + verylongwordverylongwordverylongword + "":</b></td>""

```

Can someone help me achieving this?


","**Itext 2.7.1** does not support **word-wrap : break-word** property. Following is the link for that.


<http://demo.itextsupport.com/xmlworker/itextdoc/flatsite.html>


So I have manually break the words according to the no of columns of the Table and by taking assumption of maximum letters around 60-65.Following is the code for that.



```
public String getBrokenWordsForPdfGeneration(String longString, int cutIndex){
       StringBuffer mainStringBuffer = new StringBuffer(longString);
       String returnString = """";
       while(mainStringBuffer != null && mainString.length()>0 &&                              
mainString.subString(0,mainString.length()>cutIndex?cutIndex:mainString.length()) != null){ 
          if(mainStrig.subString(0,mainString.length()>cutIndex:mainString.length()).trim().contains("" "")){
               returnString += mainStrig.subString(0,mainString.length()>cutIndex:mainString.length()).trim();
               mainStringBuffer.delete(0,mainString.length()>cutIndex:mainString.length());
          }else{
              returnString += mainStrig.subString(0,mainString.length()>cutIndex:mainString.length()).trim()+ "" "";
              mainStringBuffer.delete(0,mainString.length()>cutIndex:mainString.length());
          }
    }
    return returnString;
}

```

"
"How to return distinct values in a JSON\_ARRAYAGG
So I'm attempting to place the results of a distinct, single-column query into a JSON array so it can be used on my web server. I have it set up something like this:



```
SELECT JSON_OBJECT(
  'ArrayKey' VALUE JSON_ARRAYAGG( col ) 
) AS jsonResult
FROM(SELECT DISTINCT column_name AS col
       FROM tbl_name);

```

However, when this query returns results, the array it generates in JSON contains all values from my column and ignores the DISTINCT clause in the subquery somehow. Whenever I get rid of the JSON\_ARRAYAGG clause and output the results directly, the result are unique, but somehow the command is ignored when I add it back in. I've also attempted to place the DISTINCT clause inside the JSON\_ARRAYAGG as well, like so:



```
SELECT JSON_OBJECT(
  'ArrayKey' VALUE JSON_ARRAYAGG( DISTINCT col ) 
) AS jsonResult
FROM(SELECT DISTINCT column_name AS col
       FROM tbl_name);

```

to no avail. Does anyone know what's going wrong in my code that's causing the array to output all values instead of distinct ones?


","Interesting... Looks like a bug to me. The optimizer seems to push down too eagerly.


As workaround you can use the `NO_MERGE` hint on the subquery.



```
SELECT /*+NO_MERGE(x)*/
      json_object('ArrayKey'
                  VALUE json_arrayagg(column_name)) jsonresult
      FROM (SELECT DISTINCT
                   column_name
                   FROM tbl_name) x;

```

A CTE and a `MATERIALIZE` hint seem to work too.



```
WITH cte
AS
(
SELECT /*+MATERIALIZE*/ 
       DISTINCT
       column_name
       FROM tbl_name
)
SELECT json_object('ArrayKey'
                   VALUE json_arrayagg(column_name)) jsonresult
       FROM cte;

```

[db<>fiddle](https://dbfiddle.uk/?rdbms=oracle_18&fiddle=a7e554da90b21365c14b48cec13ce380)


"
"Does a NSData image weight in memory as much as the opened image?
I am doing an upload function to my server from my iOS app. I am right now architecting the best way to do this and thinking that my approach would run into trouble.


I've over 20 pictures that I would like to send in 1 request using `AFmultiparFormData` in `AFNetworking` with also text information that is relevant to that picture.


Now... I know how to create the `NSDictionary` with lots of information inside but never created a `NSDictionary` with 20 images in `NSData` format or attempted to upload many pictures at once. 


The issue is that I don't know whether the `NSData` images would take as much memory space as an image. Thus, I could only do 2/3 at the time (if they were full screen images which they are not, but its just an arbitrary point of measurement)


Given What I said and the seudo code below, Could I do that or would I ran out of memory? otherwise I would be left to do 1 by 1.



```
// for loop that does this 20 times
UIImage* image = [UIImage imageNamed:myExercise.exercisePictureThumb];
NSData * exercisePictureThumb = UIImagePNGRepresentation(image1);

// add to dictionary recursively

```

","You should not put all of your image data into a `NSDictionary`, as that's likely to cause memory problems.


Instead, create a dictionary or array of the URLs to your files, then make use of AFNetworking's [`multipartFormRequestWithMethod:path:parameters:constructingBodyWithBlock:`](http://cocoadocs.org/docsets/AFNetworking/1.2.1/Classes/AFHTTPClient.html#//api/name/multipartFormRequestWithMethod%3apath%3aparameters%3aconstructingBodyWithBlock%3a) and the [`AFMultipartFormData protocol`](http://cocoadocs.org/docsets/AFNetworking/1.2.1/Protocols/AFMultipartFormData.html), particularly `appendPartWithFileURL:name:fileName:mimeType:error:`. This way your image data for all of the images is not all in memory at one time, but is instead streamed from disk. Huge performance increase.


"
"Casting ""pointer to data"" to ""pointer to function""
As I know such casting from *pointer to data* to *pointer to function* is not allowed in C89, C99. What about C++?


","Short answer: An implementation is not required to support such a cast.


It may allow a `reinterpret_cast` to do it, subject to certain rules (essentially: converting to object pointer and back has to yield the original pointer).


Quoth the standard, [expr.reinterpret.cast]/8:



> 
> Converting a function pointer to an object pointer type or vice versa is conditionally-supported. The meaning of such a conversion is implementation-defined, except that if an implementation supports conversions in both directions, converting a prvalue of one type to the other type and back, possibly with different cv-qualification, shall yield the original pointer value.
> 
> 
> 


Non-embedded implementations usually allow it for purposes of dynamic linking (such as `dlsym()`). Strictly speaking, though, it is not required to work.


"
"How do I set up DNS for an apex domain (no www) pointing to a Heroku app?
I already added a custom domain to my Heroku app and it works with `www.domain.com`.


I need to know how to set up the domain without `www` to resolve to the app, too.


Here are my current DNS settings:



```
$TTL 86400
@   IN SOA ns1.first-ns.de. postmaster.robot.first-ns.de. (
    2013041500   ; serial
    14400        ; refresh
    1800         ; retry
    604800       ; expire
    86400 )      ; minimum

@                        IN NS      robotns3.second-ns.com.
@                        IN NS      robotns2.second-ns.de.
@                        IN NS      ns1.first-ns.de.

@                        IN A       88.198.38.XXX
localhost                IN A       127.0.0.1
mail                     IN A       88.198.38.XXX
ftp                      IN CNAME   www
imap                     IN CNAME   www
loopback                 IN CNAME   localhost
pop                      IN CNAME   www
relay                    IN CNAME   www
smtp                     IN CNAME   www
www                      IN CNAME   appname.herokuapp.com.
@                        IN MX 10   mail

```

What are the correct settings to use so that both `example.com` and `www.example.com` would point correctly to my Heroku app?


","(Note: root, base, apex domains are all the same thing. Using interchangeably for google-foo.)


Traditionally, to point your apex domain you'd use an A record pointing to your server's IP. This solution doesn't scale and isn't viable for a cloud platform like Heroku, where multiple and frequently changing backends are responsible for responding to requests.


For subdomains (like `www.example.com`) you can use CNAME records pointing to `your-app-name.herokuapp.com`. From there on, Heroku manages the dynamic A records behind `your-app-name.herokuapp.com` so that they're always up-to-date. Unfortunately, the DNS specification does not allow CNAME records on the zone apex (the base domain). (For example, MX records would break as the CNAME would be followed to its target first.)


Back to root domains, the simple and generic solution is to not use them at all. As a fallback measure, some DNS providers offer to setup an HTTP redirect for you. In that case, set it up so that `example.com` is an HTTP redirect to `www.example.com`.


Some DNS providers have come forward with custom solutions that allow CNAME-like behavior on the zone apex. To my knowledge, we have [DNSimple's ALIAS record](http://blog.dnsimple.com/zone-apex-naked-domain-alias-that-works/) and [DNS Made Easy's ANAME record](http://www.dnsmadeeasy.com/technology/aname-records/); both behave similarly.


Using those, you could setup your records as (using zonefile notation, even tho you'll probably do this on their web user interface):



```
@   IN ALIAS your-app-name.herokuapp.com.
www IN CNAME your-app-name.herokuapp.com.

```

Remember `@` here is a shorthand for the root domain (`example.com`). Also mind you that the trailing dots are important, both in zonefiles, and some web user interfaces.


### See also:


- [Doing DNS right with Heroku](https://web.archive.org/web/20140709085736/http://www.neilmiddleton.com/doing-dns-right-with-heroku/)
- [Avoiding Naked Domains and DNS A-records](https://devcenter.heroku.com/articles/avoiding-naked-domains-dns-arecords)


### Remarks:


- Amazon's Route 53 also has an ALIAS record type, but it's somewhat limited, in that it only works to point within AWS. At the moment I would not recommend using this for a Heroku setup.
- Some people confuse DNS providers with domain name registrars, as there's a bit of overlap with companies offering both. Mind you that to switch your DNS over to one of the aforementioned providers, you only need to update your nameserver records with your current domain registrar. You do not need to transfer your domain registration.


"
"Removing the Great Lakes from US county-level maps in R
I am using R to draw US map at county level. I downloaded the shapefile for US from [GADM](https://gadm.org/download_country_v3.html). The county-level shape file is ""gadm36\_USA\_2.shp"". I then used the code below to draw map:



```
library(sf)
library(tidyverse)

us2 <- st_read(""<Path>\\gadm36_USA_2.shp"")

mainland2 <- ggplot(data = us2) +
geom_sf(aes(fill = NAME_2), size = 0.4, color = ""black"") +
coord_sf(crs = st_crs(2163), 
         xlim = c(-2500000, 2500000), 
         ylim = c(-2300000, 730000)) + guides(fill = F)

```

The Great Lakes region (shown by red arrows) is plotted rather than left blank:
[![enter image description here](https://i.stack.imgur.com/6X08H.png)](https://i.stack.imgur.com/6X08H.png)


What I want is a figure like below, where the Great Lakes region is left blank:
[![enter image description here](https://i.stack.imgur.com/RAkZ8.jpg)](https://i.stack.imgur.com/RAkZ8.jpg)


How could I identify from the ""gadm36\_USA\_2.shp"" which rows correspond to the Great Lakes region so that I may delete them?


I understand there may be other ways to obtain shapefile than GADM. I believe GADM is an excellent source that provides bourndaries worldwide. I wish to take this opportunity to better acquaint myself with data downloaded from GADM.


Of course, other methods to obtain US county-level boundary data are welcome. I noted `USAboundaries` package also provide country, state, and county level coundaries, but I am having difficulties installing associated USAboundariesData package. Any idea to draw US counties in ways other than shapefile from GADM is welcome. Thanks.


","One way is to remove every feature that is tagged with `Lake` in the existing records (currently 13 features). First, you need to find the lakes name in the attribute table as below:



```
# retrieving the name of lakes and excluding them from the sf 

all.names = us2$NAME_2
patterns = c(""Lake"", ""lake"")

lakes.name <- unique(grep(paste(patterns, collapse=""|""), all.names, value=TRUE, ignore.case = TRUE))
#[1] ""Lake and Peninsula"" ""Lake""               ""Bear Lake""          ""Lake Michigan""      ""Lake Hurron""        ""Lake St. Clair""    
#[7] ""Lake Superior""      ""Lake of the Woods""  ""Red Lake""           ""Lake Ontario""       ""Lake Erie""          ""Salt Lake""         
#[13] ""Green Lake"" 

`%notin%` <- Negate(`%in%`)
us <- us2[us2$NAME_2 %notin% lakes.name, ]

```

Then you can map the remaining features:



```
mainland2 <- ggplot(data = us) +
  geom_sf(aes(fill = NAME_2), size = 0.4, color = ""black"") +
  coord_sf(crs = st_crs(2163), 
           xlim = c(-2500000, 2500000), 
           ylim = c(-2300000, 730000)) + guides(fill = F)
mainland2

```

[![enter image description here](https://i.stack.imgur.com/jxUOw.jpg)](https://i.stack.imgur.com/jxUOw.jpg)


Another way (much easier but less flexible) is to map county features by excluding `Water body` values from `ENGTYPE_2` as below:



```
us <- us2[(us2$ENGTYPE_2) != ""Water body"",]
mainland2 <- ggplot(data = us) +
  geom_sf(aes(fill = NAME_2), size = 0.4, color = ""black"") +
  coord_sf(crs = st_crs(2163), 
           xlim = c(-2500000, 2500000), 
           ylim = c(-2300000, 730000)) + guides(fill = F)
mainland2

```

[![enter image description here](https://i.stack.imgur.com/Vs0cb.jpg)](https://i.stack.imgur.com/Vs0cb.jpg)


"
"What does <<-EOC means?
I'd like to know what the minus (-) and the EOC in the command below means. I know some languages like Perl allows you to chose any combination of character (not bound to EOF) but is that the case here? And the minus is a complete mystery for me. Thanks in advance!



```
ftp -v -n $SERVER  >> $LOG_FILE <<-EOC
            user $USERNAME $PWD
            binary
            cd $DIR1
            mkdir $dir_lock
            get $FILE
            bye
EOC

```

","That's a here-document redirection.



```
command <<-word
here-document contents
word

```

The `word` used to delimit the here-document is arbitrary. It's common, but not necessary, to use an upper-case word.


The `-` in `<<-word` has the effect that tabs will be stripped from the beginning of each line in the contents of the here-document.



```
cat <<-SERVICE_ANNOUNCEMENT
    hello
    world
SERVICE_ANNOUNCEMENT

```

If the above here-document was written with literal tabs at the start of each line, it would result in the output



```
hello
world

```

rather than



```
    hello
    world

```

Tabs before the end delimiter are also stripped out with `<<-` (but not without the `-`):



```
cat <<-SERVICE_ANNOUNCEMENT
    hello
    world
    SERVICE_ANNOUNCEMENT

```

(same output)


"
"How can I use a twig variable as a form field name
I have a form template which has a repeated block for several similar form fields. It is used as follows:



```
{# form_template.html.twig #}
{{ include ('company-select.html.twig', {field: 'companiesInclude'})
{{ include ('company-select.html.twig', {field: 'companiesExclude'})
{{ include ('company-select.html.twig', {field: 'companiesLinked'})

```

Now I have a company select template, which looks as follows:



```
{# company-select.html.twig  (simplified, the actual template is much more complex) #}
<div class=""form form-field selector"" id=""{{ field }}"">
  {{ form_label(form.field) }}
  {{ form_widget(form.field) }}    

</div>

```

The way it is, the template fails, because 'field' is not a property of the FormView class.


How can I make the template interpolate the twig variable into the form\_xxx function call as an actual field name?


","A property of an object or an array element in twig can either be accessed via the dot notation (`.`) or via the square bracket one (`[]`).


If your property happens to be a twig variable, you will need to use the later form.


So in your case:


- `form_label(form[field])`
- `form_widget(form[field])`


Your code in *company-select.html.twig* ending up being:



```
<div class=""form form-field selector"" id=""{{ field }}"">
  {{ form_label(form[field]) }}
  {{ form_widget(form[field]) }}  
</div>

```

This is, in a really simplified example, testable [here](https://twigfiddle.com/f8xa9t).


*Note that, in a more complex flavour, you can also use the [`attribute` function](https://twig.symfony.com/doc/2.x/functions/attribute.html)*



> 
> `.`, `[]`: Gets an attribute of a variable.
> 
> 
> 


*Source: <https://twig.symfony.com/doc/3.x/templates.html#other-operators>*  

*Also worth reading on the same note: <https://twig.symfony.com/doc/3.x/templates.html#variables>*


"
"Access session across subdomains (Rails 4)
Hi i have a multitenant rails 4 application that has a simple sign in solution. However each user has a subdomain that the user gets redirected to after login. 


The problem is that as they arrive at the subdomain they are not logged in anymore due to the known problem that sessions are not shared across subdomains.


I have tried several different solution to this problem, however i do not get the session to persist across subdomains. I believe this might be due to my development environment?


I have tried all answers to this question:
[Share session (cookies) between subdomains in Rails?](https://stackoverflow.com/questions/10402777/share-session-cookies-between-subdomains-in-rails/10403338#10403338)


Nothing seems to work. Is there something I'm missing here? Is it the browser or rails 4 or....? How should i approach this problem?


**Edit:**
My sessions\_store initializer:



```
Imagesite::Application.config.session_store :cookie_store, key: '_imagesite_session', :domain => ""imagesite.dev""

```

I have also tried `"".imagesite.dev""` and `:all`.


I also tried the solution described by Evan at the other question linked above.


**Examples of subdomains:** ""ole.imagesite.dev"" or ""ole2.imagesite.dev"" just basic subdomain based on what the user has entered as his/her subdomain.


","I finally solved it!


I had to set the domain when i create the auth\_token cookie. like this:



```
cookies[:auth_token] = { value: user.auth_token, domain: "".lvh.me"" }

```

and like this to delete the cookie:



```
cookies.delete(:auth_token, :domain => '.lvh.me')

```

Complete example:



```
  def create
    user = User.find_by_username(params[:username])
    user ||= User.find_by_email(params[:username])
    if user && user.authenticate(params[:password])
      # session[:user_id] = user.id
        if params[:remember_me]
        cookies.permanent[:auth_token] = { value: user.auth_token, domain: "".lvh.me"" }
      else
        cookies[:auth_token] = { value: user.auth_token, domain: "".lvh.me"" }
      end
        redirect_to root_url(:subdomain => ""#{current_user.subdomain}""), notice: ""You are now loged in.""
    else
        flash.now.alert = ""Email or password is invalid""
        render ""new""
    end
  end

  def destroy
    #session[:user_id] = nil
    cookies.delete(:auth_token, :domain => '.lvh.me')
    redirect_to root_url(:subdomain => false), notice: ""Loged out""
  end

```

"
"What is the default text color for theme.appcompat.light?
The title says it all, looking for the default color value used in TextViews in theme.appcompat.light


tried looking for it in Android studio by hitting ctrl on theme.appcompat.light but it brought me down a rabbit hole that I couldn't find the end of.


","Looks like `theme.appcompat.light` goes all the way up to `Platform.AppCompat.Light`


By default I'm just going to assume you mean the primary color.


Here's what it looks like:



```
<style name=""Platform.AppCompat.Light"" parent=""android:Theme.Light"">
    <item name=""android:windowNoTitle"">true</item>

    ...

    <!-- Text colors -->
    <item name=""android:textColorPrimary"">@color/abc_primary_text_material_light</item>
    <item name=""android:textColorPrimaryInverse"">@color/abc_primary_text_material_dark</item>
    <item name=""android:textColorSecondary"">@color/abc_secondary_text_material_light</item>
    <item name=""android:textColorSecondaryInverse"">@color/abc_secondary_text_material_dark</item>
    <item name=""android:textColorTertiary"">@color/abc_secondary_text_material_light</item>
    <item name=""android:textColorTertiaryInverse"">@color/abc_secondary_text_material_dark</item>
    <item name=""android:textColorPrimaryDisableOnly"">@color/abc_primary_text_disable_only_material_light</item>
    <item name=""android:textColorPrimaryInverseDisableOnly"">@color/abc_primary_text_disable_only_material_dark</item>
    <item name=""android:textColorHint"">@color/hint_foreground_material_light</item>
    <item name=""android:textColorHintInverse"">@color/hint_foreground_material_dark</item>
    <item name=""android:textColorHighlight"">@color/highlighted_text_material_light</item>
    <item name=""android:textColorLink"">?attr/colorAccent</item>

    ...

</style>

```

Inside you'll see `abc_primary_text_material_light`:



```
<selector xmlns:android=""http://schemas.android.com/apk/res/android"">
    <item android:state_enabled=""false"" android:color=""@color/primary_text_disabled_material_light""/>
    <item android:color=""@color/primary_text_default_material_light""/>
</selector>

```

Which last but not least defines it's color as:



```
<color name=""primary_text_default_material_light"">#de000000</color>

```

"
"Eclipse IDE support for Java 14 early-access builds
I couldn't find any resource listing any plugin for support for JDK 14 EA builds. Is there a more recent build of Eclipse Java IDE with support for JDK 14 or at least a plugin to support it from the marketplace?


","**It's too early for that**. Remember that Eclipse must not only understand Java 14, but also has its own compiler, which in Java 14 must be able to compile [Records](https://openjdk.java.net/jeps/359), for example.


- The [**`BETA_JAVA14` branch**](https://git.eclipse.org/c/jdt/eclipse.jdt.core.git/log/?h=BETA_JAVA14) has been created.
- See [**Eclipse bug 549808 tree**](https://bugs.eclipse.org/bugs/showdependencytree.cgi?id=549808&hide_resolved=1) for **what needs to be implemented** for Java 14 support.
- Add yourself to the CC list of [**Eclipse bug 551486**](https://bugs.eclipse.org/bugs/show_bug.cgi?id=551486) to receive e-mails about the progress **when builds that can be installed are available** and when there will be an [Eclipse Marketplace](https://marketplace.eclipse.org/) entry.


"
"GCC optimizations based on integer overflow
Recently I had a discussion about someone who wanted to check for signed int overflow like this `if (A + B < 2 * max(A, B))`. Lets ignore for a second that the logic itself is wrong and discuss signed integer overflow in context of C/C++. (Which I believe fully inherits this part of standard from C). 


What kinds of check that need signed integer overflow will be optimized away by current-ish GCC and which won't?


**Since the original text wasn't all that well formulated and apparently controversial I decided to change the question somewhat, but leave the original text below.**


*All examples used below were tested gcc version `4.7.2 (Debian 4.7.2-5)` and compiled using `-O3`*


Namely, it is undefined and GCC infamously uses this to perform some branch simplifications. The first example of this that comes to mind is 




```
int i = 1;
while (i > 0){
    i *= 2;
}

```

which produces an infinite loop. Another case where this kind of optimalization kicks in is 




```
if (A + 2 < A){
    /* Handle potential overflow */
}

```

where, assuming `A` is signed integral type, the overflow branch gets completely removed. 


Even more interestingly, some cases of easily *provable* integer overflow, are left untouched, such as




```
if (INT_MAX + 1 < 0){
    /* You wouldn't write this explicitly, but after static analysis the program
       could be shown to contain something like this. */
}

```

which triggers the branch that you would expect with two's complement representation. Similarly, this code leaves the conditional branches intact




```
int C = abs(A);
if (A + C < 0){
    /* For this to be hit, overflow or underflow had to happen. */
}

```

Now for the question, is there a pattern that looks roughly like `if (A + B < C)` or `if (A + B < c)`, that will be optimized away? When I was googling around before writing this, it seemed like the last snippet should be optimized away, but I cannot reproduce this kind of error in an overflow check that doesn't operate with constant explicitly.


","Many compilers will replace expressions involving signed integers or pointers with ""false"", like



```
a + 1 < a // signed integer a
p + 1 < p // Pointer p

```

when the expression can only be true in the case of undefined behaviour. On the other hand, that allows 



```
for (char* q = p; q < p + 2; ++q) ...

```

to be inlined, substituting q = p and q = p + 1, without any check, so that's a good thing. 



```
if (A + abs (A) < 0)

```

is probably too complicated for many compilers. Note that for unsigned integers there is no undefined behaviour. As a consequence, loops using unsigned 32 bit integers with 64 bit pointers tend to be slower than necessary because the wraparound behaviour must be taken into account. For unsigned 32 bit integer and 64 bit pointers, it is possible that 



```
&p [i] > &p [i+1]

```

without undefined behaviour (not with 64 bit integers or 32 bit pointers). 


"
"Regex to detect everything not between {} and then search from whats matched
I am very new to both stackoverflow and Regexes so please forgive mistakes.


I have been searching thoroughly for a Regex to match all text that is not between curly brackets {} and from that text find certain words. For example from the following string:


`$content = 'Hello world, { this } is a string with { curly brackets } and this is for testing'`


I would like the search for word `this` to return only the second occurrence of `this` because its in the area which is not inside curly brackets.
Even if I can get a Regex to match the substrings outside the curly brackets, things get simplified for me. I found this Regex `/(}([^}]*){)/` but it cannot select the parts `Hello world,` and `and this is for testing` because these are not inside `}{` and it only selects `is a string with` part.


Also I would like to ask if it is possible to combine two Regex for a single purpose like mine. For example the first Regex finds strings outside `{}` and second finds specific words that are searched for.


I want to use this Regex in php and for now I am using a function which is more like a hack. The purpose is to find specific words that are not in `{}` ,replace them reliably and write to text files.


Thanks in advance for your help.


","**(\*SKIP)(\*F)**


You're in luck, as php's PCRE regex engine has a syntax that is wonderful for this kind of task. This tidy regex work like a charm (see [demo](http://regex101.com/r/eK9gD7)):



```
{[^{}]*}(*SKIP)(*F)|\bthis\b

```

**Okay, but how does it work?**


Glad you asked. The left side of the alternation `|` matches complete `{braces}`then deliberately fails, after which the engine skips to the next position in the string. The right side matches the `this` words you want, and we know they're the right ones because they weren't matched by the expression on the left...


**How to use it in PHP**


Just the usual, something like:



```
$regex = ""~{[^{}]*}(*SKIP)(*F)|\bthis\b~"";
$count = preg_match_all($regex,$string,$matches);

```

You'll want to have a look at `$matches[0]`


**Further reading about this and similar exclusion techniques**


This situation is very similar to this question about [""regex-matching a pattern unless...""](https://stackoverflow.com/q/23589174), which, if you're interested and enjoyed `(*SKIP)` power, you might like to read to fully understand the technique and how it can be extended.


"
"How to improve AWS RDS Performance - Heavy read write updates
After searching and reading as many posts, comments and discussions I could find I did not find one specific to my issue.


I have multiple AWS EC2 deployments with a single RDS all in the same Delivery zone us-west-2(c)


I am testing the load of the instances at a fraction of what I will be expecting very soon. The issue that concerns me is the performance while pushing updates. We will be frequently acquiring updates of 1,000 records at a time and will take and compare our data and update our data as appropriate. Consequently one read and one write per entry. It is not uncommon to have 100,000 updates come to us in an hour.


Currently I have a MySQL Database on an AWS t2.medium class RDS running 5 update processes at 22% CPU and less than 1GB Memory.


Even with these low numbers the read time to search the database of 106.3K records takes 2 to 3 full seconds and the write time is another 2 seconds.


I need some thoughts on how to improve these read/write times.


Additional info: I have a replica instance running as well. The CMS driven sites (100 and growing daily) connect to the replica instance for their content.


Thanks!


","I'm going to turn my comments above into an answer.


T2 instances only get a [fraction of a CPU](http://docs.aws.amazon.com/AWSEC2/latest/UserGuide/t2-instances.html) - 10% for a t2.micro, 40% for a t2.medium. You get CPU credits that build up, but once you use them you get throttled for CPU. You also get lower network and IO performance as you have to share physical machine resources.


What I suspect is happening is your testing has run out of CPU credits so you're being throttled. You can [monitor CPU credits](http://docs.aws.amazon.com/AWSEC2/latest/UserGuide/t2-instances.html#t2-instances-monitoring-cpu-credits) in CloudWatch, this will tell you whether or not my guess is correct.


Whether I'm right or not, t2 instances aren't suitable for systems that get work coming in constantly. A C4 or other general instance would probably be your best bet. You can monitor your database memory and CPU usage in CloudWatch to help work out what database instance size you need.


"
"Object Array Declaration in Groovy
How come I cannot declare an array of `People` in Groovy as shown.
Maybe I'm lacking the deeper understanding of classes



```
class People {
    Integer id
}

class Job {
    def func() {
        People[] p = new People[10]
    }
}

```

I get an error of People[] cannot be applied to app.People[]


","The code sample you have shown does not reproduce the error you mentioned in the question above. It's broken actually and does not compile - method `func()` is missing its body. If you correct the code to e.g.



```
class People {
    Integer id
}

class Job {
    def func() {
        People[] p = new People[10]
        assert p.size() == 10
        println p
    }
}

new Job().func()​

```

you will see it produces the expected result - check it out in the Groovy web console [here](https://groovyconsole.appspot.com/edit/5154073565200384). When you run it you will see following output to the console:



```
[null, null, null, null, null, null, null, null, null, null]

```

### The difference between Groovy and Java


When it comes to array initialization there is one significant difference between Groovy and Java. In Java you can initialize an array of `People[]` like this:



```
People[] p = new People[] { new People(), new People(), /* ... */ new People() };

```

It wont work in Groovy, because Groovy reserves `{}` for closures. In Groovy you can initialize such array as:



```
People[] p = [new People(), new People(), new People()] as People[]

```

"
"PHP Array Warning: Creating default object from empty value
I have taken values from a MySQL query into an array so that I can sort them. It all works well except I am getting lots of errors 



```
 Warning: Creating default object from empty value 

```

This is so that I can sort the array based on post\_id and quality



```
$count = 0;
$searcharticles = new stdClass();
$postid = 0;
while ($row = mysqli_fetch_object($result1) ) {

    $postid     = $row->ID;
    $title      = $row->post_title;
    $titlescore = $row->titlescore;
    $quality    = $row->quality;

    $searcharticles->$postid->post_id =$postid;
    $searcharticles->$postid->title =$title;
    $searcharticles->$postid->titlescore = $titlescore;
    $searcharticles->$postid->quality =$quality;

$count ++;
} 

```

Some posts on Stackexchange indicated that I should add in 
 $searcharticles = new stdClass();


But this has not solved the problem


I am looking to remove the errors and not just suppress the warnings.


Any ideas?


","You get the warning



> 
> Warning: Creating default object from empty value 
> 
> 
> 


because you do



```
$searcharticles->$postid->post_id = $postid;

```

Your `$searcharticles` is a StdClass. You assign the property `$postId` and then immediately chain off another property `post_id`, which forces PHP to create a new default object.


To avoid it, put this before the chained method calls: 



```
$searcharticles->$postid = new StdClass;
$searcharticles->$postid->post_id =$postid;
$searcharticles->$postid->title =$title;
$searcharticles->$postid->titlescore = $titlescore;
$searcharticles->$postid->quality =$quality;

```

"
"Command to drop all columns in a database that have prefix test\_ to run
How I can run a command in phpMyAdmin which will drop all columns in a database that have the prefix `test_`.


","To drop a column from a table, use the syntax:



```
alter table <tablename> drop column <columnname>

```

To find all the columns in a table in a database that start with test\_, do the following:



```
select column_name
from INFORMATION_SCHEMA.columns
where table_name = <table_name> and
      table_schema = <schema_name> and
      left(column_name, 5) = 'test_'  -- not using ""like"" because '_' is a wildcard char

```

If you were doing this manually, I would recommend running the following query and then pasting the results in to a mysql query interface:



```
select concat('alter table ', table_name, ' drop column ', column_name)
from INFORMATION_SCHEMA.columns
where table_name = <table_name> and
      schema_name = <schema_name> and
      left(column_name, 5) = 'test_'

```

You can do something similar in code, by running the query, returning the results and then running each row as a query.


"
"How can I use Github packages Docker registry in Kubernetes dockerconfigjson?
How can I pull `docker.pkg.github.com` Docker images from within Kubernetes cluster?


Currently, the Github Docker registry requires authentication even for packages from public Github repositories.


","1. Create new Github Personal Access Token with `read:packages` scope at <https://github.com/settings/tokens/new>.
2. Base-64 encode `<your-github-username>:<TOKEN>`, ie.:



```
$ echo -n VojtechVitek:4eee0faaab222ab333aa444aeee0eee7ccc555b7 | base64
<AUTH>

```

*Note: Make sure not to encode a newline character at the end of the string.*
3. Create kubernetes.io/dockerconfigjson secret


A) Create secret manually:



```
$ echo '{""auths"":{""docker.pkg.github.com"":{""auth"":""<AUTH>""}}}' | kubectl create secret generic dockerconfigjson-github-com --type=kubernetes.io/dockerconfigjson --from-file=.dockerconfigjson=/dev/stdin

```

B) Or, create .yml file that can be used in `kubectl apply -f`:



```
kind: Secret
type: kubernetes.io/dockerconfigjson
apiVersion: v1
metadata:
  name: dockerconfigjson-github-com
stringData:
  .dockerconfigjson: {""auths"":{""docker.pkg.github.com"":{""auth"":""<AUTH>""}}}

```

*Note for GitOps: I strongly recommend not to store the above file in plain-text in your git repository. Hydrate the value in your CD pipeline or encrypt/seal the file with tools like <https://github.com/mozilla/sops> or <https://github.com/bitnami-labs/sealed-secrets>.*
4. Now, you can reference the above secret from your pod's spec definition via `imagePullSecrets` field:



```
spec:
  containers:
  - name: your-container-name
    image: docker.pkg.github.com/<ORG>/<REPO>/<PKG>:<TAG>
  imagePullSecrets:
  - name: dockerconfigjson-github-com

```


"
"How to upload captured photo (from browser - canvas) to node server?
I want to upload the captured photo in canvas from browser in javascript to nodejs server. I am not getting a proper method. Please help. Thank you so much.



```
(function(){
  var video=document.getElementById('video'), 
  canvas = document.getElementById('canvas'),
  vendorUrl = window.URL || window.webkitURL;
  context = canvas.getContext('2d');
  //photo = document.getElementById('photo');
  navigator.getMedia = navigator.getUserMedia ||
                       navigator.webkitGetUserMedia ||
                       navigator.mozGetUserMedia ||
                       navigator.msGetUserMedia;
  navigator.getMedia({
       video: true,
       audio: false
  }, function(stream){
      video.srcObject=stream;
      video.play();
  }, function(error){

  });
  document.getElementById('capture').addEventListener('click',function(){
        context.drawImage(video,0,0,400,300);
    //    photo.setAttribute('src',canvas.toDataURL('image/png')); 
       download();
   });
})();
function download() {
var download = document.getElementById(""capture"");
var image = document.getElementById(""canvas"").toDataURL(""image/png"")
    .replace(""image/png"", ""image/octet-stream"");
download.setAttribute(""href"", image);
//download.setAttribute(""download"",""archive.png"");
}

```

This code works fine to download a captured image but I am not getting a way to upload same image to node server.


","Here is the answer code.


**At client side**



```
function covertImage() {
var image = document.getElementById(""canvas"").toDataURL(""image/png"");
sendMessage(image);
}

//Sending image data to server
function sendMessage(image){
   var msg = JSON.stringify(image);
   var xhr = new XMLHttpRequest();
   xhr.open('POST', true);
   xhr.send(msg);
   alert('file is saved');
}

```

**At Server Side**



```
http.createServer(function (request, response) {
    var post='';
    if (request.method == 'POST') {
        var body = '';
        request.on('data', function (data) {
            body += data;
        });
        request.on('end', function () {
//-------------parsing data from json to string-------------------------
            post = JSON.parse(body);
            var data = post.replace(/^data:image\/\w+;base64,/, """");
            var buf = Buffer.from(data, 'base64');
            writeFileToSystem(buf);
        });
    }

//----------saving image to server side folder------------------
    function writeFileToSystem(buf)
    {
        fs.writeFile(""images/image.png"", buf, function(err) {
            console.log(""The file was saved!"");
        });
    }

}

```

"
"Cloning a github repo into an existing project?
I have a project in Eclipse that I want to put on github, so I went to github and created the repos for them, but I want to clone them right into the folder where the files are stored in my eclipse workspace. How can I do this? 


EDIT: When I try it, the github app says it can't clone because the folder isn't empty.


EDIT 2: Will this work? Changing the name in eclipse to rename the project folder, then cloning the repo to the name I want, in the workspace, then renaming the eclipse project so they merge and I can commit the new files.


","GitHub has a [guide](https://help.github.com/articles/adding-an-existing-project-to-github-using-the-command-line/) explaining how to put an existing project on GitHub.


You should not clone the repository, but add the GitHub repository as a remote to a local repository you create yourself.


1. Go to your project folder and initialize a local repository with `git init`
2. Add and commit all your project files to the local repository. (e.g. `git add .` and `git commit -m ""message""`)
3. Add the GitHub repository as a remote. `git remote add origin *github repository URL*` (Verify with `git remote -v`)
4. Push your project to GitHub with `git push origin master`.


---


If you already have committed files to the GitHub repository, it is still possible.


1. Initialize your local repository.
2. Add GitHub as the remote.
3. Pull from GitHub.
4. Add and commit your project.
5. Push you project to GitHub


"
"How to update entire column with values in list using Sqlite3
I have a dataset stored locally in an sqlite3 database. I extracted a column, performed some operations and now want to replace ALL of the values in the database column. How can I do this?


The length of the column and the list are guaranteed to be the same length. I simply want to update the table with the new values. Is there an easy way to do this all at once?


Using python 2.7


Edited to add:


myList is a pandas series backed by a numpy array of dtype 'object'. The table column, myCol is text formatted.



```
In [1]: curr.execute('UPDATE test SET myCol= ?', myList)

---------------------------------------------------------------------------
ProgrammingError                          Traceback (most recent call last)
f:\python\<ipython-input-52-ea41c426502a> in <module>()
----> 1 curr.execute('UPDATE test SET myCol = ?', myList)

ProgrammingError: Incorrect number of bindings supplied. The current statement uses 1, and there are 401125 supplied.

```

","Use the [`.executemany()` method](http://docs.python.org/2/library/sqlite3.html#sqlite3.Cursor.executemany) instead:



```
curr.executemany('UPDATE test SET myCol= ?', myList)

```

However, `myList` *must* be a sequence of tuples here. If it is not, a generator expression will do to create these tuples:



```
curr.executemany('UPDATE test SET myCol= ?', ((val,) for val in myList))

```

Demonstration (with insertions):



```
>>> import sqlite3
>>> conn=sqlite3.connect(':memory:')
>>> conn.execute('CREATE TABLE test (myCol)')
<sqlite3.Cursor object at 0x10542f1f0>
>>> conn.commit()
>>> myList = ('foo', 'bar', 'spam')
>>> conn.executemany('INSERT into test values (?)', ((val,) for val in myList))
<sqlite3.Cursor object at 0x10542f180>
>>> list(conn.execute('select * from test'))
[(u'foo',), (u'bar',), (u'spam',)]

```

Note that for an `UPDATE` statement, you *have* to have a `WHERE` statement to identify *what row* to update. An `UPDATE` without a `WHERE` filter will update *all* rows.


Make sure your data has a row identifier to go with it; in this case it may be easier to use named parameters instead:



```
my_data = ({id=1, value='foo'}, {id=2, value='bar'})
cursor.executemany('UPDATE test SET myCol=:value WHERE rowId=:id', my_data)

```

"
"Dynamically assign the getter for a dependent property in MATLAB
In Matlab, I can define a class as such:



```
classdef klass < handle
    properties(Dependent)
        prop
    end
end

```

Matlab is perfectly happy instantiating an object of this class, even without defining a getter for `prop`. It only fails when I try to access it (understandably). I'd like to set the `GetMethod` dynamically based upon the property's name.


Unfortunately, even when the property is Dependent, the `meta.property` field for `GetMethod` is still read-only. And while inheriting from `dynamicprops` could allow *adding* a property and programmatically setting its GetMethod in every instance, I don't believe it could be used to change an existing property. I may have to go this route, but as `prop` *must* exist for every object I'd prefer to simply set the getter on a class-by-class basis. Is such a thing possible?


An alternative solution could be through some sort of catch-all method. In other languages, this could be accomplished through a Ruby-like [`method_missing`](http://rubylearning.com/satishtalim/ruby_method_missing.html) or a PHP-like [`__get()`](http://uk3.php.net/manual/en/language.oop5.overloading.php#object.get). But as far as I know there's no (documented or otherwise) analog in Matlab.




---


(My use case: this class gets inherited by many user-defined subclasses, and all their dependent properties are accessed in a similar way, only changing based on the property name. Instead of asking users to write `get.*` methods wrapping a call to the common code for each and every one of their dependent properties, I'd like to set them all dynamically with anonymous function pointers containing the necessary metadata).


","Here is my proposal: create a method in the superclass called `add_dyn_prop`. This method is to be called in the subclasses instead of creating a [dependent property](http://www.mathworks.com/help/matlab/matlab_oop/property-access-methods.html#bsxanmy) the usual way.


The idea is that the superclass inherit from [`dynamicprops`](http://www.mathworks.com/help/matlab/ref/dynamicprops.html) and use [`addprop`](http://www.mathworks.com/help/matlab/ref/dynamicprops.addprop.html) to [add](http://www.mathworks.com/help/matlab/matlab_oop/dynamic-properties--adding-properties-to-an-instance.html) a new property, and set its accessor methods manually based on its name.



```
classdef klass < dynamicprops
    methods (Access = protected)
        function add_dyn_prop(obj, prop, init_val, isReadOnly)
            % input arguments
            narginchk(2,4);
            if nargin < 3, init_val = []; end
            if nargin < 4, isReadOnly = true; end

            % create dynamic property
            p = addprop(obj, prop);

            % set initial value if present
            obj.(prop) = init_val;

            % define property accessor methods
            % NOTE: this has to be a simple function_handle (@fun), not
            % an anonymous function (@()..) to avoid infinite recursion
            p.GetMethod = @get_method;
            p.SetMethod = @set_method;

            % nested getter/setter functions with closure
            function set_method(obj, val)
                if isReadOnly
                    ME = MException('MATLAB:class:SetProhibited', sprintf(...
                      'You cannot set the read-only property ''%s'' of %s', ...
                      prop, class(obj)));
                    throwAsCaller(ME);
                end
                obj.(prop) = val;
            end
            function val = get_method(obj)
                val = obj.(prop);
            end
        end
    end
end

```

now in the subclass, instead of defining a dependent property the usual way, we use this new inherited function in the constructor to define a dynamic property:



```
classdef subklass < klass
    %properties (Dependent, SetAccess = private)
    %    name
    %end
    %methods
    %    function val = get.name(obj)
    %        val = 'Amro';
    %    end
    %end

    methods
        function obj = subklass()
            % call superclass constructor
            obj = obj@klass();

            % define new properties
            add_dyn_prop(obj, 'name', 'Amro');
            add_dyn_prop(obj, 'age', [], false)
        end            
    end
end

```

The output:



```
>> o = subklass
o = 
  subklass with properties:

     age: []
    name: 'Amro'
>> o.age = 10
o = 
  subklass with properties:

     age: 10
    name: 'Amro'
>> o.name = 'xxx'
You cannot set the read-only property 'name' of subklass. 

```

Of course now you can customize the getter method based on the property name as you initially intended.




---


# EDIT:


Based on the comments, please find below a slight variation of the same technique discussed above.


The idea is to require the subclass to create a property (defined as abstract in the superclass) containing the names of the desired dynamic properties to be created. The constructor of the superclass would then create the specified dynamic properties, setting their accessor methods to generic functions (which could customize their behavior based on the property name as you requested). I am reusing the same `add_dyn_prop` function I mentioned before.


In the subclass, we are simply required to implement the inherited abstract `dynamic_props` property, initialized with a list of names (or `{}` if you dont want to create any dynamic property). For example we write:



```
classdef subklass < klass
    properties (Access = protected)
        dynamic_props = {'name', 'age'}
    end

    methods
        function obj = subklass()
            obj = obj@klass();
        end
    end
end

```

The superclass is similar to what we had before before, only now is it its responsibility to call the `add_dyn_prop` in its constructor for each of the property names:



```
classdef klass < dynamicprops        % ConstructOnLoad
    properties (Abstract, Access = protected)
        dynamic_props
    end
    methods
        function obj = klass()
            assert(iscellstr(obj.dynamic_props), ...
                '""dynamic_props"" must be a cell array of strings.');
            for i=1:numel(obj.dynamic_props)
                obj.add_dyn_prop(obj.dynamic_props{i}, [], false);
            end
        end
    end

    methods (Access = private)
        function add_dyn_prop(obj, prop, init_val, isReadOnly)
            % input arguments
            narginchk(2,4);
            if nargin < 3, init_val = []; end
            if nargin < 4, isReadOnly = true; end

            % create dynamic property
            p = addprop(obj, prop);
            %p.Transient = true;

            % set initial value if present
            obj.(prop) = init_val;

            % define property accessor methods
            p.GetMethod = @get_method;
            p.SetMethod = @set_method;

            % nested getter/setter functions with closure
            function set_method(obj,val)
                if isReadOnly
                    ME = MException('MATLAB:class:SetProhibited', sprintf(...
                      'You cannot set the read-only property ''%s'' of %s', ...
                      prop, class(obj)));
                    throwAsCaller(ME);
                end
                obj.(prop) = val;
            end
            function val = get_method(obj)
                val = obj.(prop);
            end
        end
    end
end

```

Note: I did not use [`ConstructOnLoad`](http://www.mathworks.com/help/matlab/matlab_oop/class-attributes.html?searchHighlight=ConstructOnLoad) class attribute or [`Transient`](http://www.mathworks.com/help/matlab/matlab_oop/property-attributes.html?searchHighlight=Transient) property attribute, as I am still not sure how they would affect loading the object from a saved MAT-file in regards to dynamic properties.



```
>> o = subklass
o = 
  subklass with properties:

     age: []
    name: []

>> o.name = 'Amro'; o.age = 99
o = 
  subklass with properties:

     age: 99
    name: 'Amro'

```

"
"call a partial view using @url.action click using jquery
call a partial view on @url.action. i am displaying the records using url.action and want to load the partial view when user click on the records.


here is my code on which i want to call the partial view when user click on it.



```
<td>
<a href=""@Url.Action(""Details"", new { id=item.TeamId})"">
 @Html.DisplayFor(modelItem => item.TeamName)
</a>
</td>

```

here is my Div in which i am placing the Partial view



```
<div id=""detailsPlace"" class=""dialog_content3"" style=""display:none""></div>
         @Html.Partial(""_TeamDetails"")
    </div>

```

here is my partial view which i want to render



```
@model light.ViewModels.ViewDetailTeam
@{
    var item = Model.Team;
}

    <div class=""dialogModal_header"">@Html.Label(item.TeamName)</div>
    <div class=""dialogModal_content"">
        <div class=""main-content"">
            <div class=""navi-but"">
                    @Html.Label(item.TeamName)
                </div>
                    @Html.Label(item.Description)
                </div>
            </div>
        </div>

```

and here is my controller



```
public ActionResult Details(int id)
        {

            lightCoreModel.User loggedInUser = new lightCoreModel.User();


                ViewDetailTeam viewDetailTeam = new ViewDetailTeam();
                ViewData[""DetailModel""] = viewDetailTeam;
                viewDetailTeam.Retrieve(id);
                return PartialView(""_TeamDetails"",viewDetailTeam);

        }
now i am facing this problem with pop up its showing me the following screen.

```

![enter image description here](https://i.stack.imgur.com/BRnvZ.jpg)


","You would need Ajax to do this. First, add a script block in your view with this code: 



```
<script type=""text/javascript"">
    $(function () {
        $('.details').click(function () {
            var $buttonClicked = $(this);
            var id = $buttonClicked.attr('data-id');

            $.ajax({
                url: '@Url.Action(""Details"")',
                type: 'GET',
                data: { id: id },
                success: function (partialView) {
                    $('#detailsPlace').html(partialView);
                    $('#detailsPlace').show();
                }
            });
        });
    });
</script>

```

Then change your anchor tag to this:



```
<a href=""#"" class=""details"" data-id=""@item.TeamId"">Details</a>

```

The ajax call will be fired whenever an element with the class of details is clicked. Once clicked, the Id that is stored in the data-id attribute will be passed along to the controller. When your controller passes the partial view back, the partial view will be loaded in the success function of the ajax call, and the detailsPlace will be shown, since it's display is set to none.


"
"R Sort strings according to substring
I have a set of file names like:



```
filelist <- c(""filea-10.txt"", ""fileb-2.txt"", ""filec-1.txt"", ""filed-5.txt"", ""filef-4.txt"")

```

and I would like to filter them according to the number after ""-"".


In python, for instance, I can use the `key`parameter of the sorting function:



```
filelist <- [""filea-10.txt"", ""fileb-2.txt"", ""filec-1.txt"", ""filed-5.txt"", ""filef-4.txt""]
sorted(filelist, key=lambda(x): int(x.split(""-"")[1].split(""."")[0]))

> [""filec-1.txt"", ""fileb-2.txt"", ""filef-4.txt"", ""filed-5.txt"", ""filea-10.txt""]

```

In R, I am playing with `strsplit` and `lapply` with no luck so far.


Which is the way to do it in R?


**Edit**:
File names can be many things and may include more numbers. The only fixed pattern is that the number I want to sort by is after the ""-"". Another (real) example:



```
c <- (""boards10017-51.mp4"",  ""boards10065-66.mp4"",  ""boards10071-81.mp4"",
      ""boards10185-91.mp4"", ""boards10212-63.mp4"",  ""boards1025-51.mp4"",   
      ""boards1026-71.mp4"",   ""boards10309-89.mp4"", ""boards10310-68.mp4"",  
      ""boards10384-50.mp4"",  ""boards10398-77.mp4"",  ""boards10419-119.mp4"", 
      ""boards10421-85.mp4"",  ""boards10444-87.mp4"",  ""boards10451-60.mp4"",  
      ""boards10461-81.mp4"",  ""boards10463-52.mp4"",  ""boards10538-83.mp4"",  
      ""boards10575-62.mp4"",  ""boards10577-249.mp4"")""

```

","I'm not sure of the actual complexity of your list of file names, but something like the following might be sufficient:



```
filelist[order(as.numeric(gsub(""[^0-9]+"", """", filelist)))]
# [1] ""filec-1.txt""  ""fileb-2.txt""  ""filef-4.txt""  ""filed-5.txt""  ""filea-10.txt""

```



---


Considering your edit, you may want to change the `gsub` to something like:



```
gsub("".*-|\\..*"", """", filelist)

```

Again, without a few more text cases, it's hard to say whether this is sufficient for your needs.




---


Example:



```
 x <- c(""boards10017-51.mp4"", ""boards10065-66.mp4"", ""boards10071-81.mp4"", 
     ""boards10185-91.mp4"", ""boards10212-63.mp4"", ""boards1025-51.mp4"",     
     ""boards1026-71.mp4"", ""boards10309-89.mp4"", ""boards10310-68.mp4"",     
     ""boards10384-50.mp4"", ""boards10398-77.mp4"", ""boards10419-119.mp4"",   
     ""boards10421-85.mp4"", ""boards10444-87.mp4"", ""boards10451-60.mp4"",    
     ""boards10461-81.mp4"", ""boards10463-52.mp4"", ""boards10538-83.mp4"",    
     ""boards10575-62.mp4"", ""boards10577-249.mp4"")  

x[order(as.numeric(gsub("".*-|\\..*"", """", x)))]
##  [1] ""boards10384-50.mp4""  ""boards10017-51.mp4""  ""boards1025-51.mp4""  
##  [4] ""boards10463-52.mp4""  ""boards10451-60.mp4""  ""boards10575-62.mp4"" 
##  [7] ""boards10212-63.mp4""  ""boards10065-66.mp4""  ""boards10310-68.mp4"" 
## [10] ""boards1026-71.mp4""   ""boards10398-77.mp4""  ""boards10071-81.mp4"" 
## [13] ""boards10461-81.mp4""  ""boards10538-83.mp4""  ""boards10421-85.mp4"" 
## [16] ""boards10444-87.mp4""  ""boards10309-89.mp4""  ""boards10185-91.mp4"" 
## [19] ""boards10419-119.mp4"" ""boards10577-249.mp4"" 

```

"
"byte-wise operations on a xmm register (AMD64)
How can I access only a specific part (byte) of a `128bit xmm` register? I have to loop (bytewise) over the whole space, compare byte after byte and copy it on well defined conditions\*.


Therefore I have to compare every byte of this 16-byte register with `\0` and (maybe) overwrite it.



```
xor %ecx, %ecx #var i

copy:
    cmp $0, <i-th byte of xmm1>
    je end
    mov <byte i of xmm1>, <byte i of xmm2>
    inc %ecx
    jmp copy
end:
# ...

```

How to access the i-th byte of xmm registers?


My solution has to use as less operations as possible (only the number of ops count, not the number of their clock cycles!), so I will not use sub-functions like strlen etc.


\*) I have to move all bytes, unit the the first `\0` occurrence within b, from register a to register b.


","Technically you can access bytes using `PEXTRB` but that's not recommended for this task. I would do a SIMD compare using `PCMPEQB` then `PMOVMSKB` to get the result mask, then look up the first set bit using `BSF` then create a blend mask from that. Avoid looping, use parallelism.




---


**Update**: based on [rwong's comment](https://stackoverflow.com/questions/22283118/byte-wise-operations-on-a-xmm-register-amd64/22283246?noredirect=1#comment33864094_22283118) here is a possible implementation using `pcmpistrm`:



```
3       movdqu input, %xmm1
(gdb) si
4       movdqu replace, %xmm2
(gdb)
5       movdqa %xmm1, %xmm0
(gdb)
6       pcmpistrm $0x78, %xmm1, %xmm1
(gdb) p/s $xmm1.v16_int8
$1 = ""input\000----------""
(gdb) p/s $xmm2.v16_int8
$2 = ""replacereplacere""
(gdb) si
7       pblendvb %xmm1, %xmm2
(gdb) si
8       ret
(gdb) p/s $xmm2.v16_int8
$3 = ""repla\000----------""

```

"
"Present while a presentation is in progress? Trying to display a new view after facebook login with parse.
I'm trying to present a UITableView I've already made for users to put in data and save in parse. I'm pretty sure I don't present the navigation view.


When I log in, I get the error: 



```
Checklists[4516:c07] Warning: Attempt to present <ChecklistsViewController: 0x10525e90> 
on <UINavigationController: 0x9648270> while a presentation is in progress!

```

Thanks for your help.



```
#import ""LoginViewController.h""
#import ""ChecklistsViewController.h""
#import ""SetupViewController.h""
#import <Parse/Parse.h>

@interface LoginViewController ()

@end

@implementation LoginViewController

- (id)initWithNibName:(NSString *)nibNameOrNil bundle:(NSBundle *)nibBundleOrNil
{
    self = [super initWithNibName:nibNameOrNil bundle:nibBundleOrNil];
    if (self) {
        // Custom initialization
    }
    return self;
}

- (void)viewDidAppear:(BOOL)animated
{
    [super viewDidAppear:animated];

    PFLogInViewController *login = [[PFLogInViewController alloc] init];
    login.fields = PFLogInFieldsFacebook;
    // Need to set the delegate to be this controller.
    login.delegate = self;
    login.signUpController.delegate = self; //signUpController is a property on the login view controller
    [self presentModalViewController:login animated:NO];

}

    - (void)logInViewController:(PFLogInViewController *)logInController didLogInUser:(PFUser *)user
{
    [self dismissModalViewControllerAnimated:YES];
    NSLog(@""Successfully logged in."");
    ChecklistsViewController *controller = [[ChecklistsViewController alloc] initWithStyle:UITableViewStylePlain];
    controller.modalTransitionStyle = UITableViewStylePlain;
    [self presentModalViewController:controller animated:YES];

}

```

","This method has been deprecated for a good while



```
     presentModalViewController:animated:

```

You should use this instead



```
    presentViewController:animated:completion:

```

Same goes for this



```
    dismissModalViewControllerAnimated:

```

Now we use this



```
    dismissViewControllerAnimated:completion:

```

When we don't want a completion block, we just set it to nil.


But in your case, a completion block could fix your problem... it ensures a correct sequence of events, i.e. the presenting won't take place until the dismissing is complete.



```
    - (void)logInViewController:(PFLogInViewController *)logInController 
                   didLogInUser:(PFUser *)user
{
    [self dismissViewControllerAnimated:YES completion:^{
        NSLog(@""Successfully logged in."");
        ChecklistsViewController *controller = 
              [[ChecklistsViewController alloc] initWithStyle:UITableViewStylePlain];
        controller.modalTransitionStyle = UIModalTransitionStyleCoverVertical;
        [self presentViewController:controller animated:YES completion:nil];
    }];

}

```

[NB - modalTransitionStyle was incorrect in your original code, I have changed that also. Thanks to [Daniel G](https://stackoverflow.com/users/1652402/daniel-g) for pointing this out]


"
"New lines and tabs in json\_decode() (PHP 7)
My code using `json_decode()` worked correctly with PHP 5.6. After migration to PHP 7.0, `json_decode()` returns NULL and `json_last_error()` tells me that my error is:



> 
> Control character error, possibly incorrectly encoded
> 
> 
> 


After debugging, I found out that my problem are both tabs and new line characters in string values. If I remove them both, it works. If I leave **either** new lines **or** tabs, the error occurs.


Is `json_decode()` behavior changed in PHP 7? I would like to keep tabs and new lines in my .json files for better readability. The code works if I replace tabs to \t and new lines to \n.


How can I keep new lines and tabs?


","Due to a software licensing issue, the old `json` module was replaced with the `jsond` module. You can see the discussion of this change and the attached pull request [here](https://wiki.php.net/rfc/jsond). Now, there's not much information about the changes or about workarounds for things, but I can see that all control characters inside strings (`[\0x00-\0x1F]`) trigger an error. Unfortunately for you, it seems that this behavior is correct per the [JSON Standard](http://www.ecma-international.org/publications/files/ECMA-ST/ECMA-404.pdf):



> 
> Insignificant whitespace is allowed before or after any token. The whitespace characters are: character tabulation (U+0009), line feed (U+000A), carriage return (U+000D), and space (U+0020). Whitespace is not allowed within any token, except that space is allowed in strings.
> 
> 
> 


So, in other words, literal tabs are not allowed inside JSON strings at all; they must be `\t` or `\u0009`. So, the JSON you're consuming is in direct violation of the standard. Ideally, you should get your JSON source to return standards-compliant JSON. If that won't work, you'll have to pre-process the JSON and convert tabs inside strings to `\t`.


"
"Grails unit tests: Accessing defined beans via grailsApplication
I have some (non-Grails-artifact) classes that access the service layer beans via passing around the *grailsApplication* object. However I'm having trouble unit testing the classes implemented in this way. Why doesn't the bean get registered in the main context?



```
@TestMixin(GrailsUnitTestMixin)
class ExampleTests {
  void setUp() {}

  void tearDown() {}

  void testSomething() {
    defineBeans {
      myService(MyService)
    }

    assert grailsApplication.mainContext.getBean(""myService"") != null
  }
}

```

The above code fails with:


**org.springframework.beans.factory.NoSuchBeanDefinitionException: No bean named 'myService' is defined**


What I'm trying to do is access services from plain old Java classes via the grailsApplication. This works, but not in unit test environment. Should I do it differently?



```
class POJO {
  MyService myService;

  public POJO(GrailsApplication grailsApplication) {
    myService = (MyService) grailsApplication.getMainContext().getBean(""myService"");
  }
}

```

","The answer is that in the `GrailsUnitTestMixin` the applicationContext that holds your beans is set as the `parentContext` in the `grailsApplication`



```
beans.registerBeans(applicationContext)

static void initGrailsApplication() {
...
//the setApplicationContext in DefaultGrailsApplication set's the parentContext
grailsApplication.applicationContext = applicationContext
}

```

So you can get your beans with:



```
defineBeans {
  myService(MyService)
}

assert applicationContext.getBean(""myService"")
assert grailsApplication.parentContext.getBean(""myService"")

```

**EDIT**


Today I faced the same problem, and my solution is:



```
@Before
void setup() {
  Holders.grailsApplication.mainContext.registerMockBean(""myService"", new MyService())
}

```

"
"How to make apt-cache, apt-get into one command?
In Ubuntu 11.04 or 10.04, how can I make the `apt-get`, `apt-cache` all in one `myapt` so that I can simply use one only, like in Fedora?


For example:



```
myapt search yum   # same as apt-cache search
myapt remove yum   # same as apt-get remove
myapt install yum  # same as apt-get install

```

Any idea how to do alias for this?


","You could create a script with the following content



```
#!/bin/bash

command=""$1""; shift

case $command in
  add|depends|dotty|dump|dumpavail|gencaches|madison|pkgnames|policy|rdepends|search|show|showpkg|showsrc|stats|unmet|xvcg)
    apt-cache ""$command"" ""$@""
    ;;
  autoclean|autoremove|build-dep|check|clean|dist-upgrade|dselect-upgrade|install|purge|remove|source|update|upgrade)
    apt-get ""$command"" ""$@""
    ;;
esac

```

Suppose you call it `myapt`. Then, to still having the benefit of bash completion, you need to add the following lines to `~/.bashrc`:



```
_myapt() {
  _apt_get
  tt=(""${COMPREPLY[@]}"")
  _apt_cache
  COMPREPLY+=(""${tt[@]}"")
  return 0
} &&
complete -F _myapt $filenames myapt

```

Unfortunately `$command` should precede any options, but seems that bash completion do not works for options that follow command.


"
"How to fix when angular 2 component take full width of the page?
My angular 2 components are taking full width of the page.
When I limit the component width, its margin is stretched to full width.
So I cannot bring other components beside this. How can I limit the width of components?
I tried display: inline, but no use.Still the component takes full width allowing nothing to come beside this.



```
import { Component } from '@angular/core';
import { Item }    from './../item';
import { ItemComponent }    from './ItemComponent';

@Component({
  selector: 'Item-Set',
  directives: [ItemComponent],
  template: '<div class=""singleItem"" *ngFor=""let eachItem of itemList""><Item></Item></div>'
})

export class ItemSet{
  itemList: Item[];
  constructor(){
    this.itemList=ItemList;
  }
}

```

In style sheet



```
.singleItem{
  width:300px;
}

```

","You need to add the CSS to set the width at the correct place. In your example would that be the `styles` parameter of `ItemSet` or `Item` like


[**Plunker example**](https://plnkr.co/edit/E9ukiHJ4K9eOIqjAX3gY?p=preview)


Either on the `ItemComponent` with the `:host` selector (self):



```
@Component({
  selector: 'Item',
  template: '<div>Item</div>',
  styles: [':host {display: block; border: solid 1px red; width: 300px;}']
})
export class ItemComponent{}

```

or the parent component with `.singleItem` as selector



```
@Component({
  selector: 'Item-Set',
  directives: [ItemComponent],
  styles: [`
    .singleItem{width:300px;}
    :host {display: block; border: solid 1px green;}`
  ]
  template: '<div class=""singleItem"" *ngFor=""let eachItem of itemList""><Item></Item></div>'
})
export class ItemSet{
  itemList = ['a', 'b', 'c', 'd'];
})

```

"
"Manage GCM Registration ID in Server
I learn about GCM server today. Reading contents in developer.android.com section GCM, I still not sure how to manage registration ID in third party server. What I understand is:


- android app can register and unregister its ID from/to GCM server.
- Server side can send message into devices using registration ID and  

server key.


So, how can I get device Registration ID from GCM server? Is there I need to get directly from android devices?


Thank you for your advice.


","The best place to learn about the basic information is [here](https://developer.android.com/google/gcm/client.html) at developer.android.com


You will see this sample code in their tutorial:



```
 /**
 * Registers the application with GCM servers asynchronously.
 * <p>
 * Stores the registration ID and app versionCode in the application's
 * shared preferences.
 */
private void registerInBackground() {
    new AsyncTask() {
        @Override
        protected String doInBackground(Void... params) {
            String msg = """";
            try {
                if (gcm == null) {
                    gcm = GoogleCloudMessaging.getInstance(context);
                }
                regid = gcm.register(SENDER_ID);
                msg = ""Device registered, registration ID="" + regid;

                // You should send the registration ID to your server over HTTP,
                // so it can use GCM/HTTP or CCS to send messages to your app.
                // The request to your server should be authenticated if your app
                // is using accounts.
                sendRegistrationIdToBackend();

                // For this demo: we don't need to send it because the device
                // will send upstream messages to a server that echo back the
                // message using the 'from' address in the message.

                // Persist the registration ID - no need to register again.
                storeRegistrationId(context, regid);
            } catch (IOException ex) {
                msg = ""Error :"" + ex.getMessage();
                // If there is an error, don't just keep trying to register.
                // Require the user to click a button again, or perform
                // exponential back-off.
            }
            return msg;
        }

        @Override
        protected void onPostExecute(String msg) {
            mDisplay.append(msg + ""\n"");
        }
    }.execute(null, null, null);
    ...
}

```

If I understand correctly you are confused about how to get the GCM RegistrationId to your server implementation. You can do that in the `sendRegistrationIdToBackend()` method which you will define. 


You need to create an endpoint to receive a POST request that each device will post its RegistrationID to once it has received one from GCM. Then your server implementation will be aware of all the devices that have your app installed and have registered with GCM. 


NOTE: One thing you may want to watch out for is not creating duplicate RegistrationIDs if a device ends up posting their RegistrationID twice, or a user uninstalls and then reinstalls the app but GCM gives them the same RegistrationID you can end up with duplicates stored in your database on the server. If this happens and you want to send a push to all your users then some of your users might end up getting multiple push notifications depending on how many duplicates are in the database. 


"
"How can I implement an EPUB reader in JavaScript?
I'm wondering if I can create an [EPUB](http://en.wikipedia.org/wiki/EPUB) (free and open e-book standard) reader in JavaScript. The basic requirements would be:


1. Server parts of the EPUB reader from a server API.
2. Read the EPUB data in JavaScript.
3. Render it on page.
4. Provide some extra functionality, like text highlights or page notes.


I have no information about how I could do this. I'm willing to try a prototype project. What are the steps that I could take towards implementing such a thing?


","From your comments, it seems like you are at a very early conceptual stage, and want general guidance... well, that's going to be very difficult to give, since the entire topic is quite large. But in general, what you want to do is:


1. Read in and parse an epub file using javascript.
2. Generate HTML which represents the contents (and inline images / SVG / etc).


Well, that's at a very high level, and doesn't help us much. You can break down step 1 by reading up on the epub format itself (e.g.: [wikipedia article](http://en.wikipedia.org/wiki/EPUB) and [general info](http://blog.epubbooks.com/what-is-epub)). Pretty quickly, you should notice that the format uses OCF to package together multiple files, so your first problem will be to create an OCF reader, which also means that you will need to be able to unzip the data in javascript (Florian Margaine's links should give you an idea of how others have solved this problem). At this point, I'd start looking for existing implementations in javascript, because you probably don't want to be implementing all of this from the ground up. This is all before we're even touching the actual contents of the epub file. Once you are past this point, you should be able to read in the actual contents, and attempt to translate them into HTML.


Regarding step 2, I'd start by looking at the various features provided by epub - text, CSS styling, embedded images, etc - and start attacking those one at a time, starting with whatever gives the most return for my time (probably text...).


"
"constexpr expression and variable lifetime, an example where g++ and clang disagree
Consider the simple C++11 code:



```
template<int N>
struct Foo {};

template <int N>
constexpr int size(const Foo<N>&) { return N; }

template <int N>
void use_size(const Foo<N>& foo) { constexpr int n = size(foo); }

int main()
{
    Foo<5> foo;

    constexpr int x = size(foo);  // works with gcc and clang
                                  // _but_
    use_size(foo);                // the same statement in the use_size() 
                                  // function _only_ works for gcc
}

```

I can successfuly compile it with `g++ -std=c++11 foo.cpp`


however if I use clang++, `clang++ -std=c++11 foo.cpp` I get



```
foo.cpp:15:28: error: constexpr variable 'n' must be initialized by a constant expression
void use_size(const Foo<N>& foo) { constexpr int n = size(foo); }
                                                     ~~~~~^~~~
foo.cpp:23:5: note: in instantiation of function template specialization 'use_size<5>' requested here
    use_size(foo);                // the same statement in the use_size() 
    ^
1 error generated.

```

(nb: compiler versions. I have checked the previous statement with g++ version 5.3.1 and 7.2.1 and with clang++ version 3.6.2 and 5.0.0)


**My question:** which of g++ or clang is right? What is the problem?


","My interpretation is that **clang++ is right** and **g++ is too permissive**. 


We can find a close example ([expr.const] section, page 126) in the standard <https://isocpp.org/std/the-standard> (draft can be downloaded, [attention big PDF!](http://www.open-std.org/jtc1/sc22/wg21/docs/papers/2017/n4700.pdf) ). 



```
constexpr int g(int k) { 
    constexpr int x = incr(k); 

    return x; 
}

```

where it is explained that:



> 
> error: incr(k) is not a core constant expression because 
>  lifetime of k began outside the expression incr(k)
> 
> 
> 


This is exactly what is happening in the `use_size()` function with the `foo` argument, even if the `size()` function **only** use the `N` template parameter.



```
template <int N>
constexpr int size(const Foo<N>&) { return N; }

template <int N>
void use_size(const Foo<N>& foo) { constexpr int n = size(foo); }

```

"
"What is a farm, as opposed to a grid or a cluster?
I'm trying to understand does ""farm"" mean in computing, so that, if a farm can be exemplified as a cluster, there must be something that a farm is independently of what a cluster is.


How is it different than a grid?


Do these concepts have a general meaning, where talking about for example web servers is only one common scennario, or do they differ deppending on the context completely? If so, what are the different meanings (or the most common ones)?


Also, should I be asking this somewhere else? (If so, my apologies).


","I'll have a go:


A **server farm** is a group of servers, which may or may not be clustered, that together offer a higher computing capacity for a particular goal than an individual server. An example is a Web farm, which typically consists of a number of load-balanced Web servers, each with the same content and configurations.


While a farm may be clustered, typically a **cluster** consists of servers that are configured in a failover scenario, for instance active/passive, meaning that, say, 1/2 the servers are on active duty while the rest becomes activated only when the servers on active duty are no longer accessible. So when a server in the cluster crashes, it is not a problem because another server takes over.


A **grid** then is typically a configuration of computers that work together (in parallel) to solve a complicated task, such as organic chemistry computations or playing chess against the world champion. The work is divided into parts and each computer takes on a part and reports its findings to the main node which then synthesizes the final result.


I hope this helps.


Some links:


Wikipedia definition of server farm: <http://en.wikipedia.org/wiki/Server_farm>


Wikipedia definition of computing cluster: <http://en.wikipedia.org/wiki/Cluster_(computing)>


Wikipedia definition of grid computing: <http://en.wikipedia.org/wiki/Grid_computing>


"
"How to check if an array has an element at the specified index?
I know there is array\_key\_exists() but after reading the documentation I'm not really sure if it fits for this case:


I have an $array and an $index. Now I want to access the $array, but don't know if it has an index matching $index. I'm not talking about an associative array, but an plain boring normal numerically indexed array.


Is there an safe way to figure out if I would really access an $array element with the given $index (which is an integer!)?


PHP may not care if I access an array with an index out of bounds and maybe just returns NULL or so, but I don't want to even attempt to code dirty, so I *want* to check if the array has the key, or not ;-)


","You can use either the language construct [`isset`](http://php.net/isset), or the function [`array_key_exists`](http://php.net/array_key_exists) : numeric or string key doesn't matter : it's still an associative array, for PHP.


`isset` should be a bit faster *(as it's not a function)*, but will return `false` if the element exists and has the value `NULL`.


  

For example, considering this array :



```
$a = array(
    123 => 'glop', 
    456 => null, 
);

```

And those three tests, relying on `isset` :



```
var_dump(isset($a[123]));
var_dump(isset($a[456]));
var_dump(isset($a[789]));

```

You'll get this kind of output :



```
boolean true
boolean false
boolean false

```

Because :


- in the first case, the element exists, and is not `null`
- in the second, the element exists, but is `null`
- and, in the third, the element doesn't exist


  

On the other hand, using `array_key_exists` like in this portion of code :



```
var_dump(array_key_exists(123, $a));
var_dump(array_key_exists(456, $a));
var_dump(array_key_exists(789, $a));

```

You'll get this output :



```
boolean true
boolean true
boolean false

```

Because : 


- in the two first cases, the element exists -- even if it's `null` in the second case
- and, in the third, it doesn't exist.


"
"Git completions for alias functions
I have a few aliased functions such as `gb` for `git branch` and `gco` for `git checkout` in my .zshrc. This works great when I remember the full branch name I'm creating, deleting, checking out, etc. However, I noticed that completions no longer seem to work. Previously, I could do 



```
$ git checkout m<TAB>

```

and it would autocomplete master if that was the name of a branch. Now, however, I get the following error when I use:



```
$ gco m<TAB>
_git:15: parse error: condition expected: 1

```

I'm not sure why this is occurring. It appears there's possibly a missing argument, but I'm not sure why.


*Edit:*


I'm setting an alias for `git branch` and `git checkout` in my .zshrc file like this:



```
alias gco='git checkout'
alias gb='git branch'

```

","After a little more digging around I found a solution to the same problem for bash that looks to work just as well for zsh. After defining the two functions `__define_git_completion` and `__git_shortcut` you would call `__git_shortcut` to set up your alias and the completion in one call.



```
__define_git_completion () {
eval ""
    _git_$2_shortcut () {
        COMP_LINE=\""git $2\${COMP_LINE#$1}\""
        let COMP_POINT+=$((4+${#2}-${#1}))
        COMP_WORDS=(git $2 \""\${COMP_WORDS[@]:1}\"")
        let COMP_CWORD+=1

        local cur words cword prev
        _get_comp_words_by_ref -n =: cur words cword prev
        _git_$2
    }
""
}

__git_shortcut () {
    type _git_$2_shortcut &>/dev/null || __define_git_completion $1 $2
    alias $1=""git $2 $3""
    complete -o default -o nospace -F _git_$2_shortcut $1
}

__git_shortcut  ga    add
__git_shortcut  gb    branch            # I use this one.
__git_shortcut  gba   branch -a
__git_shortcut  gco   checkout          # and I use this one, too.
__git_shortcut  gci   commit -v
__git_shortcut  gcia  commit '-a -v'
__git_shortcut  gd    diff
__git_shortcut  gdc   diff --cached
__git_shortcut  gds   diff --stat
__git_shortcut  gf    fetch
__git_shortcut  gl    log
__git_shortcut  glp   log -p
__git_shortcut  gls   log --stat

```

- [View Gist from the original answer here.](https://github.com/bronson/dotfiles/blob/731bfd951be68f395247982ba1fb745fbed2455c/.bashrc#L81%3aL114)
- [Found the answer on this thread.](http://git.661346.n2.nabble.com/Bash-tab-completion-for-git-fetch-alias-is-broken-on-Git-1-7-7-1-td6980366.html)


"
"Not able to fetch json data from a ""SINGLE-TYPE"" content-type in strapi
I used strapi as my backend for all my data flow, I created single-type content, this single type has different components attached to it. After configuring all the single-types, I tried to fetch the data through rest Api, But I didn't got the json data as a response.


`This is my json response`


[![This was the json response](https://i.stack.imgur.com/0PTmx.png)](https://i.stack.imgur.com/0PTmx.png)


`This is my single-type structure`


[![This is my single-type structure](https://i.stack.imgur.com/RB5Xu.png)](https://i.stack.imgur.com/RB5Xu.png)


","***So to get back all the data from the ""single-type"" the following procedure are:***


- After creating the **single-Type** content, Go to the settings of the strapi admin page and then click the **Role** option.
- Click on the public tab.
- locate **permissions** tab and search for the *single-type* you created and click on the single-type tab.
- Then you will find the permissions for that single-type, tick on **find** option, there are other option too, tick it as per your needs.
- After that you will see a api endpoint on the right side of the screen for that particular single-type.
- Copy that endpoint.
- Use your browsers or any api client like *Postman* etc to test the api endpoint.


***Now the Main Trick comes here***   



To get the result of the end point you have to type something like this:


`http://localhost:1337/api/""your single-type name""?populate=*`   



This above code will list all the components and all its data inside of it will be displayed in the json response.   



**NOTE:**


If there is any **image data or any other media files inside the component** it won't be displayed, so to to get that also along with other data you need to type:


`http://localhost:1337/api/""your single-type name""?populate[your component name which contains that image data][populate]=*`


*When you type the api endpoint url your ***single-type*** name should be without double quotes*


Follow this link : [***For More Information regarding how to handle the rest api in strapi***](https://docs.strapi.io/developer-docs/latest/developer-resources/database-apis-reference/rest/populating-fields.html#field-selection)


"
"is there a way to modify an attribute of a DOM element directly with jquery?
Does anybody know a way to shorten this:



```
$('.el').attr('src', $('.el').attr('src').replace('/subpath1/', '/subpath2/'))

```

","Do it with a callback function where you can get the old attribute value as the second argument.



```
$('.el').attr('src',function(i, oldSrc){
  return oldSrc.replace('/subpath1/', '/subpath2/');
})

```



---


With [ES6 arrow function](https://developer.mozilla.org/en/docs/Web/JavaScript/Reference/Functions/Arrow_functions)



```
$('.el').attr('src', (i, s) => s.replace('/subpath1/', '/subpath2/'))

```



---


You can reduce the code length using capturing group regex.



```
$('.el').attr('src', (i, s) => s.replace(/(\/subpath)1\//, '$12/'))

```

"
"Delegates.observable() is not getting notified when using MutableList
I'm trying to use `Delegates.observable()` in order to get notified when a change on a `MutableList` has happened.



```
val items: MutableList<Item> by Delegates.observable(startingItems) {
        _, old, new ->
        Log.e(""observable"", ""${old.size}  -> ${new.size}"")
    }

```

but nothing is happening whenever i try to remove nor add something to the list. I mean there's no trace of the log in the Logcat as it's supposed to be there.


","The [docs](https://kotlinlang.org/docs/reference/delegated-properties.html) state:



> 
> **Observable**
> 
> 
> `Delegates.observable()` takes two arguments: the initial value and a handler for modifications. **The handler gets called every time we assign to the property** (after the assignment has been performed). It has three parameters: a property being assigned to, the old value and the new one.
> 
> 
> 


In your case, you do not assign to `items`, you only add to the existing instance. The callback never gets invoked.


Suggestion: Use a mutable property with a read-only `List` and reassign it when a new element is being added:



```
var items: List<String> by Delegates.observable(mutableListOf()) { _, old, new ->
    println(""changed"")
}

//add like this:
user.items += ""new val""

```

The `plus` operator does not call `add` on the list but creates a new instance with all old elements plus the new one.


"
"What is the difference between ""?"" and ""= null"" in PHP function parameters
I was doing some research but wasn't able to find an answer (probably beacause I did not searched it right)


Consider this piece of code:



```
public function foo(?array $optionalParam);

```

And then this one:



```
public function foo(array $optionalParam = null);

```

What differs between them? Using PHPstorm I noticed that when I use the `?`, it creates a PHPdoc and mark the variable type as `type|null`. But when I call the function without that argument, PHP screams on my face ""you kidding me? where is $optionalParam"". In the other side, I managed to use with no problems the `=null` option.


Sorry if this question is too simple, but i did not find any answers online.


","First of all, the `?` goes before the type, not after... other than this:


Using



```
public function foo(?array $optionalParam);

```

you are forced to pass something, that can be either `null` or an `array`, infact:



```
<?php
function foo(?array $optionalParam){
  echo ""test"";
}

foo(); // doesn't work
foo(null); // works
foo([]); // works

```

where instead using



```
public function foo(array $optionalParam = null);

```

will accept `null`, an `array`, or 0 parameters



```
<?php
function foo(array $optionalParam = null){
  echo ""test"";
}

foo(null); // works
foo(); // work
foo([]); // works

```

"
"Type safe heterogeneous container pattern to store lists of items
I'm trying to implement a type-safe heterogeneous container to store lists of heterogeneous objects.


I have seen several exameples of type-safe heterogeneous container pattern ([link](https://stackoverflow.com/questions/416540/java-map-with-values-limited-by-keys-type-parameter)) but all of them store a single object of a type. 


I have tryed to implement it as follows:



```
public class EntityOrganizer {  

    private Map<Class<?>, List<Object>> entityMap = new HashMap<Class<?>, List<Object>>();

    public <T> List<T> getEntities(Class<T> clazz) {
        return entityMap.containsKey(clazz) ? entityMap.get(clazz) : Collections.EMPTY_LIST;
    }

    private <T> void addEntity(Class<T> clazz, T entity) {
        List<T> entityList = (List<T>) entityMap.get(clazz);
        if(entityList == null) {
            entityList = new ArrayList<T>();
            entityMap.put(clazz, (List<Object>) entityList);
        }
        entityList.add(entity);
    }   
}

```

But the problem is this code is full of unchecked casts. Can someone help with a better way of implementing this?


Many thanks


","The question is, what is ""unchecked cast""?


Sometimes casts are provably safe, unfortunately the proof is beyond javac's capability, which does only limited static analysis enumerated in the spec. But the programmer is smarter than javac.


In this case, I argue that these are ""checked casts"", and it's very appropriate to suppress the warning.


See 2 other related examples:


[Heterogeneous container to store genericly typed objects in Java](https://stackoverflow.com/questions/6139325/heterogeneous-container-to-store-genericly-typed-objects-in-java/6139572#6139572)


[Typesafe forName class loading](https://stackoverflow.com/questions/6197802/typesafe-forname-class-loading/6198090#6198090)


"
"Git-svn merge two SVN branches
I have read numerous SO questions and blogs on `git-svn` and merging. Some of them (notably `git-svn` man page) warns *against* using merge anywhere near the branch I plan to `dcommit` from.


Others, like this [SO answer](https://stackoverflow.com/a/4238528/465345), state that merging is fine as long as I use **one-off** feature branches that I delete before `dcommitting`.


I have a setup with two long-living SVN (and git) branches:


- **trunk** (git-svn: `svn/trunk`, git: `master`) - the stable branch
- **branches/devel** (git-svn: `svn/devel`, git: `devel`) - the main development branch from which I fork off feature branches (and I can live with rebasing them back into `devel` instead of merging).


![Snapshot of current state in Source Tree](https://i.stack.imgur.com/EwZNb.png)


In the picture above, **(1)** shows past SVN history, where `branches/devel` has been merged into `trunk`.


**(2)** shows that I have successfully `dcommitted` my current development back into SVN.


**The question:** can I use `git-svn` to merge **two SVN branches** so that history shows the merge point (as SVN itself can do)? In other words: what happens if I were to `dcommit` from `master` as shown in **(3)**?


Will this mess up my SVN (or git) history? Or will this just plain forget that `devel` was merged into `master` at all?


**EDIT:** If `git` just forgets that `devel` was merged into `master`, is there a difference between `dcommiting` from point **(3)** and applying all the commits as a single **patch**?


","Digging some more into the matter, I found out that git supports setting `svn:mergeinfo` property on the SVN branch when `dcommit`'ing:



```
git svn dcommit --mergeinfo ""/branches/somebranch:1-3""

```

**NB!** the `svn:mergeinfo` is **overwritten** with whatever is given on the command-line, so be careful to list *previous merges too*.


While more recent git version added the config parameter to automatically set this property:



```
config key: svn.pushmergeinfo

```

I had some troubles with the automatic mergeinfo - for one reason or the other GIT calculated it wrong and I couldn't get it to work.


**SOLUTION: [git-merge-svn](https://gist.github.com/borgand/5441827)**


To automate the process, I wrote a shell script [git-merge-svn](https://gist.github.com/borgand/5441827) which can be used to merge two SVN branches with correct `svn:mergeinfo` set on the dcommit.


The script handles both situations:


- the branch is not merged in git - will do `git merge` beforehand
- the branches have been already merged in git (but not in SVN) - will traverse until previous ancestor for the merged commit revisions.


With this script I was able to produce these merges solely on git-side and retain the merge info so that GIT graph shows the log nicely:


![git-merge-svn result](https://i.stack.imgur.com/uXa2u.png)


**Example usage:**


1. Make some commits on *devel6*
2. `dcommit` *devel6* to SVN (**required** to get SVN revision numbers for the commits)
3. check out *testtunk6* - yes, I know I made a typo in the name ;-)
4. `git merge-svn devel6`


The last commant outputs:



```
% git merge-svn devel6
About to do an SVN merge: devel6 -> testtunk6

* NEW MERGE COMMIT
|\
| * devel6 [7b71187] (r102)
* | testtunk6 [0682a45] (r101)
 \|
  * [273d6d6] (r100)


STEP 1: GIT merge
Executing:
  git merge devel6

Continue? (y/n) [n]: y
Merge made by the 'recursive' strategy.
 testfile | 1 +
 1 file changed, 1 insertion(+)

STEP 2: SVN dcommit

executing:
git svn dcommit --mergeinfo
/idp/branches/devel:9-32,35-41 /idp/branches/devel6:89 /idp/branches/devel6:94 /idp/branches/devel6:93 /idp/branches/devel6:96 /idp/branches/devel6:97 /idp/branches/devel6:99 /idp/branches/devel6:100 /idp/branches/devel6:102

Continue? (y/n) [n]: y
Committing to https://my.svn.host/svn/auth/idp/branches/testtunk6 ...
  M testfile
Committed r103
  M testfile
Found merge parent (svn:mergeinfo prop): 7b71187fc371d3f86658c5850009e63be88157ac
r103 = 87759323cbadd38bac78087d57b6870a926287e7 (refs/remotes/svn/testtunk6)
No changes between 3fb2168cfbbe605fbd810d76513443203a85a549 and refs/remotes/svn/testtunk6
Resetting to the latest refs/remotes/svn/testtunk6

```

"
"In Firebase, how can I query the most recent 10 child nodes?
I'm using `childByAutoId()` to generate my children. Each child looks like:



```
{
    user_id: 1
}

```

I'd like to get the last 10 most recently added, sorted by time DESC. What's the easiest way to do this?


","The answer is that you need to use a bit of reverse logic, and also store a timestamp key:value pair within each node as a negative value. I omitted the user\_id: 1 to keep the answer cleaner.


Here's the Firebase structure



```
""test"" : {
    ""-KFUR91fso4dEKnm3RIF"" : {
      ""timestamp"" : -1.46081635550362E12
    },
    ""-KFUR9YH5QSCTRWEzZLr"" : {
      ""timestamp"" : -1.460816357590991E12
    },
    ""-KFURA4H60DbQ1MbrFC1"" : {
      ""timestamp"" : -1.460816359767055E12
    },
    ""-KFURAh15i-sWD47RFka"" : {
      ""timestamp"" : -1.460816362311195E12
    },
    ""-KFURBHuE7Z5ZvkY9mlS"" : {
      ""timestamp"" : -1.460816364735218E12
    }
  }

```

and here's how that's written out to Firebase; I just used a IBAction for a button to write out a few nodes:



```
let testRef = self.myRootRef.childByAppendingPath(""test"")

let keyRef = testRef.childByAutoId()

let nodeRef = keyRef.childByAppendingPath(""timestamp"")

let t1 = Timestamp

nodeRef.setValue( 0 - t1) //note the negative value

```

and the code to read it in



```
    let ref = self.myRootRef.childByAppendingPath(""test"")
    ref.queryOrderedByChild(""timestamp"").queryLimitedToFirst(3).observeEventType(.ChildAdded, withBlock: { snapshot in
        print(""The key: \(snapshot.key)"") //the key
    })

```

and I declared a little function to return the current Timestamp



```
var Timestamp: NSTimeInterval {
    return NSDate().timeIntervalSince1970 * 1000
}

```

and the output



```
The key: -KFURBHuE7Z5ZvkY9mlS
The key: -KFURAh15i-sWD47RFka
The key: -KFURA4H60DbQ1MbrFC1

```

As you can see, they are in reverse order.


Things to note:


1. Writing out your timestamp as negative values
2. When reading in use .queryLimitedToFirst instead of last.


On that note, you can also just read the data as usual and add it to an Array then then sort the array descending. That puts more effort on the client and if you have 10,000 nodes may not be a good solution.


"
"MySQLi query returns only one row
This code only returns one row, but it should return 2 rows. I have tried the SQL in phpMyAdmin and it perfectly returned 2 rows. What am I doing wrong here?



```
$request_list_result = $mysqli->query(""
SELECT buddy_requester_id, buddy_reciepient_id, user_id, user_fullname FROM sb_buddies
JOIN sb_users ON buddy_requester_id=user_id
WHERE buddy_status='0' AND buddy_reciepient_id='"". get_uid() .""'"");

$request_list_row = $request_list_result->fetch_array();

echo $request_list['user_fullname'];

```

By the way, the code above is getting process to profile.php via following script:



```
$index = new Template('views/template/one.php', array(
'subtitle' => 'Dashboard',
'stylesheets' => array('/assets/css/profile.css'),
'scripts' => array('/assets/js/dashboard.js'),
'sidebar' => 'sidebar.php',
'content' => 'views/profile.php',
'errors' => $errors,
'successes' => $successes,
'request_list' => $request_list_row //right here
), true);

```

","You need to loop through the results (as TheSmose mentioned)



```
while ($request_list_row = $request_list_result->fetch_array()) {
    echo $request_list['user_fullname'];
}

```

**AND** you need to send the resulting array `$request_list` to the template rather than `$request_list_row`.


**Change this**



```
'request_list' => $request_list_row //right here

```

**to this**



```
'request_list' => $request_list //right here

```

If you want more than just `user_fullname` in your template (and you don't have PHP >= 5.3 required for `mysqli_result::fetch_all`), then you will need to build up your own array inside the loop.


I don't know what your template code expects, but you could try



```
while ($request_list_row = $request_list_result->fetch_array()) {
    echo $request_list[] = $request_list_row;
}

```

"
"What are the .NET Standard versioning rules?
[.NET Standard](https://learn.microsoft.com/en-us/dotnet/articles/standard/library) prescribes an API that all .NET Platforms must implement. What are its versioning rules? Is it `breaking.adding`, in which `1.4` adds to and remains backward compatible with `1.3` whereas `2.x` is not backward compatible with `1.x`?


The documentation is not clear on this. [Some Microsoft docs](https://learn.microsoft.com/en-us/dotnet/articles/standard/library) indicate pure backward compatibility:



> 
> Given a .NET Standard Library version, **you can use libraries that target that same or lower version.** (emphasis added)
> 
> 
> 


Now that 2.0 is out, the above doesn't seem correct. That being said, [the release blog post](https://blogs.msdn.microsoft.com/dotnet/2016/09/26/introducing-net-standard/) said: 



> 
> From a library targeting .NET Standard you’ll be able to reference [libraries targeting] .NET Standard, **if their version is lower or equal to the version you’re targeting.** (emphasis added)
> 
> 
> 


That same blog post contradicted itself by saying: 



> 
> In order to allow .NET Framework 4.6.1 to support .NET Standard 2.0, we had to remove all the APIs from .NET Standard that were introduced in .NET Standard 1.5 and 1.6.
> 
> 
> 


Now that 2.0 is out, what are the versioning rules? It appears to be `breaking.adding`. Where has MSFT documented this? 


","I've explained this in a bit more detail in our [On.NET episode on .NET Standard](https://channel9.msdn.com/Shows/On-NET/Immo-Landwerth-Net-Standard#time=6m58s).


Generally, this is how .NET Standard works:


1. .NET Standard will version linearly, with the intention of not making breaking changes between versions. In other words, you can think of the API surface of .NET Standard as concentric circles, where higher versions have more APIs.
2. A specific version of a .NET platform will implement a specific version of .NET Standard.
3. When choosing a .NET Standard version to target consider this trade-off:
	- The higher the version number, the more APIs you can use
	- The lower the version number, the more .NET platforms support it


So why is there this talk about breaking changes? The short answer is because we made a mistake when defining .NET Standard 1.x and didn't take platform reach into consideration. You should ignore .NET Standard 1.5 and 1.6 and avoid taking a dependency on them. If you do that, .NET Standard 2.0 is a strict superset of .NET Standard 1.4.


For more details, read the section *.NET Standard 2.0 breaking change: adding .NET Framework 4.6.1 compatibility* in [my blog post on .NET Standard](https://blogs.msdn.microsoft.com/dotnet/2016/09/26/introducing-net-standard/).


**Update**. After a lot of community feedback we decided not to perform this breaking change. More details around this decision is listed in the [.NET Standard FAQ](https://github.com/dotnet/standard/blob/master/docs/faq.md#how-does-net-standard-versioning-work).


"
"Print binary matches only with sed?
Let's first have a binary test file:



```
echo -e '\x00\x01\x00\x0a\x00\x0f\x32\x7a\xb0\x00\x00\x01' > test.bin

hexdump -C test.bin 
# 00000000  00 01 00 0a 00 0f 32 7a  b0 00 00 01 0a           |......2z.....|
# 0000000d

```

Now let's see if I can match the bytes' sequence 0x0f 0x32 0x7a with `sed`:



```
sed -n 's/\(\x0f\x32\x7a\)/\1/p' test.bin | hexdump -C
# 00000000  00 0f 32 7a b0 00 00 01  0a                       |..2z.....|
# 00000009

```

That works as expected - the printed match is from the last linefeed 0x0a to the ~~next one~~ end. Now, I want to print the match only - first I try to filter out with `.*` regex at start:



```
sed -n 's/.*\(\x0f\x32\x7a\)/\1/p' test.bin | hexdump -C
# 00000000  0f 32 7a b0 00 00 01 0a                           |.2z.....|
# 00000008

```

That works - now let's do the same, but also for the trailing part:



```
sed -n 's/.*\(\x0f\x32\x7a\).*/\1/p' test.bin | hexdump -C
# 00000000  0f 32 7a b0 00 00 01 0a                           |.2z.....|
# 00000008

```

Well, that does *not* work - only the heading part is removed - but the trailing part keeps going, even if I also terminated my `sed` match pattern with `.*` ??!


What is going on here - and how can I get `sed` to print out only the bytes 0x0f 0x32 0x7a on output (taking into account that ~~`hexdump`~~ `sed` will add the final linefeed 0x0a, when it prints a match)?


","Interesting. Here's a simpler repro case:



```
echo -en '\xff\x80' | sed -n 's/\xff.*/!/p' | hexdump -C

```

The above prints `21 80` which is `!\x80`. The `\x80` can be a larger ASCII code too, but it cannot be smaller: `\x7F` has `sed` doing the expected thing, printing only the `!`.


Also check out what this does:



```
echo -en '\xff\x80' | sed -n 's/\xff./!/p' | hexdump -C

```

It prints nothing at all.


So the question becomes, what's special about `\x80` and higher? Well, [UTF-8](http://en.wikipedia.org/wiki/UTF-8) of course! In UTF-8, having the first bit of a code point set indicates more bytes are coming. And `sed` never finds them, so it never interprets the character at all.


If you want to ""fix"" it, tell `sed` to use the ""good old"" C locale:



```
LC_ALL=C sed ...

```

Then you get your expected output:



```
echo -e '\x00\x01\x00\x0a\x00\x0f\x32\x7a\xb0\x00\x00\x01' |
  LC_ALL=C sed -n 's/.*\(\x0f\x32\x7a\).*/\1/p' |
  hexdump -C

00000000  0f 32 7a 0a                                       |.2z.|

```

"
"error: value ""5.1.2"" for key ""Version"" in group ""Desktop Entry"" is not a known version
I have built the latest version of Paraview from source and I am trying to add the desktop file to the dash. 


Running



```
sudo desktop-file-install /opt/ParaView5.1.2/Applications/ParaView/paraview.desktop

```

Produces the error



```
/usr/share/applications/paraview.desktop: error: value ""5.12"" for key ""Version"" in group ""Desktop Entry"" is not a known version
Error on file ""/opt/ParaView5.1.2/Applications/ParaView/paraview.desktop"": Failed to validate the created desktop file

```

The desktop file is 



```
[Desktop Entry]
Version=5.1.2
Type=Application
Name=ParaView
Comment=Parallel visualization application
TryExec=paraview
Icon=paraview

```

","The `Version` entry in a `.desktop` file doesn't refer to the version of the target program. It's the version of the desktop file specification that this desktop file conforms to. As explained in the [Freedesktop.org specification](https://specifications.freedesktop.org/desktop-entry-spec/latest/ar01s05.html) (emphasis mine):



> 
> Version of the Desktop Entry Specification that the desktop entry conforms with. Entries that confirm with this version of the specification should use 1.0. **Note that the version field is not required to be present.**
> 
> 
> 


So, most importantly, you can just remove the `version` entry, it isn't needed. If you really want it though, it will have to be a valid desktop specification version like `1.0`. 


"
"Invariant Violation: requireNativeComponent: ""RNCSafeAreaProvider"" was not found in the UIManager
Implementing React Navigate in the project installing the package `npm install @react-navigation/native @react-navigation/stack` shows the following error.


[![THIS IS MY ERROR](https://i.stack.imgur.com/Q1A5O.png)](https://i.stack.imgur.com/Q1A5O.png)


Can anyone help me with this issue?


","**TL;DR**


Try upgrading your expo CLI with `npm install -g expo-cli` and running `expo update` to fix the dependencies.


**My case**


I experienced this issue today in an expo app while trying to setup React Navigation.


I noticed this message in the console when launching expo



```
Some of your project's dependencies are not compatible with currently installed expo package version:
 - react-native-reanimated - expected version range: ~1.7.0 - actual version installed: ^1.9.0
 - react-native-screens - expected version range: ~2.2.0 - actual version installed: ^2.8.0
 - react-native-safe-area-context - expected version range: 0.7.3 - actual version installed: ^2.0.0
 - @react-native-community/masked-view - expected version range: 0.1.6 - actual version installed: ^0.1.10
Your project may not work correctly until you install the correct versions of the packages.
To install the correct versions of these packages, please run: expo install [package-name ...]

```

Updating expo and running expo update to set the dependencies to something compatible with my SDK seemed to solve the issue for me



```
npm install -g expo-cli
expo update

```

I noticed that some dependencies were reverted to older versions that now seem to work. E.g. (in package.json):



```
-""react-native-safe-area-context"": ""^2.0.0"",
+""react-native-safe-area-context"": ""0.7.3"",

```

Even when the warning message explicitly instructed me to run the `expo install` for each problematic dependency, doing so would lead to the same version that was throwing the error.


"
"How to search for users by ID and customData in Stormpath
I'm thinking about using Stormpath with it's Java API as a user storage.


Looks good, except that I can't search for users.


For example, I get this error:



```
Exception in thread ""main"" com.stormpath.sdk.resource.ResourceException: HTTP 400, Stormpath 2105 (http://docs.stormpath.com/errors/2105): Account ID is not a supported query property.

```

when I execute this query:



```
HashMap<String, Object> queryParams = Maps.newHashMap();
queryParams.put(""ID"", ""4mPXXXXXXXXXX"");
searchResult = application.getAccounts(queryParams);

```

Searching for a user by e-mail however works. I get the same error when I try to search a user by a login-token I stored in the customData property.


It looks like what I want to do is not possible as it looks like the only properties you can query are e-mail and username. But why would they provide this functionality if it didn't work. What am I missing?


","There is an impedance mismatch between common relational database behaviors and those in a REST API. Querying by id, while common in relational databases, is not idiomatic behavior for REST APIs (or most HTTP-based web sites). The URL (href) is the canonical 'pointer' to a resource on the web. In other words, in REST APIs, the canonical identifier *is* the href. Any token inside an href (any internal 'id', special characters, whatever) should be opaque to REST clients and ignored by clients entirely. URLs are king in HTTP and REST.


That being the case, the Stormpath SDK tries to be true to RESTful best practices, so you can obtain any Stormpath resource by using the `client.getResource` method, which accepts an href and the type of object you expect that href to represent:



```
String href = ""https://api.stormpath.com/v1/accounts/"" + id;
Account account = client.getResource(href, Account.class);

```

That said, there's nothing wrong with wanting this to be more conveniently represented in the client API, for example, `client.getAccount(String id)` if you want to keep the notion of IDs. If so, please open a new [feature request](https://github.com/stormpath/stormpath-sdk-java/issues) and we'll be very happy to consider it.


As for queryable `Account` properties, those are [documented here](http://docs.stormpath.com/rest/product-guide/#search-accounts). Stormpath will soon make data in Custom Data searchable as well. While Stormpath feature timelines are never announced, it is the company's highest engineering priority and it should be out soon.


One workaround that is helpful for some people is to store data they want to search in Account Fields that you don't use in your applications. For example, you could use the 'middle name' field to store, say, favorite color. This would only be temporary until custom data search is available. HTH!


"
"How can I change a Boolean property in Lit-Element to hide content on my component?
I have a custom component that should hide content when I set a boolean property to false. Every other property gets reflected except that one. I must be doing something wrong.



```
static get properties(){
      title: {
        type: String,
        attribute: 'title',
      },

      titleEnable: {
        type: Boolean,
        attribute: 'title-enable',
      },
}

constructor() {
    super();
    this.titleEnable=true;
    this.title=""default"";
}

render(){

    return html`                
        ${this.titleEnable 
        ? html`
        <p class=""title"" ?hidden=""${!this.titleEnable}"">${this.title}</p>
        `
        : html ``} 
    ` 
}

```

If I use that component like `<my-component></my-component>` in an HTML file it shows: default as expected.


If I use it like this: `<my-component title=""New Title""></my-component>` it displays: New Title as expected.


BUT if I try to hide it `<my-component title-enable=""false""></my-component>` the boolean just doesn't change. I've tried !title-enable, title-enable='false"", .titleEnable=false and all the variants you can imagine. What pisses me the most is that whenever I set in the constructor 'this.titleEnable=false' and I happen to just declare the variable WITHOUT value on the tag and it takes it as TRUE an ""default"" appears. `<my-component title-enable></my-component>` I am completely lost.


","Ok, this is a tricky one. You need to handle it differently by passing some object as below:



```
  static get properties() {
    return {
      titleConfig: {
        type: Object,
        attribute: 'title-config'
      }
    }
  }

  render() {
    return html`                
      ${this.titleConfig.titleEnable
        ? html`
      <p class=""title"" ?hidden=""${!this.titleConfig.titleEnable}"">${this.titleConfig.title}</p>
      `
        : html``} 
  `
  }


```

In **HTML**:



```
<my-component title-config='{""title"":""shan"", ""titleEnable"": false}'></my-component>

```

Now, the question is why it is `true` every time?



> 
> **Answer:** For a Boolean property to be configurable from markup, it must default to false. If it defaults to true, you cannot set it to false from markup, since the presence of the attribute, with or without a value, equates to true. This is the standard behavior for attributes in the web platform.
> 
> 
> 


It is taken from [polymer doc](https://polymer-library.polymer-project.org/1.0/docs/devguide/properties#configuring-boolean-properties).


*So, by creating an attribute `title-enable`, the `HTML` considers this attribute as `true`*


It's really a bummer for someone who starts working on Polymer or LitElement . 


"
"Why do componentWillUpdate() and componentWillMount() not trigger render function when setState used in these methods?
We can call setState() inside every component lifecycle methods.


**Why do we not call it inside componentWillUpdate() and componentWillMount() ?**


**Why don't these methods trigger render function when put setState inside those ?**


Can anyone explain it in detail please ?


Thank you.


","**setState() in componentWillMount()**



> 
> `componentWillMount()` is invoked immediately before mounting occurs. It
>  is called before `render()`, therefore setting state in this method will
>  not trigger a re-rendering.
> 
> 
> 


`setState()` in `componentWillMount()` will not trigger a **re-render**, but we have never called `render()` at this point yet. So, setting state here will prepare the state object before we enter the first `render()` pass.


Actually, for preparing state object for our very first `render()`, a better way is to use the `constructor()` instead of calling `setState` in `componentWillMount`.


More reading:


- [official doc](https://facebook.github.io/react/docs/react-component.html#componentwillmount)
- [a more comprehensive unofficial doc](https://developmentarc.gitbooks.io/react-indepth/content/life_cycle/birth/premounting_with_componentwillmount.html)




---


**setState() in componentWillUpdate()**



> 
> Note that you cannot call `this.setState()` here. If you need to update
>  state in response to a prop change, use `componentWillReceiveProps()`
>  instead.
> 
> 
> 


`componentWillUpdate()` is called every time a re-render is required, such as when `this.setState()` is called. The reason we do not call `this.setState()` here is that the method triggers another `componentWillUpdate()`. If we trigger a state change in `componentWillUpdate()` we will end up in an infinite loop.


More reading:


- [Official doc](https://facebook.github.io/react/docs/react-component.html#componentwillupdate)
- [comprehensive unofficial doc](https://developmentarc.gitbooks.io/react-indepth/content/life_cycle/update/tapping_into_componentwillupdate.html)


"
"XML-driven GUIs and performance
Reading the [Online Developer Guide page on XML layouts](http://developer.android.com/guide/topics/ui/declaring-layout.html), I've found the following statement:



> 
> Your UI descriptions are external to your application code, which means that you can modify or adapt it without having to modify your source code and recompile.
> 
> 
> 


I know about the many advantages of XML layouts and resources, but since the XML files are placed inside the APK, I think there's no real way to modify the GUI without repackaging. I mean, most of us are using eclipse ADT plugin and ANT to package the apps, so there's no real gain in not compiling the class files (since after modifying the resource files, a developer will have to repackage the app and produce a new APK file). Because of this, there is no way to redeploy just the GUI on devices without redeploying the entire APK.


If this supposition were true, then XML files are going to be the same during the lifetime of an APK file. I guess these files (specially layout ones) have to be parsed and processed at runtime (before an activity's onCreate) which would be less efficient than building the GUI programmatically (like in Swing). Layout is not usually a bottleneck, but if I'm correct, I see here an small waste of time that could be better used (for instance, with animations).


Reading the same page, it states:



> 
> When you compile your application, each XML layout file is compiled into a View resource.
> 
> 
> 


Inspecting one of my APKs, I've been looking for a precompiled file inside `classes.dex` and there's nothing but my java classes and the `R.class` file. The layout XML files are inside the `/res/layout folder`, and there's a file named `resources.arsc` which seems to contain info about other resources (strings, icons names), but nothing related to Views, I think (correct me if I'm wrong).


My questions:  



- Are XML layout files precompiled, and to which file?
- If not, is there a compilation option to fully precompile layout info into a file for faster load time? (ideally, this would be a class file)
- If not, is there a way to create that file in the first execution and cache it on the application installation folder, so that subsequent runtime executions can read that file instead of parsing the XML and have a faster load time?


Thanks in advance.


","If you have a gander at the class documentation for [LayoutInflater](http://developer.android.com/reference/android/view/LayoutInflater.html), you'll notice they say:



> 
> For performance reasons, view inflation relies heavily on pre-processing of XML files that is done at build time. Therefore, it is not currently possible to use `LayoutInflater` with an `XmlPullParser` over a plain XML file at runtime; it only works with an `XmlPullParser` returned from a compiled resource (`R.something file`.)
> 
> 
> 


So yes, layout files are indeed precompiled to some degree, ~~and, judging by the above excerpt, it would be in the `R$layout.class` output file (but I'm not 100% sure on that).~~ The precompiled layout file is located in your compiled APK package, as `/res/layout/<layout_id>.xml`. You'll notice if you extract it and open it in a text editor that most plain-text XML elements have been mapped to some binary form.


It would probably be the same kind of compression you can see in `AndroidManifest.xml` files that get packaged into the APKs.


"
"Upgrading from Gradle 5.1.1 to 6.0.1 breaks Kotlin multiplatform build
I've upgraded my Kotlin multiplatform project to use Gradle `6.0.1`:



```
gradle wrapper --gradle-version 6.0.1 --distribution-type all

```

and now my build is breaking. It doesn't recognize the common modules I add as a dependency to my project:



```
dependencies {
    commonMainApi(""mygroup:mylib:$myversion"")
}

```

I'm using the Kotlin DSL and this project is also a multiplatform project. What I get is a wall of text detailing all the imports which can't be resolved (all supposed to be imports from a `common` module in my dependency to a `common` module in my project).


The only thing I did was to upgrade to Gradle `6.0.1`. If I restore the previous state my build is OK. What am I doing wrong?


","This is most likely connected with the fact that Gradle 6.0+ doesn't eagerly request the `*.module` metadata files from a repository unless the `*.pom` of the module contains a special marker, which was absent in `*.pom`s published with older Gradle versions (pre-5.3, I believe)


Those `*.module` metadata files are needed to properly interpret a single dependency as both the common code metadata used to analyze the common sources of your project and the platform-specific artifacts that your targets are built against. Without that, the dependency is resolved to the library's root module which has no artifacts at all.


To fix this on the consumer side, you can make Gradle request those `*.module` metadata files by adding this statement to the repository declaration in your build scripts:



```
repositories {
    jcenter { 
        metadataSources { 
            gradleMetadata()
            mavenPom() 
        }
    }
    // or, if you are using a custom Maven repository:
    maven(""https://my.repo.com"") { 
        metadataSources { 
            gradleMetadata()
            mavenPom() 
        } 
    }
}

```

In the Gradle docs: [Supported metadata sources](https://docs.gradle.org/current/userguide/declaring_repositories.html#sec:supported_metadata_sources)




---


**UPD:** JitPack seems to remove the Gradle module metadata marker (`<!-- do_not_remove: published-with-gradle-metadata -->`) from the POMs, which results in Gradle not requesting `*.module` metadata files. A similar workaround as described above can also be used.


"
"How to inject ApplicationContext from Component?
I want to use ApplicationContext instance inside MyComponent class. When I try to autowire, I am getting null pointer exception when Spring initializing my components on startup.


Is there any way to autowire ApplicationContext inside MyComponent class?



```
@SpringBootApplication
public class SampleApplication implements CommandLineRunner {

    @Autowired
    MyComponent myComponent;

    @Autowired
    ApplicationContext context;  //Spring autowires perfectly at this level

    public static void main(String[] args) {
        SpringApplication.run(SampleApplication.class, args);
    }

}


@Component
public class MyComponent{

    @Autowired
    ApplicationContext ctx;

    public MyComponent(){
        ctx.getBean(...) //throws null pointer
    }

}

```

","If you want to make sure, that your dependency (bean) is initialized and ready in the constructor you should use constructor injection, not field injection: 



```
    @Autowired
    public MyComponent(ApplicationContext ctx){
        ctx.getBean(...) // do something
    }

```

Another approach is to use [`@PostConstruct`](https://docs.oracle.com/javaee/7/api/javax/annotation/PostConstruct.html) like below: 



```
@Component
public class MyComponent {

    @Autowired
    ApplicationContext ctx;

    public MyComponent(){

    }

    @PostConstruct
    public void init() {
        ctx.getBean(...); // do something
    }

}

```

You get `NPE` because spring needs to first create the bean (`MyComponent`) and then set the field's value, so in the constructor, the value of your field is `null`.


"
"How do I have no default web browser, so the system asks what browser to use when I click on a link?
I use Internet Explorer for some things and Chrome for others. I'd like to have a choice when I click on a link of which browser to open the link in. Is this possible in Windows? I am currently running Windows 8, although there should be similarities between versions.


To the best of my knowledge this isn't possible.


","I found the following as the simplest solution. Warning: Do not do this if you are not a local administrator.


Modify the following registry keys:


- `[HKEY_CURRENT_USER\Software\Microsoft\Windows\Shell\Associations\UrlAssociations\http\UserChoice]`
- `[HKEY_CURRENT_USER\Software\Microsoft\Windows\Shell\Associations\UrlAssociations\https\UserChoice]`


Change `ProgId` to be a nonsensical value.


Remove your local admin account from permissions on the UserChoice folder. (You will still retain admin privileges as long as you are in the administrator's group)


When clicking on a http/https link, Windows will look at that registry key to see what program to run. When it can't finds it, it brings up a list of all registered http handlers for you to choose from. Normally it would set whatever one you choose to that registry key, but it fails due to the permission being removed. It would work if it had elevated permissions, but that web browser selection screen does not have elevated permissions.


A hack, but it's working for me, and I believe it is safe.


"
"How to swipe through different ionic tabs
first post ever here but would really appreciate some help or advice on this:


I am currently building a project using the ionic framework and after building a functional version I decided that having the ability to swipe between tabs to show the separate sections of the app.


I built the app using the tab template that ionic offers so each page is shown through the ion-nav-view element and is a template called through state changes declared in the app.js file (see below):



```
angular.module('starter', ['ionic', 'starter.controllers', 'starter.services'])

.run(function($ionicPlatform) {
  $ionicPlatform.ready(function() {
    // Hide the accessory bar by default (remove this to show the accessory bar above the keyboard
    // for form inputs)
    if (window.cordova && window.cordova.plugins && window.cordova.plugins.Keyboard) {
      cordova.plugins.Keyboard.hideKeyboardAccessoryBar(true);
    }
    if (window.StatusBar) {
      // org.apache.cordova.statusbar required
      StatusBar.styleLightContent();
    }
  });
})

.config(function($stateProvider, $urlRouterProvider) {

  // setup an abstract state for the tabs directive
    .state('tab', {
    url: ""/tab"",
    abstract: true,
    templateUrl: ""templates/tabs.html""
  })

  // Each tab has its own nav history stack:

  .state('tab.dash', {
    url: '/dash',
    views: {
      'tab-dash': {
        templateUrl: 'templates/tab-dash.html',

      }
    }
  })

  .state('tab.notes', {
      url: '/notes',
      views: {
        'tab-notes': {
          templateUrl: 'templates/tab-notes.html',
          controller: 'noteController'
        }
      }
    })

  .state('tab.todos', {
    url: '/todos',
    views: {
      'tab-todos': {
        templateUrl: 'templates/tab-todos.html',
        controller: 'todoController'
      }
    }
  })

  .state('tab.doodles', {
    url: '/doodles',
    views: {
      'tab-doodles': {
        templateUrl: 'templates/tab-doodles.html',
      }
    }
  })

  // if none of the above states are matched, use this as the fallback
  $urlRouterProvider.otherwise('/tab/dash');

});

```

What I want to know is; Is there a way I can allow the user to swipe left and right to switch between the different pages?


Is it even possible? If so, should it be when there is a need to scroll as well?


I hope this is enough detail, if not I would be happy to provide as much as I can. Thanks for listening!


","Yes, this is possible. I played around with the tabs template and came to the following result:



```
<ion-content on-swipe-right=""goBack()"" on-swipe-left=""goForward()"">

```

And in each controller you will need the corresponding functions:



```
.controller('MyCtrl', function ($scope, $ionicTabsDelegate) {

    $scope.goForward = function () {
        var selected = $ionicTabsDelegate.selectedIndex();
        if (selected != -1) {
            $ionicTabsDelegate.select(selected + 1);
        }
    }

    $scope.goBack = function () {
        var selected = $ionicTabsDelegate.selectedIndex();
        if (selected != -1 && selected != 0) {
            $ionicTabsDelegate.select(selected - 1);
        }
    }
})

```

I don't know if this is best practice and very robust. Like I said, I just played around a little after reading the [docs](http://ionicframework.com/docs/api/service/$ionicTabsDelegate/).


I hope I gave you an idea of how it could work.


"
"When an instance of a class is returned, do all properties get returned?
Let's say I have a class called `Customer` with some properties:



```
public class Customer
{
    public string CustAddr = GetCustAddr(); //Long procedure returning address

    public double CustAcct = GetCustAcct(); //Long procedure returning account

    public string CustInvoice = GetCustInvoice(); //Long procedure returning invoice
}

```

This class is returned via a function:



```
public Customer GetData(string query)
{
    Customer cust = new Customer();
    //set the values for each of the properties
    return cust;
}

```

Right now it is returning the whole class and I can use it like this:



```
lblDisplay = GetData(""QUERY"").CustAddr.ToString();

```

However, suppose each property takes a lot of time to compute. If I only want the CustAddr value, it still computes and has `CustAcct` and `CustInvoice` available for me.


How do I alter my function to **only** return the property I'm looking for, aside from breaking up my class into individual procedures to call? For example, I could just:



```
lblDisplay = GetCustAddr().ToString();

```

but that's not what I'm looking for. I think it's better to have all my data in an organized structure instead of a bunch of different procedures.


","This is a great candidate for lazy initialization. Here I present an example using properties (you have fields).



```
public class Customer
{
    public Lazy<string> CustAddr { get; private set; }
    public Lazy<double> CustAcct { get; private set; }
    public Lazy<string> CustInvoice { get; private set; }

    public Customer()
    {
        CustAddr = new Lazy<string>(GetCustAddr);
        CustAcct = new Lazy<double>(GetCustAcct);
        CustInvoice = new Lazy<string>(GetCustInvoice);
    }
}

```

However, I would be remiss if I didn't point out that you should not use floating-point types (`float`, `double`) to store monetary values. Consider instead using `decimal`.


"
"How to use setTimeout on vuex action?
I would like to empty the alert state, so that the alert is not displayed, I don't really know how to trigger removeAlert action x seconds after addMovieToFavourites or removeMovieToFavourites executed, code below, thanks.


alert.js



```
const state = {
  alert: null
}
const mutations = {
  SET_ALERT(state, alert) {
    state.alert = alert
  },

  REMOVE_ALERT(state) {
    state.alert = null
  }
}
const actions = {
  setAlert({ commit }, alert) {
    commit('SET_ALERT', alert)
  },
  removeAlert({ commit }) {
    commit('REMOVE_ALERT')
  }
}

```

media.js
adding/ removing triggers the alert message



```
const actions = {
  addMovieToFavourites({ commit, dispatch }, movie) {
    commit('ADD_FAVOURITE', movie)
    const alert = {
      type: 'success',
      message: 'Added to favourites!'
    }
    dispatch('alert/setAlert', alert, { root: true })
  },
  removeMovieFromFavourites({ commit, dispatch }, movie) {
    commit('REMOVE_FAVOURITE', movie)
    const alert = {
      type: 'success',
      message: 'Removed from favourites!'
    }
    dispatch('alert/setAlert', alert, { root: true })
  },
}

```

alert.vue



```
<script>
import { mapActions, mapState } from 'vuex'
export default {
  name: 'Alert',
  data() {
    return {
      timeout: null
    }
  },
  mounted() {
    this.timeout = setTimeout(() => this.removeAlert(), 3000)
  },

  beforeDestroy() {
    clearTimeout(this.timeout)
  },
  computed: {
    alertTypeClass() {
      return `alert-${this.alert.type}`
    },

    ...mapState('alert', ['alert'])
  },

  methods: mapActions('alert', ['removeAlert'])
}
</script>

```

","Add it and remove it from the `media` actions:



```
addMovieToFavourites({ commit, dispatch }, movie) {
  commit('ADD_FAVOURITE', movie)
  const alert = {
    type: 'success',
    message: 'Added to favourites!'
  }

  // Add alert
  dispatch('alert/setAlert', alert, { root: true })

  // Remove alert
  setTimeout(() => {
    dispatch('alert/removeAlert', { root: true })
  }, 3000)
}

```

If all alerts work this way, you could automatically trigger the removal from every alert by queueing it in the `setAlert` action:



```
const actions = {
  setAlert({ commit }, alert) {
    commit('SET_ALERT', alert);
    setTimeout(() => {
      commit('REMOVE_ALERT');
    }, 3000);
  },
  ...
}

```

Then you wouldn't need the `removeAlert` action.


You may also need some alert management or `clearTimeout` to handle multiple alerts within 3 seconds. As it's written, a previous alert's removal could mean that a later alert isn't shown for a full 3 seconds.


"
"Object.assign—override nested property
I have an Object `a` like that:



```
const a = {
  user: {
   …
   groups: […]
   …
  }
}

```

whereby there are a lot more properties in `a.user`


And I would like to change only the `a.user.groups` value. If I do this:



```
const b = Object.assign({}, a, {
  user: {
    groups: {}
  }
});

```

`b` doesn't have any other Property except `b.user.groups`, all others are deleted. Is there any ES6 way to only change the nested property, without loosing all the other, with `Object.assign`?


","After some trying I could find a solution that looks pretty nice like that:



```
const b = Object.assign({}, a, {
  user: {
    ...a.user,
    groups: 'some changed value'
  }
});

```

To make that answer more complete here a tiny note:



```
const b = Object.assign({}, a)

```

is essentially the same as:



```
const b = { ...a }

```

since it just copies all the properties of `a` (`...a`) to a new Object. So the above can written as:



```
 const b = {
   ...a,          //copy everything from a
   user: {        //override the user property
      ...a.user,  //same sane: copy the everything from a.user
      groups: 'some changes value'  //override a.user.group
   }
 }

```

"
"Selenium+firefox: empty execute\_script arguments
I'm trying to set `textarea` value using javascript instead send\_keys() method.


As documentation says, I should be able to pass webelement to execute\_script as parameter and refer to this parameter via `arguments` array. However, I've checked in firefox js console, that `arguments` is Object and it doesn't matter what i put as execute\_script argument - `arguments` is always an empty object. 



```
>>>> web = webdriver.Firefox()
>>>> web.get(""http://somepage.com"")
>>>> element = web.find_element_by_tag_name(""textarea"")
>>>> web.execute_script(""return typeof(arguments)"", element)
u'object'
>>> web.execute_script(""return arguments"",element)
[]

```

Anyone has any experience with similarly subject? How can I put webElement as argument for javascript?


Using Firefox 35.0, selenium 2.44.0.


","Here is the relevant bug: [Firefox 35: Passing arguments to executeScript isn't working](https://code.google.com/p/selenium/issues/detail?id=8390&q=2.44%20arguments&colspec=ID%20Stars%20Type%20Status%20Priority%20Milestone%20Owner%20Summary).


Which was fixed in [selenium 2.45](https://pypi.python.org/pypi/selenium) which was released today, **upgrade selenium package**:



```
pip install --upgrade selenium

```



---


Old answer:


I was able to reproduce the problem using `selenium==2.44.0` and `Firefox 35.0`:



```
>>> element = web.find_element_by_tag_name('textarea')
>>> web.execute_script(""return arguments"",element)
[]

```

**Downgrading to `Firefox 34.0.5`** solved the issue:



```
>>> element = web.find_element_by_tag_name('textarea')
[<selenium.webdriver.remote.webelement.WebElement object at 0x1022d1bd0>]

```

"
"What mathematics is needed for a lunar lander game?
I'd like to build a game to learn `cocos2d`. `Lunar lander` is the first exercise coming in my mind. Any pointer/source code/tutorial of the physics calculations required will be appreciated. Thanks!


","You'll need stuff like this: 


1. Newton's laws of motion in 2D.
2. Ability to change effect of gravity. 9.8 m/s^2 is the right acceleration on earth, but you should be able to change this to the appropriate value for Mars, moon, Jupiter, etc.
3. Ability to turn thrusters off and on to counteract the effect of gravity. Not a very interesting game if you don't, because every one ends in a crash.
4. Way to relate duration of thruster fire with fuel consumption. If you don't manage fuel well you crash.
5. Initial conditions (e.g., height above surface, initial velocity, initial fuel, etc.)


You'll start with initial conditions and loop over a number of time steps. At the end of each step you'll check the position and velocity. If the y-position above the surface is zero or negative you'll have landed. If the velocity is greater than a critical y-value you'll have a crash; less than the critical value means a safe, soft landing. 


You'll solve Newton's equations of motion numerically. In your case it's four coupled, first order ordinary differential equations: rate of change of velocity in x- and y-directions and rate of change of position in x- and y-directions. If you have the thrusters in place you'll add another equation for conservation of mass for the fuel. 


You can eliminate two equations if you assume that there are no x-components: the lunar lander moves perpendicular to the surface, the thruster force only has a non-zero component in the vertical direction. If that's true, you're down to three equations.


You'll do time stepping, so it'll be good to read up integration techniques like explicit Euler or implicit 5th order Runge-Kutta. 


A challenging problem - not trivial. Good luck.


"
"SAX Feature Not Supported in Spring Boot 2.6.1
Spring boot 2.6.1 supports logback 1.2.7


I've added the following dependencies in project



```
        <dependency>
            <groupId>ch.qos.logback</groupId>
            <artifactId>logback-classic</artifactId>
        </dependency>
        <dependency>
            <groupId>ch.qos.logback</groupId>
            <artifactId>logback-core</artifactId>
        </dependency>

```

But the build fails with below exception:



```
Caused by: javax.xml.parsers.ParserConfigurationException: SAX feature 'http://xml.org/sax/features/external-general-entities' not supported.
    at oracle.xml.jaxp.JXSAXParserFactory.setFeature(JXSAXParserFactory.java:272)
    at ch.qos.logback.core.joran.event.SaxEventRecorder.buildSaxParser(SaxEventRecorder.java:82)
    ... 44 more

```

In SaxEventRecorder.class, it fails while building the SAX parser



```
    private SAXParser buildSaxParser() throws JoranException {
        try {
            SAXParserFactory spf = SAXParserFactory.newInstance();
            spf.setValidating(false);
            //spf.setFeature(""http://apache.org/xml/features/disallow-doctype-decl"", true);
            spf.setFeature(""http://xml.org/sax/features/external-general-entities"", false);
            spf.setFeature(""http://xml.org/sax/features/external-parameter-entities"", false);
            spf.setNamespaceAware(true);
            return spf.newSAXParser();
        } catch (Exception pce) {
            String errMsg = ""Parser configuration error occurred"";
            addError(errMsg, pce);
            throw new JoranException(errMsg, pce);
        }
    }

```

Is there a way to override the implementation?


Thanks :)


","Had the same issue during an upgrade. It wasn't a Spring issue, and you don't want to override the implementation yourself.


`SAXParserFactory` is an abstract class - it can have multiple implementations. The `newInstance` method picks the last/top implementation on the classpath.


I had found that I had an ""extra"" implementation on my classpath due to a dependency having their own implementation of `SAXParserFactory` rather than the 'typical' implementation that Logback expects. And this ""extra"" `SAXParserFactory` did not support the feature that Logback is trying to enable.


I had to explicitly declare a xerces implementation higher in my dependencies in order for a ""real"" implementation to take precedence over the other ""extra"" implementation on the classpath.


So my advice is for you to see what implementations of this abstract class you have (I could easily see this using the IntelliJ IDE), then manage/re-arrange your dependencies so that a proper implementation supporting the feature has a higher precedence on your classpath.


"
"Haskell read lines of file
I want to read a whole file into a string and then use the function `lines` to get the lines of the string. I'm trying to do it with these lines of code:



```
main = do
   args <- getArgs
   content <- readFile (args !! 0)
   linesOfFiles <- lines content

```

But I'm getting the following error by compiling ad it fails:



```
Couldn't match expected type `IO t0' with actual type `[String]'
In the return type of a call of `lines'
In a stmt of a 'do' block: linesOfFiles <- lines content

```

I thought by binding the result of `readFile` to content it will be a `String` DataType, why isn't it?


","
> 
> I thought by binding the result of readFile to content it will be a String DataType, why isn't it?
> 
> 
> 


It is a `String` indeed, that's not what the compiler complains about. Let's look at the code:



```
main = do
   args <- getArgs
   content <- readFile (args !! 0)

```

Now `content` is, as desired, a plain `String`. And then `lines content` is a `[String]`. But you're using the monadic binding in the next line



```
   linesOfFiles <- lines content

```

in an `IO ()` do-block. So the compiler expects an expression of type `IO something` on the right hand side of the `<-`, but it finds a `[String]`.


Since the computation `lines content` doesn't involve any `IO`, you should bind its result with a `let` binding instead of the monadic binding,



```
   let linesOfFiles = lines content

```

is the line you need there.


"
"Ant Propertyregex part of file path
I am looping list of files inside base directory using for of Ant-Contrib library. I want to get the part of file path after base directory but without filename.


For example my base Directory is : C:\projects\Dev\Main\Sample Game\js
I have lot of js scripts and inside this folder and subfolders name like



```
C:\projects\Dev\Main\Sample Game\js\simple\welcome\ss.js
C:\projects\Dev\Main\Sample Game\js\hard\welcome\cc.js
C:\projects\Dev\Main\Sample Game\js\easy\welcome\ee.js

```

I want to get only `\simple\welcome\`. I am using the below code.


<http://pastebin.com/TpqXBb27>



```
<for param=""filename"">
    <path id=""project.fileset"">
        <fileset dir=""${basedir}/js"" includes=""/*"">
            <include name=""**/*.js"" />
        </fileset>
    </path>
    <sequential>
        <basename property=""file.@{filename}"" file=""@{filename}""/>
        <propertyregex property=""currentdirectory"" 
            input=""@{filename}""
            regexp=""${basedir}/js//([^//]*)//${file.@{filename}}"" 
            select=""\1""
            casesensitive=""false""
            override=""true"" />
        <echo message=""FullPath:@{filename}"" />
        <echo message=""Directory:${currentdirectory}"" />
    </sequential>
</for>

```

I don't know what regex I should provide there... I tried with regex `${basedir}/js//([^//]*)//${file.@{filename}}`, but getting the below error



```
java.util.regex.PatternSyntaxException: Illegal/unsupported escape sequence near index 3
C:\projects\Dev\Main\Sample Game\js\simple\welcome\ss.js

```

Please give some suggestion for this regex


","This sounds like more of a job for `pathconvert` rather than `propertyregex`



```
<for param=""filename"">
    <path id=""project.fileset"">
        <fileset dir=""${basedir}/js"" includes=""/*"">
            <include name=""**/*.js"" />
        </fileset>
    </path>
    <sequential>
        <basename property=""file.@{filename}"" file=""@{filename}""/>
        <!-- dirname strips off the trailing file name, leaving the full dir -->
        <dirname property=""dir.@{filename}"" file=""@{filename}""/>
        <pathconvert property=""currentdirectory.@{filename}"">
            <file file=""${dir.@{filename}}"" />
            <!-- pathconvert strips off the leading ${basedir}/js -->
            <map from=""${basedir}/js"" to="""" />
        </pathconvert>
        <echo message=""FullPath:@{filename}"" />
        <echo message=""Directory:${currentdirectory.@{filename}}"" />
    </sequential>
</for>

```

"
"Offline Map in iOS
I have a big need to do an offline map for my app, as it is made mostly for Thailand, where internet connection is often hard to come by. I am using OpenStreetMap right now for my `MKTileOverlay` but am having issues implementing it for offline use. I have found a tutorial that says to subclass MKTileOverlay. So, in my ViewController where the map is I have:



```
 -(void) viewWillAppear:(BOOL)animated {
CLLocationCoordinate2D coord = {.latitude =  15.8700320, .longitude =  100.9925410};
MKCoordinateSpan span = {.latitudeDelta =  3, .longitudeDelta =  3};
MKCoordinateRegion region = {coord, span};
[mapView setRegion:region];
}
- (void)viewDidLoad {
    [super viewDidLoad];
    self.title = @""Map"";
    NSString *template = @""http://tile.openstreetmap.org/{z}/{x}/{y}.png"";
    self.overlay = [[XXTileOverlay alloc] initWithURLTemplate:template];
    self.overlay.canReplaceMapContent = YES;
    [mapView addOverlay:self.overlay level:MKOverlayLevelAboveLabels];


}
- (MKOverlayRenderer *)mapView:(MKMapView *)mapView rendererForOverlay:(id)overlay {

        return [[MKTileOverlayRenderer alloc] initWithTileOverlay:overlay];

}

```

In my subclass of MKTileOverlay, I have:



```
- (NSURL *)URLForTilePath:(MKTileOverlayPath)path {
    return [NSURL URLWithString:[NSString stringWithFormat:@""http://tile.openstreetmap.org/{%ld}/{%ld}/{%ld}.png"", (long)path.z, (long)path.x, (long)path.y]];
}

- (void)loadTileAtPath:(MKTileOverlayPath)path
                result:(void (^)(NSData *data, NSError *error))result
{
    if (!result) {
        return;
    }
    NSData *cachedData = [self.cache objectForKey:[self URLForTilePath:path]];
    if (cachedData) {
        result(cachedData, nil);
    } else {
        NSURLRequest *request = [NSURLRequest requestWithURL:[self URLForTilePath:path]];
        [NSURLConnection sendAsynchronousRequest:request queue:self.operationQueue completionHandler:^(NSURLResponse *response, NSData *data, NSError *connectionError) {
            result(data, connectionError);
        }];
    }
}

```

The issue is that NOTHING gets loaded at all, unless I comment out the code in the subclass. Where am I messing up?


","I am also interested in loading the map in offline mode. 


1. First of all I downloaded the tiles into my local system (computer) with the help of **gmapcatcher**.


You can get some information about the [gmapcatcher here](https://code.google.com/p/gmapcatcher/). And you can [download this application here](https://code.google.com/p/gmapcatcher/downloads/list).


One more important thing if you want to download the files **check offline** and there are many ways that can save the files. For this scenario select **OSM** in setting window (located to the left side where you will specify the path to download in the setting window)
2. Simply add that created tiles folder to the project.
3. And I don' know objective-c so I am implementing in the swift and I posted that code only.


Here is the code 



```
import UIKit
import MapKit
class ViewController: UIViewController, MKMapViewDelegate {
    @IBOutlet weak var mapView: MKMapView!

    override func viewDidLoad() {
        super.viewDidLoad()
        self.mapView.delegate = self

        let baseURL = NSBundle.mainBundle().bundleURL.absoluteString
        let urlTemplate = baseURL.stringByAppendingString(""OSM_sat_tiles/{z}/{x}/{y}.png"")
        //OSM_sat_tiles is the folder name which has the tiles.
        let layer = MKTileOverlay(URLTemplate: urlTemplate)
        layer.canReplaceMapContent = true
        self.mapView.addOverlay(layer)
    }

    func mapView(mapView: MKMapView, rendererForOverlay overlay: MKOverlay) -> MKOverlayRenderer {
        if overlay is MKTileOverlay {
            let renderer = MKTileOverlayRenderer(overlay:overlay)
            renderer.alpha = 0.8
            return renderer
        }
        return MKTileOverlayRenderer(overlay: overlay)
    }
}

```

I will go brief on the code.


In the **viewDidLoad** delegate method I am creating *MKTileOverlay* by giving path of the tiles folder. Since I have placed in the project it will be in the bundle.


And in the **rendererForOverlay** delegate method I am returning the 
*MKTileOverlayRenderer*.


"
"Python os.environ throws key error?
I'm accessing an environment variable in a script with `os.environ.get` and it's throwing a `KeyError`. It doesn't throw the error from the Python prompt. This is running on OS X 10.11.6, and is Python 2.7.10.


What is going on?



```
$ python score.py
Traceback (most recent call last):
  File ""score.py"", line 4, in <module>
    setup_logging()
  File ""/score/log.py"", line 29, in setup_logging
    config = get_config()
  File ""/score/log.py"", line 11, in get_config
    environment = os.environ.get('NODE_ENV')
  File ""/System/Library/Frameworks/Python.framework/Versions/2.7/lib/python2.7/UserDict.py"", line 23, in __getitem__
    raise KeyError(key)
KeyError: 'NODE_ENV'
$ python -c ""import os; os.environ.get('NODE_ENV')""
$

```

**As requested, here's the source code for `score.py`**



```
from __future__ import print_function

from log import get_logger, setup_logging
setup_logging()
log = get_logger('score')

```

**And here's `log.py`**



```
import json
import os
import sys

from iron_worker import IronWorker
from logbook import Logger, Processor, NestedSetup, StderrHandler, SyslogHandler

IRON_IO_TASK_ID = IronWorker.task_id()

def get_config():
  environment = os.environ.get('NODE_ENV')
  if environment == 'production':
    filename = '../config/config-production.json'
  elif environment == 'integration':
    filename = '../config/config-integration.json'
  else:
    filename = '../config/config-dev.json'

  with open(filename) as f:
    return json.load(f)

def setup_logging():
  # This defines a remote Syslog handler
  # This will include the TASK ID, if defined
  app_name = 'scoreworker'
  if IRON_IO_TASK_ID:
    app_name += '-' + IRON_IO_TASK_ID

  config = get_config()

  default_log_handler = NestedSetup([
    StderrHandler(),
    SyslogHandler(
      app_name,
      address = (config['host'], config['port']),
      level = 'ERROR',
      bubble = True
    )
  ])
  default_log_handler.push_application()

def get_logger(name):
  return Logger(name)

```

","Try running:



```
find . -name \*.pyc -delete

```

To delete your `.pyc` files. 


Researching your problem I came across [this question](https://stackoverflow.com/q/26861856/3642398), where a user was experiencing the same thing: `.get()` seemingly raising a `KeyError`. In that case, it was caused, according to [this accepted answer](https://stackoverflow.com/a/26906934/3642398), by a `.pyc` file which contained code where a `dict` value was being accessed by key (i.e., `mydict['potentially_nonexistent_key']`), while the traceback was showing the code from the updated `.py` file where `.get()` was used. I have never heard of this happening, where the traceback references current code from a `.py` file, but shows an error raised by an outdated `.pyc` file, but it seems to have happened at least once in the history of Python...


It is a long shot, but worth a try I thought.


"
"How to retrieve all the documents in couchdb by given key?
Our couch DB url is <http://10.1.5.88:5984/_utils/database.html?testdata/>


I want to retrieve all the documents for a given key. I'm not sure whether we can search by passing the key, as a query parameter, in the above url.


Please tell me know how to display the query results via the browser.


For example: get all documents where key=""countryname"".


Thanks.


","You will not be able to query the database using ..?testdata&key=""countryname"".


To query the database to get all documents where `key=""countryname""`, you will have to first create a `view` in the database. See the CouchDB documentation at <http://docs.couchdb.org/en/1.6.1/couchapp/views/intro.html>.


I have provided an example below, if it helps:


Suppose your database contains the following documents:



```
{
  ""_id"": ""708ae78bd79fa9802c827e03d3000fe3"",
  ""_rev"": ""1-f3408ffe54a9b9156dca5c9ad56ce396"",
  ""name"": ""AFGHANISTAN"",
  ""code"": ""AF""
}

{
  ""_id"": ""708ae78bd79fa9802c827e03d3001767"",
  ""_rev"": ""1-9952b25ea39460b9923de505945efe3a"",
  ""name"": ""ALAND ISLANDS"",
  ""code"": ""AX""
}

{
  ""_id"": ""708ae78bd79fa9802c827e03d30030a5"",
  ""_rev"": ""1-c1f3c9b2310f07f2985aa3fa363a281f"",
  ""name"": ""ALBANIA"",
  ""code"": ""AL""
}

```

Create a design document `""_design/countries""` with a view `""country""` in the database.



```
{
  ""_id"": ""_design/countries"",
  ""_rev"": ""1-b30408713ded172df62502b45c959563"",
  ""views"": {
    ""country"": {
      ""map"": ""function(doc) {\n  emit(doc.name, null);\n}""
    }
  },
  ""language"": ""javascript""
}

```

You can now query the database like this: `http://127.0.0.1:5984/countries/_design/countries/_view/country?key=%22ALBANIA%22&include_docs=true`


This will return you the document for Albania.



```
{
  ""total_rows"": 3,
  ""offset"": 2,
  ""rows"": [
    {
      ""id"": ""708ae78bd79fa9802c827e03d30030a5"",
      ""key"": ""ALBANIA"",
      ""value"": null,
      ""doc"": {
        ""_id"": ""708ae78bd79fa9802c827e03d30030a5"",
        ""_rev"": ""1-c1f3c9b2310f07f2985aa3fa363a281f"",
        ""name"": ""ALBANIA"",
        ""code"": ""AL""
      }
    }
  ]
}

```

"
"Compare varchar with char
I am trying to figure out what Postgres does when comparing `varchar` with `char`.


Here is one of my tests:



```
test=# select 'a'::character varying = 'a     '::character;
?column?
----------
 t

test=# select 'ab'::character varying = 'ab     '::character;
?column?
----------
 f

```

This looks like a bug to me. Does anyone know what is going on here? Are there good documents on this topic?


","Not a bug at all.  

Casting the string literal `'ab '` to `character`, this is what you get:



```
a

```

[Per documentation:](http://www.postgresql.org/docs/current/interactive/datatype-character.html)



> 
> `character` without length specifier is equivalent to `character(1)`.
> 
> 
> 


`'a'::character(1)` will then be coerced to `varchar` (`character varying`) to test for equality with `'a'::varchar` or `'ab'::varchar` and yield `TRUE` or `FALSE` respectively.


Basically, there is hardly any good reason to use `character` *at all*. It's a legacy type that has outlived its usefulness. Just use `text` or `varchar`.


- [Any downsides of using data type ""text"" for storing strings?](https://stackoverflow.com/questions/20326892/any-downsides-of-using-data-type-text-for-storing-strings/20334221#20334221)


"
"Why does dd from /dev/random give different file sizes?
I am running the following command on an ubuntu system:



```
dd if=/dev/random of=rand bs=1K count=2

```

However, every time I run it, I end up with a file of a different size. Why is this? How can I generate a file of a given size filled with random data?


","You're observing a combination of the peculiar behavior of `dd` with the peculiar behavior of Linux's `/dev/random`. Both, by the way, are rarely the right tool for the job.


Linux's `/dev/random` returns data sparingly. It is based on the assumption that the entropy in the pseudorandom number generator is extinguished at a very fast rate. Since gathering new entropy is slow, `/dev/random` typically relinquishes only a few bytes at a time.


`dd` is an old, cranky program initially intended to operate on tape devices. When you tell it to read one block of 1kB, it attempts to read one block. If the read returns less than 1024 bytes, tough, that's all you get. So `dd if=/dev/random bs=1K count=2` makes two `read(2)` calls. Since it's reading from `/dev/random`, the two `read` calls typically return only a few bytes, in varying number depending on the available entropy. See also [When is dd suitable for copying data? (or, when are read() and write() partial)](https://unix.stackexchange.com/questions/17295/when-is-dd-suitable-for-copying-data-or-when-are-read-and-write-partial)


Unless you're designing an OS installer or cloner, you should never use `/dev/random` under Linux, always `/dev/urandom`. The `urandom` man page is somewhat misleading; `/dev/urandom` is in fact suitable for cryptography, even to generate long-lived keys. The only restriction with `/dev/urandom` is that it must be supplied with sufficient entropy; Linux distributions normally save the entropy between reboots, so the only time you might not have enough entropy is on a fresh installation. Entropy does not wear off in practical terms. For more information, read [Is a rand from /dev/urandom secure for a login key?](https://security.stackexchange.com/questions/3936/is-a-rand-from-dev-urandom-secure-for-a-login-key) and [Feeding /dev/random entropy pool?](https://security.stackexchange.com/questions/89/feeding-dev-random-entropy-pool/7074#7074).


Most uses of `dd` are better expressed with tools such as `head` or `tail`. If you want 2kB of random bytes, run



```
head -c 2k </dev/urandom >rand

```

With older Linux kernels, you could get away with



```
dd if=/dev/urandom of=rand bs=1k count=2

```

because `/dev/urandom` happily returned as many bytes as requested. But this is no longer true since kernel 3.16, [it's now limited to 32MB](https://bugs.debian.org/cgi-bin/bugreport.cgi?bug=775328).


In general, when you need to use `dd` to extract a fixed number of bytes and its input is not coming from a regular file or block device, you need to read byte by byte: `dd bs=1 count=2048`.


"
"Delete files older than given time from directory
I have created list of files to store some attribute after they are opened once in my app.
Every time the file is opened those attributes are changed so I made them be deleted and created again.


I have created all the files files using 



```
File file =new File(getExternalFilesDir(null),
                currentFileId+"""");
if(file.exists()){
           //I store the required attributes here and delete them
           file.delete();
}else{
          file.createNewFile();
}

```

I want to delete all of those files here older than a week as those stored attributes won't be required any more.
What'd be the appropriate method to do so?


","This should do the trick. It will create a calendar instance to 7 days ago and compare if the modified date of the file is before that time. If it is that means that the file is older than 7 days.



```
    if(file.exists()){
        Calendar time = Calendar.getInstance();
        time.add(Calendar.DAY_OF_YEAR,-7);
        //I store the required attributes here and delete them
        Date lastModified = new Date(file.lastModified());
        if(lastModified.before(time.getTime())) {
            //file is older than a week
            file.delete();
        }
    }else{
        file.createNewFile();
    }

```

If you want to get all the files in a dir you can use this and then iterate the result and compare each file.



```
public static ArrayList<File> getAllFilesInDir(File dir) {
    if (dir == null)
        return null;

    ArrayList<File> files = new ArrayList<File>();

    Stack<File> dirlist = new Stack<File>();
    dirlist.clear();
    dirlist.push(dir);

    while (!dirlist.isEmpty()) {
        File dirCurrent = dirlist.pop();

        File[] fileList = dirCurrent.listFiles();
        for (File aFileList : fileList) {
            if (aFileList.isDirectory())
                dirlist.push(aFileList);
            else
                files.add(aFileList);
        }
    }

    return files;
}

```

"
"Interface field in record
Can I rely on the fact, that an interface field in a record is always initialized to `nil`?



```
TMyRec = record  
  FGuard : IInterface;
  FObject : TObject;
  procedure CheckCreated;
end;

```

This would allow me to write:



```
procedure TMyCheck.CheckCreated;
begin
if (FGuard = nil) then
  begin
  FObject := TObject.Create;
  FGuard := TGuard.Create (FObject);
  end;
end;

```

(for automatic lifetime management)


I know that interface fields are initialized to `nil` but is that also true when contained in a record?


","Yes you can rely on that.


All reference-counted variables:


- Strings;
- Dynamic arrays;
- Variants;
- Interfaces;
- Nested records containing those kind of variables.


are initialized to `nil` when a `record` is allocated, if you use `New` or a dynamic array - even locally on the stack. Of course, if you use a plain `GetMem` or work with pointers, you'll have to initialize it by yourself (e.g. using a `FillChar`).


If you are curious, there is an hidden call to the following procedure of System.pas:



```
procedure _InitializeRecord(p: Pointer; typeInfo: Pointer);

```

This will fill all the reference-counted variables memory to 0, but won't set the other members of the `record`. In fact, in a `class` instance, the whole field memory is initialized with 0, including all members - for a `record`, the initialization is only for reference-counted types.


Note that in some cases, I've found out that this initialization was not properly generated if you use the `object` type instead of `record` - [at least under Delphi 2009-2010](http://blog.synopse.info/post/2011/01/29/record-and-object-issue-in-Delphi-2010). So if your code has some `object` type declaration, you may better switch to `record` (and loose inheritance), or explicitly call `FillChar`.


If you are curious, here is an optimized version I wrote in asm - available in [our enhanced RTL](http://blog.synopse.info/post/2010/01/18/Enhanced-System-Run-Time-for-Delphi-7-and-Delphi-2007).



```
procedure _InitializeRecord(p: Pointer; typeInfo: Pointer);
// this procedure is called at most object creation -> optimization rocks here!
asm
        { ->    EAX pointer to record to be initialized }
        {       EDX pointer to type info                }
        MOVZX   ECX,[EDX+1]                  { type name length }
        PUSH    EBX
        PUSH    ESI
        PUSH    EDI
        MOV     EBX,EAX                     // PIC safe. See comment above
        LEA     ESI,[EDX+ECX+2+8]           { address of destructable fields }
        MOV     EDI,[EDX+ECX+2+4]           { number of destructable fields }
@@loop:
        mov edx,[esi]    // type info
        mov eax,[esi+4]
        mov edx,[edx]
        add esi,8
        add eax,ebx      // data to be initialized
        movzx ecx,[edx]  // data type
        cmp ecx,tkLString
        je @@LString
        jb @@err
        cmp ecx,tkDynArray
        je @@DynArray
        ja @@err
        jmp dword ptr [ecx*4+@@Tab-tkWString*4]
        nop; nop; nop // align @@Tab
@@Tab:  dd @@WString,@@Variant,@@Array,@@Record
        dd @@Interface,@@err
@@LString:
@@WString:
@@Interface:
@@DynArray: // zero 4 bytes in EAX
        dec edi
        mov dword ptr [eax],0
        jg @@loop
        POP     EDI
        POP     ESI
        POP     EBX
        RET
@@Variant: // zero 16 bytes in EAX
        xor ecx,ecx
        dec edi
        mov [eax],ecx
        mov [eax+4],ecx
        mov [eax+8],ecx
        mov [eax+12],ecx
        jg @@loop
        jmp @@exit
@@err:
        MOV     AL,reInvalidPtr
        POP     EDI
        POP     ESI
        POP     EBX
        JMP     Error
@@Array:
@@Record: // rarely called in practice
        mov ecx,1
        call _InitializeArray
        dec edi
        jg @@loop
@@exit:
        POP     EDI
        POP     ESI
        POP     EBX
end;

```

"
"Where is std::copy\_exception defined?
The book *[C++ Concurrency in Action](http://www.cplusplusconcurrencyinaction.com/)* by Anthony Williams states in *4.2.4 Saving an exception for the future* that it is possible to store an exception directly without throwing using `std::copy_exception`. However I seem to be unable to find the standard library header where `std::copy_exception` is defined. Where can I find it?


","**tl;dr**: `std::copy_exception` was renamed to `std::make_exception_ptr` in `<exception>` for the final C++11 standard.




---


The committee decided that the name `copy_exception` (probably copied into the standard from [`boost::copy_exception`](http://www.boost.org/doc/libs/1_63_0/libs/exception/doc/copy_exception.html)) was misleading for the following reasons.


The `copy_exception` function returns an `exception_ptr` to a copy of its argument, as if



```
template <class E>
exception_ptr copy_exception(E e) {
    try {
        throw e;
    } catch (...) {
        return current_exception();
    }
}

```

When called with an `exception_ptr` as argument, the function would return another `exception_ptr` pointing to a copy of the `exception_ptr` given as argument, instead of pointing to what the `exception_ptr` argument points to. Because the name `copy_exception` was misleading for this case, the function was renamed to `std::make_exception_ptr` for the final C++11 standard. See the [C++ Standard Library Defect Report 1130](http://www.open-std.org/jtc1/sc22/wg21/docs/lwg-defects.html#1130) for details and discussion on this issue.


The `std::make_exception_ptr` function is defined in `<exception>`.


"
"Prepending a path on make
In my makefile I have an object variable. I need to prepend obj/ to every .o file. How would I do this?



```
CC=g++
CFLAGS=-C -Wall
LDFLAGS=-lsqlite3 -lpthread -ldl
SOURCES=main.cpp Database.cpp actionInit.cpp TileSet.cpp Player.cpp
OBJECTS=$(SOURCES:.cpp=.o)

EXECUTABLE=mahjong-counter

all: bin $(OBJECTS) $(EXECUTABLE)

bin:
    mkdir -p bin

%.o: %.cpp
    $(CC) $(LDFLAGS) $< -c -o $@

$(EXECUTABLE): $(OBJECTS)
    $(CC) $(LDFLAGS) $(OBJECTS) -o $(EXECUTABLE)

clean:
    rm $(OBJECTS)

```

","You want CXX, not CC. CC is for C compiler, not the C++ compiler
In any case, I believe the following should work:



```
CXX=g++
CXXFLAGS=-C -Wall
LDFLAGS=-lsqlite3 -lpthread -ldl

OBJ_DIR = obj
BIN_DIR = bin
EXECUTABLE=mahjong-counter

SOURCES= main.cpp Database.cpp actionInit.cpp TileSet.cpp Player.cpp
OBJECTS= $(SOURCES:%.cpp=$(OBJ_DIR)/%.o)

all: dirs $(OBJECTS) $(EXECUTABLE)

$(EXECUTABLE): $(OBJECTS)
    $(CC) $(LDFLAGS) $(OBJECTS) -o $(EXECUTABLE)

$(OBJ_DIR)/%.o: %.cpp
        $(CXX) $(CXXFLAGS) $< -o $@

dirs:
    mkdir -p $(BIN_DIR)
    mkdir -p $(OBJ_DIR)

.PHONY: dirs all

```

"
"Remove similar items from List
The following data is returned from an SQL View:



```
Name        CandidateID         Filled
Tom Jones   1003436             2014-05-09  07:13:53.087
Tom Jones   1003436             2014-05-09  07:13:18.957
Ed  Harris  1421522             2014-05-09  08:17:20.234

```

I only want the one Tom Jones record with the latest Filled time. How can I achive this in C#/LINQ while getting or after getting data from server?


","Maybe something like this:



```
var q = from n in table
        group n by new {n.CandidateID,n.Name} into g
        select new 
        {
            CandidateID = g.Key.CandidateID, 
            Name = g.Key.Name, 
            Filled = g.Max(t=>t.Filled)
        };

```

**Test class**



```
class Foo
{
    public string Name { get; set; }
    public int CandidateID { get; set; }
    public DateTime Filled { get; set; }
}

```

**Test case**



```
var ls=new List<Foo>
    {
        new Foo(){Name=""Tom Jones"",CandidateID=1003436,
                  Filled=DateTime.Parse(""2014-05-09  07:13:53.087"")},
        new Foo(){Name=""Tom Jones"",CandidateID=1003436,
                  Filled=DateTime.Parse(""2014-05-09  07:13:18.957"")},
        new Foo(){Name=""Ed  Harris"",CandidateID=1421522,
                  Filled=DateTime.Parse(""2014-05-09  08:17:20.234"")}
    };
    var q = 
        (from n in ls
        group n by new {n.CandidateID,n.Name} into g
        select new 
        {
            CandidateID = g.Key.CandidateID, 
            Name = g.Key.Name, 
            Filled = g.Max(t=>t.Filled)
        });

```

**Output**



```
CandidateID     Name           Filled 
1003436         Tom Jones      09/05/2014 7:13:53 AM 
1421522         Ed  Harris     09/05/2014 8:17:20 AM 

```

"
"I uninstalled Unity by mistake
I'm running Ubuntu 14.04. I uninstalled Unity from the Software Center by mistake. How do I reinstall it?


","Open a terminal and execute the following commands :  

To open the terminal on desktop press `Ctrl`+`Alt`+`T`.



```
sudo apt-get update
sudo apt-get install --reinstall ubuntu-desktop
sudo apt-get install unity  

```

Or ... perform the commands from a virtual console.  

To access the virtual console press `Ctrl`+`Alt`+`F1`.


Enter user name and password and then execute the commands.  

After the installation is finished - I recommend to restart the PC. 



```
sudo reboot  

```

**Update - addressing your follow up LXDE question :**  

When the login screen appears select Unity and log in.  

If you want to **uninstall the LXDE desktop** - execute : 



```
sudo apt-get purge lxde

```

"
"How to use argparse arguments as function names
I want to implement the example from the Argparse intro:



```
import argparse

parser = argparse.ArgumentParser(description='Process some integers.')
parser.add_argument('integers', metavar='N', type=int, nargs='+',
                   help='an integer for the accumulator')
parser.add_argument('--sum', dest='accumulate', action='store_const',
                   const=sum, default=max,
                   help='sum the integers (default: find the max)')

args = parser.parse_args()
print(args.accumulate(args.integers))

```

But in my case, I'd like to have a wider range of possible function names to choose from. 



```
def a(): ...
def b(): ...
def c(): ...

parser.add_argument('-f', '--func', 
                     choices=[a, b, c], 
                     required=True,
                     help=""""""Choose one of the specified function to be run."""""")
parser.func()

```

Using it like that doesn't work as intended, I get



```
$ python program.py -f=a
program.py: error: argument -f/--func: invalid choice: 'a' (choose from 
<function a at 0x7fa15f32f5f0>, 
<function b at 0x7fa15f32faa0>, 
<function c at 0x7ff1967099b0>)

```

I know I could solve it with basic string arguments and flow control, but it would be a lot less cluttered and easier to maintain if I could use parser arguments directly as function names.


","You can use basic string arguments and minimize the clutter by using a `dict` to get the actual function from its name.


I'm on Python 2.6.6, so I can't use `argparse`, but this code should give you the general idea:



```
#!/usr/bin/env python

def a(): return 'function a'
def b(): return 'function b'
def c(): return 'function c'

func_list = [a, b, c]
func_names = [f.func_name for f in func_list]
funcs_dict = dict(zip(func_names, func_list))

f = funcs_dict['b']
print f()

```

**output**



```
function b

```

So you can pass `func_names` to `argparse` and compactly retrieve the desired function using `funcs_dict`.




---


In more recent Python versions, you should use `f.__name__` instead of `f.func_name`. (That will also work in Python 2.6, but I was unfamiliar with the newer syntax when I wrote this answer).


"
"Space filling with circles of unequal size
Here is my problem:


- I have a bunch of circles that I need to display inside a canvas.
- There are an arbitrary number of circles, each with a predefined radius.
- The summed area of circles is always smaller than the area of the canvas.


I want to position the circles so that they take up the maximal space available inside the canvas, without touching each other. My goal is to achieve a visually pleasing effect where the circles appear well distributed inside the canvas. I don't know if this is really ""space filling"", as my goal is not to *minimize* the distance between elements, but rather to *maximize* it.


Here is an example of what I am trying to achieve:


![Circles](https://i.stack.imgur.com/9Qcbc.jpg)


My first ""brute force"" idea was the following:


1. For each circle: calculate the shortest distance between its border and each other circle's border; sum all of these distances, call that X.
2. Calculate the sum of all X's.
3. Randomly change the distances between the circles.
4. Redo 1-3 for a preset number of iterations and take the maximal value obtained at step (2).


However, this does not seem elegant; I'm sure there is a better way to do it. Is there any existing algorithm to achieve such a layout? Is there any existing library that I could use (JavaScript or Ruby) to achieve this?


**Edit**


Here is a [Javascript version](https://gist.github.com/2089431) of the accepted answer, which uses Raphael to draw the circles.


","I would try to insert sphere after sphere (largest first). Each one is added in the largest available space, with some random jitter.


One relatively easy way to find (more or less) the largest available space, is to imagine a grid of points on your view and store for each grid point (in a 2D array) the closest distance to any item: edge or sphere, whichever is closest. This array is updated as each new sphere is added. 


To add a new sphere, just take the grid point with highest distance and apply some random jitter (you actually know how much you can jitter, because you know the distance to the closest item). (I would randomize not more than (d-r)/2 where d is the distance in the array and r is the radius of the sphere to add. 


Updating this array after adding another circle is no rocket science: you calculate for each grid point the distance to newly added sphere and replace the stored value if that was larger.


It is possible that your grid is too coarse, and you can't add any more circle (when the 2D array contains no distances larger than the radius of the circle to add). Then you have to increase (e.g. double) the grid resolution before continuing. 


Here are some result of this implementation (it took me about 100 lines of code)


- 100 Circles of varying size


![100 circles of varying size](https://i.stack.imgur.com/LqfN6.png)


- 500 Circles of varying size


![500 circles of varying size](https://i.stack.imgur.com/MLR4P.png)


- 100 Circles of same size


![enter image description here](https://i.stack.imgur.com/CcfAL.png)


And here is some rough C++ code (just the algorithm, don't expect this to compile)



```
    // INITIALIZATION

    // Dimension of canvas
    float width = 768;
    float height = 1004;

    // The algorithm creates a grid on the canvas
    float gridSize=10;

    int gridColumns, gridRows;
    float *dist;

    void initDistances()
    {
      // Determine grid dimensions and allocate array
      gridColumns = width/gridSize;
      gridRows = height/gridSize;

      // We store a 2D array as a 1D array:
      dist = new float[ gridColumns * gridRows ];

      // Init dist array with shortest distances to the edges
      float y = gridSize/2.0;
      for (int row=0; row<gridRows; row++)
      {
        float distanceFromTop = y;
        float distanceFromBottom = height-y;
        for (int col=0; col<gridColumns; col++)
        {
          int i = row*gridColumns+col;
          dist[i]=(distanceFromTop<distanceFromBottom?distanceFromTop:distanceFromBottom);
        }
        y+=gridSize;
      }
      float x = gridSize/2.0;
      for (int col=0; col<gridColumns; col++)
      {
        float distanceFromLeft = x;
        float distanceFromRight = width-x;
        for (int row=0; row<gridRows; row++)
        {
          int i = row*gridColumns+col;
          if (dist[i]>distanceFromLeft) dist[i] = distanceFromLeft;
          if (dist[i]>distanceFromRight) dist[i] = distanceFromRight;
        }
        x+=gridSize;
      }
    }

    void drawCircles()
    {
      for (int circle = 0; circle<getNrOfCircles(); circle++)
      {
        // We assume circles are sorted large to small!
        float radius = getRadiusOfCircle( circle ); 

        // Find gridpoint with largest distance from anything
        int i=0;
        int maxR = 0;
        int maxC = 0;
        float maxDist = dist[0];

        for (int r=0; r<gridRows; r++) 
          for (int c=0; c<gridColumns; c++)
          {
            if (maxDist<dist[i]) {
              maxR= r; maxC= c; maxDist = dist[i];
            }
            i++;
          }

        // Calculate position of grid point
        float x = gridSize/2.0 + maxC*gridSize;
        float y = gridSize/2.0 + maxR*gridSize;

        // Apply some random Jitter
        float offset = (maxDist-radius)/2.0;
        x += (rand()/(float)RAND_MAX - 0.5) * 2 * offset;
        y += (rand()/(float)RAND_MAX - 0.5) * 2 * offset;


        drawCircle(x,y,radius);


        // Update Distance array with new circle;
        i=0;
        float yy = gridSize/2.0;
        for (int r=0; r<gridRows; r++)
        {
          float xx = gridSize/2.0;
          for (int c=0; c<gridColumns; c++)
          {
            float d2 = (xx-x)*(xx-x)+(yy-y)*(yy-y);

            // Naive implementation
            // float d = sqrt(d2) - radius;
            // if (dist[i]>d) dist[i] = d;

            // Optimized implementation (no unnecessary sqrt)
            float prev2 = dist[i]+radius;
            prev2 *= prev2;
            if (prev2 > d2)
            {
              float d = sqrt(d2) - radius;
              if (dist[i]>d) dist[i] = d;
            }



            xx += gridSize;
            i++;
          }
          yy += gridSize;
        }
      }
    }

```

"
"Powershell add text to files if does not exist
I would like to create a Powershell script that adds text to \*.jsp \*.sh \*.cmd files. I would also like it to check if the text already exist in that file and if it does then it skips the file. So far I have found Select-String which will find the text in the file. How do I then use this information to add text to top of the files that do not have the text in them? I also found Add-Content which seems to add the content I want but I would like to do it at the beginning and have some logic to not just keep re-adding it every time I run the ps1.



```
Select-String -Pattern ""TextThatIsInMyFile"" -Path c:\Stuff\batch\*.txt

Add-Content -Path ""c:\Stuff\batch\*.txt"" -Value ""`r`nThis is the last line""

```

","Very similar to what @MikeWise has, but a little better optimized. I have it pulling the data and making the provider filter the files returned (much better than filtering afterwards). Then I pass it to a `Where` statement using `Select-String`'s `-quiet` parameter to provide boolean $true/$false to the Where. That way only file that you want to look at are even looked at, and only those missing the text you need are altered.



```
Get-ChildItem ""C:\Stuff\Batch\*"" -Include *.jsp,*.sh,*.cmd -File | 
    Where{!(Select-String -SimpleMatch ""TextThatIsInMyFile"" -Path $_.fullname -Quiet)} |
    ForEach{
        $Path = $_.FullName
        ""TextThatIsInMyFile"",(Get-Content $Path)|Set-Content $Path
    }

```

**Edit:** As you discovered the `\*` does not work with `-Recursive`. Use the following if you need to recurse:



```
    Get-ChildItem ""C:\Stuff\Batch"" -Include *.jsp,*.sh,*.cmd -File -Recurse

```

"
"Dereference arbitrary memory location in C
I'm trying to debug a program I've written. According to the debugger a particular `void *` holds the value 0x804b008. I'd like to be able to dereference this value (cast it to an `int *` and get it's value).


I'm getting a Segmentation Error with this code. (The program with the `void *` is still running in the background btw - it's 'paused')



```
#include <stdio.h>

int main() {
    int* pVal = (int *)0x804b008;
    printf(""%d"", *pVal);
}

```

I can see why being able to deference any point in memory could be dangerous so maybe that's why this isn't working.


Thank you!


","Your program (running in the debugger) and this one won't be running in the same virtual memory space; accessing that pointer (even if it were a valid one) won't give you any information.


Every program running on your machine has its own logical address space. Your operating system, programming language runtime, and other factors can affect the actual literal values you see used as pointers for any given program. But one program *definitely* can't see into another program's memory space, barring of course software debuggers, which do some special tricks to support this behaviour.


In any case, your debugger should let you see whatever memory you want while your program is paused - assuming you have a valid address. In gdb `x/x 0x804b008` would get you what you want to see.


For more information:


1. Wikipedia article on [Virtual Memory](http://en.wikipedia.org/wiki/Virtual_memory).
2. [`gdb` documentation](http://www.gnu.org/software/gdb/documentation/)


"
"Trying to understand exceptions in C#
I do not really use any try/catches in my code ever but I'm trying to break that habit and now get in to using exceptions.


I figure the most important place to have it in my application would be reading a file and I'm trying to implement that now but I'm unsure of the ""best-practices"" for doing so. Currently I'm doing something like this:



```
private void Parse(XDocument xReader)
{
    IEnumerable<XElement> person = xReader.Descendants(""Person"").Elements();

    foreach (XElement e in person)
        personDic[e.Name.ToString()] = e.Value;

    if (personDic[""Name""] == null || personDic[""Job""] == null || personDic[""HairColor""] == null)
        throw new KeyNotFoundException(""Person element not found."");
}

```

But I am unsure if this is correct. I have this for handling it:



```
try
{
    personsReader.Read(filename, persons);
}
catch (KeyNotFoundException e)
{
    MessageBox.Show(e.Message);

    return;
}

// Do stuff after reading in the file..

```

However when showing e.Message it just shows the generic KeyNotFoundException error message and not by custom error message. Also I'm not sure if in general I am going about this whole ""exception handling stuff"" properly. I do return in the catch because if the file is not read successfully obviously I just want to pretend like the user never tried to open a file and let him try again with another file.


Am I doing this properly? Again I am fairly new to using exceptions and I want to make it sure I got it down right before continuing on and applying this to the rest of my program.


Also, why do people say not to do `catch (Exception e)`? It seems like in this case I would want to do that because regardless of what error occurs when reading in a file, if there is an error, I want to stop reading the file, display the error message, and return. Wouldn't that always be the case? I can understand not wanting to handle Exception e if you would want to handle something differently based on the exception but in this case wouldn't I want to just handle the base exception class in case anything goes wrong?


","You should catch exceptions when you can *handle* the condition and do something useful. Otherwise you should let it bubble up the call stack and perhaps someone above you can handle it. Some apps have unhandled exception handlers to handle it at the outer most layer but in general, unless you know you have some useful way to handle it, let it go.


In your case, you're handling not being able to read a resource and informing the user. You're handling it. Concerning the generic exception, one thing you can do is catch and re-throw a better exception. If you do that, make sure you incude the root cause exception as the inner exception. You can also trace or log the details if appropriate.



```
throw new MyGoodExceptionType (""Could not read file"", e);  // e is caught inner root cause.

```

Now the UI shows a good error and perhaps the inner root cause is in a log etc...


Some typical mistakes:


- **Handling exceptions deep in the stack in a generic library method:** Remember that a common library function may get called in many different code paths. You likely don't have the context whether it should be handled and whether it's appropriate to handle it. the caller higher in the stack likely has context and knows whether it's safe to handle. Typically that means higher layers of code decide to handle. In lower layers, typically you let them flow.
- **Swallowing Exception:** Some code catches exceptions (especially lower in the stack) and then the root condition just evaporates making it maddening to debug. Once agan, if you can handle it, do so. If not, let it go.
- **Exceptions should be exceptional:** Don't use excpetions for flow control. For example, if you're reading a resource, don't try and read and then catch the exception and make a decision point. Instead, call ifexists, check bool and make decisions in your code. this especially helps when you set the debugger to break on exceptions. You should be able to run clean and if the debugger breaks, it should be a real issue. Having the debugger break constantly when debugging is problematic. I personally like throwing exceptions very rarely and always try to avoid for flow control.


Hope that helps. 


"
"How do you use globbing in perl for a one-liner with many files, avoiding xargs/find/etc
When there are too many matching files, shells like `bash` break if you include a glob pattern on the commandline like



```
perl -pi -e 's/hi/bye/' too_many_files*

```

You can work around this with `xargs`, gnu `parallel`, or `find`, but for complex commands, these can be difficult to get right in terms of quoting, and they can also be less efficient than running `perl` once.


Is there a way to use `perl`'s built-in globbing support for something like this? (which does not work)



```
perl -pi -e 's/hi/bye/' 'manyfiles*' # <-- Does not work.

```

","As noted [in this answer](https://stackoverflow.com/a/27942616/411282), you can use a `BEGIN` block to have `perl` (rather than the shell) expand the file list:


slightly modified from the original:



> 
> Leave globing to perl instead of bash which has limitations,
> 
> 
> 
> ```
> perl -pi -e 'BEGIN{ @ARGV = glob(pop) } s/#//g' ""*""
> 
> ```
> 
> or when there are spaces in globed directory,
> 
> 
> 
> ```
> perl -MFile::Glob=bsd_glob -pi -e 'BEGIN{ @ARGV = bsd_glob(pop) } s/#//g' ""*""
> 
> ```
> 
> 


For more about `glob` vs `bsd_glob`, see [this post](https://stackoverflow.com/a/32262481/411282).


(This is intentionally duplicated as I had trouble finding the answer quickly with search terms I had in mind.)


"
"Front-matter defaults in VuePress config
I'm trying to switch my documentation site from GitBook to Vuepress and got stuck with front-matter variables. In GitBook, you just add variables in the config and then use them anywhere on the page as `{{ book.variable_name }}`. In Vuepress, at glance, things seem to be trickier.


I need to configure several variables that are used across the whole site, so adding them to each page would be a complete nightmare. The documentation tells nothing about how to configure front-matter variables but has a link to the Jekyll site. On the Jekyll site, I found [this article](https://jekyllrb.com/docs/configuration/front-matter-defaults/) that is exactly what I want to achieve. The problem is that I have no idea how to use this info in config.


Any help is highly appreciated. I asked this question in the [official repo](https://github.com/vuejs/vuepress/issues/1377) but that didn't help.


","To define some variables that you could access anywhere in your site, you can add them to your theme config.


If you haven't already got one, create a `config.js` file at `.vuepress/config.js`. 


This file should export an object.


You want to add a `themeConfig: {}` to this.


Properties you set on the `themeConfig` object will be available throughout your site on `$themeConfig`.



```
//- .vuepress/config.js

module.exports = {
  themeConfig: {
    //- Define your variables here
    author: 'Name',
    foo: 'bar'
  }
}

```


```
  {{ $themeConfig.author }} //- 'Name'
  {{ $themeConfig.foo }} //- 'bar

```

You can also make this easy to override locally / per page, by using global computed functions. (This could also provide a cleaner way to access the variables)


Adding an `enhanceApp.js` file in the same place as `config.js`, will give you access to the Vue instance - where you can define a mixin for all components.


You can define some computed properties in this mixin that first check for a value in the pages frontmatter data and then fall back to the value set in the themeConfig. Allowing you to set some default values that can be locally overridden per page.



```
//- .vuepress/enhanceApp.js

export default ({ Vue }) => {
  Vue.mixin({
    computed: {
      author() {
        const { $themeConfig, $frontmatter } = this
        return $frontmatter.author || $themeConfig.author
      },
      foo() {
        const { $themeConfig, $frontmatter } = this
        return $frontmatter.foo || $themeConfig.foo
      }
    }
  })
}


```


```
  {{ author }}  //- 'Name'
  {{ foo }} //- 'bar

```

[Vuepress config docs](https://v1.vuepress.vuejs.org/guidebasic-config.html#app-level-enhancements)
[Vuepress app level enhancement](https://v1.vuepress.vuejs.org/guidebasic-config.html#app-level-enhancements)


"
"Is chrome tabId unique across windows
I need to know if the chrome tabId is unique across all the open windows. Incognito and normal. Is it guaranteed that non of the open tabs in all the windows will have the same tabId ?


I searched the documentation but could not find any reliable answer. 


","Yes, the tab ID is unique within a *browser session*. It's also mentioned in the documentation of [`chrome.tabs`](http://code.google.com/chrome/extensions/tabs.html#type-Tab):



> 
> `Tab  
> 
>  ( object )  
> 
>      id ( integer )  
> 
>         The ID of the tab. 
>  **Tab IDs are unique within a browser session**.`
> 
> 
> 


If you still don't believe it, create an extension which has the `tabs` permission, and the right to run in an incognito window. Then run the following code in the background page:



```
// Create incognito window
chrome.windows.create({incognito: true, url:'about:blank'}, showTabId);
// Create normal window
chrome.windows.create({incognito: false, url:'about:blank'}, showTabId);

function showTabId(_window) {
    console.log(_window.tabs[0].id);        // Or alert, whatever.
    chrome.tabs.remove(_window.tabs[0].id); // Closes tab & window, user-friendly
}

```

The logged numbers are increasing (if you consider two numbers as a too small sample, run the [`chrome.windows.create`](http://code.google.com/chrome/extensions/windows.html#method-create) method in a loop, until you believe it).


"
"Default value for ansible hosts in playbook?
My sample playbook (for ansible 2.1) is:



```
---
# This is sample playbook.

- name: add sample_role
  hosts: '{{ target }}'
  become: true
  become_user: root
  roles:
    - sample role

```

When i run it, everything is fine:



```
ansible-playbook -i staging test_playbook.yml --extra-vars ""target=192.168.15.29""

```

How can i omit targets?


I've tried something like this:



```
  hosts: '{{ target | default(all) }}'

```

or 



```
  hosts: '{{ target | default(hostvars) }}'

```

then run:



```
ansible-playbook -i staging test_playbook.yml

```

, but unfortunately:



> 
> ERROR! 'all' is undefined
> 
> 
> 


or:



> 
> ERROR! 'hostvars' is undefined
> 
> 
> 


","If you simply add the `all` without quote then it take it as variable, so you can fixed it like this:


`hosts: '{{ target | default(""all"") }}'`


As far `hostvars` it's a reserved word in ansible and can't use it for this purpose I think. It's value will be something like this:



```
ok: [localhost] => {
    ""msg"": {
        ""localhost"": {
            ""ansible_check_mode"": false,
            ""ansible_version"": {
                ""full"": ""2.1.0.0"",
                ""major"": 2,
                ""minor"": 1,
                ""revision"": 0,
                ""string"": ""2.1.0.0""
            },
            ""group_names"": [],
            ""groups"": {
                ""all"": [
                    ""localhost""
                ],
                ""ungrouped"": []
            },
            ""inventory_dir"": null,
            ""inventory_file"": null,
            ""inventory_hostname"": ""localhost"",
            ""inventory_hostname_short"": ""localhost"",
            ""omit"": ""__omit_place_holder__41c600ef78930ed8b38e6eed4e5b5ab51199729e"",
            ""playbook_dir"": ""/Users/xyz/test-ansible""
        }
    }
}

```

Hope that help you.


"
"Does yieldIfContendedSafely() lose the benefits of a transaction?
I have a long-running operation which I perform in a background thread. As it is important for the operation to either complete successfully or not at all, I am wrapping the entire operation in a transaction.


Aspects of the UI need read-only access to the database during this time. To avoid blocking the UI, I am experimenting with inserting calls to `db.yieldIfContendedSafely()` in the main loop of the background operation.


This does what I want in that the UI is no longer blocked, but it's not completely clear to me if this is risking a loss of data integrity.


The javadoc for `yieldIfContendedSafely()` says:



> 
> Temporarily end the transaction to let other threads run. The
>  transaction is assumed to be successful so far. Do not call
>  setTransactionSuccessful before calling this. When this returns a new
>  transaction will have been created but not marked as successful. This
>  assumes that there are no nested transactions (beginTransaction has
>  only been called once) and will throw an exception if that is not the
>  case.
> 
> 
> 


Does this mean that my long-running operation is actually being committed to the database in separate chunks, or is the *overall* transaction maintaining enough state to commit the whole lot in one go at the end, thus preserving data-integrity?


","
> 
> Does this mean that my long-running operation is actually being committed to the database in separate chunks
> 
> 
> 


Yes. Within `yieldIfContendedSafely()`, Android calls `setTransactionSuccessful()`, `endTransaction()`, and begins a new transaction -- committing your statements in the process. There is no mechanism to rollback the ""real"" transaction after it ends.


This behavior only occurs if there is another thread waiting on the database, otherwise `yieldIfContendedSafely()` does nothing. 


I checked this with the following scenario. I started two threads: one inserted data into a table using a transaction, another read data out of the same table. The transaction didn't call `setTransactionSuccessful()` so normally everything is rolled back at the end, leaving the table empty. I added a call to `yieldIfContendedSafely()`, and afterwards the table was not empty and had data from the transaction.


"
"run j2me app on android?
is it possible to run java SE or Java ME app on android ???


","If you search for ""J2ME runner android"" on Google, you'll find some hits for projects which claim to allow you to run J2ME projects on Android.


e.g. <http://www.netmite.com/android/>


I'm not sure what you mean by Java SE in this case. Technically Dalvik is a lot closer to J2SE than to J2ME. A large set of the APIs are there, and code will likely cross-compile without issue. You can even include a lot of J2SE JARs in your project and use the build tools to re-compile them from bytecode to DEX.


<http://en.wikipedia.org/wiki/Dalvik_(software)>


The infrastructure is vastly different, however, and you're not going to get an application written/compiled for a desktop running a standard JVM to work on Android without changes and recompiling.


"
"Does being a Senior Developer using one language qualify you to be a Senior Developer using another language?
If I work as a Developer in one language (e.g. Java) and work my way up to Senior Developer, would that qualify me to be a Senior Developer for a position using another language (e.g. Ruby)?


","The best way to answer this is to look at what the difference between a Developer and a Senior Developer. Assuming that it's not just a time served thing, generally I'd expect both Developers and Senior Developers to be able to:


- Write code competently in the languages required by the role
- Diagnose and fix bugs
- Write unit tests
- Follow standards and reasonable best practice (version control, documentation)
- Have a broad basic technical competent
- Act in a professional manner


In addition I'd expect a Senior Developer to:


- Mentor other members of staff in best practice
- Be and acknowledge reference point for at least some of the languages being used by the team
- Actively research and champion new areas of best practice
- Take technical ownership of more complex issues / areas of code and provide solid solutions


So, the question then becomes do you fulfil the extended criteria for your second (or third or fourth) language? I'd suggest that so long as you're technically competent enough in the language you're moving to then yes as most of the Senior Developer stuff tends to be transferable.


"
"Check if pull needed in Git
How do I check whether the remote repository has changed and I need to pull?


Now I use this simple script:



```
git pull --dry-run | grep -q -v 'Already up-to-date.' && changed=1

```

But it is rather heavy.


Is there a better way? The ideal solution would check all the remote branches, and return names of the changed branches and the number of new commits in each one.


","First use [`git remote update`](https://git-scm.com/docs/git-remote), to bring your remote refs up to date. Then you can do one of several things, such as:


1. `git status -uno` will tell you whether the branch you are tracking is ahead, behind or has diverged. If it says nothing, the local and remote are the same.
2. `git show-branch *master` will show you the commits in all of the branches whose names end in 'master' (eg *master* and *origin/master*).


If you use `-v` with `git remote update` (`git remote -v update`) you can see which branches got updated, so you don't really need any further commands.


However, it looks like you want to do this in a script or program and end up with a true/false value. If so, there are ways to check the relationship between your current *HEAD* commit and the head of the branch you're tracking, although since there are four possible outcomes you can't reduce it to a yes/no answer. However, if you're prepared to do a `pull --rebase` then you can treat ""local is behind"" and ""local has diverged"" as ""need to pull"", and the other two (""local is ahead"" and ""same"") as ""don't need to pull"".


You can get the commit id of any ref using `git rev-parse <ref>`, so you can do this for *master* and *origin/master* and compare them. If they're equal, the branches are the same. If they're unequal, you want to know which is ahead of the other. Using `git merge-base master origin/master` will tell you the common ancestor of both branches, and if they haven't diverged this will be the same as one or the other. If you get three different ids, the branches have diverged.


To do this properly, eg in a script, you need to be able to refer to the current branch, and the remote branch it's tracking. The bash prompt-setting function in `/etc/bash_completion.d` has some useful code for getting branch names. However, you probably don't actually need to get the names. Git has some neat shorthands for referring to branches and commits (as documented in `git rev-parse --help`). In particular, you can use `@` for the current branch (assuming you're not in a detached-head state) and `@{u}` for its upstream branch (eg `origin/master`). So `git merge-base @ @{u}` will return the (hash of the) commit at which the current branch and its upstream diverge and `git rev-parse @` and `git rev-parse @{u}` will give you the hashes of the two tips. This can be summarized in the following script:



```
#!/bin/sh

UPSTREAM=${1:-'@{u}'}
LOCAL=$(git rev-parse @)
REMOTE=$(git rev-parse ""$UPSTREAM"")
BASE=$(git merge-base @ ""$UPSTREAM"")

if [ $LOCAL = $REMOTE ]; then
    echo ""Up-to-date""
elif [ $LOCAL = $BASE ]; then
    echo ""Need to pull""
elif [ $REMOTE = $BASE ]; then
    echo ""Need to push""
else
    echo ""Diverged""
fi

```

*Note:* older versions of git didn't allow `@` on its own, so you may have to use `@{0}` instead.


The line `UPSTREAM=${1:-'@{u}'}` allows you optionally to pass an upstream branch explicitly, in case you want to check against a different remote branch than the one configured for the current branch. This would typically be of the form *remotename/branchname*. If no parameter is given, the value defaults to `@{u}`.


The script assumes that you've done a `git fetch` or `git remote update` first, to bring the tracking branches up to date. I didn't build this into the script because it's more flexible to be able to do the fetching and the comparing as separate operations, for example if you want to compare without fetching because you already fetched recently.


"
"WPF Datagrid selecteditem = null in MVVM
I'm trying to work with a datagrid using the MVVM pattern. The problem is that whenever I change the VM property which is binded to SelectedItem to null, the View doesn't ""deselect"" the currently selected item. This is my binding in xaml:



```
<DataGrid Grid.Column=""0"" Grid.Row=""0"" 
    ItemsSource=""{Binding Path=Users}"" 
    AutoGenerateColumns=""False"" 
    CanUserAddRows=""False"" 
    IsReadOnly=""True"" 
    SelectedItem=""{Binding Path=SelectedUser, Mode=TwoWay}"">

```

The SelectedItem binding works from the view to the VM thus in the SelectedUser property I always have the selected object. The problem is that in the VM I'm doing some stuff which sometimes changes the SelectedUser property to null so I would expect the datagrid to deselect the row as well. Instead, it remains selected and if I try to click on the same row, the property doesn't update. If I click on any other row, the property changes as expected.


Is there a way to make the datagrid deselect if it's binded property is set to null? Also I'm looking for a MVVM solution as I don't want to write code behind. I can solve this by writing code behind so don't waste time offering such solutions :)


l.e.: this is my property in the VM:



```
public RPLUser SelectedUser
        {
            get
            {                
                return selectedUser;
            }
            set
            {
                selectedUser = value;
                OnPropertyChanged(""SelectedUser"");
            }
        }

```

Thanks in advance!


","I recommend to check the `Output Window` in visual studio and see if any `Binding` is failing. 


Are you sure when you select something, the selection updates into the `SelectedUser` property? 


Did u put a breakpoint in setter of `SelectedUser` and see that it is hitting when you select something on the datagrid?


The reasons for this `Binding` to break could be many ... 


1. `SelectedUser` is of different type than individual `Users`.
2. `SelectedUser` does not match by reference with any items in `Users`.
3. How and where are you setting null?


The following code in my case works perfectly fine...



```
    <tk:DataGrid MaxHeight=""200"" AutoGenerateColumns=""False""
                 ItemsSource=""{Binding}""
                 SelectedItem=""{Binding MySelItem,
                                        ElementName=MyDGSampleWindow,
                                        Mode=TwoWay}""
                 IsReadOnly=""True"">
        <tk:DataGrid.Columns>
            <tk:DataGridTextColumn Header=""Key""
                                   Binding=""{Binding Key,
                                                     Mode=OneWay}""/>
            <tk:DataGridTextColumn Header=""Value""
                                   Binding=""{Binding Value,
                                                     Mode=OneWay}""/>
        </tk:DataGrid.Columns>
    </tk:DataGrid>

```

When I set `MyDGSampleWindow.MySelItem` as null, the datagrid propertly deselects. Perhaps you might need to give us more input on how are you actually setting the value as null.


"
"How to read a slot from an object reference, without evaluating the pointer
I wanted to get the value of an attribute from a pointer, inside another object, but accessing it without evaluating the reference got me the error



```
When attempting to read the slot's value (slot-value), the slot
POS is missing from the object *NODE-1*.

```

Here is a piece of code to simulate the error:



```
(defclass node ()
  ((pos 
    :initarg :pos
    :initform '(0 0)
    :accessor pos)))

(defclass edge ()
  ((vertices
    :initarg :vertices
    :accessor vertices)))

(defparameter *node-1* (make-instance 'node))
(defparameter *node-2* (make-instance 'node :pos '(100 100)))
(defparameter *edge-1* (make-instance 'edge :vertices '(*node-1* *node-2*)))

```

After that, evaluating this expression throws an error



```
(slot-value (car (slot-value *edge-1* 'vertices)) 'pos)

```

But this one has the desired behavior



```
(slot-value (eval (car (slot-value *edge-1* 'vertices))) 'pos)

```

I already know that `eval` is used for ugly hacks, that's why I'm trying to find a clever way to do what I need.


","
> 
> When attempting to read the slot's value (slot-value), the slot
>  `POS` is missing from the object `*NODE-1*`
> 
> 
> 


`*node-1*` is not a CLOS instance. It's a symbol. `slot-value` needs a CLOS instance. Thus trying to compute the slot-value of a symbol makes no sense.


**Side note: `objects` in Common Lisp**


In the error message, the term *object* `*NODE-1*` means a specific symbol object at runtime. Symbols are objects, too. They are not CLOS objects, that is they are not CLOS instances of a CLOS class. But generally, even a symbol or a string are in Common Lisp considered to be *objects*.


**Why is it a symbol?**


You set the `vertices` slot to be the value of `'(*node-1* *node-2*)`. That's a literal list of two symbols, since you have quoted the list.



```
CL-USER 152 > '(*node-1* *node-2*)
(*NODE-1* *NODE-2*)

CL-USER 153 > (mapcar #'type-of *)
(SYMBOL SYMBOL)

```

**One uses objects directly**


If you want to compute a list of actual values of those symbols as variables, you need to compute a list of computed values:



```
CL-USER 154 > '((+ 1 2))
((+ 1 2))

CL-USER 155 > (list (+ 1 2))
(3)

CL-USER 160 > (list *node-1* *node-2*)
(#<NODE 402032A2C3> #<NODE 402032B98B>)

CL-USER 161 > (mapcar #'type-of *)
(NODE NODE)

```

The function `list` creates a fresh list with its arguments as contents.


**Getting the value of a symbol: use `SYMBOL-VALUE`**



```
(symbol-value '*node-1*)

```

"
"Obtain field (td) values by clicking on table row
I need your help.


I have the following table as detailed below. What I would like to do is to click on one of the rows in the table (headers excluded) and obtain the column values of the selected row.


How do I populate the textboxes with the information from the selected row in the table?


No libraries please.



```
<!DOCTYPE html>

<html>

<head>

</head>

<body>


<table id=""myTable"" cellspacing=""1"">             
    <thead>> 
        <tr> 
            <th>first name</th> 
            <th>last name</th> 
            <th>age</th> 
            <th>total</th> 
            <th>discount</th> 
            <th>diff</th> 
        </tr> 
    </thead> 
    <tbody> 
        <tr> 
            <td>peter</td> 
            <td>parker</td> 
            <td>28</td> 
            <td>9.99</td> 
            <td>20.3%</td> 
            <td>+3</td> 
        </tr> 
        <tr> 
            <td>john</td> 
            <td>hood</td> 
            <td>33</td> 
            <td>19.99</td> 
            <td>25.1%</td> 
            <td>-7</td> 
        </tr> 
        <tr> 
            <td>clark</td> 
            <td>kent</td> 
            <td>18</td> 
            <td>15.89</td> 
            <td>44.2%</td> 
            <td>-15</td> 
        </tr> 
        <tr> 
            <td>bruce</td> 
            <td>almighty</td> 
            <td>45</td> 
            <td>153.19</td> 
            <td>44%</td> 
            <td>+19</td> 
        </tr> 
        <tr> 
            <td>bruce</td> 
            <td>evans</td> 
            <td>56</td> 
            <td>153.19</td> 
            <td>23%</td> 
            <td>+9</td> 
        </tr> 
    </tbody> 
</table>

Firstname is:<input type=""text"" id=""firstname"" />
<br>
Lastname is:<input type=""text"" id=""lastname"" />
<br>
Age is:<input type=""text"" id=""age"" />
<br>
Total is:<input type=""text"" id=""total"" />
<br>
Discount is:<input type=""text"" id=""discount"" />
<br>
Diff is:<input type=""text"" id=""diff"" />
</body>

</html>

```

","Here is one example how to use [event delegation](https://stackoverflow.com/questions/1687296/what-is-dom-event-delegation) in javascript (it's especially convenient for dynamically added content) . The event handler is attached to TABLE element's onclick attribute:



```
function run(){
    document.getElementById('myTable').onclick = function(event){
        event = event || window.event; //for IE8 backward compatibility
        var target = event.target || event.srcElement; //for IE8 backward compatibility
        while (target && target.nodeName != 'TR') {
            target = target.parentElement;
        }
        var cells = target.cells; //cells collection
        //var cells = target.getElementsByTagName('td'); //alternative
        if (!cells.length || target.parentNode.nodeName == 'THEAD') { // if clicked row is within thead
            return;
        }
        var f1 = document.getElementById('firstname');
        var f2 = document.getElementById('lastname');
        var f3 = document.getElementById('age');
        var f4 = document.getElementById('total');
        var f5 = document.getElementById('discount');
        var f6 = document.getElementById('diff');
        f1.value = cells[0].innerHTML;
        f2.value = cells[1].innerHTML;
        f3.value = cells[2].innerHTML;
        f4.value = cells[3].innerHTML;
        f5.value = cells[4].innerHTML;
        f6.value = cells[5].innerHTML;
    }
}

```

JSFIDDLE: [code](http://jsfiddle.net/Xp54u/2/) / [result page](http://jsfiddle.net/Xp54u/2/show)


The alternative solution, where the event handler is directly attached to each TR element can look like this: 



```
function run() {
    var t = document.getElementById('myTable');
    var rows = t.rows; //rows collection - https://developer.mozilla.org/en-US/docs/Web/API/HTMLTableElement
    for (var i=0; i<rows.length; i++) {
        rows[i].onclick = function () {
            if (this.parentNode.nodeName == 'THEAD') {
                return;
            }
            var cells = this.cells; //cells collection
            var f1 = document.getElementById('firstname');
            var f2 = document.getElementById('lastname');
            var f3 = document.getElementById('age');
            var f4 = document.getElementById('total');
            var f5 = document.getElementById('discount');
            var f6 = document.getElementById('diff');
            f1.value = cells[0].innerHTML;
            f2.value = cells[1].innerHTML;
            f3.value = cells[2].innerHTML;
            f4.value = cells[3].innerHTML;
            f5.value = cells[4].innerHTML;
            f6.value = cells[5].innerHTML;
        };
    }
}

```

JSFIDDLE: [code](http://jsfiddle.net/Xp54u/3/) / [result page](http://jsfiddle.net/Xp54u/3/show)


"
"Separate documentation from code
I don't know if it's possible but I'm wondering if there is a way to keep the code and the documentation in separate files but still work the same as it does normally with inline documentation.


","Yes, you can keep XML documentation comments in external files and include them in your code files using the `<include>` tag.


From [the MSDN documentation](http://msdn.microsoft.com/en-us/library/9h8dy30z.aspx):



> 
> The tag lets you refer to comments in another file that describe the types and members in your source code. This is an alternative to placing documentation comments directly in your source code file. By putting the documentation in a separate file, you can apply source control to the documentation separately from the source code. One person can have the source code file checked out and someone else can have the documentation file checked out.
> 
> 
> 


For example, you might have a file named `xml_include_tag.doc`, containing the following documentation comments:



```
<MyDocs>

    <MyMembers name=""test"">
        <summary>
            The summary for this type.
        </summary>
    </MyMembers>

    <MyMembers name=""test2"">
        <summary>
            The summary for this other type.
        </summary>
    </MyMembers>

</MyDocs>

```

And you would include this documentation in your code file like so:



```
/// <include file='xml_include_tag.doc' path='MyDocs/MyMembers[@name=""test""]/*' />
class Test
{
    static void Main()
    {
    }
}

/// <include file='xml_include_tag.doc' path='MyDocs/MyMembers[@name=""test2""]/*' />
class Test2
{
    public void Test()
    {
    }
}

```

"
"Flutter & AndroidX incompatibility How to set a dependency manually
I'm getting error(s) while compiling due to AndroidX incompatibility:



> 
> 
> > 
> > Android dependency 'androidx.vectordrawable:vectordrawable' has different version for the compile (1.0.0) and runtime (1.0.1) classpath. You should manually set the same version via DependencyResolution
> > 
> > 
> > 
> 
> 
> 


following [-> this post <-](https://stackoverflow.com/questions/49295415/you-should-manually-set-the-same-version-via-dependencyresolution) I've added some code to build.gradle



```
allprojects {

configurations.all {
    resolutionStrategy.force""androidx.vectordrawable:vectordrawable:1.0.0"",
}
repositories {
    google()
    jcenter()
}

```

next run gave me this other error



> 
> 
> > 
> > Android dependency 'androidx.core:core' has different version for the compile (1.0.0) and runtime (1.0.1) classpath. You should manually set the same version via DependencyResolution
> > 
> > 
> > 
> 
> 
> 


I tried to add this 



> 
> ""androidx.vectordrawable:vectordrawable:1.0.0"",""androidx.core:core:1.0.0"",
> 
> 
> 


but I'm probably doing it wrong, since I get the classical ""unexpected bla bla bla""


any suggestion?


thanks in advance


[edit] i've tried this old trick as well, but didn't work
(also downgrading the packages as required [HERE](https://flutter.dev/docs/development/packages-and-plugins/androidx-compatibility#avoiding-androidx))



```
rootProject.allprojects {
    project.configurations.all {
        resolutionStrategy.eachDependency { details ->
            if (details.requested.group == 'androidx.core') {
                details.useVersion ""1.0.1""
            }
            if (details.requested.group == 'androidx.lifecycle') {
                details.useVersion ""2.0.0""
            }
            if (details.requested.group == 'androidx.versionedparcelable') {
                details.useVersion ""1.0.0""
            }
        }
    }
}

```

now returns a different error 



> 
> 
> > 
> > Android dependency 'androidx.appcompat:appcompat' has different version for the compile (1.0.0) and runtime (1.0.2) classpath. You should manually set the same version via DependencyResolution
> > 
> > 
> > 
> 
> 
> 


","1. In android/gradle/wrapper/gradle-wrapper.properties change the line starting with distributionUrl like this: distributionUrl=https://services.gradle.org/distributions/gradle-4.10.2-all.zip


2.In android/build.gradle, replace:



```
dependencies {
    classpath 'com.android.tools.build:gradle:3.2.1'
}

```

by



```
dependencies {
  classpath 'com.android.tools.build:gradle:3.3.0'
}

```

3.In android/gradle.properties, append



```
android.enableJetifier=true
android.useAndroidX=true

```

4.In android/app/build.gradle:


Under android {, make sure compileSdkVersion and targetSdkVersion are at least 28.


5.Replace all deprecated libraries with the AndroidX equivalents. For instance, if you’re using the default .gradle files make the following changes:


In android/app/build.gradle



```
testInstrumentationRunner ""android.support.test.runner.AndroidJUnitRunner""

```

by



```
testInstrumentationRunner ""androidx.test.runner.AndroidJUnitRunner""

```

Finally, under dependencies replace



```
androidTestImplementation 'com.android.support.test:runner:1.0.2'
androidTestImplementation 'com.android.support.test.espresso:espresso-core:3.0.2'

```

by



```
androidTestImplementation 'androidx.test:runner:1.1.1'
androidTestImplementation 'androidx.test.espresso:espresso-core:3.1.1'

```

"
"Error: Excepiton in System.Threading.ThreadAbortException: Thread was being aborted
While downloading template i m getting following error message.


I have tried instead of Response.Flush(); with Response.End();. But getting same error. 



```
Error: Excepiton in Download:System.Threading.ThreadAbortException: Thread was being aborted.
at System.Threading.Thread.AbortInternal()
at System.Threading.Thread.Abort(Object stateInfo)
at System.Web.HttpResponse.End()

```

Any idea to avoid above exception


Code 



```
private void DownloadFile(string filePath, string downloadFileName)
{
    Response.ContentType = ""application/ms-excel"";
    Response.AddHeader(""content-disposition"", ""attachment; filename="" + downloadFileName);
    Response.TransmitFile(filePath);
    // Response.Flush();
    Response.End();
}

```

Thanks in advance..


","As answered here :- [How to Avoid Response.End() ""Thread was being aborted"" Exception during the Excel file download](https://stackoverflow.com/questions/20988445/how-to-avoid-response-end-thread-was-being-aborted-exception-during-the-exce)



> 
> Replace this : `HttpContext.Current.Response.End();`
> 
> 
> With this :
> 
> 
> 



```
HttpContext.Current.Response.Flush(); // Sends all currently buffered output to the client.
HttpContext.Current.Response.SuppressContent = true;  // Gets or sets a value indicating whether to send HTTP content to the client.
HttpContext.Current.ApplicationInstance.CompleteRequest(); // Causes ASP.NET to bypass all events and filtering in the HTTP pipeline**

```


> 
> chain of execution and directly execute the EndRequest event.
> 
> 
> 


And answered here :- [ASP.NET exception ""Thread was being aborted"" causes method to exit](https://stackoverflow.com/questions/1014439/asp-net-exception-thread-was-being-aborted-causes-method-to-exit)



> 
> This is a ThreadAbortException; it's a special exception that is
>  automatically rethrown at the end of every catch block, unless you
>  call Thread.ResetAbort().
> 
> 
> ASP .Net methods like Response.End or Response.Redirect (unless you
>  pass false) throw this exception to end processing of the current
>  page; your someFunctionCall() is probably calling one of those
>  methods.
> 
> 
> ASP .Net itself handles this exception and calls ResetAbort to
>  continue processing.
> 
> 
> 


"
"Is everything from imported go modules included in my application?
Is everything from the imported Go packages included in my application or is the unused code discarded? Something like Javascript's 'treeshaking'?


Eg. If I have created for myself utility functions dealing with Gin and Gorm (or with some other 3rd party functionality) - would it make sense to package my utility functions also in separate libraries or just bundle them in one?


","""Tree-shaking"" is for JavaScript, where you're delivering the code to the runtime in source form, and removing unused bits of the source requires active effort (and doing so in a way that the result still *works* requires some smart analysis).


In compiled languages, especially when using static linking (which is nearly always the case for Go), not linking in unused functions is just that... if nothing ever refers to a function, then there's no reason to emit it. It's pretty much default behavior. The linker has to know where all of the calls are to insert the correct addresses, so it doesn't have to do significant extra work to notice when a function is never called.



> 
> Eg. If I have created for myself utility functions dealing with Gin and Gorm - would it make sense to have these functions in 2 separate libraries? Or is it completely harmless to have them in 1 library and have the Gorm utility functions (and therefore Gorm dependencies) imported even if my application has no DB part?
> 
> 
> 


Putting them in one library is a silly idea just from a dependencies point of view. If a program uses your library, and never uses gorm, then why should it require gorm in its build process? Why should a failure to fetch or build gorm have the potential to break the build of a program that never uses it? You're not paying a runtime or binary-size cost, you're paying a maintainability cost.


"
"Objective-c How to do validation on textField
I would like to prevent the user from not entering any data as well as entering only spaces. So basically there must at least be one character without a space. Then I would also like to remove any spaces at the beginning of the word so the first letter is a character that is not a space.


**edit**


the user must enter something and if the user enters a few spaces before it then I want to trim those spaces. I also want to prevent the user from just entering spaces.


**Example**


if the user enter's a name and surname like "" James Dean"" I would like to take the first space away but not the second space between James and Dean.


","Set your `UIViewController` to be your destination `UITextField`'s `delegate` and implement this method:



```
- (BOOL)textField:(UITextField *)textField shouldChangeCharactersInRange:(NSRange)range replacementString:(NSString *)string {

  // verify the text field you wanna validate
  if (textField == _nameTextField) {

    // do not allow the first character to be space | do not allow more than one space
    if ([string isEqualToString:@"" ""]) {
      if (!textField.text.length)
        return NO;
      if ([[textField.text stringByReplacingCharactersInRange:range withString:string] rangeOfString:@""  ""].length)
        return NO;
    }

    // allow backspace
    if ([textField.text stringByReplacingCharactersInRange:range withString:string].length < textField.text.length) {
      return YES;
    }

    // in case you need to limit the max number of characters
    if ([textField.text stringByReplacingCharactersInRange:range withString:string].length > 30) {
      return NO;
    }

    // limit the input to only the stuff in this character set, so no emoji or cirylic or any other insane characters
    NSCharacterSet *set = [NSCharacterSet characterSetWithCharactersInString:@""abcdefghijklmnopqrstuvwxyzABCDEFGHIJKLMNOPQRSTUVWXYZ1234567890 ""];

    if ([string rangeOfCharacterFromSet:set].location == NSNotFound) {
      return NO;
    }

  }

  return YES;
}

```

"
"How can I unshorten a URL?
I want to be able to take a shortened or non-shortened URL and return its un-shortened form. How can I make a python program to do this?


Additional Clarification:


- Case 1: shortened --> unshortened
- Case 2: unshortened --> unshortened


e.g. `bit.ly/silly` in the input array should be `google.com` in the output array   

e.g. `google.com` in the input array should be `google.com` in the output array


","Send an HTTP HEAD request to the URL and look at the response code. If the code is 30x, look at the `Location` header to get the unshortened URL. Otherwise, if the code is 20x, then the URL is not redirected; you probably also want to handle error codes (4xx and 5xx) in some fashion. For example:



```
# This is for Py2k.  For Py3k, use http.client and urllib.parse instead, and
# use // instead of / for the division
import httplib
import urlparse

def unshorten_url(url):
    parsed = urlparse.urlparse(url)
    h = httplib.HTTPConnection(parsed.netloc)
    h.request('HEAD', parsed.path)
    response = h.getresponse()
    if response.status/100 == 3 and response.getheader('Location'):
        return response.getheader('Location')
    else:
        return url

```

"
"How to force the update of a gem with bundler?
I have a private git server hosting a gem we developed. The gem got some commits, but the version did not actually change. 


How can I force bundler to update my gem even if the version hasn't changed?


I tried ""bundler update mygemname"" but it did not update anything.


Thanks


","that will not work - there is no ""force"" option - you will have to modify your .gemspec file and increase the version number, then do `gem build ...`, and `bundle install`


It is critical for bundler to be able to read the version number from your gem, which was introduced in the `.gemspec` file. It's confusing not only to bundler or `gem update` , but also confusing to people if you forget to update the version number in the .gemspec file. They would end up having gem-files lying around, and not be able to tell which versions they are, e.g. which one is newer? (of course you could use md5-sum, but that's beside the point here).


The best thing to do is to correct the mistake in the .gemspec file, and re-release the gem.


Check the bundler source code for available options:


e.g.: `bundler-1.0.15/lib/bundler/cli.rb`


(search for `desc ""install""`)


"
"Why is this not a constant expression?
In this trivial example, `test2` fails to compile even though `test1` succeeds, and I don't see why that is the case. If `arr[i]` is suitable for a return value from a function marked `constexpr` then why can it not be used as a non-type template argument?



```
template<char c>
struct t
{ 
    static const char value = c;
};

template <unsigned N>
constexpr char test1(const char (&arr)[N], unsigned i)
{
    return arr[i];
}

template <unsigned N>
constexpr char test2(const char (&arr)[N], unsigned i)
{
    return t<arr[i]>::value;
}

int main()
{
    char a = test1(""Test"", 0); //Compiles OK
    char b = test2(""Test"", 0); //error: non-type template argument 
                               //is not a constant expression
}

```

Edit: This makes no difference:



```
template<char c>
struct t
{ 
    static const char value = c;
};

template <unsigned N>
constexpr char test1(const char (&arr)[N])
{
    return arr[0];
}

template <unsigned N>
constexpr char test2(const char (&arr)[N])
{
    return t<arr[0]>::value;
}

int main()
{
    char a = test1(""Test""); //Compiles OK
    char b = test2(""Test""); //error: non-type template argument 
                            //is not a constant expression
}

```

","Short answer: there are no `constexpr` function parameters in `C++11/14`.


Longer answer: in `test1()`, if `i` is not a compile-time constant, the function is still usable at run-time. But in `test2()`, it cannot be known to the compiler whether `i` is a compile-time constant, and yet it is required for the function to compile.


E.g. the following code for `test1` will compile



```
int i = 0;    
char a = test1(""Test"", i); // OK, runtime invocation of test1()

constexpr int i = 0;
constexpr char a = test1(""Test"", i); // also OK, compile time invocation of test1()

```

Let's simply your `test2()` to 



```
constexpr char test3(unsigned i)
{
    return t<i>::value;
}

```

This will not compile for `test3(0)` because inside `test3()`, it cannot be proven that `i` is an **unconditional** compile-time expression. You would need `constexpr` function parameters to be able to express that.


## Quote from the Standard


**5.19 Constant expressions [expr.const]**



> 
> 2 A conditional-expression e is a core constant expression unless the
>  evaluation of e, following the rules of the abstract machine (1.9),
>  would evaluate one of the following expressions:
> 
> 
> — an id-expression that refers to a variable or data member of
>  reference type unless the reference has a preceding initialization and
>  either  
> 
>  — it is initialized with a constant expression or
> 
> 
> — it is a non-static data member of an object whose lifetime began within the evaluation of e;
> 
> 
> 


This section has the following code example corresponding to your question:



```
constexpr int f1(int k) {
    constexpr int x = k; // error: x is not initialized by a
                         // constant expression because lifetime of k
                         // began outside the initializer of x
    return x;
}

```

Because `x` in the above example is not a constant expression, it means that you can't instantiate templates with either `x` or `k` inside `f1`.


"
"C# Epoch time to date and timezone convertion
what is good way to convert Epoch time to date, but using specific timezone


I need to check two dates. One date i have in database and another in x website.


I know that both dates are exactly the same in relation of yyyy-MM-dd


**in example:**


**Database date is:** 1398891600 = 2014-05-01


**X Website date is:** 1398902400000 = 2014-05-01


C# code ""converter""



```
private string DateChecker_inEPOCH(string epochdate)
        {
            if (epochdate.Length < 13)
            {
                epochdate = epochdate + ""000"";
            }
            DateTime dt = new DateTime(1970, 1, 1, 0, 0, 0).AddMilliseconds(Convert.ToDouble(epochdate));

            string dateFormated = dt.ToString(""yyyy-MM-dd"");

            return dateFormated;

        }

```

Issue is that when i convert using my method one db date is 2014-04-30 and website date is 2014-05-01.


I tested dates using <http://www.epochconverter.com/>


","
> 
> Database date is: 1398891600 = 2014-05-01
> 
> 
> X Website date is: 1398902400000 = 2014-05-01
> 
> 
> 


EpochConverter.com gives me the following results:


1398891600000 -> Wed, 30 Apr 2014 21:00:00 GMT


1398902400000 -> Thu, 01 May 2014 00:00:00 GMT


So your code seems to be consistent with EpochConverter.com.


With regards to timezones, I would make sure all dates are stored/retrieved using UTC. This will save you a lot of headaches when it comes to converting to/from different timezones. In this case, assuming the epoch was stored relative to a UTC date, you would simply do the following:



```
DateTime dt = new DateTime(1970, 1, 1, 0, 0, 0, DateTimeKind.Utc).AddMilliseconds(Convert.ToDouble(epochdate));

```

You can then use the following if you need to display a local time to the end user:



```
utcDate.ToLocalTime();

```

"
"regex not removing underscore from pattern
I was trying to code for allowing certain special characters in a string by using java.util.regex.Matcher and java.util.regex.pattern but this is not removing underscore from the same. I'm new here. I need help on this.
Code extract below:



```
  // String to be scanned to find the pattern.
  String line = ""This order was _:$ placed for QT3000! OK?"";
  String pattern = ""[^\\w\\s\\-?:().,'+\\/]"";
  String s = null;

  // Create a Pattern object
  Pattern r = Pattern.compile(pattern);

  // Now create matcher object.
  Matcher m = r.matcher(line);
  s= m.replaceAll("""");
  System.out.println(""Output: "" + s);

```

Expected: This order was : placed for QT3000 OK?
Actual : This order was \_: placed for QT3000 OK?


","The `\w` pattern matches underscores and `[^\w]` matches any char but letters, digits and an underscore. 


Replace with `\p{Alnum}`:



```
String pattern = ""[^\\p{Alnum}\\s?:().,'+/-]"";

```

Note I put the hyphen at the end of the character class so as not to escape it and remove the escaping `\` from the `/` as it is not a special regex metacharacter.


See the [Java regex demo](https://www.regexplanet.com/share/index.html?share=yyyyu0f0uyr).


The `[^\\p{Alnum}\\s?:().,'+/-]` pattern matches any char but:


- `\p{Alnum}` - alphanumeric `[a-zA-Z0-9]`
- `\s` - whitespaces
- `?` - a question mark
- `:` - a colon
- `(` - a `(` symbol
- `)` - a `)` symbol
- `.` - a dot
- `,` - a comma
- `'` - a single quotation mark
- `+` - a plus
- `/` - a forward slash
- `-` - a hyphen.


"
"Using PHP 5.5's password\_hash and password\_verify function
Say I wanted to store a password for a user, would this be the right way to do it with PHP 5.5's `password_hash()` function (or this version for PHP 5.3.7+: <https://github.com/ircmaxell/password_compat>)?



```
$options = array(""cost"" => 10, ""salt"" => uniqid());
$hash = password_hash($password, PASSWORD_BCRYPT, $options);

```

Then I would do:



```
mysql_query(""INSERT INTO users(username,password, salt) VALUES($username, $hash, "" . $options['salt']);

```

To insert into database.


Then to verify:



```
$row = mysql_fetch_assoc(mysql_query(""SELECT salt FROM users WHERE id=$userid""));
$salt = $row[""salt""];
$hash = password_hash($password, PASSWORD_BCRYPT, array(""cost"" => 10, ""salt"" => $salt));

if (password_verify($password, $hash) {
    // Verified
}

```

","Ignoring the issues with your database statements for now, I'll answer the question regarding `password_hash`.


In short, no, that is not how you do it. You do not want to store the salt alone, you should be storing both the hash and salt, and then using both to verify the password. `password_hash` returns a string containing both.


The `password_hash` function returns a string that contains both the hash and the salt. So:



```
$hashAndSalt = password_hash($password, PASSWORD_BCRYPT);
// Insert $hashAndSalt into database against user

```

Then to verify:



```
// Fetch hash+salt from database, place in $hashAndSalt variable
// and then to verify $password:
if (password_verify($password, $hashAndSalt)) {
   // Verified
}

```

Additionally, as the comments suggest, if you're interested in security you may want to look at `mysqli` (`ext/mysql` is deprecated in PHP5.5), and also this article on SQL injection: <http://php.net/manual/en/security.database.sql-injection.php>


"
"What is connections per host mongodb?
I followed [this](https://www.youtube.com/watch?v=jLUczMkXudc) tutorial and there is configuration **connections per host**.


What is this?


","connectionsPerHost are the amount of physical connections a single Mongo client instance (it's singleton so you usually have one per application) can establish to a mongod/mongos process. At time of writing the java driver will establish this amount of connections eventually even if the actual query throughput is low (in order words you will see the ""conn"" statistic in mongostat rise until it hits this number per app server).


There is no need to set this higher than 100 in most cases but this setting is one of those ""test it and see"" things. Do note that you will have to make sure you set this low enough so that the total amount of connections to your server do not exceed
Found here [How to configure MongoDB Java driver MongoOptions for production use?](https://stackoverflow.com/questions/6520439/how-to-configure-mongodb-java-driver-mongooptions-for-production-use)


"
"whats the difference between @spawn fetch and @sync @sync in julia
to do multiple work asyncronously (sorting vectors, and function calculation mainly (where the calculations is compute or memory bound,at the moment, i can write those operations in the following ways:


1. using `Threads.@spawn`



```
_f1 = Threads.@spawn f1(x)
_f2 = Threads.@spawn f2(x)
_f3 = Threads.@spawn f3(x)
y1 = fetch(_f1)
y2 = fetch(_f2)
y3 = fetch(_f3)

```

There is also this pattern (seems cleaner):



```
@sync begin
  @async y1 = f1(x)
  @async y2 = f2(x)
  @async y3 = f3(x)
end

```

for this particular use case, which one is preferred?


","Julia has the following support for parallelization (also see <https://docs.julialang.org/en/v1/manual/parallel-computing/>)


1. generating SIMD assembly code using `@simd` and `@inbounds` macros (read <https://docs.julialang.org/en/v1/base/simd-types/> for more details)
2. Green threading (co-routines) - these are not actual threads but allow one task to wait for another, typically where the other tasks do not consume CPU on the current thread. Good examples include waiting for intensive I/O operations or orchestration of distributed processes. Green threads are very light and there can be thousands of them but they all are executed within a single (calling) system thread.
3. Threads via the `Threads` module. This allows to spawn actual system threads. The advantage (compared to the next scenario in this list) is that all threads share the same process memory.
4. Multiprocessing/distributed computing via `Distributed` module. The nice thing here is that you can use the same API moving from a single machine to a cluster. In any HPC computing scenario this is the first choice to consider
5. GPU computing


In your post you are considering green threads vs real threads. The rule is simple:


- if your functions `f1`, `f2`, `f3` do mostly I/O and are not CPU intensive (e.g. download files from internet) or wait for other processes to complete - use green threading (`@sync`/`@async`)
- otherwise use `Threads` (or `Distributed` computing). This is particularly important in case of CPU-intensive jobs where green threads will not give you any performance boost as they utilize a single system thread.


"
"How to add a ""Done"" button with in the DatePicker Through StoryBoard?
In my app i want to display UIDatePicker when user click on button.and that date save into UITextFiled.I done this things.
My problem is when date picker appears there is no done button,How can add that done button.
Upto now i tried.



```
- (IBAction)pickerAction:(id)sender
{
datePicker.datePickerMode=UIDatePickerModeDate;
datePicker.hidden=NO;
datePicker.date=[NSDate date];
[datePicker addTarget:self action:@selector(TextTitle:) forControlEvents:UIControlEventValueChanged];
[self.view addSubview:datePicker];
NSDateFormatter * df = [[NSDateFormatter alloc] init];
[df setDateFormat:@""M-d-yyyy""];
selectedDate.text=[df stringFromDate:datePicker.date];
} 

-(void)TextTitle:(id)sender
{
NSDateFormatter *df = [[NSDateFormatter alloc] init];
[df setDateFormat:@""M-d-yyyy""];
selectedDate.text = [NSString stringWithFormat:@""%@"",
                      [df stringFromDate:datePicker.date]];
 }

```

How can i add done button with this code.
please help me.


","Answer is 



```
datePicker=[[UIDatePicker alloc]init];
datePicker.datePickerMode=UIDatePickerModeDate;
[TextField1 setInputView:datePicker];

UIToolbar *toolBar=[[UIToolbar alloc]initWithFrame:CGRectMake(0, 0, self.view.frame.size.width, 44)];
[toolBar setTintColor:[UIColor grayColor]];
UIBarButtonItem *doneBtn=[[UIBarButtonItem alloc]initWithTitle:@""Done"" style:UIBarButtonItemStyleBordered target:self action:@selector(ShowSelectedDate)];
UIBarButtonItem *space=[[UIBarButtonItem alloc]initWithBarButtonSystemItem:UIBarButtonSystemItemFlexibleSpace target:nil action:nil];
[toolBar setItems:[NSArray arrayWithObjects:space,doneBtn, nil]];

[TextField1 setInputAccessoryView:toolBar];

```

"
"Using two wireless NICs at the same time on Windows 10 desktop PC
I have read similar questions in the past (i.e. [How do I use my dual band Wi-Fi card to connect to 5GHz and 2.4GHz simultaneously on Windows?](https://superuser.com/questions/774256/how-do-i-use-my-dual-band-wi-fi-card-to-connect-to-5ghz-and-2-4ghz-simultaneousl) and [Connect multiple WiFi networks simultaneously](https://superuser.com/questions/139980/connect-multiple-wifi-networks-simultaneously)), however they don't address my question adequately.


I am attempting to understand if I can use 2 separate wireless cards on my Desktop PC to connect to the same wireless router at the same time, specifically as so:


- WiFi Adapter 1 using 5Ghz 802.11ac to Router
- WiFi Adapter 2 using 2.4Ghz 802.11n to Router (same router)


Currently I have them set to 2 different SSIDs, although that certainly can be changed if needed.


To be clear, I want to use both networks to boost speeds, that's the intent here.


","This is no different than connecting 2 wired Ethernet ports on the same computer to the same network... it requires what is called interface bonding, NIC teaming, or link aggregation, which is only officially supported by Windows Server OS for Microsoft products (it worked in Windows 10 up to build 10240, officially broken now), but this also requires a compatible switch or router that supports some kind of networking teaming or balancing protocol, which many enterprise grade switches and routers do, but virtually no consumer products do, and to my knowledge there is no wireless router or AP out there capable of this itself. Although using 2 switch ports that support link aggregation to two unique APs could potentially do this with a compatible OS, which Windows 10 is not (EDIT: Microsoft has confirmed that Windows 10 should have this capability, but it is broken in all builds of Win10 after 10240, with no ETA on fix at this time). 


Honestly, your best bet is to just use the 5Ghz 802.11ac connection and not bother with the second wireless connection unless it is connecting to a different network, or switch to a hardwired connection. 


I can give multiple links if you wish to confirm all this information, but nitty gritty of this is boils down to this from a Microsoft engineer (Jeffrey Tippet) when discussing this in a [Microsoft's Social discussion channel](https://social.technet.microsoft.com/Forums/office/en-US/92cbe1d3-aa6e-4099-a415-dc6d3e4ba8e4/using-wireless-in-nic-teaming?forum=winserverPN), here is the summary: *The NIC Teaming feature is technically implemented using the MAC layer (layer 2), and the MAC layer is different for each media type. You cannot team WLAN/WWAN NICs because we did not build the extra support for that; it wouldn't make sense to spend time building that support when NIC Teaming is squarely aimed as a server feature.*


"
"Direct-initializing an object inside a condition
It is possible to define and copy-initialize a variable inside the condition of an `if` statement :



```
if(int i = 17) { ... }

```

This also works with user-defined types, given that they overload `operator bool` :



```
if(Foo f = 42)      { ... }
if(Foo f = Foo(43)) { ... }

```

Why can't I use direct-initialization, like the following ?



```
if(Foo f(51)) { ... }

```

GCC emits `error: expected primary-expression before 'f'`.


[Live on Coliru](http://coliru.stacked-crooked.com/a/57e93d0e82350720)


Is there a reason other than ""because the grammar says so"" ? And how can I work around it ?


I'm working with VC++03, where `Foo` :


- is a RAII-sensitive object, for which I took care *not* to define a copy constructor
- is a template taking arguments from the user
- has a two-parameters constructor


... so I'd rather avoid copying it or repeating its type.


Note : Although my actual problem is with C++03, I'm (academically) interested about answers in C++11.


","In C++03, one could solely use the copy-initialization syntax:



> 
> *selection-statement*:   
>         `if (` *condition* `)` *statement*   
>         […]
> 
> 
> *condition*:  
>         *expression*  
>         *type-specifier-seq **declarator*** `=`
> *assignment-expression*
> 
> 
> 


Since C++11, list-initialization was added:



> 
> *condition*:  
> 
>          *expression*  
> 
>         *attribute-specifier-seq*opt *decl-specifier-seq **declarator*** `=` *initializer-clause*  
> 
>         *attribute-specifier-seq*opt *decl-specifier-seq **declarator
>  braced-init-list***
> 
> 
> 


The syntax of direct-initialization, i.e. `Foo f(…)`, was presumably avoided for the same reason it was disallowed for non-static data member initializers: Ambiguities, in particular the ""most vexing parse"". 


"
"How to add moq as a dependency in dotnet Core?
I have the following dependencies in my dotnet core application:



```
 ""dependencies"": {
        ""xunit"": ""2.2.0-beta2-build3300"",
        ""dotnet-test-xunit"": ""2.2.0-preview2-build1029"",
        ""Moq"": ""4.0.10827""
    },

```

And no matter what version i download of Moq its simply just not supported, it says: 



```
Package Moq 4.0.10827 is not compatible with netcoreapp1.0 (.NETCoreApp,Version=v1.0). Package Moq 4.0.10827 supports:
      - net35 (.NETFramework,Version=v3.5)
      - net40 (.NETFramework,Version=v4.0)
      - sl4 (Silverlight,Version=v4.0)
    One or more packages are incompatible with .NETCoreApp,Version=v1.0.

```

But i read on this blogpost: [Moq on .NET Core](http://dotnetliberty.com/index.php/2016/02/22/moq-on-net-core/) that it was possible, i have the nuget plugin in studio code, so it autocompletes packages, i just cannot find any package when i write `moq.netcore` Maybe i am asking more of an approach to finding out if such a plugin actually exists, more than an answer, because right now i can't see on nuget if packages are supported in dotnet Core, how do you guys check if it has support? and do you only look for packages on [Nuget.org](https://www.nuget.org/) ?


Thanks


EDIT: Solution project.json:



```
{
    ""version"": ""1.0.0-*"",
    ""testRunner"": ""xunit"",
    ""dependencies"": {
        ""xunit"": ""2.2.0-beta2-build3300"",
        ""dotnet-test-xunit"": ""2.2.0-preview2-build1029"",
        ""moq"": ""4.6.38-alpha""
    },
    ""frameworks"": {
        ""netcoreapp1.0"": {
            ""dependencies"": {
                ""Microsoft.NETCore.App"": {
                    ""type"": ""platform"",
                    ""version"": ""1.0.0""
                }
            }
        }
    }
}

```

","I guess current stable version in Nuget is 4.5.23 and in your code you have mentioned it should be 4.0.10827, may be that is causing problem.


As shown in your error, Moq 4.0.10827 is not compatible with netcoreapp1.0 and it only supports till 4.0, Look [here](https://www.nuget.org/packages/Moq/4.6.38-alpha) for more details on versions of Moq


I have also wrote blog on Moq in .Net core, which is [here](https://neelbhatt40.wordpress.com/2016/10/15/moq-in-asp-net-core/). But I make changes in this everyday due to new changes.


**Edit:** As per DenLilleMand:



> 
> 4.6.38-alpha works - but e.g. 4.5.3 doesn't work, that complains that Moq 4.5.3 supports net45 and one or more packages are incompatible
>  with .NETCoreApp V1.0.
> 
> 
> 


"
"How load balancer works in RabbitMQ
I am new to RabbitMQ, so please excuse me for trivial questions:


1) In case of clustering in RabbitMQ, if a node fails, load shift to another node (without stopping the other nodes). Similarly, we can also add new fresh nodes to the existing cluster without stopping existing nodes in cluster. Is that correct?


2) Assume that we start with a single rabbitMQ node, and create 100 queues on it. Now producers started sending message at faster rate. To handle this load, we add more nodes and make a cluster. But queues exist on first node only. How does load balanced among nodes now? And if we need to add more queues, on which node we should add them? Or can we add them using load balancer.


Thanks In Advance


","
> 
> 1) In case of clustering in RabbitMQ, if a node fails, load shift to another node (without stopping the other nodes). Similarly, we can also add new fresh nodes to the existing cluster without stopping existing nodes in cluster. Is that correct?
> 
> 
> 


If a node on which the queue was created fails, rabbitmq will elect a new master for that queue in the cluster as long as mirroring for the queue is enabled. Clustering provides HA based on a policy that you can define.



> 
> 2) Assume that we start with a single rabbitMQ node, and create 100 queues on it. Now producers started sending message at faster rate. To handle this load, we add more nodes and make a cluster. But queues exist on first node only. How does load balanced among nodes now? 
> 
> 
> 


The load is not balanced. The distributed cluster provides HA and not load balancing. Your requests will be redirected to the node in the cluster on which the queue resides.



> 
> And if we need to add more queues, on which node we should add them? Or can we add them using load balancer.
> 
> 
> 


That depends on your use case. Some folks use a round robin and create queues on separate nodes. 


### In summary


- For HA use [mirroring](https://www.rabbitmq.com/ha.html) in the cluster.
- To balance load across nodes, use a [LB to distribute across Queues](http://insidethecpu.com/2014/11/17/load-balancing-a-rabbitmq-cluster/comment-page-1/).
- If you'd like to load balance the queue itself take a look at [Federated Queues](https://www.rabbitmq.com/federated-queues.html). They allow you to fetch messages on a downstream queue from an upstream queue.


"
"Check if element is visible in DOM
Is there any way that I can check if an element is visible in pure JS (no jQuery) ?


So, given a DOM element, how can I check if it is visible or not? I tried:



```
window.getComputedStyle(my_element)['display']);

```

but it doesn't seem to be working. I wonder which attributes should I check. It comes to my mind:



```
display !== 'none'
visibility !== 'hidden'

```

Any others that I might be missing?


","According to [this MDN documentation](https://developer.mozilla.org/en-US/docs/Web/API/HTMLElement.offsetParent), an element's `offsetParent` property will return `null` whenever it, or any of its parents, is hidden via the display style property. Just make sure that the element isn't fixed. A script to check this, if you have no `position: fixed;` elements on your page, might look like:



```
// Where el is the DOM element you'd like to test for visibility
function isHidden(el) {
    return (el.offsetParent === null)
}

```

On the other hand, if you *do* have position fixed elements that might get caught in this search, you will sadly (and slowly) have to use [`window.getComputedStyle()`](https://developer.mozilla.org/en-US/docs/Web/API/Window.getComputedStyle). The function in that case might be:



```
// Where el is the DOM element you'd like to test for visibility
function isHidden(el) {
    var style = window.getComputedStyle(el);
    return (style.display === 'none')
}

```

Option #2 is probably a little more straightforward since it accounts for more edge cases, but I bet its a good deal slower, too, so if you have to repeat this operation many times, best to probably avoid it.


"
"Ada: how to solve ""Circular Unit Dependency""?
Suppose I have two records: *Person* and *Animal*. Each record is in a separate package.


Package persons:



```
with animals;
use animals;

package persons is 

    type person is record
     ...
     animalref: animalPOINTER;
     ...
    end record;

    type personPOINTER is access person;

end persons;

```

Package animals:



```
with persons;
use persons;

package animals is 
    type animal is record
     ...
     ownerref:  personPOINTER;
     ...
    end record;

    type animalPOINTER is access animal;

end animals;

```

I have Circular Unit Dependency here, and compiler produces fatal error.


Does anyone have a pattern to address such issue ?


Thanks!


","You need `limited with`, which was introduced to address exactly this problem. See the [Rationale for Ada 2005, section 4.2](http://www.adaic.org/resources/add_content/standards/05rat/html/Rat-4-2.html).


`Animals` and `Persons` are symmetric (my editor has adjusted the layout and casing; I’ve added one record component to each so the demo program, below, can print something):



```
limited with Animals;
package Persons is

   --  One of the few things you can do with an incomplete type, which
   --  is what Animals.Animal is in the limited view of Animals, is to
   --  declare an access to it.
   type AnimalPOINTER is access Animals.Animal;

   type Person is record
      Name : Character;
      Animalref : AnimalPOINTER;
   end record;

end Persons;

limited with Persons;
package Animals is

   type PersonPOINTER is access Persons.Person;

   type Animal is record
      Name : Character;
      Ownerref : PersonPOINTER;
   end record;

end Animals;

```

The demo program has the full view of `Animals` and `Persons`. This example is pretty clumsy; you may be able to organise things better by adding subprograms to `Animals` and `Persons`. Note that the *body* of `Animals` can (and must) `with Persons;` if it needs to use anything in `Persons`.



```
with Ada.Text_IO; use Ada.Text_IO;
with Animals;
with Persons;
procedure Animals_And_Persons is
   A : Persons.animalPOINTER := new Animals.Animal;
   P : Animals.PersonPOINTER := new Persons.Person;
begin
   A.all := (Name => 'a', Ownerref => P);
   P.all := (Name => 'p', Animalref => A);
   Put_Line (P.Name & "" owns "" & P.Animalref.Name);
   Put_Line (A.Name & "" is owned by "" & A.Ownerref.Name);
end Animals_And_Persons;

```

which when compiled and run gives



```
$ ./animals_and_persons 
p owns a
a is owned by p

```

"
"mypy: how to ignore ""missing attribute"" errors in mixins
I cannot get mypy to work properly with mixins: it keeps complaining that my mixins reference missing attributes. Consider this example:



```
class PrintValueMixin:
    """"""A mixin that displays values""""""

    def print_value(self) -> None:
        print(self.value)


class Obj(PrintValueMixin):
    """"""An object that stores values. It needs a mixin to display its values""""""

    def __init__(self, value):
        self.value = value


instance = Obj(1)
instance.print_value()

```

If I run mypy on this file, I get an error:



```
error: ""PrintValueMixin"" has no attribute ""value""

```

Of course it has no attribute ""value"". It is a mixin, it should not have its own attributes!


So how do I make mypy understand this?


","I think this is a sign of an imperfectly designed class hierarchy. Mixins should not depend on stuff that are in classes inheriting them. I know this is against duck typing, but we are in ""static"" typing realm and rules are more strict here.


If you want to get rid of the issue without refactoring the code you can do the following:



```
class PrintValueMixin:
    """"""A mixin that displays values""""""
    value: int   # or whatever type it has

    def print_value(self) -> None:
        print(self.value)                                

```

Now, the error is gone. It's because mypy sees `value` as a class attribute. Mind that it's uninitialized – `value` doesn't have any object bind to it. Thus, this has no real consequence in runtime and you won't use it by mistake.


"
"Remove LCD Stand for Wall Mounting - FSM-270YG
Based on Jeff Atwood's [post on Coding Horror](http://www.codinghorror.com/blog/2012/07/the-ips-lcd-revolution.html), I ordered one of these monitors, and I've been absolutely loving it. However, I recently (i.e. today) took the next step in monitor-y goodness and fastened the sucker to an articulated wall mount.


Unfortunately, I can't figure out how to remove the stand. The flat portion comes off with a single screw, but the leg it fastens to has no apparent method of removing it.


Has anyone figured out a trick for removing these, so they don't just stick out below the screen? Should I remove the screws from the backside of the screen, and look for an internal connection to remove? Or just give up and live with it? (After all, it's a great display, it's floating in the air in front of me, and the stand leg is only a minor annoyance).


","The arm is mounted from the inside by two screws. In order to get to them, you'll need to remove the 4 screws on the back side along the bottom. This will free up the face plate which you can pry off- GENTLY- by slipping a razor blade or flat head screwdriver inside of the seam. There are several tabs where the plastic is locked into place. Too much force and you'll break something.


Remove the front face with the screen facing up. You'll see that the LCD component has a metal bracket kept in place by four tabs which are a part of the outer casing, two on each side. You can use a screwdriver to gently pry them out.


With the monitor free, gently slide it until the metal mounting plate is exposed. Only slide it so that you can access the screws and no more, and further and you might detach or even break the ribbons.


Unscrew the plate. The inside plate will be free as well as the mounting arm underneath.


Gently reassemble: make sure the LCD is snapped into the 4 mounting tabs, and then replace the outer face until all it's tabs are snapped into place and there is no gap in the seam.


Turn over, replace the screws, and you should be good to go.


(I have just completed this process myself without incident or blemish, and the monitor is now mounted to my wall, looking fine.)


"
"C# byte addition operation
Could someone point me, why here:



```
Byte b = 100; 
b = (Byte)(b+200);

```

I have to use explicit type conversion. But here



```
Byte b = 100;
b += 200;

```

I don't need to do this?


Does compiler generate different IL code for this two cases? And which case is better?


","Because the standard permits it (see the second case below):



> 
> ### 14.14.2 Compound assignment
> 
> 
> An operation of the form `x` *op*`=` `y` is processed by applying binary operator overload resolution (§14.2.4) as if the operation was written `x` *op* `y`. Then,
> 
> 
> - If the return type of the selected operator is implicitly convertible to the type of `x`, the operation is evaluated as `x = x` *op* `y`, except that `x` is evaluated only once.
> - Otherwise, if the selected operator is a predefined operator, if the return type of the selected operator is explicitly convertible to the type of `x`, and if `y` is implicitly convertible to the type of `x` or the operator is a shift operator, then the operation is evaluated as `x = (T)(x`  *op* `y)`, where `T` is the type of `x`, except that `x` is evaluated only once.
> - Otherwise, the compound assignment is invalid, and a compile-time error occurs.
> 
> 
> 


The IL code should be essentially identical in this case. Of course, if evaluating `b` has side effects, it will be evaluated twice in the `b = (byte)b + 200` case and only once when using compound assignment.


"
"input textbox hidden behind keyboard on android Chrome
I have a mobile web page of the following format:


header - logo etc - absolute positioned 


content - scrollable, absolute positioned


footer, 40px absolute positioned


The content area has multiple input boxes. On Chrome on Android, 
tapping on an input box at the bottom of the page causes the input
to not be in view when the soft keyboard pops up. The page
does scroll to try and move the box up, but not enough to actually
be visible. It ends up hiding behind the footer. 


Any idea how to fix this? This occurs only on Chrome-Android. Safari on iOS
and IE on Windows Phone and other mobile browsers work just fine.


","I was hoping for a CSS only solutions, but I don't think there is one. Gaunt Face's answer is a good way of doing it. Unfortunately in my case, this would require a few changes and has the possibility of breaking automation among other things (since the url has #targets added to it).


I solved this by changing the position type in 2 click handlers.


I have a click handler for any input/textarea field. In that click handler, I change the position style of the form container div to be static. Note this will push the footer to the bottom if the container field has a scrollbar. This, in my case is not a problem as when the keyboard pops up, only a couple of fields are visible. The user cannot visibly see any difference. 


I have a second handler for when the user clicks out of an input field - page click handler - where I restore the the position type to absolute, making the page appear just as it was before. 


This has one unintended consequence - The scroll position is lost. I fixed this by getting the offset of the input field and calling scrollTop on the parent div with that offset, thus restoring the position. 


I've seen a few questions about this, but did not find an answer. I hope this helps someone. 


"
"What is the best way to load all navigation properties and their children in EF 4.1
I am trying to get a `List<EntityType>`, each item has about 15 navigation properties that I need to load, some of navigations properties have a navingation properties that I need to load too, this is a sample code



```
class AA
{
  public EntityReference<B> Bobj
  {
     get;
     set;
  }
  public EntityCollection<C> CCollection
  {
     get;
     set;
  }
}
class B
{
  public EntityCollection<X> XCollection
  {
     get;
     set;
  }
}
class C
{
  public EntityReference<Y> YObj
  {
     get;
     set;
  }
}

List<AA> AList = Dbcontext.AAs.Where(a => a.ID==4).ToList();

```

I want each item in `List<AA>` has been loaded with its `BObj` with its `XCollection` and and all `CCollection` items with its `YObj` of each.


I tried to load them using `Include` and `Load` but I think there is an **elegant** way to do that 


also I tried to set `LazyLoadingEnabled` to `false` but that does't load the full graph I don't know why?


so please I need the best way to load the full graph of my `EntityType`


I am using **Database first** approach `EF 4.1` and `C# 4.0`


","Here's a neat trick I've picked up:



```
List<AA> AList = Dbcontext.AAs.Where(a => a.ID==4)
    .Select(a => new {a, a.Bobj, a.CCollection})
    .ToList().Select(o => o.a).ToList();

Dbcontext.Bs.Where(b => b.A.ID==4)
    .Select(b => new {b, b.XCollection})
    .ToList();
Dbcontext.Cs.Where(c => c.A.ID==4)
    .Select(c => new {c, c.YObj})
    .ToList();

```

Entity Framework will figure out how these various objects are connected. You'll be making a manageable number of round-trips (3 in this case), and the SQL query results produced will have relatively little repeated data.


"
"When we should use do?
Sometimes the program shows an error if I don't use do. But it runs well without do sometimes. 
ex:



```
countdown ::Int -> IO ()
countdown x =  if x <= 0 
               then putStrLn ""The End.""
               else putStrLn (show (x))

```

runs well 
But 



```
countdown ::Int -> IO ()
countdown x =  if x <= 0 
               then putStrLn ""The End.""
               else putStrLn (show (x)) 
                    countdown (x-1)

```

shows error


","**Short answer:** line breaks don't mean ""next statement"" in Haskell, like they do in many mainstream languages, like Python, Ruby, and recently even JavaScript.




---


**Long answer**


First, let's get rid of the `if`. It only clouds the issue:



```
countdown1 :: Int -> IO ()
countdown1 x = putStrLn (show x)

countdown2 :: Int -> IO ()
countdown2 x = putStrLn (show x)
               countdown (x-1)

```

Notice that the type of your function is `IO ()`. That's the type of what the function must ultimately calculate. Or ""produce"", if you will. This means that whatever is on the right of the equality sign must be of type `IO ()`.



```
countdown x = .........
              ^       ^
              +-------+
                  \
                   the type of whatever is here must be `IO ()`

```

The first version, `countdown1`, does satisfy this, because the expression `putStrLn (show x)` is, indeed, of type `IO ()`


The second version, `countdown2`, on the other hand, looks very strange to the compiler. It looks like you're trying to call `putStrLn`, and you're trying to pass it three parameters: `(show x)`, `countdown`, and `(x-1)`. 


Don't let the newline between `(show x)` and `countdown` confuse you: as mentioned above, newlines don't mean ""next statement"" in Haskell. This is because in Haskell there is no such thing as a ""statement"" in the first place. Everything is an expression.


But wait! If newlines don't count as ""next statement"", then how the hell do I tell Haskell to perform several actions in order? Like, for example, first do `putStrLn`, and then do `countdown`?


Well, this is where monads come in. Monads (of which `IO` is a prime example) were specifically brought into Haskell for this very reason: to express order of things. And the primary operation for that is *""bind""*, which in Haskell exists in the form of an operator `>>=`. This operator takes a monadic value (such as `IO Int` or `IO ()`) on the left, and takes a function that returns a monadic value (such as `Int -> IO String`) on the right, and ""glues"" them together. The result is a new monadic action, which consists of the two input actions executed one after the other.


Applying this to your example with `putStrLn` and `countdown`, it would look like this:



```
putStrLn (show x) >>= (\y -> countdown (x-1))
^               ^     ^                     ^
+---------------+     +---------------------+
    \                        \
     first monadic value      a function that takes the result of the
                              first action as parameter and returns
                              the second action

```

But this is a bit inconvenient. Sure, you can glue together two actions, maybe even three. But after a while this becomes very messy. (at first I had an example here, but then decided to do without; just trust me: it does get messy)


So to relieve the mess, the language now offers syntactic sugar in the form of the `do` notation. Inside the `do` notation, newline *does*, in fact, mean ""next statement"", and these consecutive ""statements"" get desugared into a sequence of calls to the `>>=` operator, giving them the semantics of ""executing in order"". Something like this:



```
do
  y <- f x
  z <- h y            ====>      f x >>= (\y -> h y >>= (\z -> g x y z))
  g x y z

```

So to the naked eye the `do` notation does look roughly equivalent to multiple-line program in Python, Ruby, or JavaScript. But underneath it all, it's still a pure functional program, where everything is an expression, and the order of (non-pure, effectful) operations is explicitly controlled.




---


So, to summarize: you need to use `do` in your program to express the order - first `putStrLn`, and then `countdown`:



```
countdown :: Int -> IO ()
countdown x = if x <= 0 
              then putStrLn ""The End.""
              else do
                     putStrLn (show (x)) 
                     countdown (x-1)

```

But you *don't* need to use `do` when there is just one operation, so there is no order to speak of.


And if you don't want to use `do` for whatever reason, you can desugar it manually into the equivalent `>>=` call:



```
countdown :: Int -> IO ()
countdown x = if x <= 0 
              then putStrLn ""The End.""
              else putStrLn (show (x)) >>= \_ -> countdown (x-1)

```

"
"Unity3d How to detect taps on android?
Currently I am struggling to work out how to create a simple bit of code that will tell me if a tap has happened, allowing me to get the position of taps. Having looked through many different articles I am rather stumped as `Touch.tapCount` that is often suggested only works for ios. Also, i have tried to use `Touch.deltaTime` and/or `Touch.deltaPosition` to detect a tap but have failed.


I define a tap as having


- A very short period between the finger initially touching and finally exiting the phone
- Little or no movement


Thanks for reading this, I hope its clear and precise and if any detail or assistance is required on my behalf in order for your to answer feel free to ask. Any assistance is gratefully received. Thanks.


Note - I work in C#


","First you have to find out if your Android device indeed can register several touches. If you have a newer device, this shouldn't be a problem. I'm going to assume that your device can, and if it can not - you'll soon find out soon enough.


let's start with the update method.



```
void Update() {
// Nothing at the moment
}

```

What we first want to do is register touches. We can do this by putting a foreach inside, checking for touches in Input.touches. Like this:



```
    void Update() {
    foreach (Touch touch in Input.touches) {

    }
}

```

By doing this, we are always checking how many touches there are currently on the screen. What we can now do is check by fingerId, and if fingerId == 0, 1, 2... run some code. Here's what we got now:



```
    void Update() {
        foreach (Touch touch in Input.touches) {

        if (touch.fingerId == 0) {
            // Finger 1 is touching! (remember, we count from 0)
        }

        if (touch.fingerId == 1) {
            // finger 2 is touching! Huzzah!
        }
    }
}

```

We're great so far! What we now want to do is detect the motion we want. In our case, we wanted taps, right? That should work perfectly with TouchPhase Began, and Ended. There's also TouchPhase.Moved, but we don't need that now.



```
if (touch.fingerId == 0) {

    if (Input.GetTouch(0).phase == TouchPhase.Began) {
        Debug.Log(""First finger entered!"");
    }

    if (Input.GetTouch(0).phase == TouchPhase.Ended) {
        Debug.Log(""First finger left."");
    }
}

```

Here we are checking the phase of the corresponding finger. If you run that now, you should be able to see the messages in the console whenever your first touch enters, as well as leaves the screen. This can be done with several touches, so here's the 'whole' script:



```
    void Update() {
    foreach (Touch touch in Input.touches) {

        if (touch.fingerId == 0) {
            if (Input.GetTouch(0).phase == TouchPhase.Began) {
                Debug.Log(""First finger entered!"");
            }
            if (Input.GetTouch(0).phase == TouchPhase.Ended) {
                Debug.Log(""First finger left."");
            }
        }

        if (touch.fingerId == 1) {
            if (Input.GetTouch(1).phase == TouchPhase.Began) {
                Debug.Log(""Second finger entered!"");
            }
            if (Input.GetTouch(1).phase == TouchPhase.Ended) {
                Debug.Log(""Second finger left."");
            }
        }
    }
}

```

I'm hoping this will help you. I'm fairly new at this myself, so if we're lucky - someone with more experience can come and help. I'm confident that this could be written a lot cleaner. Just remember that if you build it, you can't see the console messages. Check out Unity Remote if you haven't already. Good luck! :)


"
"Is there a way to animate changing a UILabel's textAlignment?
I am using iOS 7 and I am trying to move a label that is centered off to the left of my view. Currently I am trying to do this by changing how my UILabel is aligned, and I am trying to animate it in the process. I am currently calling the following:



```
[UIView animateWithDuration:0.5
                 animations:^{
                      self.monthLabel.textAlignment = NSTextAlignmentLeft;
                 } completion:nil];

```

But this just jumps the label to the new alignment. Is there a way to animate this, or a better way to adjust the label to make this happen?


","Set the UILabel frame size to exactly contain the text and center the UILabel in your view. 



```
self.monthLabel.text = @""February"";
[self.monthLabel sizeToFit];
self.monthLabel.center = parentView.center; // You may need to adjust the y position

```

Then set the alignment which should not affect the layout since there will be no extra space.



```
self.monthLabel.textAlignment = NSTextAlignmentLeft;

```

Next, animate the UILabel frame size so it slides over where you want it.



```
[UIView animateWithDuration:0.5
             animations:^{
                  CGRect frame = self.monthLabel.frame;
                  frame.origin.x = 10;
                  self.monthLabel.frame = frame;
             } completion:nil];

```

"
"Flutter - Use Image.Network with cookies
I want to get an avatar image from my Rest api but I have to pass a cookie, otherwise I have a 401 unauthorized.


Is it possible to use something like an Image.Network to do that ?


","These is a header parameter that I missed. 


Here is what I do :



```
 Map<String, String> headersMap = {
  'Cookie' : 'jwt-cookie=eyJhbGciOiJIUzUxMiJ9.eyJqdGkiOiI0IiwiaXNzIjoiMSIsInN1YiI6InRtYSIsImlhdCI6MTU1NjExNTY2MCwiZXhwIjoxNTU2NzIwNDYwfQ.DQMV59lTlGSgVN_viwlUaJIxZNO_Sru0gQT31EnKZEdD533OR9VUCRYaj5pY8ist48zRUmn6HXs4M_oWkkzm7A'
};

```

I set my cookie in a map called headersMap, then I can set my headers in the headers parameter like that :



```
Image.network(baseURL + userService.currentUser.id.toString(), headers: headersMap,)

```

For more customisation I just found this [plugin](https://pub.dartlang.org/packages/flutter_advanced_networkimage)


Thanks to @pskink for the solution.


"
"How to disable url changes with Angular routing
My Angular 7 app needs to be rendered in an iFrame. Due to the peculiar nature of the system, I need to prevent the URL from changing, or maybe I should say that I need Angular not to expect the URL to change.


I've found posts that show how to prevent it on specific routing calls, but the URL is updated when the app loads too.


The objective is to maintain the URL like: website.com/applicaition, even when the route changes such that there is no update to the window.location. 


<https://website.com/application> [desired for every view]
<https://website.com/application/home> [NOT desired for ANY view]


Thanks,
Wayne


","If you don't wish to have the URL updated, at all, you need to do two things.


First, add:



```
{ initialNavigation: false }

```

to the router settings like this:



```
const routes: Routes = [
  { path: 'home', component: HomeComponent },
  { path: '**', component: HomeComponent }
];

@NgModule({
  imports: [RouterModule.forRoot(routes, { initialNavigation: false })],
  exports: [RouterModule]
})

```

Unfortunately, I couldn't find a way to update this globally, so you will need to set the following for each navigation point in the app.


With router-links:



```
<a routerLink=""/home"" routerLinkActive=""active"" skipLocationChange=""true"" >Home</a>

```

With ts navigation calls:



```
router.navigate(['/home'], { skipLocationChange: true });

```

(Ref: <https://angular.io/api/router/NavigationExtras>)


I hope this helps!


"
"Parsing Python Module Docstrings
Is it possible to parse module-level docstrings with the AST?


I am working on a python documenter [here](http://github.com/mvanveen/Tacco) and visiting the module tokens and grabbing the documentation does not yield the module-level docstring. So far, I've had to resort to importing the module and grabbing its `__doc__` or using `inspect` to grab the documentation.


I looked into the [pydoc module](http://docs.python.org/library/pydoc.html) source for clues as to how other documenters parse docstrings, and discovered that pydoc ends up having to do basically the same thing as my documenter in order to grab the module-level strings.


Am I missing something? Is the only way to parse module-level docstrings through actually importing the module, or is it possible to parse the docstrings out of the AST directly?


","Maybe I miss-understand the question, but can't you just do this (python 2.7.1)?


test file:



```
""""""
DOC STRING!!
""""""

def hello():
    'doc string'
    print 'hello'

hello()

```

Interactive session:



```
>>> M = ast.parse(''.join(open('test.py')))
>>> ast.get_docstring(M)
'DOC STRING!!'

```

You can also walk through the ast, looking for the slot the doc string would be in.



```
>>> M._fields
('body',)
>>> M.body
[<_ast.Expr object at 0x10e5ac710>, <_ast.FunctionDef object at 0x10e5ac790>, <_ast.Expr object at 0x10e5ac910>]
>>> # doc would be in the first slot
>>> M.body[0]._fields
('value',)
>>> M.body[0].value
<_ast.Str object at 0x10e5ac750>
>>> # it contains a string object, so maybe it's the doc string
>>> M.body[0].value._fields
('s',)
>>> M.body[0].value.s
'\nDOC STRING!!\n'

```

"
"Foreach returns only one output
Okay so I am stuck here and I know there would be a simple solution, but I can't figure it out. So the thing is that I have a function which creates columns in a database if not exists. I got the function to work but I want to add the functionality in the function of detecting the column type (VARCHAR, INT etc) and creating columns accordingly. I am trying to get this done by `gettype` keyword. Now, `$productData` is in array so I need to apply foreach to get them separated and then I get the value type, the problem is that in the end foreach loop, it gives me only string. Where `$productData` has 4 strings and 1 integer .. 



```
$array       = (array) $productData;
$arrayValues = array_values( $array );
$i = 0;

foreach ( $arrayValues as $key => $value ) {
    $types = array( ""id-$i"" => gettype( $value ) );
    $i++;
}
var_dump($types);

```

Can somebody help me out there please? 


Thanks ..


","You need to initialise the array only once and add elements to it later:



```
$array       = (array) $productData;
$arrayValues = array_values( $array );
$i = 0;

$types = array();
foreach ($arrayValues as $key => $value) {
    $types[""id-$i""] = gettype($value);
    $i++;
}
var_dump($types);

```

Be careful with `gettype()` as it will fail on bojects. You need to make sure that each `$value` is an internal type


**Update:**
The `gettype()` will return `""object""` for all objects. So it is a safe way to determine the type of a variable. Although in case you need the class of the object you need to use `get_class()` in case `gettype()` returns `""object""`


"
"Why does /var/log/wtmp becomes so huge? How to inspect wtmp files?
Just noticed some 640MB wtmp file in a virtual container (Ubuntu Hardy).



```
# last -n 10000 -f /var/log/wtmp.1|wc -l
384
# ls -hl /var/log/wtmp.1
-rw-rw-r-- 1 root utmp 641M 21. Sep 07:49 /var/log/wtmp.1

```

logrotate was not installed (I just did that and forced rotating).


Are there records in there not being displayed by `last` (which should show the last 1000 entries, but apparently there are only 384).


From quickly skimming the `wtmp/utmp` man page, it does not look like a single entry should use about 1,6MB.


Is there another program besides `last` to inspect these files?


","`logrotate` was a good idea. 


Like any regular file, wtmp could have been ""sparse"" (cf. lseek(2) ""holes"" and `ls -s`) which can show a extreme file size that actually occupies little disk. How did the hole get there, if it was a hole? `getty(8)` and friends could have had a bug. Or a system crash and fsck repair could have caused it.


If you are looking to see the raw contents of wtmp, `od` or `hd` are good for peeking at binaries and have the happy side effect of showing long runs of empty as such.


Unless it recurs, I wouldn't give it much more thought. A marginally competent intruder would do a better job than that, the contents aren't all that interesting, and little depends on them. 


"
"Build Python as UCS-4 via pyenv
I run into this issue [ImportError numpy/core/multiarray.so: undefined symbol: PyUnicodeUCS2\_AsASCIIString](https://stackoverflow.com/q/17443166/1391441) installing Python in a [pyenv-virtualenv](https://github.com/yyuu/pyenv-virtualenv) environment.


In my case, it happens with the `matplotlib` package instead of `numpy` (as in the above question), but it's basically the same issue.


The answer given in that question is a simple:



> 
> Rebuild NumPy against a Python built as UCS-4.
> 
> 
> 


I don't know how to do this. In [this other question](https://stackoverflow.com/a/19575808/1391441) it is said that one has to use:



```
./configure --enable-unicode=ucs4

```

but I don't know how to use that command along with [pyenv](https://github.com/yyuu/pyenv).


This issue is also mentioned in `pyenv`'s repo [issue list](https://github.com/yyuu/pyenv/issues/518), and a solution given in [a comment](https://github.com/yyuu/pyenv/issues/518#issuecomment-199827456). Sadly (for me) I can not understand how to apply the fix explained in said comment.


So my question basically is: how do I build Python as UCS-4 via `pyenv`?


","Installing `python` with `pyenv` with `ucs2`:



```
$ export PYTHON_CONFIGURE_OPTS=--enable-unicode=ucs2
$ pyenv install -v 2.7.11
  ...
$ pyenv local 2.7.11
$ pyenv versions
  system
* 2.7.11 (set by /home/nwani/.python-version)
$ /home/nwani/.pyenv/shims/python
Python 2.7.11 (default, Aug 13 2016, 13:42:13) 
[GCC 4.8.4] on linux2
Type ""help"", ""copyright"", ""credits"" or ""license"" for more information.
>>> import sysconfig
>>> sysconfig.get_config_vars()['CONFIG_ARGS']
""'--prefix=/home/nwani/.pyenv/versions/2.7.11' '--enable-unicode=ucs2' '--libdir=/home/nwani/.pyenv/versions/2.7.11/lib' 'LDFLAGS=-L/home/nwani/.pyenv/versions/2.7.11/lib ' 'CPPFLAGS=-I/home/nwani/.pyenv/versions/2.7.11/include '""

```

Installing `python` with `pyenv` with `ucs4`:



```
$ pyenv uninstall 2.7.11
pyenv: remove /home/nwani/.pyenv/versions/2.7.11? y
$ export PYTHON_CONFIGURE_OPTS=--enable-unicode=ucs4
$ pyenv install -v 2.7.11
...
$ pyenv local 2.7.11
$ pyenv versions
  system
* 2.7.11 (set by /home/nwani/.python-version)
$ /home/nwani/.pyenv/shims/python
Python 2.7.11 (default, Aug 13 2016, 13:49:09) 
[GCC 4.8.4] on linux2
Type ""help"", ""copyright"", ""credits"" or ""license"" for more information.
>>> import sysconfig
>>> sysconfig.get_config_vars()['CONFIG_ARGS']
""'--prefix=/home/nwani/.pyenv/versions/2.7.11' '--enable-unicode=ucs4' '--libdir=/home/nwani/.pyenv/versions/2.7.11/lib' 'LDFLAGS=-L/home/nwani/.pyenv/versions/2.7.11/lib ' 'CPPFLAGS=-I/home/nwani/.pyenv/versions/2.7.11/include '""

```

"
"Comparing image on how much degrees tilted
I have an problem as seen below the one on top is the original image the one on the bottom is titlted to the left 90 degrees. Is there a way to calculate the change in angle. I have tried Hough's Transform, I can detect the location, however I cannot detect the angel that is changed. Is there an alternative to detect the degreeschanged.


![enter image description here](https://i.stack.imgur.com/GpPp2.png)


![enter image description here](https://i.stack.imgur.com/tGfkT.png)


","I've done this before using a Symmetric Phase-Only Matched Filter. It was actually for fingerprint recognition, allowing rotation and scaling. Unfortunately, it's pretty complicated. You need to be comfortable with how to compute an FFT and with similar-level math. I did not do it in C#, but in MATLAB (and in synthesizable Verilog, but that's another story altogether). I would recommend using MATLAB or a similar math package first to be sure you have the algorithm right.


Hopefully, though, someone has already implemented this algorithm in a .NET image processing library somewhere.


The paper I used was the following. It is more math-based than algorithmic, so it will take some work to convert it to code:



> 
> Qin-Sheng Chen; Defrise, M.; Deconinck, F.
>  [""Symmetric phase-only
>  matched filtering of Fourier-Mellin
>  transforms for image registration and
>  recognition,""](http://ieeexplore.ieee.org/stamp/stamp.jsp?tp=&arnumber=387491&isnumber=8785)
>  Pattern Analysis and
>  Machine Intelligence, IEEE
>  Transactions on , vol.16, no.12,
>  pp.1156-1168, Dec 1994
> 
> 
> **Abstract**: Presents a new method to
>  match a 2D image to a translated,
>  rotated and scaled reference image.
>  The approach consists of two steps:
>  the calculation of a Fourier-Mellin
>  invariant (FMI) descriptor for each
>  image to be matched, and the matching
>  of the FMI descriptors. The FMI
>  descriptor is translation invariant,
>  and represents rotation and scaling as
>  translations in parameter space. The
>  matching of the FMI descriptors is
>  achieved using symmetric phase-only
>  matched filtering (SPOMF). The
>  performance of the FMI-SPOMF algorithm
>  is the same or similar to that of
>  phase-only matched filtering when
>  dealing with image translations. The
>  significant advantage of the new
>  technique is its capability to match
>  rotated and scaled images accurately
>  and efficiently. The innovation is the
>  application of SPOMF to the FMI
>  descriptors, which guarantees high
>  discriminating power and excellent
>  robustness in the presence of noise.
>  This paper describes the principle of
>  the new method and its discrete
>  implementation for either image
>  detection problems or image
>  registration problems. Practical
>  results are presented for various
>  applications in medical imaging,
>  remote sensing, fingerprint
>  recognition and multiobject
>  identification 
> 
> 
> 


"
"How to customize Swagger endpoint url?
I have multiple web api projects (microservices), and I want to expose them using only one *swagger-ui* link.
I'll call each web api project **EndpointA** and **EndpointB** for the sake of this post.


I've created a *swagger-ui* project, and I've added each endpoint to this project.


My *swagger-ui* project `Startup.cs`



```
public void Configure(IApplicationBuilder app, IHostingEnvironment env)
{
    app.UseSwaggerUI(c => {
        c.SwaggerEndpoint(""/EndpointA/swagger/v1/swagger.json"", ""EndpointA"");
        c.SwaggerEndpoint(""/EndpointB/swagger/v1/swagger.json"", ""EndpointB"");
    });
}

```

My endpoints A/B `Startup.cs`



```
public void ConfigureServices(IServiceCollection services)
{
    services.AddSwaggerGen(c =>
    {
        c.SwaggerDoc(""v1"", new Info
        {
            Version = ""v1"",
            Title = ""EndpointA"", // or ""EndpointB"",
        });
    });
}

```



---


My problem is that each of the `swagger.json` files doesn't contains the endpoint prefix. Instead of being `/EndpointA/Controller/Action`, the paths are `/Controller/Action`, which are not valid.




---


I have tried to set a custom url prefix in my swagger ui project like this `c.RoutePrefix = ""EndpointA"";`. It works for my *EndpointA*, but I can only set one for all endpoints so *EndpointB* won't works.


I also tried to set a prefix directly in each endpoint using `SwaggerGen`, but I don't know how / if it's possible.


","I ended up using the [`UsePathBaseExtensions.UsePathBase(IApplicationBuilder, PathString)` Method](https://learn.microsoft.com/en-us/dotnet/api/microsoft.aspnetcore.builder.usepathbaseextensions.usepathbase?view=aspnetcore-2.2) to register a base path to my api.



```
public void Configure(IApplicationBuilder app, IHostingEnvironment env)
{
    app.UsePathBase(new PathString(""/EndpointA""));
    app.UseMvc();
}

```



---


An alternate solution would have been to use the swagger base path document filter like so:



```
public void ConfigureServices(IServiceCollection services)
{
     services.AddSwaggerGen(c =>
     {
         c.DocumentFilter<BasePathFilter>();
     });
 }

```

And the filter:



```
public class BasePathFilter: IDocumentFilter
{
    public void Apply(SwaggerDocument swaggerDoc, DocumentFilterContext context)
    {
        swaggerDoc.BasePath = ""/EndpointA"";
    }
}

```

"
"What is the use of the View's setTag method?
I am implementing a `ListView` using custom adapters. To improve the performance I am returning the newly inflated `View` if `convertView` is null, else returning the recycled `View`.


Now in this process I used the *ViewHolder* pattern and used the `setTag()` method of the `View`. Initially I thought this is some kind of bookmarking but I've not completely understood the use of it. I've checked different blog post where they just used this in the code.


Can someone please explain me the use of the `setTag()` method?


","Basically you can store any kind of object as tag (and cast it back when calling `getTag`). This can be a simple ID or some complex data. It's some information which you associate with this view.


In the case of lists and the *view holder* pattern it's a simple object which contains references to views of the tagged view (group). So you don't have to call `findViewById` every time when you're updating the content of the view. It's just an performance optimization.


*Can we store data of list item in the view tag?*  

No. Because of view recycling you have (e.g.) 10 views which are reused for 1000 list items. Storing data in the tag makes no sense here. It's better to use an custom data object to store the list item state (probably the same array which contains the displayed data) or you persist it right away on list item change.


See also [setTag documentation](http://developer.android.com/reference/android/view/View.html#setTag%28java.lang.Object%29).


"
"Host multiple asp.net core web application under a single linux server
I am new to asp.net core. I read the whole Microsoft official document and able to host the application in Linux Apache server. But I want to host multiple asp.net core web applications under a single IP Address. Please, anyone have the solution post here.


Thanks in Advance


","The [official document](https://learn.microsoft.com/en-us/aspnet/core/host-and-deploy/linux-apache?view=aspnetcore-2.2#configure-apache) shows us a way to use Apache as a reverse proxy :



```
<VirtualHost *:*>
    RequestHeader set ""X-Forwarded-Proto"" expr=%{REQUEST_SCHEME}
</VirtualHost>

<VirtualHost *:80>
    ProxyPreserveHost On
    ProxyPass / http://127.0.0.1:5000/
    ProxyPassReverse / http://127.0.0.1:5000/
    ServerName www.example.com
    ServerAlias *.example.com
    ErrorLog ${APACHE_LOG_DIR}helloapp-error.log
    CustomLog ${APACHE_LOG_DIR}helloapp-access.log common
</VirtualHost>

```

Basically, this configuration will make Apache listens on `*:80` and proxy any HttpRequest whose `ServerName` equals `www.example.com` to `http://127.0.0.1:5000/`.


This is how Apache used as an proxy to ASP.NET Core works.


As for your question, suppose you have two asp.net core web applications:


1. the first one is called `WebApp1` and listens on `0.0.0.0:5000` .
2. and the other is called `WebApp2` and listens on `0.0.0.0:6000` .


Your Apache server listens on `0.0.0.0:80`. For any incoming http request,


- when `Host` equals `www.webapp1.org`, proxy this request to `0.0.0.0:5000`
- when `Host` equals `www.webapp2.org`, proxy this request to `0.0.0.0:6000`


So you could add two proxies :


proxy 1 :



```
<VirtualHost *:80>
    ProxyPreserveHost On
    ProxyPass / http://127.0.0.1:5000/
    ProxyPassReverse / http://127.0.0.1:5000/
    ServerName www.webapp1.org
    ServerAlias *.webapp1.org
    ErrorLog ${APACHE_LOG_DIR}webapp1-error.log
    CustomLog ${APACHE_LOG_DIR}webapp1-access.log common
</VirtualHost>

```

proxy 2 :



```
<VirtualHost *:80>
    ProxyPreserveHost On
    ProxyPass / http://127.0.0.1:6000/
    ProxyPassReverse / http://127.0.0.1:6000/
    ServerName www.webapp2.org
    ServerAlias *.webapp2.org
    ErrorLog ${APACHE_LOG_DIR}webapp1-error.log
    CustomLog ${APACHE_LOG_DIR}webapp2-access.log common
</VirtualHost>

```

"
"ARC, worth it or not?
When I moved to Objective C (iOS) from C++ (and little Java) I had hard time understanding memory management in iOS. But now all this seems natural and I know retain, autorelease, copy and release stuff. After reading about ARC, I am wondering is there more benefits of using ARC or it is just that you dont have to worry about memory management. Before moving to ARC I wanted to know how worth is moving to ARC.


1. XCode has ""Convert to Objective C ARC"" menu. Is the conversion is that simple (nothing to worry about)?
2. Does it help me in reducing my apps memory foot-print, memory leaks etc (somehow ?)
3. Does it has much testing impact on my apps ?
4. What are non-obvious advantages?
5. Any Disadvantage of moving to it?


","Here's my specific take on ARC:



> 
> 1) XCode has ""Convert to Objective C ARC"" menu. Is the conversion is that simple (nothing to worry about)?
> 
> 
> 


It's simple. It works. Use it. As Kevin Low points out though, you will need to go through and fix up the bits where you use Core Foundation objects. That will just require a healthy lashing of [`__bridge` or `__bridge_transfer`](http://clang.llvm.org/docs/AutomaticReferenceCounting.html#objects.operands.casts) though.



> 
> 2) Does it help me in reducing my apps memory foot-print, memory leaks etc (somehow ?)
> 
> 
> 


Nope, not really. OK, sort of. It will help reduce memory leaks where you have coded incorrectly previously. It won't reduce memory footprint.



> 
> 3) Does it has much testing impact on my apps ?
> 
> 
> 


None whatsoever.



> 
> 4) What are non-obvious advantages?
> 
> 
> 


The future. There'll be more to come on the bonus that the compiler taking an intricate knowledge of how objects are reference counted gives. For example ARC provides the lovely [`objc_retainAutoreleasedReturnValue`](http://www.galloway.me.uk/2012/02/how-does-objc_retainautoreleasedreturnvalue-work/) optimisation already, which is very nice.



> 
> 5) Any Disadvantage os moving to it?
> 
> 
> 


None whatsoever.


Please take my word for it and start using ARC. There's no reason (IMO) not to, thus the advantages definitely out-weigh the disadvantages!


For an in-depth look at how ARC works to perhaps help convince you that it's good, please take a look at my blog posts entitled ""A look under ARC's hood"" - [here](http://www.galloway.me.uk/2012/01/a-look-under-arcs-hood-episode-1/), [here](http://www.galloway.me.uk/2012/01/a-look-under-arcs-hood-episode-2/), [here](http://www.galloway.me.uk/2012/02/a-look-under-arcs-hood-episode-3/) & [here](http://www.galloway.me.uk/2012/02/a-look-under-arcs-hood-episode-4/).


"
"Token Authentication or OAuth 2 for Django Rest Framework API (Or Both?)
I'm setting up a new API using Django REST Framework, and I'm a bit confused how to set up authentication. The API I'm setting up is consumed by the public, whom I want to have the most flexibility possible.


Out of the box, DRF provides Basic Authentication, Session Authentication and Token Authentication. Using another package, you can also add OAuth2 authentication.


If I set up my API to have Token Authentication -- which is included out of the box -- why do I need to set up OAuth2? I've read a bit about this, but since DRF's Token Auth is fairly custom, I haven't been able to find much comparing the two approaches.


From what I can tell, OAuth is just an advanced form of Token Auth, so why bother installing and configuring it?


","I was fleetingly familiar with OAuth2, but had to do a little homework to think this one through. The advantage of OAuth over the token auth that is built in is that OAuth can be negotiated by a computer, provides a distinct token per client, and can easily be revoked.


So, if I set up OAuth as an authentication on my API, that'd let you delegate a third party to use your data. For example, Twitter has OAuth, so you can set up a third party client -- Tweetie or whatever -- with an OAuth token that it can use to access your data. However, if you stop using Tweetie, or you stop trusting Tweetie, you can revoke it's OAuth access without resetting your password and thus revoking access from every client that's configured.


So in my case, I have a mostly read API, and there aren't (yet) resources that are private to particular users. As a result, we have no need for OAuth, since we don't anticipate people needing to give their passwords to third party clients.


"
"Error in mlogit: Error in solve.default(H, g[!fixed]) : system is computationally singular: reciprocal condition number = 3.4767e-18
I have a conjoint survey data in long format. the first few rows look like this:



```
 ID alt choice size tar length brand flavor gender age yr_smoke num_smoke job
 1  1   1     no    1   1      1     1      1      2  35       10        20   1
 2  1   2     no    1   1      1     1      1      2  35       10        20   1
 3  1   3     no    1   1      1     1      1      2  35       10        20   1
 4  1   4     no    1   1      1     1      1      2  35       10        20   1
 5  1   5     no    1   1      1     1      1      2  35       10        20   1
 6  1   6     no    1   1      1     1      1      2  35       10        20   1

```

I used mlogit.data as:



```
data_mlogit_ct1_test2 <- mlogit.data(data_mlogit_ct1_test1,choice=""choice"",
shape=""long"",alt.var=""alt"")

```

The first few rows of the transformed data looks like:



```
   ID alt choice size tar length brand flavor gender age yr_smoke num_smoke job
   1.1  1   1  FALSE    1   1      1     1      1      2  35       10        20   1
   1.2  1   2  FALSE    1   1      1     1      1      2  35       10        20   1
   1.3  1   3  FALSE    1   1      1     1      1      2  35       10        20   1
   1.4  1   4  FALSE    1   1      1     1      1      2  35       10        20   1
   1.5  1   5  FALSE    1   1      1     1      1      2  35       10        20   1
   1.6  1   6  FALSE    1   1      1     1      1      2  35       10        20   1

```

Now I am trying to fit mlogit as:



```
 fit_mlogit_ct11 <- mlogit(choice~size+tar+length+brand+flavor, 
                           data_mlogit_ct1_test1,shape=""long"",chid.var=""ID"",
                           alt.var=""alt"",method=""bfgs"",heterosc=TRUE,tol=10)

```

I got this error:


Error in solve.default(crossprod(attr(x, ""gradi"")[, !fixed])) : 
 Lapack routine dgesv: system is exactly singular


Then I did:



```
  fit_mlogit_ct11 <- mlogit(choice~size+tar+length+brand+flavor, data_mlogit_ct1_test2)

```

Now I got:


Error in solve.default(H, g[!fixed]) : 
 system is computationally singular: reciprocal condition number = 3.4767e-18


Can somebody please help me?


","I think the problem is because you did not define right your individual specific variables. According to the package manual, when you build your models with mlogit function you can use : | to separate the alternative specific with individual specific variables. For example:



```
fit_mlogit_ct11 <- mlogit(choice~1|size+tar+length+brand+flavor, 
                           data_mlogit_ct1_test1,shape=""long"",chid.var=""ID"",
                           alt.var=""alt"",method=""bfgs"",heterosc=TRUE,tol=10)

```

Which simply says that all your variables are individual specific and your take only the intercept for the alternative specific. But indeed as it has already mentioned above your data look odd since they look the same, since you have only individual specific variables. This leads eventually to singularity (your variables correlate with each other with 1). 


"
"Better Way to Parse XML Without Knowing The Schema Beforehand
Using SQL Server 2005 I need to parse all values from XML WITHOUT knowing the element names beforehand. 


Here is what I have:



```
declare @xml xml

--REMEMBER THIS XML COULD CONTAIN ANYTHING. THIS IS AN EXAMPLE
set @xml = '
<row xmlns:xsi=""http://www.w3.org/2001/XMLSchema-instance"">
  <id>1</id>
  <name>asdf</name>
  <siteaddress1>asdf</siteaddress1>
  <siteaddress2>asdf</siteaddress2>
  <sitecity>asdf</sitecity>
  <sitestateid>4</sitestateid>
  <sitezip>333333</sitezip>
  <phone>asdfsadf</phone>
  <epaid>34343</epaid>
</row>'

declare @rows int
declare @i int
declare @sql nvarchar(max)

set @i = 0
set @rows = (SELECT Tbl.Col.value('count(*)', 'int') FROM @xml.nodes('//row') Tbl(Col))
set @sql = ''

while (@i <= @rows)
begin
    set @i = @i + 1
    set @sql = @sql + 'select Tbl.Col.value(''*[' + cast(@i as varchar(10)) + ']'', ''varchar(max)'') 
                       FROM @x.nodes(''//row'') Tbl(Col)
                       union '
end

set @sql = substring(@sql,0,len(@sql)-5)

exec sp_executesql @sql,N'@x xml', @x = @xml;

```

This works but it's ugly and not set based. There has to be a better way.


EDIT. In my case the schema will always be



```
<row xmlns:xsi=""http://www.w3.org/2001/XMLSchema-instance"">
...
...
...
</row>

```

","Here you go:



```
DECLARE @xml XML

set @xml = '
<row xmlns:xsi=""http://www.w3.org/2001/XMLSchema-instance"">
  <id>1</id>
  <name>asdf</name>
  <siteaddress1>asdf</siteaddress1>
  <siteaddress2>asdf</siteaddress2>
  <sitecity>asdf</sitecity>
  <sitestateid>4</sitestateid>
  <sitezip>333333</sitezip>
  <phone>asdfsadf</phone>
  <epaid>34343</epaid>
</row>'

select r.value('fn:local-name(.)', 'nvarchar(50)') as 'elementName',
    r.value('.','nvarchar(max)') AS 'value'
FROM
    @xml.nodes('//row/*') AS records(r)

```

If you just want the distinct element names you just have to do:



```
select distinct r.value('fn:local-name(.)', 'nvarchar(50)') as t FROM
    @xml.nodes('//row/*') AS records(r)

```

"
"nodejs + multer + angularjs for uploading without redirecting
I am using Nodejs + Multer +angularjs for uploading files on the server.  
  

i have a simple HTML file: 



```
<form action=""/multer"" method=""post"" enctype=""multipart/form-data"">
<input type=""file"" id=""photo"" name=""photo""/> 
<button id=""Button1"">Upload</button>
</form>

```

Nodejs part:



```
var multer  = require('multer');

var storage = multer.diskStorage({
    destination: function (req, file, cb) {
        cb(null, './uploads/')
    },
    filename: function (req, file, cb) {
        cb(null, file.originalname)
    }
})

app.post('/multer', upload.single('photo'), function (req, res) {
    res.end(""File uploaded."");
});

```

**this works perfectly** and the file is successfully uploaded.   
  

but this redirect me to ""/multer"" after uploading the file (because of the form element).  

**How do i stay on the same page??**..possibly using angularjs  

so i tried this:  

  

making a HTML angular file:



```
<section data-ng-controller=""myCtrl"">
<input type=""file"" id=""photo"" name=""photo""/> 
<button id=""Button1"" ng-click=""f()"">Upload</button>
</section>

```

and a Angularjs controller:



```
angular.module('users').controller('myCtrl',[$scope,function($scope){
    $scope.f=function(){
        var photo = document.getElementById('photo');
        var file = photo.files[0];
        if (file) {

            //code to make a post request with a file object for uploading?????

            //something like..
            //$http.post('/multer', file).success(function(response) {
            //console.log(""success"");
            //});
        }
    }
}]);

```

  

**CAN SOMEONE HELP ME WITH THE CODE FOR MAKING A POST REQUEST WITH A FILE OBJECT FOR UPLOADING USING MULTER FROM ANGULARJS CONTROLLER ?**
  

  

thanks


","Angularjs directive:  




```
angular.module('users').directive('fileModel', ['$parse', function ($parse) {
return {
    restrict: 'A',
    link: function(scope, element, attrs) {
        var model = $parse(attrs.fileModel);
        var modelSetter = model.assign;

        element.bind('change', function(){
            scope.$apply(function(){
                modelSetter(scope, element[0].files[0]);
            });
        });
    }
};
}]);

```

Angular html file:  




```
<input type=""file"" file-model=""myFile""/><br><br>
<button ng-click=""uploadFile()"">Upload</button>

```

Angularjs Controller:  




```
$scope.uploadFile = function(){

        var file = $scope.myFile;
        var uploadUrl = ""/multer"";
        var fd = new FormData();
        fd.append('file', file);

        $http.post(uploadUrl,fd, {
            transformRequest: angular.identity,
            headers: {'Content-Type': undefined}
        })
        .success(function(){
          console.log(""success!!"");
        })
        .error(function(){
          console.log(""error!!"");
        });
    };

```

Nodejs server route file:  




```
var multer  = require('multer');
var storage = multer.diskStorage({
    destination: function (req, file, cb) {
        cb(null, './uploads/')
    },
    filename: function (req, file, cb) {
        cb(null, file.originalname+ '-' + Date.now()+'.jpg')
    }
});
var upload = multer({ storage: storage });

app.post('/multer', upload.single('file'));

```

Enjoy!


"
"How to change the status bar notification icons' color/tint in android (marshmallow and above 23+)?
As the title says, how do I change the status bar icons' color to have a dark tint instead of the default white.


**FROM**


[![enter image description here](https://i.stack.imgur.com/yna6Y.jpg)](https://i.stack.imgur.com/yna6Y.jpg)


**TO**


[![dark status bar](https://i.stack.imgur.com/esQqV.jpg)](https://i.stack.imgur.com/esQqV.jpg)


","For the status bar icons' to have a dark tint instead of the default white, add the following tag in your `styles.xml` (or more precisely in `values-v23/styles.xml`) file:



```
<item name=""android:windowLightStatusBar"" tools:targetApi=""23"">true</item>

```

You can also change the flag at runtime by setting it to any `View`:



```
View yourView = findViewById(R.id.your_view);

if (Build.VERSION.SDK_INT >= Build.VERSION_CODES.M) {
    if (yourView != null) {
        yourView.setSystemUiVisibility(View.SYSTEM_UI_FLAG_LIGHT_STATUS_BAR);
    }
}

```

If you want to reset the changes, clear the flag like this:



```
yourView.setSystemUiVisibility(0);

```

"
"ExceptionHandling Middleware vs filter aspnetcore webapi 2.1
I wonder if somebody can clarify this one .I find the usage confusing.


Both links and video did NOT answer my question


I am aware of links like
[asp.net core middleware vs filters](https://stackoverflow.com/questions/42582758/asp-net-core-middleware-vs-filters)


and even a video about it


However for the purpose of the webApi alone does it make sense to implement both?


My understanding is that they cover different parts of the system.


Is there a project (github)or link that clearly explain how to use them.


Can I use both in the same webApi ?


Any samples?


thanks


","You can use both in your project but if your project only has MVC APIs and you are concerned only about capturing unhandled exceptions from your code then using a filter or middleware won't make much of difference and I would say use a filter.


However, if you are more concerned about errors that may occur outside of the MVC context or your code, for example, you want to capture an error in the Routing middleware then use a middleware.


Another major reason to use a filter would be that you get the MVC context in the filter. So if you have a logic like that if the exception occurred when the call was made to POST /orders then you want to log a different exception and if it was in another action then do something else. In such a case use a Filter.


In summary, I try to follow a flow like this


**AM I CONCERNED ABOUT ERRORS IN THE ASP.NET FRAMEWORK** - Use Middleware


**AM I CONCERNED ONLY ABOUT ERRORS IN MY CODE** - Use Filter


**DO I NEED TO CAPTURE THE GLOBAL ERRORS IN THE ASP.NET FRAMEWORK AS WELL AS I HAVE UNIQUE LOGIC BASED ON WHERE THE ERROR OCCURRED** - Use Both


"
"Invoke jQuery Unobtrusive validation on Backbone.js Views
Validations messages are not getting displayed when dynamically parsing html to be validated with jQuery unobtrusive validation plugin. Js fiddle to display the problem is below


**Jsfiddle**


<http://jsfiddle.net/R92Yn/2/>



```
render: function () {
    var tmpl = _.template(this.template);
    $.validator.unobtrusive.parse(tmpl);
    this.$el.html(tmpl);
    return this;
},
validateForm: function () {
    //$(this.$el).validate();
}

```

I am trying to display a error message that usernaem is required when clicking validate button. But nothing happens, also after you blur out a js error is visible in the console. So how do i make validation work along with Backbone.js views


","The jQuery unobtrusive validation plugin (which is originally built for ASP.Net MVC3) requires that your `input`s are placed inside an `form` element. So if don't have `form` the plugin won't work.


So modify your template to include a form:



```
<script type=""text/template"" id=""loginTemplate"">
    <form id=""myform"">
        <label>Username</label>
        <input type=""text"" name=""username"" id=""username"" data-val=""true"" 
               data-val-required=""user name required""/>
        <span data-valmsg-for=""username"" data-valmsg-replace=""true""></span>
        <input type=""button"" id=""submit"" value=""Validate""/>
    </form>
</script>

```

Note: I've also change `data-val-required=""true""` to `data-val-required=""user name required""` because with the `data-val-required` you specify the error message.


In order to the unobtrusive validation work you need to call `$.validator.unobtrusive.parse` after your Backbone view has been attached to the DOM. To do this [there are multiple ways exits](https://stackoverflow.com/questions/9145680/calling-javascript-on-rendering-views-in-backbone-js-post-render-callback). 


One is to use the `_.defer()` function:



```
render: function() {
        var tmpl = _.template(this.template);            
        this.$el.html(tmpl);
        _.defer( function(){ $.validator.unobtrusive.parse(tmpl); });            
        return this;
}

```

Then you can manually trigger the unobtrusive validation with the following code



```
validateForm: function() {
        $(""#myform"").data('unobtrusiveValidation').validate();
}

```

And finally here is a working [JSFiddle](http://jsfiddle.net/wJV6g/).


"
"Vaadin - Expand components in Grid Layout in Vaadin
I am doing my project in Vaadin 6. In that, I have integrated the components in a Grid layout. I have attached a image with this. It resembles my Project layout. It is similar to eclipse. So, The layout will have two side panels on left and right. and a center and bottom panel. 
And the red color in the image are buttons. When I press them the panel will be minimized. I want the center panel to expand and occupy the minimized panel's area. I have integrated the panels in grid layout. Can anyone tell me how to expand a component in the grid layout.


When I minimize 'R'; 'C' & 'B' have to occupy it's area.


When I minimize 'B'; 'C' have to occupy it's area.


When I minimize 'L'; 'C' & 'B' have to occupy it's area.
![layout](https://i.stack.imgur.com/Xdk78.jpg)


","Here a working example without additional add-on. You can use horizontal and vertical layout. 



```
public class TestUI extends UI {

    Panel L = new Panel();
    Panel C = new Panel();
    Panel B = new Panel();
    Panel R = new Panel();

    Button bL = new Button(""Toggle L"", new CloseButtonHandler(L));
    Button bC = new Button(""Toggle C"", new CloseButtonHandler(C));
    Button bB = new Button(""Toggle B"", new CloseButtonHandler(B));
    Button bR = new Button(""Toggle R"", new CloseButtonHandler(R));

    @Override
    protected void init(VaadinRequest request) {
        L.setWidth(""80px"");
        L.setHeight(""100%"");
        L.setContent(new Label(""L""));
        R.setWidth(""80px"");
        R.setHeight(""100%"");
        R.setContent(new Label(""R""));
        B.setHeight(""80px"");
        B.setWidth(""100%"");
        B.setContent(new Label(""B""));
        C.setHeight(""100%"");
        C.setHeight(""100%"");
        C.setContent(new Label(""C""));

        VerticalLayout vl = new VerticalLayout();
        vl.addComponent(C);
        vl.addComponent(B);
        vl.setExpandRatio(C, 1);
        vl.setSizeFull();

        HorizontalLayout hl = new HorizontalLayout();
        hl.addComponent(L);
        hl.addComponent(vl);
        hl.addComponent(R);
        hl.setExpandRatio(vl, 1);
        hl.setSizeFull();

        CssLayout root = new CssLayout();
        root.addComponent(bL);
        root.addComponent(bC);
        root.addComponent(bB);
        root.addComponent(bR);
        root.addComponent(hl);
        root.setSizeFull();

        setContent(root);
    }

    private class CloseButtonHandler implements ClickListener {
        private Panel layout;

        public CloseButtonHandler(Panel layout) {
            this.layout = layout;
        }

        @Override
        public void buttonClick(ClickEvent event) {
            layout.setVisible(!layout.isVisible());
        }
    }
}

```

The example is for Vaadin 7 but the concept should work with Vaadin 6 too. 


"
"Is there a way to specify a range of valid values for a function argument with type hinting in python?
I am a big fan of the type hinting in python, however I am curious if there is a way to specify a valid range of values for a given parameter using type hinting.


What I had in mind is something like



```
from typing import *

def function(
        number: Union[float, int],
        fraction: Float[0.0, 1.0] = 0.5 # give a hint that this should be between 0 and 1,
):
    return fraction * number


```

I can imagine one can enforce this with an assertion, or perhaps specify what the valid range of values is within the docstring, but it feels like having something like Float[0.0, 1.0] would look more elegant.


","Python 3.9 introduced [`typing.Annotated`](https://docs.python.org/3/library/typing.html#typing.Annotated):



```
In [75]: from typing import *

In [76]: from dataclasses import dataclass

In [77]: @dataclass
    ...: class ValueRange:
    ...:     min: float
    ...:     max: float
    ...:

In [78]: def function(
    ...:         number: Union[float, int],
    ...:         fraction: Annotated[float, ValueRange(0.0, 1.0)] = 0.5
    ...: ):
    ...:     return fraction * number
    ...:

```

Like any other type hint it does not perform any runtime checks:



```
In [79]: function(1, 2)
Out[79]: 2

```

However you can implement your own runtime checks. The code below is just an example, it does not cover all cases and probably an overkill for your simple function:



```
In [88]: import inspect

In [89]: @dataclass
    ...: class ValueRange:
    ...:     min: float
    ...:     max: float
    ...:
    ...:     def validate_value(self, x):
    ...:         if not (self.min <= x <= self.max):
    ...:             raise ValueError(f'{x} must be in range [{self.min}, {self.max}]')
    ...:

In [90]: def check_annotated(func):
    ...:     hints = get_type_hints(func, include_extras=True)
    ...:     spec = inspect.getfullargspec(func)
    ...:
    ...:     def wrapper(*args, **kwargs):
    ...:         for idx, arg_name in enumerate(spec[0]):
    ...:             hint = hints.get(arg_name)
    ...:             validators = getattr(hint, '__metadata__', None)
    ...:             if not validators:
    ...:                 continue
    ...:             for validator in validators:
    ...:                 validator.validate_value(args[idx])
    ...:
    ...:         return func(*args, **kwargs)
    ...:     return wrapper
    ...:
    ...:

In [91]: @check_annotated
    ...: def function_2(
    ...:         number: Union[float, int],
    ...:         fraction: Annotated[float, ValueRange(0.0, 1.0)] = 0.5
    ...: ):
    ...:     return fraction * number
    ...:
    ...:

In [92]: function_2(1, 2)
---------------------------------------------------------------------------
ValueError                                Traceback (most recent call last)
<ipython-input-92-c9345023c025> in <module>
----> 1 function_2(1, 2)

<ipython-input-90-01115cb628ba> in wrapper(*args, **kwargs)
     10                 continue
     11             for validator in validators:
---> 12                 validator.validate_value(args[idx])
     13
     14         return func(*args, **kwargs)

<ipython-input-87-7f4ac07379f9> in validate_value(self, x)
      6     def validate_value(self, x):
      7         if not (self.min <= x <= self.max):
----> 8             raise ValueError(f'{x} must be in range [{self.min}, {self.max}]')
      9

ValueError: 2 must be in range [0.0, 1.0]

In [93]: function_2(1, 1)
Out[93]: 1

```

"
"What is the best way to not use allowmainthreadqueries() using Android Room?
I always use `allowmainthreadqueries()` in Android Room. But many people say it is not recommend to do it.
With this function, Room is very easy to use.
But if i should change my app to not using this function, How do I do that?
Is there an effective way to do it?


","Yes, It is not a good approach to query DB on Main Thread. There are different ways, to keep the DB operation off the main(UI) thread, Mainly


1. Kotlin Coroutines. <https://developer.android.com/codelabs/kotlin-android-training-coroutines-and-room#0>
2. RxJava(RxAndroid) <https://medium.com/androiddevelopers/room-rxjava-acb0cd4f3757>
3. LiveData <https://developer.android.com/codelabs/android-training-livedata-viewmodel#0>
4. Executors
5. AsycTask


Out of the above, the best option would always depend upon what is your use case.
But the first three are the best options to use. since they provide a callback on the result of a query.


"
"How to stop HttpURLConnection connect on Android
I use AsyncTask to connect an URLPath as below code. 



```
@Override
public void onCreate(Bundle savedInstanceState) {
    super.onCreate(savedInstanceState);
    setContentView(R.layout.left_list);

    Button btn = (Button)findViewById(resid);
    btn.setOnClickListener(new OnClickListener() {
        @Override
        public void onClick(View v) {
            //how to stop connect
        }
    });
    new Connecting().execute();
}


class Connecting extends AsyncTask<String, String, String> {
    @Override
    protected void onPreExecute() {
            super.onPreExecute();
        //do something
    }

        @Override
        protected String doInBackground(String... aurl) {
            try {
                URL url = new URL(URLPath);
                connection = (HttpURLConnection)url.openConnection();
                connection.setConnectTimeout(30000);
                connection.setReadTimeout(30000);
                connection.setDoInput(true);
                connection.setUseCaches(false);
                connection.connect();
                is = connection.getInputStream();

            }
            catch (Exception e) {
                e.printStackTrace();
            }
            return null;
        }

        @Override
        protected void onPostExecute(String unused) {
            super.onPostExecute(unused);
        //access InputStream is
        }
}

```

connection may spend much time in connect.  

While it connecting, I want to set the Bbutton btn is pressed to stop the connection connect.  

How can I set in setOnClickListener method? 


","Wrap the code that uses `HttpURLConnection` inside a `Future`, which you can cancel at will. You can also use the `Future`'s timeout feature, since the timeouts in `HttpURLConnection` are not reliable.


Define an executor and future within your networking class:



```
final ExecutorService executor = Executors.newCachedThreadPool(Executors.defaultThreadFactory());
Future<MyResult> future;

```

Then wrap your networking code inside the `Future` like this:



```
future = executor.submit(new Callable<MyResult>() {
    @Override
    public MyResult call() throws IOException {
        HttpURLConnection connection = (HttpURLConnection) new URL(url).openConnection();
        try {
            // .. use the connection ...
        } finally {
            connection.disconnect();
        }
    }
});

MyResult result = future.get(TIMEOUT, TimeUnit.SECONDS);

```

When you want to cancel, you can make this call from any thread:



```
future.cancel(true);

```

If the future times out or is canceled, the connection task may continue to block on a `HttpURLConnection` method, but your application won't be held back. You should still set appropriate timeouts on the connection to speed up freeing of threads and network sockets.


"
"Vegetable factory
I'm new to Java. I'm trying to build a simple hierarchy of classes. 



```
abstract class Vegetable
{
    public Vegetable(double weight)
    {
        this.weight = weight;
        ...
    }

    ...
}

class Tomato extends Vegetable
{
    public Tomato(double weight)
    {
            super(weight);
        ...

    }
    ...
}

class Cucumber extends Vegetable
{
    public Cucumber(double weight)
    {
            super(weight);
        ...

    }
    ...
}

```

I have some abstract methods in class `Vegetable`, which are overridden in `Tomato` and `Cucumber`.
I want to implement logic of buying vegetables. I've created the following class for it: 



```
public class VegetableFactory
{
    public static ArrayList<Vegetable> buy(int num, Class vegetableClass)
    {
        ArrayList<Vegetable> res = new ArrayList<Vegetable>(num);
        for (int i = 0; i < num; i++)
        {
            double weight = (Math.random() * 200) + 100;
            Vegetable v = null;
            try
            {
                v = (Vegetable) vegetableClass.getDeclaredConstructor(Double.TYPE).newInstance(weight);
            } catch (InstantiationException e)
            {
                e.printStackTrace();
            } catch (IllegalAccessException e)
            {
                e.printStackTrace();
            } catch (InvocationTargetException e)
            {
                e.printStackTrace();
            } catch (NoSuchMethodException e)
            {
                e.printStackTrace();
            }
            res.add(v);
        }
        return res;
    }
}  

```

Is it a good way of doing that? And is it ideologically correct to name it factory? Any suggestions of improvement are highly appreciated.


","I would use an enum type instead of having to pass a class in. This way, it's impossible to pass something bad in, and you can catch errors at compile time instead of runtime.



```
enum VegetableType { CARROT, TOMATO, CUCUMBER };

```

Then your buy method would be simply



```
static ArrayList<Vegetable> buy(int num, VegetableType vegType) {
    ArrayList<Vegetable> veggies = new ArrayList<Vegetable>(n); //n is the initial capacity - size is still zero.
    for(int i=0; i < num; ++i) {
        double weight = (Math.random() * 200) + 100;

        switch(vegType) {
        case VegetableType.CARROT:
            veggies.add(new Carrot(weight));
            break;
        case VegetableType.TOMATO:
            veggies.add(new Tomato(weight));
            break;
        case VegetableType.CUCMBER:
            veggies.add(new Cucumber(weight));
            break;
        }
    }
    return veggies;
 }

```

Now of course the problem is that you end up having a big `switch` statement in the middle in your loop. Generally, while ugly I'd say that this is sufficient - you get some more control over each of the individual constructors, so you're not constrained to the same set of parameters.


The other thing that you could do is have a factory method on the `VegetableType` to create an instance:



```
enum VegetableType {
    CARROT { Vegetable createInstance(int weight) { return new Carrot(weight); }},
    TOMATO { Vegetable createInstance(int weight) { return new Tomato(weight); } },
    CUCUMBER { Vegetable createInstance(int weight) { return new Cucumber(weight); } };

    abstract Vegetable createInstance(int weight);
}

```

Then your list becomes



```
static ArrayList<Vegetable> buy(int num, VegetableType vegType) {
    ArrayList<Vegetable> veggies = new ArrayList<Vegetable>();
    for(int i=0; i < num; ++i) {
        double weight = (Math.random() * 200) + 100;
        veggies.add(vegType.createInstance(weight));
    }
    return veggies;
 } 

```

Finally, if you want to get fancy, `java.lang.reflect` has the `Constructor` class:


<http://docs.oracle.com/javase/1.5.0/docs/api/java/lang/reflect/Constructor.html>


You can then use this to get the constructor to the `Vegetable` type of your choice, and then call construct repeatedly in the loop.


"
"Ansible | Set\_fact with condition
I have the list:



```
val_list:
    - val: ""D""
    - val: ""A""
    - val: ""B""
    - val: ""C""

```

I would like to set variable true if list contain `val: B` without creating register variable task, for example:



```
- set_fact:
    containB: false

- set_fact:
    containB: true
  when: item.val=='B'
  with_items
    - ""{{ val_list }}""

```

I would like to convert to something like that:



```
- set_fact: 
    containB: true if item.val=='B' | default(false)
#or containB: true if any val in val_list is 'B'
  with_items
    - ""{{ val_list }}""

```

","1. extract values from the list with the `map` filter
2. check that 'B' is an element of the resulting list


In a nutshell:



```
- hosts: localhost

  vars:
    val_list:
      - val: ""D""
      - val: ""A""
      - val: ""B""
      - val: ""C""

    list_contains_B: ""{{ 'B' in (val_list | map(attribute='val')) }}""

  tasks:

    - name: Does list contain B ?
      debug:
        var: list_contains_B

```

Which gives:



```
PLAY [localhost] ***********************************************************************************************************************************************************************************************************************

TASK [Gathering Facts] *****************************************************************************************************************************************************************************************************************
ok: [localhost]

TASK [Does list contain B ?] ***********************************************************************************************************************************************************************************************************
ok: [localhost] => {
    ""list_contains_B"": true
}

PLAY RECAP *****************************************************************************************************************************************************************************************************************************
localhost                  : ok=2    changed=0    unreachable=0    failed=0    skipped=0    rescued=0    ignored=0 

```

"
"(Rational) Unified Process vs Waterfall Model
I was reading about software development model and life cycle, where I learned about waterfall model and Unified process. However both the process involves requirement gathering, design phase, development testing and deployment (inception, elaboration, construction and transition phase in unified process). 


Can any one help me out with the difference between the two? 


","You haven't specified 'which' Unified Process or 'which' Waterfall Process - there were many variants of both, so some of the comparison will be lost in generalization.


Taking for example the [Rational Unified process](https://en.wikipedia.org/wiki/IBM_Rational_Unified_Process), which differs from waterfall processes in that the disciplines (Analysis, Design, Coding, Testing etc) are done iteratively and concurrently, whereas in [waterfall processes](https://en.wikipedia.org/wiki/Waterfall_model), the disciplines are generally done sequentially (e.g. Coding only starts once Requirements have been finalized and Design has been accepted).


In RUP, the term `phases` (inception, elaboration, construction, transition) are NOT dedicated to a single discipline or single deliverable - RUP phases are all [multi disciplinary](https://en.wikipedia.org/wiki/File:Development-iterative.gif) - e.g. although Inception is *primarily* about requirements and analysis; some design and prototype coding are also encouraged to reduce risk and improve estimates for future phases, and even in the construction phase, further analysis may be required.


RUP uses the term 'generation' to indicate another full cycle of development e.g. for a ""version 2"" of a project, where new work on generation 2 would begin at the inception phase.


Another major difference was that RUP pushed the concept of Visual Models (especially UML) as deliverable artifacts which described the requirements, high and class level design (and in some cases code generation was possible from detailed UML models), whereas Waterfall artifacts were typically very document heavy (e.g. the [ESA / IEEE processes](http://codecourse.sourceforge.net/materials/The-Waterfall-Lifecycle-Model.html))


Another difference was in the approach to commercial engagement. Waterfalls typically promoted the concept of a 'contractual' Software Requirement or Software Specification document, which defined the deliverable (functionally, and non-functionally), and from which a project budget or fixed price transaction would be based. Instead, RUP promoted budgeting at a per-phase basis, e.g. and the effort / cost for the following phase would be known / iterated / improved as one of the deliverables of the preceding phase had been delivered.


In many software development operations, [Agile processes](https://en.wikipedia.org/wiki/Agile_software_development) have superseded both Waterfall and RUP, although many of the artifacts and learnings of Waterfall and RUP remain. Agile has the primary benefit of breaking work down into much smaller chunks (typically 2 week sprints, instead of months-long RUP phases or year-long waterfall projects). This quick turnaround allows features to be delivered on a cost vs priority basis, is better adapted to ever-changing requirements, and allows obstacles to success to be identified much sooner than either Waterfall or RUP. Agile also cuts out a lot of the waste - lets face it, only a small percentage of developers ever read detailed specification documents or pore over detailed UML diagrams.


"
"NDK debugging: ndk-gdb fails to pull app\_process. Who and when creates the app\_process binary?
While investigating a problem with hitting breakpoints in native code, I decided to check if ndk-gdb works OK. I've deleted app\_process, started Java debugging and ran `ndk-gdb --force`. Guess what, app\_process is not created. `ndk-gdb --verbose` output has this line :



```
## COMMAND: adb_cmd pull /system/bin/app_process obj/local/armeabi-v7a/app_process
remote object '/system/bin/app_process' not a file or directory
Pulled app_process from device/emulator.

```

I've uninstalled the app from the target device, rebooted the device and repeated. Still no app\_process. So, I wonder what's the problem here, and I also wonder when is this file created, and by what process / script.


","As mentioned before, Android 5.0 has `/system/bin/app_process` as a symlink to `/system/bin/app_process32`. Since a symlink cannot be pulled with `adb pull`, the `ndk-gdb` script will not be able to work as-is.


Change `ndk-gdb` as follows to support api21 changes as well as backwards compatibility for < api21:



```
# Get the app_server binary from the device
APP_PROCESS=$APP_OUT/app_process
if [ ""$API_LEVEL"" -lt ""21"" ] ; then
    run adb_cmd pull /system/bin/app_process `native_path $APP_PROCESS`
    log ""Pulled app_process from device/emulator to $APP_PROCESS""
else
    run adb_cmd pull /system/bin/app_process32 `native_path $APP_PROCESS`
    log ""Pulled app_process32 from device/emulator to $APP_PROCESS""
fi

```

EDIT: or:



```
# Get the app_server binary from the device
APP_PROCESS=$APP_OUT/app_process
APP_PROCESS_DEVICE=app_process32
if [ ""$API_LEVEL"" -lt ""21"" ] ; then
    APP_PROCESS_DEVICE=app_process
fi
run adb_cmd pull /system/bin/$APP_PROCESS_DEVICE `native_path $APP_PROCESS`
log ""Pulled $APP_PROCESS_DEVICE from device/emulator to $APP_PROCESS""

```

Original Change is also here: <http://pastebin.com/YfxNs06U>. Note that this change renames `app_process32` to `app_process` when it is pulled to your development machine to support having a single debugging config in Eclipse.


"
"Combine a list of matrices to a single matrix by rows
Let's say I have a list of matrices (all with the same number of columns). How would I append / combine these matrices by row ('row bind', `rbind`) to get a single matrix?


Sample:



```
> matrix(1, nrow=2, ncol=3)
     [,1] [,2] [,3]
 [1,]    1    1    1
 [2,]    1    1    1
> matrix(2, nrow=3, ncol=3)
     [,1] [,2] [,3]
[1,]    2    2    2
[2,]    2    2    2
[3,]    2    2    2
> m1 <- matrix(1, nrow=2, ncol=3)
> m2 <- matrix(2, nrow=3, ncol=3)

```

Now we can have many matrices in a list, let's say we have only two:



```
l <- list(m1, m2)

```

I would like to achieve something like:



```
> rbind(m1, m2)
     [,1] [,2] [,3]
[1,]    1    1    1
[2,]    1    1    1
[3,]    2    2    2
[4,]    2    2    2
[5,]    2    2    2

```

I can easily do it on 2 matrices but I am not sure how to do it with a list of matrices.


","Use `do.call(rbind,...)`



```
> m1 <- matrix(1, nrow=2, ncol=3)
> m2 <- matrix(2, nrow=3, ncol=3)
> l <- list(m1, m2)
> do.call(rbind, l)
     [,1] [,2] [,3]
[1,]    1    1    1
[2,]    1    1    1
[3,]    2    2    2
[4,]    2    2    2
[5,]    2    2    2

```

You may also be interested in the `rbind.fill.matrix()` function from the ""plyr"" package, which will also let you bind matrices with differing columns, filling in with `NA` where necessary.



```
> m1 <- matrix(1, nrow=2, ncol=3)
> m2 <- matrix(2, nrow=3, ncol=4)
> l <- list(m1, m2)
> library(plyr)
> rbind.fill.matrix(l)
     1 2 3  4
[1,] 1 1 1 NA
[2,] 1 1 1 NA
[3,] 2 2 2  2
[4,] 2 2 2  2
[5,] 2 2 2  2

```

"
"Why no i++ in Scala?
I just wonder why there is no `i++` to increase a number. As what I know, languages like Ruby or Python doesn't support it because they are dynamically typed. So it's obviously we cannot write code like `i++` because maybe `i` is a string or something else. But Scala is statically typed - the compiler absolutely can infer that if it is legal or not to put `++` behind a variable.


So, why doesn't `i++` exist in Scala?


","Scala doesn't have `i++` because it's a functional language, and in functional languages, operations with side effects are avoided (in a purely functional language, no side effects are permitted at all). The side effect of `i++` is that `i` is now 1 larger than it was before. Instead, you should try to use immutable objects (e.g. `val` not `var`).


Also, Scala doesn't really need `i++` because of the control flow constructs it provides. In Java and others, you need `i++` often to construct `while` and `for` loops that iterate over arrays. However, in Scala, you can just say what you mean: `for(x <- someArray)` or `someArray.foreach` or something along those lines. `i++` is useful in imperative programming, but when you get to a higher level, it's rarely necessary (in Python, I've never found myself needing it once).


You're spot on that `++` *could* be in Scala, but it's not because it's not necessary and would just clog up the syntax. If you really need it, say `i += 1`, but because Scala calls for programming with immutables and rich control flow more often, you should rarely need to. You certainly could define it yourself, as operators are indeed just methods in Scala.


"
"SQL query to find rows
I have a table `mytable` in ORACLE database having only one column `Roll_no` like following:



```
Roll_no
----------
CD01 0001
CD01 0002
CD01 0004
CD01 0005
CD01 0008

```

here `CD01` is fixed, after `CD01` one space is there and then numbers are written `0001` and onwards.


Please see here that `CD01 0003`, `CD01 0006`, `CD01 0007` are missing and these are my required output.


Required output:



```
Roll_no
----------
CD01 0003
CD01 0006
CD01 0007

```

How can I write a SQL query to find out these missing alphanumeric characters? Any ideas please


","You could generate the full list of number and then with a NOT EXISTS (or NOT IN) filter the existing records out. For example:



```
SELECT new_roll_no
  FROM (SELECT 'CD01 ' || to_char(rownum, 'fm0000') new_roll_no FROM dual 
        CONNECT BY LEVEL <= (SELECT to_number(MAX(substr(roll_no, 6))) FROM T))
 WHERE new_roll_no NOT IN (SELECT roll_no FROM T)
 ORDER BY 1

```

Here's how it works: 


- CONNECT BY LEVEL is a bit of a trick query, it will generate rows up to the level you indicate (in this case I chose the max from your table). I think this query first appeared on the [askTom](http://asktom.oracle.com) site and is now quite widespread.
- I've generated 8 rows, I will next use the TO\_CHAR function to make a column look like your primary key (see the [format model](http://docs.oracle.com/cd/E11882_01/server.112/e26088/sql_elements004.htm#SQLRF51073) documentation of the [TO\_CHAR](http://docs.oracle.com/cd/E11882_01/server.112/e26088/functions201.htm#i79330) function).
- I then compare these generated values to your actual table value (NOT IN is an ANTI-JOIN)


"
"Timed Metadata with AVPlayer
Can you control how `AVURLAsset` loads the data it requires? I need to add a custom HTTP header (`ICY-MetaData=1`) and can't seem to figure out how.


","Apparently, `AVPlayer` automatically requests metadata from Icecast. The code below works perfectly.



```
class ViewController: UIViewController {
    var Player: AVPlayer!
    var PlayerItem: AVPlayerItem!

    override func viewDidLoad() {
        super.viewDidLoad()

        PlayerItem = AVPlayerItem(URL: NSURL(string: ""http://live.machine.fm/aac""))
        PlayerItem.addObserver(self, forKeyPath: ""timedMetadata"", options: nil, context: nil)
        Player = AVPlayer(playerItem: PlayerItem)
        Player.play()
    }

    override func observeValueForKeyPath(keyPath: String, ofObject object: AnyObject, change: [NSObject : AnyObject], context: UnsafeMutablePointer<Void>) -> Void {

        if keyPath != ""timedMetadata"" { return }

        var data: AVPlayerItem = object as AVPlayerItem

        for item in data.timedMetadata as [AVMetadataItem] {
            println(item.value)
        }
    }

    override func didReceiveMemoryWarning() {
        super.didReceiveMemoryWarning()
        // Dispose of any resources that can be recreated.
    }
}

```

"
"How does Windows Aero prevent screen tearing?
When Windows Aero is on, there isn't any screen tearing, as opposed to when using the Basic theme without Aero. 


Why, though, do games with vertical sync seem to have a much greater input lag than when doing something on the desktop? It even seems to sync windowed games with V-sync off, better than they would be if they were using whatever method is built into the game.


Why and how does this happen, or appear to happen?


","Actually, the reason Aero can prevent tearing without having to force VSYNC on in an application is because it is a compositing window manager. It draws the desktop and all windows asynchronously with respect to any running application.


That is to say, when you enable compositing in Windows Vista / 7 (it is ***always*** enabled in Windows 8), it will draw all of the windows using a copy of the last image that was swapped from back to front. The compositor (Microsoft calls it DWM, **Desktop Window Manager**) composites everything with VSYNC enabled, and because it draws using a copy of the last **fully-swapped** front buffer image for each window it never displays partially drawn frames. It does, however, have an unfortunate side effect of preventing really old software that was designed to draw into the front buffer exclusively from working correctly - but modern software does not do this.


In a nutshell, the DWM adds an additional layer of protection against tearing. One that will allow any application running on the system to draw at a rate irrespective of the display's refresh rate and still prevent tearing. On such a system, in windowed mode the only thing enabling VSYNC in OpenGL or Direct3D is good for is actually throttling back CPU/GPU utilization. 


This is why adaptive VSYNC was created, the idea there is not to penalize applications that cannot sustain the display's refresh rate by forcing them into a lower factor of the refresh rate (e.g. 60 Hz --> 30, 20, 15, 10, 12, 6, 5, 4, 3, 2, 1) but to limit applications that are drawing faster than the monitor can display images from using an excessive amount of CPU / GPU power.


"
"What does it mean when psql has blank database access privileges?
When I run `\l+` in `psql` to get a list of all databases and their access privileges, I see that some of the databases have blank access privileges. What does it mean when the `Access privileges` column is blank in `psql`?


If there is some sort of default, how can I see what that default is?



```
me=# \l+
                                                                      List of databases
    Name     |    Owner    | Encoding |   Collate   |    Ctype    |   Access privileges   |  Size   | Tablespace |                Description                 
-------------+-------------+----------+-------------+-------------+-----------------------+---------+------------+--------------------------------------------
 me          | me          | UTF8     | en_US.UTF-8 | en_US.UTF-8 |                       | 7239 kB | pg_default | 
 postgres    | postgres    | UTF8     | en_US.UTF-8 | en_US.UTF-8 |                       | 7151 kB | pg_default | default administrative connection database
 template0   | postgres    | UTF8     | en_US.UTF-8 | en_US.UTF-8 | =c/postgres          +| 7041 kB | pg_default | unmodifiable empty database
             |             |          |             |             | postgres=CTc/postgres |         |            | 
 template1   | postgres    | UTF8     | en_US.UTF-8 | en_US.UTF-8 | =c/postgres          +| 7041 kB | pg_default | default template for new databases
             |             |          |             |             | postgres=CTc/postgres |         |            | 
(4 rows)

```

","
> 
> *What does it mean when the Access privileges column is blank*
> 
> 
> 


[Quote from the manual:](https://www.postgresql.org/docs/current/ddl-priv.html#PRIVILEGE-ABBREVS-TABLE)



> 
> If the ""Access privileges"" column is empty for a given object, it means the object has default privileges (that is, its privileges column is null). Default privileges always include all privileges for the owner, and can include some privileges for PUBLIC depending on the object type, as explained above
> 
> 
> 


The default privileges are also documented there



> 
> PostgreSQL grants privileges on some types of objects to PUBLIC by default when the objects are created. No privileges are granted to PUBLIC by default on tables, table columns, sequences, foreign data wrappers, foreign servers, large objects, schemas, or tablespaces. For other types of objects, the default privileges granted to PUBLIC are as follows: CONNECT and TEMPORARY (create temporary tables) privileges for databases; EXECUTE privilege for functions and procedures; and USAGE privilege for languages and data types (including domains).
> 
> 
> 


"
"How can we stop a running java process through Windows cmd?
I am a newbie in cmd, so please allow me to ask a stupid question: How can we stop a running Java process through Windows cmd?


For example, if we start Jetty (a mini web server) with the following command:



```
start javaw -jar start.jar

```

How do we find the process and stop it later?


Obviously the following command does not work:



```
stop javaw -jar start.jar

```

","It is rather messy but you need to do something like the following:



```
START ""do something window"" dir
FOR /F ""tokens=2"" %I in ('TASKLIST /NH /FI ""WINDOWTITLE eq do something window""' ) DO SET PID=%I
ECHO %PID%
TASKKILL /PID %PID%

```

Found this on [this page](https://web.archive.org/web/20110731025432/http://it.toolbox.com/blogs/database-solutions/kill-a-process-in-windows-bat-19875). (archived)


(This kind of thing is much easier if you have a UNIX / LINUX system ... or if you run Cygwin or similar on Windows.)


"
"Change time format for a specific application
I want to use a specific locale for some application but not for the rest of my system. How do I do this?


","# Theoretical part


You can change the locale or only parts of it for a specific process before you create it by changing its environment.


Check it out by launching `gnome-calculator` via



```
env LC_NUMERIC=de_DE.UTF-8 gnome-calculator

```

if you currently use a period as your decimal point and via



```
env LC_NUMERIC=en_IE.UTF-8 gnome-calculator

```

if you currently use a comma as you decimal point.


The `gnome-calculator` process will use the locale stated before its call.


Note that this doesn't persist if you close the application and open it via the Dash or just call `gnome-calculator` in a terminal. In fact, it is process-specific and you can use several instances of `gnome-calculator`, some of them using periods and some of them using commas as their decimal points. Check it out by running `gnome-calculator` in a different terminal. It will use your normal settings.


# Finding the right locale


You don't need to find a single locale which matches your needs for everything. Instead, it's sufficient to find one that matches what you want in a specific localization category. Here, we only care about the time format which can be changed by manipulating the environment variable `LC_TIME`.


You're probably interested in the international time format ([ISO 8601](https://en.wikipedia.org/wiki/ISO_8601)) or the time format of some country of which you know the time format. For the former, use `en_DK.UTF-8`. For the latter, use the 2-letters abbreviation of the language in lowercase letters, an underscore, the 2-letters abbreviation for the county in capital letters, and then `.UTF-8`. For example, `en_IE.UTF-8` is Irish English, `de_DE.UTF-8` is German German, `de_CH.UTF-8` is Swiss German, and `fr_CH.UTF-8` is Swiss French.


# Practical part


Now that you learned a bit about how cool Linux is, let's get to the practical part.


You probably launch your application via the Dash. If you do so, you use a desktop file. Find your desktop file. It's either in `/usr/share/applications` or in `~/.local/share/applications`. The former is system-wide, the latter is user-specific.


I'll use Firefox as the example application. It's desktop file is `/usr/share/applications/firefox.desktop`.


After you found the desktop file you want to manipulate, copy it to your user's desktop file folder:



```
cp /usr/share/applications/firefox.desktop ~/.local/share/applications

```

You can now manipulate the copy such that your changes only affect the user-local version.


Open it in a text editor. It doesn't matter whether you use a command line one or a graphical one. Normal Ubuntu comes with GEdit as its default graphical text editor:



```
gedit ~/.local/share/applications/firefox.desktop

```

If it doesn't work because you use a flavor of Ubuntu which doesn't come with GEdit, just use `nano`:



```
nano ~/.local/share/applications/firefox.desktop

```

Find the line which begins with `Exec=`. For Firefox, it's `Exec=firefox %u`. Then take whatever comes after `Exec=` and get it into this format with the locale you want instead of `en_DK.UTF-8`:



```
Exec=env LC_TIME=en_DK.UTF-8 firefox %u

```

Of course, you need to substitute `en_DK.UTF-8` by whatever locale you want to use.


Save the file.


After you made your changes, you need to make the desktop file executable. The file you copied was executable but that property isn't copied, so you need to set it again for the new file. Do do so, run this command:



```
chmod +x ~/.local/share/applications/firefox.desktop

```

Now, restart your application.


Dash usually updates the desktop files automatically but on slow systems, it might take a while. Or at least it did in some version. If it doesn't work instantly, log out and back in again.


"
"How can I protect an accessor only IEnumerable property from mutation?
Suppose I have this:



```
public class MyClass
{
    private readonly List<string> _strings = new List<string>();

    public IEnumerable<string> MyStrings 
    { 
        get 
        { 
            return _strings; 
        }
    }
}

```

How can I prevent the following from actually adding to the member variable list of string?



```
var theClass = new MyClass();
((List<string>)theClass.MyStrings).Add(""I do not want to be able to add this!"");

```

","Create a new type that ""wraps"" the sequence. While there are many ways of doing this, one fairly simple ""hack"" is to just use `Select` on it with the identity transformation:



```
return _strings.Select(x=>x); 

```

This will create a new object with it's own iterator that, internally, has a reference to the `_strings` list. Due to that extra layer of indirection (and with the extra bonus of the type being a type internal to the language that is not valid to use outside of the language definition), the caller can no longer cast their way to the list.


If you want something that's a bit clearer to read that makes the intentions a bit more known, you can just make a new method for it:



```
public static IEnumerable<T> AsSequence<T>(this IEnumerable<T> sequence)
{
    return sequence.Select(x => x);
}

```

(Feel free to name it whatever is clearest to you.)


"
"How Does A Compiler Work?
*Note: I am surprised that this hasn't been asked before, and if it has I could not find it in a search.*


I've been on tons of websites, I've read tons of articles, and I have heard tons of explanations. Most of them were good, but they were all either to broad or too complicated or just plain bad. So my question is , how does a compiler work?


If this is a difficult, broad question, please tell me. But if not, please answer the question.


","A compiler is a program that translates the source code for another program from a programing language into executable code.


The source code is typically in a high-level programming language (e. g. Pascal, C, C++, Java, Perl, C#, etc.). The executable code may be a sequence of machine instructions that can be executed by the CPU directly, or it may be an intermediate representation that is interpreted by a virtual machine (e. g. Java byte code).


In short, a compiler converts a program from a human-readable format into a machine-readable format.


As to how a compiler works, that is indeed complicated. There are books and university courses on the subject. I will attempt to briefly outline the main stages of the process, but this will be a very cursory overview.


1. Lexing - break up the text of the program into ""tokens"". The tokens are the ""words"" of the programming language, such as identifiers (keywords, variable names, function names, etc.) or operators (=, \*, &, etc.).
2. Parsing - convert the sequence of tokens into a parse tree, which is a data structure representing various language constructs: type declarations, variable declarations, function definitions, loops, conditionals, expressions, etc.
3. Optimization - evaluate constant expressions, optimize away unused variables or unreachable code, unroll loops if possible, etc.
4. Translate the parse tree into machine instructions (or JVM byte code).


Again, I stress that this is a very brief description. Modern compilers are very smart, and, consequently, very complicated.


"
"Using list preference in Android
I have a text to speech application where the user can select a language and also select a male or female voice. The problem is that for each language there are different strings used to called the male and female voice but in my preference I only have two options (male and female).



```
<string-array name=""Language"">
    <item>English (US)</item>
    <item>English (UK)</item>
    <item>French (France)</item>
    <item>Spanish (Spain)</item>
    <item>Italian</item>
</string-array>

<string-array name=""languageAlias"">
    <item>""en-US""</item>
    <item>""en-GB""</item>
    <item>""fr-FR""</item>
    <item>""es-ES""</item>
    <item>""it-IT""</item>
</string-array>

<string-array name=""Voice"">
    <item>Male</item>
    <item>Female</item>
</string-array>

<string-array name=""VoiceAlias"">
    <item>""usenglishmale""</item>
    <item>""usenglishfemale""</item>
    <item>""ukenglishmale""</item>
    <item>""ukenglishfemale""</item>
    <item>""eurfrenchmale""</item>
    <item>""eurfrenchfemale""</item>
    <item>""eurspanishmale""</item>
    <item>""eurspanishfemale""</item>
    <item>""euritalianmale""</item>
    <item>""euritalianfemale""</item>        
</string-array>

```

I'm trying to find a way to only reference the relevant male and female voiceAlias depending on the language selected. Is it possible to do this here or do I have to write some code which changes the values of the voiceAlias array depending on the language selected? 


Thanks in Advance


","Ok, you can accomplish this with two `ListPreference`s and an `OnPreferenceChangeListener` for each. First the XML:



```
<ListPreference 
    android:key=""language_preference""
    android:title=""Language""
    android:entries=""@array/Language""
    android:entryValues=""@array/languageAlias""/>

<ListPreference 
    android:key=""gender_preference""
    android:title=""Gender""
    android:entries=""@array/Voice""
    android:entryValues=""@array/VoiceData""/>

```

Let's make a new entry in res/values/array.xml:



```
<string-array name=""VoiceData"">
    <item>0</item>
    <item>1</item>
</string-array>

```

And now in your extention of `PreferenceActivity`, we're going to take the string values which persist in your `SharedPreferences` and from them create a completely new entry in the `SharedPreferences` which gets its value from ""VoiceAlias"".



```
SharedPreferences shareprefs = getPreferenceManager().getSharedPreferences();
Resources resources = YourContext.getResources();

private void makeVoiceData() {
    String languageData = shareprefs.getString(""language_preference"", ""en-US"");
    int genderData = Integer.parseInt(shareprefs.getString(""gender_preference"", ""0""));
    String[] voiceAlias = resources.getStringArray(R.array.VoiceAlias);

    int a = 0
    String[] languageAlias = resources.getStringArray(R.array.languageAlias);
    for (a ; a < languageAlias.length ; a++) {
        if (languageAlias[a].equals(languageData)) {
            break;
        }
    }

    shareprefs.putString(""VoiceAlias"", voiceAlias[(2 * a) + genderData]);
}

ListPreference language_preference = getPreference(""language_preference"");
ListPreference gender_preference = getPreference(""gender_preference"");

language_preference.setOnPreferenceChangeListener(new OnPreferenceChangeListener() {
    public boolean onPreferenceChanged(Preference preference, Object newValue) {
        shareprefs.putString(""language_preference"", (String) newValue);
        makeVoiceData();
    }
});

gender_preference.setOnPreferenceChangeListener(new OnPreferenceChangeListener() {
    public boolean onPreferenceChanged(Preference preference, Object newValue) {
        shareprefs.putString(""gender_preference"", (String) newValue);
        makeVoiceData();
    }
});

```

"
"Macro calling a function works in interpreter, fails in compiler (SBCL + CMUCL)
As suggested in [a macro-related question I recently posted to SO](https://stackoverflow.com/questions/8115097/lisp-can-a-macro-be-recursive/), I coded a macro called ""fast"" via a call to a function [(here is the standalone code in pastebin)](http://pastebin.com/ZXrNtjxJ):



```
(defun main ()
  (progn
    (format t ""~A~%"" (+ 1 2 (* 3 4) (+ 5 (- 8 6))))
    (format t ""~A~%"" (fast (+ 1 2 (* 3 4) (+ 5 (- 8 6)))))))

```

This works in the REPL, under both SBCL and CMUCL:



```
$ sbcl
This is SBCL 1.0.52, an implementation of ANSI Common Lisp.
...
* (load ""bug.cl"")
22
22

$

```

Unfortunately, however, the code no longer compiles:



```
$ sbcl
This is SBCL 1.0.52, an implementation of ANSI Common Lisp.
...
* (compile-file ""bug.cl"")
...
;   during macroexpansion of (FAST (+ 1 2 ...)). Use *BREAK-ON-SIGNALS* to
;   intercept:
;
;    The function COMMON-LISP-USER::CLONE is undefined.

```

So it seems that by having my macro ""fast"" call functions (""clone"",""operation-p"") at compile-time, I trigger issues in Lisp compilers (verified in both CMUCL and SBCL).


Any ideas on what I am doing wrong and/or how to fix this?


","Some remarks about your code.


- multiple tests of an object for equality can be replaced by `MEMBER`
- backquote with a following comma does nothing. You can just remove that.
- you can ensure that your functions are available for a macro by a) moving these functions to an additional file and compile/load that before use of the macro, by b) using `EVAL-WHEN`to inform the compiler to evaluate the definition of the functions or by c) adding the functions to the macro as local functions


Example:



```
(defmacro fast (&rest sexpr)
  (labels ((operation-p (x)
             (member x '(+ - * /)))
           (clone (sexpr)
             (if (consp sexpr)
                 (destructuring-bind (head . tail) sexpr
                   (if (operation-p head)
                       `(the fixnum (,head ,@(clone tail)))
                     (cons (clone head) (clone tail))))
               sexpr)))
    (car (clone sexpr))))

```

Note that this and your version of `FAST` are not complete code walkers. They recognize only simple function calls (and not the other Lisp constructs like LAMBDA, LET, FLET, LABELS, etc.).


"
"What does editor.cursorBlinking do?
The documentation says that the options for editor.cursorBlinking are 'blink', 'smooth', 'phase', 'expand' and 'solid'. It does not describe what each of these means.


What does each of these do?


","The extra cursor animations idea is detailed [here](https://github.com/Microsoft/vscode/issues/7923), and it points to a [JSFiddle showing the five new options](https://jsfiddle.net/rxercv6v/). The following styles show what each does:



```
@keyframes blink  {   0%,  49% { opacity: 0; }
                     51%, 100% { opacity: 1; }
}
@keyframes smooth {   0%,  20% { opacity: 0; }
                     60%, 100% { opacity: 1; }
}
@keyframes phase  {   0%,  20% { opacity: 0; }
                     90%, 100% { opacity: 1; }
}
@keyframes expand {   0%,  20% { transform: scaleY(0); }
                     80%, 100% { transform: scaleY(1); }
}

```

Basically:


- `blink` spends 49% in a given state, 2% transitioning, and 49% in the other state. Transitioning, in this and most below involves just changing the opacity. Because this (and the others) specify `alternate` (the animation runs in alternating directions), this means 98% in a given state and 2% transitioning to the other state.
- `smooth` spends 60% in a given state, 40% transitioning to the other state.
- `phase` spends 30% in a given state, 70% transitioning.
- `expand` spends 40% in a given state, 60% transitioning. The transition in this case is not *opacity* but the scaling on the y-axis, meaning it grows from, and shrinks to, the vertical midpoint.
- `solid` means no blinking at all, the cursor is always there.


In order to make the answer self-contained, I've copied the relevant stuff from that JSFiddle to below (compressing, and better labelling the different styles). Just run the code snippet to get a more visual representation:




---





```
body { font-family: ""Menlo"", ""Monaco"", ""Courier New"", ""Courier""; font-size: 9pt; background-color: black; color: #fff; }
.line { line-height: 18px; background-color: #333; overflow: hidden; margin: 2px; vertical-align: middle; }
.cursor { position: relative; top: 2px; width: 5px; height: 14px; background-color: #eee; display: inline-block; margin-left: 1px; margin-right: 1px; }

@keyframes blink { 0%, 49% { opacity: 0; } 51%, 100% { opacity: 1; } }
@keyframes smooth { 0%, 20% { opacity: 0; } 60%, 100% { opacity: 1; } }
@keyframes phase { 0%, 20% { opacity: 0; } 90%, 100% { opacity: 1; } }
@keyframes expand { 0%, 20% { transform: scaleY(0); } 80%, 100% { transform: scaleY(1); } }

.blink { animation: blink 0.5s ease-in-out 0s infinite alternate; }
.smooth { animation: smooth 0.5s ease-in-out 0s infinite alternate; }
.phase { animation: phase 0.5s ease-in-out 0s infinite alternate; }
.expand { animation: expand 0.5s ease-in-out 0s infinite alternate; }
```


```
<div class=""line""><div class=""cursor blink""></div>-blink</div>
<div class=""line""><div class=""cursor smooth""></div>-smooth</div>
<div class=""line""><div class=""cursor phase""></div>-phase</div>
<div class=""line""><div class=""cursor expand""></div>-expand</div>
<div class=""line""><div class=""cursor solid""></div>-solid</div>
```






"
"Performant Entity Serialization: BSON vs MessagePack (vs JSON)
Recently I've found ***[MessagePack](http://www.msgpack.org)***, an alternative **binary** serialization format to Google's [Protocol Buffers](https://developers.google.com/protocol-buffers/) and [JSON](http://json.org/) which also outperforms both.


Also there's the [BSON](http://bsonspec.org/) serialization format that is used by MongoDB for storing data.


Can somebody elaborate the *differences and the dis-/advantages of BSON vs MessagePack*?




---


Just to complete the list of performant binary serialization formats: There are also [**Gobs**](http://blog.golang.org/2011/03/gobs-of-data.html) which are going to be the successor of Google's Protocol Buffers. However in contrast to all the other mentioned formats those are not language-agnostic and rely on [Go's built-in reflection](http://golang.org/pkg/reflect/) there are also Gobs libraries for at least on other language than Go.


","// Please note that I'm author of MessagePack. This answer may be biased.


**Format design**


1. Compatibility with JSON


In spite of its name, BSON's compatibility with JSON is not so good compared with MessagePack.


BSON has special types like ""ObjectId"", ""Min key"", ""UUID"" or ""MD5"" (I think these types are required by MongoDB). These types are not compatible with JSON. That means some type information can be lost when you convert objects from BSON to JSON, but of course only when these special types are in the BSON source. It can be a disadvantage to use both JSON and BSON in single service.


MessagePack is designed to be transparently converted from/to JSON.
2. MessagePack is smaller than BSON


MessagePack's format is less verbose than BSON. As the result, MessagePack can serialize objects smaller than BSON.


For example, a simple map {""a"":1, ""b"":2} is serialized in 7 bytes with MessagePack, while BSON uses 19 bytes.
3. BSON supports in-place updating


With BSON, you can modify part of stored object without re-serializing the whole of the object. Let's suppose a map {""a"":1, ""b"":2} is stored in a file and you want to update the value of ""a"" from 1 to 2000.


With MessagePack, 1 uses only 1 byte but 2000 uses 3 bytes. So ""b"" must be moved backward by 2 bytes, while ""b"" is not modified.


With BSON, both 1 and 2000 use 5 bytes. Because of this verbosity, you don't have to move ""b"".
4. MessagePack has RPC


MessagePack, Protocol Buffers, Thrift and Avro support RPC. But BSON doesn't.


These differences imply that MessagePack is originally designed for network communication while BSON is designed for storages.


**Implementation and API design**


1. MessagePack has type-checking APIs (Java, C++ and D)


MessagePack supports static-typing.


Dynamic-typing used with JSON or BSON are useful for dynamic languages like Ruby, Python or JavaScript. But troublesome for static languages. You must write boring type-checking codes.


MessagePack provides type-checking API. It converts dynamically-typed objects into statically-typed objects. Here is a simple example (C++):



```
    #include <msgpack.hpp>

    class myclass {
    private:
        std::string str;
        std::vector<int> vec;
    public:
        // This macro enables this class to be serialized/deserialized
        MSGPACK_DEFINE(str, vec);
    };

    int main(void) {
        // serialize
        myclass m1 = ...;

        msgpack::sbuffer buffer;
        msgpack::pack(&buffer, m1);

        // deserialize
        msgpack::unpacked result;
        msgpack::unpack(&result, buffer.data(), buffer.size());

        // you get dynamically-typed object
        msgpack::object obj = result.get();

        // convert it to statically-typed object
        myclass m2 = obj.as<myclass>();
    }

```

2. MessagePack has IDL


It's related to the type-checking API, MessagePack supports IDL. (specification is available from: <http://wiki.msgpack.org/display/MSGPACK/Design+of+IDL>)


Protocol Buffers and Thrift require IDL (don't support dynamic-typing) and provide more mature IDL implementation.
3. MessagePack has streaming API (Ruby, Python, Java, C++, ...)


MessagePack supports streaming deserializers. This feature is useful for network communication. Here is an example (Ruby):



```
    require 'msgpack'

    # write objects to stdout
    $stdout.write [1,2,3].to_msgpack
    $stdout.write [1,2,3].to_msgpack

    # read objects from stdin using streaming deserializer
    unpacker = MessagePack::Unpacker.new($stdin)
    # use iterator
    unpacker.each {|obj|
      p obj
    }

```

"
"""send-mail: Applet not found"" when using mail command on alpine linux
I am using alpine linux running in a docker container to run a webapp. The webapp is supposed to send emails using the mail command from the command line. 


To support this, I have installed mail as follows:



```
apk add mailx

```

When I try to send a message I get the following error: 



```
bash-4.3# mail foo@bar.com
Subject: test
EOT
Null message body; hope that's ok
bash-4.3# send-mail: applet not found

```

Not sure what I am doing wrong. Any help is highly appreciated. 


","There is no default MTA available in Alpine. Besides mailx, you will also need to install and configure a MTA (Mail Transfer Agent), such as postfix:



```
# apk add postfix
(1/3) Installing db (5.3.28-r0)
(2/3) Installing libsasl (2.1.26-r8)
(3/3) Installing postfix (3.1.3-r0)
Executing postfix-3.1.3-r0.pre-install
Executing busybox-1.25.1-r0.trigger
OK: 8 MiB in 16 packages

```

Start postfix: 



```
~ # postfix start
postfix/postfix-script: warning: not owned by root: /var/spool/postfix/.
postfix/postfix-script: warning: not owned by root: /var/spool/postfix/pid
postfix/postfix-script: starting the Postfix mail system

```

Send the test message: 



```
~ # mail -s ""Test"" test@boo.com
.
EOT
Null message body; hope that's ok

```

"
"AngularJS $scope.model is undefined
AngularJS `$scope.model` is undefined when I initialize it via Javascript



```
<button ngf-select=""upload($file)"">Выбрать</button>
                <span id=""photoSpan""></span>
                <input type=""text""  id=""photo"" name=""uploadPhoto"" ng-model=""uploadPhoto"">

```

Here is my initialization code:



```
document.getElementById(""photo"").value = fileName;
                $(""#photoSpan"").text(resp.config.data.file.name);
                console.log('Success ' + resp.config.data.file.name + 'uploaded. Response: ' + resp.data);

```

So



```
createExpositionResource.photo = $scope.uploadPhoto;
        console.log('--------------'+$scope.uploadPhoto);

```

returns `undefined`.


What am I doing wrong?


Thank you!


","
> 
> What am I doing wrong?
> 
> 
> 


Using Jquery: Have a look at this question: [""Thinking in AngularJS"" if I have a jQuery background?](https://stackoverflow.com/questions/14994391/thinking-in-angularjs-if-i-have-a-jquery-background?rq=1)


You should replace



```
document.getElementById(""photo"").value = fileName;

```

by



```
$scope.uploadPhoto = fileName;

```

Short explanation : If you update an angular variable (eg : `document.getElementById(""photo"").value = fileName;`), angular won't know that you did it. There is worse with your code. For Angular, what's in `$scope` is his ""trusted data source"". Which means that Angular could also erase the value you put by using Jquery.


"
"iphone: delete row from sqlite database
I have tried many ways to achieve this biut unable to delete row from sqlite, please help me to correct following piece of code



```
dbPath = [self applicationDocumentsDirectory];
NSLog(@""%@"", dbPath);
dbPath=[dbPath stringByAppendingPathComponent:@""database""];
dbPath=[dbPath stringByAppendingPathComponent:@""OFFENDERSDB.sqlite""];

NSLog(@""database path --> %@"", dbPath);

if (sqlite3_open([dbPath UTF8String], &database) == SQLITE_OK) {
    const char *sqlStatement = (""DELETE * from OFFENDERS_LIST where ID=%d "",2);

    sqlite3_stmt *compiledStatement;

    if (sqlite3_prepare(database, sqlStatement, -1, &compiledStatement, NULL) == SQLITE_OK) {

        NSLog(@""offender deleted"");
    }
    sqlite3_finalize(compiledStatement);
}

```

Thanx in advance


","have you tried using:



```
sqlite3_exec(database, sqlStatement...

```

with your DELETE statement?


**ALSO.... it's DELETE FROM.... not DELETE \* FROM....**


Snagged the following code from another place out on the internet for a more complete example...



```
sqlite3 *db;
int rc;

rc = sqlite3_open( ""C:\\MyDatabase"", &db );

if ( rc )
{
        sqlite3_close(db);
}
else //Database connection opened successfuly
{
        char *zErrMsg = 0;

        rc = sqlite3_exec( db, ""DELETE FROM yourTable"", NULL, NULL, &zErrMsg );

        if( rc != SQLITE_OK )
        {
                sqlite3_free( zErrMsg );
        }

        sqlite3_close(db);
}

```

"
"Summing the prices from Transaction files
I need to cut down the run time of this query. Currently it's taking 45 minutes. Is there something I can change in the table or the query to allow this to run faster?



```
SELECT SUM(B.PRICE),C.ST_FILE
FROM
    TAD1D.ST_EUCM AS B,
    TAD1D.ST_DSA_TRAN  AS C,
    TAD1A.TIBHLD T
WHERE C.ST_TRAN_FILE=B.ST_TRAN_FILE
    AND C.ST_IND='Y' 
    AND C.ST_SRC_SYS_CD='HANN'
    AND C.ST_TBL_NAME='ST_EUCM'
    AND B.PARTNUMB=T.ACCT_NUM
    AND B.FVSRVCCODE=T.SRV_CD
GROUP BY C.ST_FILE
EXCEPT
SELECT sum(A.ACCUM_C),D.ST_FILE from
                    TAD1A.TIBHLD_DLY_VAL A,
                    TAD1D.ST_DSA_TRAN  AS D
            where A.SRC_SYS_CD = 'HANN'
                    AND D.ST_SRC_SYS_CD=A.SRC_SYS_CD
                    AND D.ST_TBL_NAME='ST_EUCM'
                    AND DATE(A.HLD_VAL_DT)=DATE(D.ST_FILE)
                    GROUP BY D.ST_FILE
                    WITH UR --

```

There are no indexes on either table. Please provide a tip/solution in the event i am able to change the table structure or add an index (I may not).


","**Nitpicks**


There is a slight amount of inconsistency in your capitalization of keywords. `SUM()` / `sum()` and `WHERE` / `where`. Otherwise, it's pretty consistent throughout!




---


# Aliases


You use the following aliases: `B` `C` `T`. There is no meaning or explanation of what those mean. You should use meaningful aliases, especially where your table names don't really help, such as in your case.




---


# Old-style `JOIN`


This:



```
FROM
    TAD1D.ST_EUCM AS B,
    TAD1D.ST_DSA_TRAN  AS C,
    TAD1A.TIBHLD T
WHERE C.ST_TRAN_FILE=B.ST_TRAN_FILE
    AND C.ST_IND='Y' 
    AND C.ST_SRC_SYS_CD='HANN'
    AND C.ST_TBL_NAME='ST_EUCM'
    AND B.PARTNUMB=T.ACCT_NUM
    AND B.FVSRVCCODE=T.SRV_CD

```

...is [deprecated pre-ANSI-92 syntax](http://www.straightpathsql.com/archives/2012/11/sql-server-join-syntax-its/). It should not be used, rather you should favor explicit `JOIN` syntax, like such:



```
FROM
    TAD1D.ST_EUCM AS B
INNER JOIN 
    TAD1D.ST_DSA_TRAN  AS C
    ON C.ST_TRAN_FILE = B.ST_TRAN_FILE
INNER JOIN 
    TAD1A.TIBHLD T
    ON B.PARTNUMB = T.ACCT_NUM
    AND B.FVSRVCCODE = T.SRV_CD
WHERE
    AND C.ST_IND = 'Y' 
    AND C.ST_SRC_SYS_CD = 'HANN'
    AND C.ST_TBL_NAME = 'ST_EUCM'

```

Notice I added spaces around your `=` operators as it makes the code easier to read. You would also want to use similar `JOIN` syntax in your subquery.




---


# Indexes


You stated:



> 
> There are no indexes on either table. Please provide a tip/solution in the event i am able to change the table structure or add an index (I may not).
> 
> 
> 


And to that, I say resoundingly: ***YES!*** Add indexes!


But make sure even before you do that, look carefully at your query execution plan and look at where the most expensive steps are. That should give you a clue as to what is happening. I suspect there are a few nasty nested loops in there that are bogging it down, and depending on your RDBMS there are some optimizations available, and there may be ways to change the execution manually.


"
"Remove trailing zeros from a string
I would like to remove trailing zero's according to the following rules:


- The input number will always have 4 decimal places
- The output number should always have between 2 and 4 decimals
- If the last two digits after the decimal is a zero, then remove both of them
- If the last digit after the decimal is a zero, then remove it


Examples:


- 1.0000 -> 1.00
- 1.4000 -> 1.40
- 1.4100 -> 1.41
- 1.4130 -> 1.413
- 1.4136 -> 1.4136


I have considered the following:



```
if (substr($number, -2) == '00') return substr($number, 0, 4);
if (substr($number, -1) == '0') return substr($number, 0, 5);
return $number;

```

Are there any better ways of doing this?


","I think this should work:



```
return preg_replace('/0{1,2}$/', '', $number);

$strings = array ('1.4000', '1.4100', '1.4130', '1.4136', '1.4001', '1.0041');
foreach ($strings as $number) {
  echo ""$number -> "" . preg_replace('/0{0,2}$/', '', $number) . ""\n"";
}

```

Produces:



```
1.4000 -> 1.40
1.4100 -> 1.41
1.4130 -> 1.413
1.4136 -> 1.4136
1.4001 -> 1.4001
1.0041 -> 1.0041

```

"
"Sharing of footnote between different part of tables using flextable
I need to create table with same footnote being placed in both header and body of the table, I cannot figure out how to make it happen using `flextable`, what I can create is something as below:



```
library(flextable)
library(dplyr)
library(tidyr)

data(iris)
iris %>%
  as_tibble %>%
  gather(.,key = variable,value = value,-Species) %>%
  group_by(Species,variable) %>%
  summarise(value=formatC(mean(value),digits = 2,format = 'f')) %>%
  ungroup %>%
  spread(.,key = variable,value = value) %>%
  flextable %>%
  footnote(.,part = 'header',i = 1,j = c(2:5),
           value = as_paragraph(c('Rounded to two decimal places')),
           ref_symbols = c('*'),
           inline=FALSE) %>%
  footnote(.,part = 'body',i = c(1:3),j = 1,
           value = as_paragraph(c('Rounded to two decimal places')),
           ref_symbols = c('*'),
           inline=FALSE)

```

Currently I created two footnotes with the same statement for header and body, I wonder if I can merge the two statements into one.


Thanks!


","(I did not imagine footnotes would be repeated when this function has been implemented but) by using `merge_v`, you can merge them if identical:



```
library(flextable)
library(dplyr)
library(tidyr)

data(iris)
iris %>%
  as_tibble %>%
  gather(.,key = variable,value = value,-Species) %>%
  group_by(Species,variable) %>%
  summarise(value=formatC(mean(value),digits = 2,format = 'f')) %>%
  ungroup %>%
  spread(.,key = variable,value = value) %>%
  flextable %>%
  footnote(.,part = 'header',i = 1,j = c(2:5),
           value = as_paragraph(c('Rounded to two decimal places')),
           ref_symbols = c('*'),
           inline=FALSE) %>%
  footnote(.,part = 'body',i = c(1:3),j = 1,
           value = as_paragraph(c('Rounded to two decimal places')),
           ref_symbols = c('*'),
           inline=FALSE) %>% 
  merge_v(part = ""footer"")


```

[![enter image description here](https://i.stack.imgur.com/6MUIz.png)](https://i.stack.imgur.com/6MUIz.png)


"
"Running Ethernet over two phone lines
I am trying to run Ethernet through my house along existing phone wires. I had originally planned to use the phone wires to pull the cat5e cable, but it appears the phone lines are stapled to studs somewhere. From questions like [this one](https://superuser.com/questions/153281/convert-home-phone-wiring-to-ethernet), I know I can use the cable to get 100M at best because there are only 3 pairs, and gigabit requires 4 pairs. 


However, I am fortunate enough to have two phone lines that run parallel that I can use. This gives me a total of 6 pairs. I know the phone wires still have higher crosstalk, but would combining the two help? Could I get gigabit speeds? How should I match the phone line pairs to the Ethernet pairs?


","Your existing voice grade phone wiring is probably no better than Cat3, if even that. So you'd probably struggle to get 10BASE-T to work, and 100 mbps and Gigabit Ethernet is probably out of the question.


10BASE-T (10mbps) was the last widely-available flavor of Ethernet that was designed to work over Cat3 voice grade wiring. So assuming your existing phone wiring is even Cat3 (note: untwisted telephone line cords are not), and if you have two pair, you could possibly use them for 10mbps Ethernet.


What's commonly called ""100BASE-T"" is actually ""100BASE-TX"", which requires two pair of **Cat5**. There were two competing IEEE standards for 100 mbps Ethernet over unshielded twisted pair (UTP) wiring, but they were never widely deployed: 100BASE-T4 (required 4 pairs of Cat3) and 100BASE-T2 (required 2 pairs of Cat3). I don't think you can find any 100BASE-T4 or 100BASE-T2 NICs or hubs/switches anymore, because you could barely find them when they were new in the 1990's.


Gigabit Ethernet (1000BASE-T) requires 4 pairs of Cat5.


The IEEE stopped caring about voice grade wiring after 100BASE-TX was a huge market success while 100BASE-T{2,4} flopped.


I suppose it's possible that given a short enough run of Cat3 that you could get lucky and happen to get 100BASE-TX or 1000BASE-T to work, but it seems unlikely to me. I think there's a decent chance you could make 10BASE-T work, but even that isn't guaranteed without knowing for sure that you have good Cat3 end-to-end, and you're wiring it like Ethernet (point to point) and not like home phone wiring (which often has branches or multiple jacks on a long passive bus).


"
"How to animate border transition expand with CSS?
I'm trying to create a CSS transition where the `border-bottom` of `a href` element expands on `:hover` to the width of the link.


What I could find is a CSS solution where the `background width` is animated:
<http://jsfiddle.net/mfn5nefb/82/>


But that's not what I want, because after click on a navigation tab I want to leave the border as is. So I'd have to directly animate the border, instead of the background.



```
<!-- assume the border-bottom is only applied to an active=clicked navigation tab -->
<h1 style=""border-bottom: 3px solid green;"">CSS IS AWESOME</h1>

h1 {
    color: #666;
    position: relative;
    display: inline-block;
}

h1:after {
    position: absolute;
    left: 50%;
    content: '';
    height: 40px;
    height: 5px;
    background: #f00;
    transition: all 0.5s linear;
    width: 0;
    bottom: 0;  
}

h1:hover:after {
    width: 270px;
    margin-left: -135px;
}

```

Here you see the ""active"" link gets a green border. And on hover I'd like to animate the other tabs, but the border itself. Currently the background is animated, which appears *above* the border and thus looks misaligned.


","You can still achieve this by using a pseudo-element (with `background`) and expand it on `hover`. All that is required is to set the value for the `bottom` property as the inverse of expected `border-width` and also set the `height` of the pseudo-element to be the same as the `border-width`.





```
h1 {
  color: #666;
  position: relative;
  display: inline-block;
}
h1:after {
  position: absolute;
  left: 50%;
  content: '';
  height: 3px;
  background: #f00;
  transition: all 0.5s linear;
  width: 0;
  bottom: -3px;
}
h1:hover:after {
  width: 100%;
  left: 0px;
}
```


```
<!-- assume the border-bottom is only applied to an active=clicked navigation tab -->
<h1 style=""border-bottom: 3px solid green;"">Tab1</h1>
<h1>Tab2</h1>
```








---


Another way to achieve the same effect using the `border` property itself would be to use `transform: scale` like in the below snippet. Here the `scaleX(0)` makes the original width of the element as 0 and on `hover` it is transitioned to full width using `scaleX(1)`. Since, default `transform-origin` is at `50%` in X-axis, the border looks as though it is expanding from the center.





```
h1 {
  color: #666;
  position: relative;
  display: inline-block;
}
h1:after {
  position: absolute;
  left: 0%;
  top: 0%;
  content: '';
  height: 100%;
  transition: all 0.5s linear;
  width: 100%;
  border-bottom: 3px solid red;
  transform: scaleX(0);
}
h1:hover:after {
  transform: scale(1);
}
```


```
<!-- assume the border-bottom is only applied to an active=clicked navigation tab -->
<h1 style=""border-bottom: 3px solid green;"">Tab1</h1>
<h1>Tab2</h1>
```






"
"Calling CICS program from JCL batch
Is it possible to call a CICS program from JCL batch?


If yes how? If no why not?


","Yes. Technically, a COBOL batch program is communicating with CICS, and you execute the COBOL batch program from JCL. You can read more in the CICS Transaction Server manual.



> 
> The external CICS® interface is an application programming interface that enables a non-CICS program (a client program) running in MVS™ to call a program (a server program) running in a CICS region and to pass and receive data by means of a communications area. The CICS application program is invoked as if linked-to by another CICS application program.
> 
> 
> 


[Introduction to the external CICS interface](http://publib.boulder.ibm.com/infocenter/cicsts/v2r3/index.jsp?topic=/com.ibm.cics.ts23.doc/dfhtm/dfhtm4a.htm)


I believe that your CICS program cannot use maps (must be conversational), but it's been a while since I've worked with CICS, so I might be mistaken.


"
"Snap application doesn't see files from another partition
I have installed KeePassXC snap package on Ubuntu 17.10. I have multiple linux systems installed on the same drive, so I have made most of my home folders like ~/Documents created as links to a separate ext4 partition (users have same uid so permissions are correct). I tried to open my passwords' database which I have inside Documents, but keepassxc cannot see any files inside Documents. If I put the file in my home folder then it sees it.


Are snap packages able to find files outside from my home folder? Any way to make KeePassXC (and the rest snap packages) access the files at my separate ext4 partition (maybe some configuration somewhere)?


","Snaps are typically confined. By default snaps can access nothing but their own writable areas, but they can gain permissions by way of interfaces. There are two interfaces that provide the ability to access files beyond the default: the `home` interface, and the `removable-media` interface. I assume KeePassXC uses the `home` interface, which is how you can access `~/Documents`. Perhaps it even uses the `removable-media` interface, but that only covers stuff mounted into `/media` or `/mnt`. If your other partitions are mounted elsewhere, the snap will not have permission to follow the symlinks.


The ways to make this work, ordered by difficulty:


1. Mount the partitions into your home directory rather than symlinking them.
2. Mount the partitions into `/media` and ensure KeePassXC uses the `removable-media` interface. This may be an upstream change, but a small one.
3. Change KeePassXC into a classic snap, thereby making it no longer strictly confined. This is a larger change and upstream may not be interested in it.


**Update:** The KeyPassXC `snapcraft.yaml` is [here](https://github.com/keepassxreboot/keepassxc/blob/develop/snapcraft.yaml). Looks like it already has the `removable-media` interface. All you need to do to follow (2) is to connect it (it's not automatically connected):



```
$ sudo snap connect keepassxc:removable-media

```

"
"How to distinguish between snap and normal application in ""Activities""
If I have, say, Firefox installed both as a normal package and as a snap, they will appear in the *Activities* application listing and also two items will be returned by a search for ""Firefox"". Both show the same name and icon. How do I know which one to choose?


I know I can easily do it from the command line. But I would like to know what is the intended way from the Ubuntu GUI. Or, if that is a not yet implemented feature, if there is a GUI approach to this.


","You may create two new custom `.desktop` launchers for each in `~/.local/share/applications/`.


First create a file called `firefox.desktop` (for the default version) in `~/.local/share/applications/`. Open it with a text editor and put the following lines and save the file.



```
[Desktop Entry]
Comment=default Firefox browser
Terminal=false
Name=Firefox (default)
Exec=firefox %u
Type=Application
Icon=/path/to/icon/for/default/firefox
NoDisplay=false

```

Put the full path to the image (png or svg are preferable) you want to use as default Firefox's icon in place of `/path/to/icon/for/default/firefox` or simply put `Icon=firefox` if you want the default Firefox icon as per your icon theme.


Alternatively, copy the `firefox.desktop` file from `/usr/share/applications/` to `~/.local/share/applications/` and make the changes to `Name=` and `Icon=` fields as you wish.


Similarly, create another `.desktop` file, say `firefox-snap.desktop` (for the snap version) in `~/.local/share/applications/`. Open it with a text editor and put the following lines and save the file.



```
[Desktop Entry]
Comment=snappy Firefox browser
Terminal=false
Name=Firefox (snap)
Exec=/snap/bin/firefox
Type=Application
Icon=/path/to/icon/for/snap/firefox
NoDisplay=false

```

Like before put the path to image file you want to use as the icon of snappy Firefox.


Now in Activities overview you'll see two Firefox launchers titled ""*Firefox (default)*"" and ""*Firefox (snap)*"". They will also have different icons if you have used different icons in their associated `.desktop` files.


**P.S.** You may use a GUI application like `alacarte` (aka ""Main Menu"") to create the launchers instead of manually creating and editing the `.desktop` files.


"
"Can't list some folders
I'm getting an ""Access denied"" exception when trying to list some folders on my PC, such as 'c:\Documents and Settings'. My application is running from an administrator account and with elevated privileges (`<requestedExecutionLevel level=""requireAdministrator"" uiAccess=""false"" />`) 


What else I need to do to access such folders?


","Running as administrator is necessary, but not enough, you also need to enable the [Backup](https://msdn.microsoft.com/en-us/library/windows/desktop/bb530716.aspx) privilege, as implicitely documented here: [Junction Points](https://msdn.microsoft.com/en-us/library/windows/desktop/bb968829.aspx) 



> 
> %SystemDrive%\Documents and Settings directory is now stored in the %SystemDrive%\Users directory
>  [...] Backup applications must be capable of backing up and restoring
>  junction points.
> 
> 
> 


So here is an example that works:



```
using (Privilege p = new Privilege(Privilege.Backup))
{
    foreach (string path in Directory.GetFileSystemEntries(@""c:\documents and settings""))
    {
        Console.WriteLine(path);
    }
}

```

Note there is no public .NET class to enable privilege programmatically to my knowledge, but there is one internal in *System.Security.AccessControl* (mscorlib) that I have exposed using a thin wrapper. It's based on the fully public Windows [AdjustTokenPrivileges function](https://msdn.microsoft.com/en-us/library/windows/desktop/aa375202.aspx).



```
/// <summary>
/// Defines a Windows privilege.
/// </summary>
public sealed class Privilege : IDisposable
{
    private static Type _privilegeType;
    private object _privilege;

    static Privilege()
    {
        _privilegeType = typeof(string).Assembly.GetType(""System.Security.AccessControl.Privilege"", false); // mscorlib
    }

    /// <summary>
    /// Initializes a new instance of the <see cref=""Privilege""/> class and enable the privilege.
    /// </summary>
    /// <param name=""name"">The privilege name.</param>
    public Privilege(string name)
        : this(name, true)
    {
    }

    /// <summary>
    /// Initializes a new instance of the <see cref=""Privilege"" /> class.
    /// </summary>
    /// <param name=""name"">The privilege name.</param>
    /// <param name=""enable"">if set to <c>true</c> the privilege is enabled.</param>
    public Privilege(string name, bool enable)
    {
        if (name == null)
            throw new ArgumentNullException(""name"");

        if (_privilegeType == null)
            throw new NotSupportedException();

        _privilege = _privilegeType.GetConstructors()[0].Invoke(new object[] { name });
        if (enable)
        {
            Enable();
        }
    }

    /// <summary>
    /// Disable this privilege from the current thread. 
    /// </summary>
    public void Revert()
    {
        _privilegeType.InvokeMember(""Revert"", BindingFlags.Public | BindingFlags.Instance | BindingFlags.InvokeMethod, null, _privilege, null);
    }

    /// <summary>
    /// Gets a value indicating whether Revert must be called.
    /// </summary>
    /// <value>
    ///   <c>true</c> if Revert must be called; otherwise, <c>false</c>.
    /// </value>
    public bool NeedToRevert
    {
        get
        {
            return (bool)_privilegeType.InvokeMember(""NeedToRevert"", BindingFlags.Public | BindingFlags.Instance | BindingFlags.GetProperty, null, _privilege, null);
        }
    }

    /// <summary>
    /// Enables this privilege to the current thread. 
    /// </summary>
    public void Enable()
    {
        _privilegeType.InvokeMember(""Enable"", BindingFlags.Public | BindingFlags.Instance | BindingFlags.InvokeMethod, null, _privilege, null);
    }

    /// <summary>
    /// Performs application-defined tasks associated with freeing, releasing, or resetting unmanaged resources.
    /// </summary>
    public void Dispose()
    {
        if (NeedToRevert)
        {
            Revert();
        }
    }

    /// <summary>
    /// The SE_ASSIGNPRIMARYTOKEN_NAME privilege.
    /// </summary>
    public const string AssignPrimaryToken = ""SeAssignPrimaryTokenPrivilege"";

    /// <summary>
    /// The SE_AUDIT_NAME privilege.
    /// </summary>
    public const string Audit = ""SeAuditPrivilege"";

    /// <summary>
    /// The SE_BACKUP_NAME privilege.
    /// </summary>
    public const string Backup = ""SeBackupPrivilege"";

    /// <summary>
    /// The SE_CHANGE_NOTIFY_NAME privilege.
    /// </summary>
    public const string ChangeNotify = ""SeChangeNotifyPrivilege"";

    /// <summary>
    /// The SE_CREATE_GLOBAL_NAME privilege.
    /// </summary>
    public const string CreateGlobal = ""SeCreateGlobalPrivilege"";

    /// <summary>
    /// The SE_CREATE_PAGEFILE_NAME privilege.
    /// </summary>
    public const string CreatePageFile = ""SeCreatePagefilePrivilege"";

    /// <summary>
    /// The SE_CREATE_PERMANENT_NAME privilege.
    /// </summary>
    public const string CreatePermanent = ""SeCreatePermanentPrivilege"";

    /// <summary>
    /// The SE_CREATE_SYMBOLIC_LINK_NAME privilege.
    /// </summary>
    public const string CreateSymbolicLink = ""SeCreateSymbolicLinkPrivilege"";

    /// <summary>
    /// The SE_CREATE_TOKEN_NAME privilege.
    /// </summary>
    public const string CreateToken = ""SeCreateTokenPrivilege"";

    /// <summary>
    /// The SE_DEBUG_NAME privilege.
    /// </summary>
    public const string Debug = ""SeDebugPrivilege"";

    /// <summary>
    /// The SE_ENABLE_DELEGATION_NAME privilege.
    /// </summary>
    public const string EnableDelegation = ""SeEnableDelegationPrivilege"";

    /// <summary>
    /// The SE_IMPERSONATE_NAME privilege.
    /// </summary>
    public const string Impersonate = ""SeImpersonatePrivilege"";

    /// <summary>
    /// The SE_INC_BASE_PRIORITY_NAME privilege.
    /// </summary>
    public const string IncreaseBasePriority = ""SeIncreaseBasePriorityPrivilege"";

    /// <summary>
    /// The SE_INCREASE_QUOTA_NAME privilege.
    /// </summary>
    public const string IncreaseQuota = ""SeIncreaseQuotaPrivilege"";

    /// <summary>
    /// The SE_INC_WORKING_SET_NAME privilege.
    /// </summary>
    public const string IncreaseWorkingSet = ""SeIncreaseWorkingSetPrivilege"";

    /// <summary>
    /// The SE_LOAD_DRIVER_NAME privilege.
    /// </summary>
    public const string LoadDriver = ""SeLoadDriverPrivilege"";

    /// <summary>
    /// The SE_LOCK_MEMORY_NAME privilege.
    /// </summary>
    public const string LockMemory = ""SeLockMemoryPrivilege"";

    /// <summary>
    /// The SE_MACHINE_ACCOUNT_NAME privilege.
    /// </summary>
    public const string MachineAccount = ""SeMachineAccountPrivilege"";

    /// <summary>
    /// The SE_MANAGE_VOLUME_NAME privilege.
    /// </summary>
    public const string ManageVolume = ""SeManageVolumePrivilege"";

    /// <summary>
    /// The SE_PROF_SINGLE_PROCESS_NAME privilege.
    /// </summary>
    public const string ProfileSingleProcess = ""SeProfileSingleProcessPrivilege"";

    /// <summary>
    /// The SE_RELABEL_NAME privilege.
    /// </summary>
    public const string Relabel = ""SeRelabelPrivilege"";

    /// <summary>
    /// The SE_REMOTE_SHUTDOWN_NAME privilege.
    /// </summary>
    public const string RemoteShutdown = ""SeRemoteShutdownPrivilege"";

    ///// <summary>
    ///// The SE_RESERVE_PROCESSOR_NAME privilege.
    ///// </summary>
    //public const string ReserveProcessor = ""SeReserveProcessorPrivilege"";

    /// <summary>
    /// The SE_RESTORE_NAME privilege.
    /// </summary>
    public const string Restore = ""SeRestorePrivilege"";

    /// <summary>
    /// The SE_SECURITY_NAME privilege.
    /// </summary>
    public const string Security = ""SeSecurityPrivilege"";

    /// <summary>
    /// The SE_SHUTDOWN_NAME privilege.
    /// </summary>
    public const string Shutdown = ""SeShutdownPrivilege"";

    /// <summary>
    /// The SE_SYNC_AGENT_NAME privilege.
    /// </summary>
    public const string SyncAgent = ""SeSyncAgentPrivilege"";

    /// <summary>
    /// The SE_SYSTEM_ENVIRONMENT_NAME privilege.
    /// </summary>
    public const string SystemEnvironment = ""SeSystemEnvironmentPrivilege"";

    /// <summary>
    /// The SE_SYSTEM_PROFILE_NAME privilege.
    /// </summary>
    public const string SystemProfile = ""SeSystemProfilePrivilege"";

    /// <summary>
    /// The SE_SYSTEMTIME_NAME privilege.
    /// </summary>
    public const string SystemTime = ""SeSystemtimePrivilege"";

    /// <summary>
    /// The SE_TAKE_OWNERSHIP_NAME privilege.
    /// </summary>
    public const string TakeOwnership = ""SeTakeOwnershipPrivilege"";

    /// <summary>
    /// The SE_TCB_NAME privilege.
    /// </summary>
    public const string TrustedComputingBase = ""SeTcbPrivilege"";

    /// <summary>
    /// The SE_TIME_ZONE_NAME privilege.
    /// </summary>
    public const string TimeZone = ""SeTimeZonePrivilege"";

    /// <summary>
    /// The SE_TRUSTED_CREDMAN_ACCESS_NAME privilege.
    /// </summary>
    public const string TrustedCredentialManagerAccess = ""SeTrustedCredManAccessPrivilege"";

    /// <summary>
    /// The SE_UNDOCK_NAME privilege.
    /// </summary>
    public const string Undock = ""SeUndockPrivilege"";

    /// <summary>
    /// The SE_UNSOLICITED_INPUT_NAME privilege.
    /// </summary>
    public const string UnsolicitedInput = ""SeUnsolicitedInputPrivilege"";
}

```

"
"Why Jenkins says ""Server rejected the 1 private key(s)"" while launching the agent?
I am successfully able to connect to remote machine using SSH but when I am launching the agent from Jenkins it throws the following error:



```
ERROR: Server rejected the 1 private key(s) for user1 (credentialId:xxxxx-xxxx-xxxx-xxxx-xxxxxxxxxxxx/method:publickey)
[01/19/17 05:35:15] [SSH] Authentication failed.
hudson.AbortException: Authentication failed.
    at hudson.plugins.sshslaves.SSHLauncher.openConnection(SSHLauncher.java:1219)
    at hudson.plugins.sshslaves.SSHLauncher$2.call(SSHLauncher.java:714)
    at hudson.plugins.sshslaves.SSHLauncher$2.call(SSHLauncher.java:709)
    at java.util.concurrent.FutureTask.run(FutureTask.java:266)
    at java.util.concurrent.ThreadPoolExecutor.runWorker(ThreadPoolExecutor.java:1142)
    at java.util.concurrent.ThreadPoolExecutor$Worker.run(ThreadPoolExecutor.java:617)
    at java.lang.Thread.run(Thread.java:745)
[01/19/17 05:35:15] Launch failed - cleaning up connection
[01/19/17 05:35:15] [SSH] Connection closed.

```

I can establish SSH connection from master machine to the node machine using `user1`, however when I am trying to launch the agent using `user1` from jenkins it is rejecting the private key. Is there any solution to overcome this issue?


","I solve this issue following below steps:


**From the target slave node's console**


1. Switch to the `root` user:



```
sudo su

```

2. Add a jenkins user with the home `/var/lib/jenkins` (Note: I am keeping my home directory in `/var/lib/jenkins`):



```
useradd -d /var/lib/jenkins jenkins

```

**From the Jenkins Master** 


Copy the `/var/lib/jenkins/.ssh/id_rsa.pub` key from the **Jenkins** user on the master 


**From the target slave node's console**


1. Create an authorized\_keys file for the Jenkins user



```
mkdir /var/lib/jenkins/.ssh
touch /var/lib/jenkins/.ssh/authorized_keys

```

2. Paste the key from the Jenkins master into the file vim. Save with `:wq!`
3. Make sure the files have correct owner and permission.



```
chown -R jenkins /var/lib/jenkins/.ssh
chmod 600 /var/lib/jenkins/.ssh/authorized_keys
chmod 700 /var/lib/jenkins/.ssh

```

"
"HTML rowspan 1,5 (one and a half)
Let assume I have following table:



```
<table border=""2"">
<tr>
            <td>abcsss</td>
            <td>sdf</td>
            <td>def</td>
            <td>def</td>
            <td rowspan=""1,5"">defsssss</td>
</tr>
        <tr>
            <td>abcsss</td>
            <td>sdf</td>
            <td>def</td>
            <td>def</td>
</tr>
    <tr>
            <td>abcsss</td>
            <td>sdf</td>
            <td>def</td>
            <td>def</td>
            <td rowspan=""1,5"">defsssss</td>
</tr>
</table>

```

I know that code rowspan=""1,5"" isn't working but how would I make it work? Is there some css or html trick to make field like there were rowspan=""1,5"" working?


","
```
<tr>
            <td rowspan=""2"">abcsss</td>
            <td rowspan=""2"">sdf</td>
            <td rowspan=""2"">def</td>
            <td rowspan=""2"">def</td>
            <td rowspan=""3"">defsssss</td>
</tr>

```




```
<table border=""2"">
  <tr>
    <td rowspan=""2"">abcsss</td>
    <td rowspan=""2"">sdf</td>
    <td rowspan=""2"">def</td>
    <td rowspan=""2"">def</td>
    <td rowspan=""3"">defsssss</td>
  </tr>
  <tr></tr>
  <tr>
    <td rowspan=""2"">abcsss</td>
    <td rowspan=""2"">sdf</td>
    <td rowspan=""2"">def</td>
    <td rowspan=""2"">def</td>
  </tr>
  <tr>
    <td rowspan=""3"">defsssss</td>
  </tr>

  <tr>
    <td rowspan=""2"">abcsss</td>
    <td rowspan=""2"">sdf</td>
    <td rowspan=""2"">def</td>
    <td rowspan=""2"">def</td>

  </tr>
</table>
```






"
"Why C++11 implicitly deletes my default constructor?
I wanted to make my class uncopyable using C++11, before I privately inherited my class from `boost::noncopyable`. In C++11 I implemented following:



```
class Foo
{
public:
    Foo(const Foo&) = delete;
    Foo& operator=(const Foo&) = delete;
};

```

With this change compilation of a client code (with VS 2013) gives me following error:



```
..\main.cpp(9): error C2512: 'Foo' : no appropriate default constructor available

```

My client code is pretty simple:



```
int main()
{
    Foo foo;
}

```

Is there any C++11 rule which implicitly *deletes default constructor* in my case ?


","The default constructor for `Foo` is not deleted, it is simply not defined. An explicitly defaulted/deleted constructor is a *user-declared* constructor, and this causes an implicit declaration of a default constructor to be inhibited. `Foo` has a *user-declared* constructor, the explicitly deleted copy constructor.


From N3337, *§12.1/5* **[class.ctor]**



> 
> A *default* constructor for a class `X` is a constructor of class `X` that can be called without an argument. **If there is no user-declared constructor for class X, a constructor having no parameters is implicitly declared as defaulted** (8.4). `...`
> 
> 
> 




---


Note that if the default constructor was indeed deleted, the error message would probably have said as much. For instance, if I change the definition of `Foo` to 



```
class Foo
{
public:
    int &i;
};

```

the presence of the data member of reference type implicitly deletes the default constructor. Now, [clang and gcc produce the error messages](http://coliru.stacked-crooked.com/a/eaa1af70f80dfbdc)



> 
> error: call to implicitly-deleted default constructor of 'Foo'
> 
> 
> 


and 



> 
> error: use of deleted function 'Foo::Foo()'
> 
> 
> 


Compare these to the error messages produced by your [original example](http://coliru.stacked-crooked.com/a/a0698e5c9b16db9a)



> 
> error: no matching constructor for initialization of 'Foo'  
> 
>  error: no matching function for call to 'Foo::Foo()'
> 
> 
> 




---


To fix your example, you'll need to explicitly default the default constructor



```
Foo() = default;

```

"
"Cannot delete rows from a temporal history table
I've recently discovered temporal tables in SQL Server. I'd like to start using this functionality. However the biggest hurdle is not being able to delete records from it. Due to GDPR compliance this is an absolute must.


Deleting records from a history table obviously leads to the error: 



> 
> Cannot delete rows from a temporal history table
> 
> 
> 


So to be able to delete records from a history table I have to disable `SYSTEM_VERSIONING`, then delete, and then re-enable `SYSTEM_VERSIONING`. Unless there's another way I'm not aware of?


Since it's not possible to use GO's in a stored procedure/`SqlCommand`, how can I ensure deleting a history record does not mess with other transactions e.g. updates sent to the temporal table during deleting records from the history table will still result in records being added to the history table? 


I've tried creating a stored procedure to wrap it in one transaction but this fails because the `ALTER TABLE` statement disabling the `SYSTEM_VERSIONING` is not executed yet leading to the same error.



```
CREATE PROCEDURE [dbo].[OrderHistoryDelete]
     (@Id UNIQUEIDENTIFIER)
AS
BEGIN
    BEGIN TRANSACTION

    ALTER TABLE [dbo].[Order] SET ( SYSTEM_VERSIONING = OFF )
    -- No GO possible here obviously.

    DELETE FROM [dbo].[OrderHistory] WITH (TABLOCKX) 
    WHERE [Id] = @Id

    ALTER TABLE [dbo].[Order] SET ( SYSTEM_VERSIONING = ON (HISTORY_TABLE = [dbo].[OrderHistory]))

    COMMIT TRANSACTION
END
GO

```

","If you make the `DELETE` dynamic, your stored procedure will successfully `ALTER` the table, `DELETE` the records in question, and then `ALTER` it back.



```
CREATE PROCEDURE [dbo].[OrderHistoryDelete]
     (@Id UNIQUEIDENTIFIER)
AS
BEGIN
    DECLARE @sql VARCHAR(MAX)

    BEGIN TRANSACTION

        ALTER TABLE [dbo].[Order] SET ( SYSTEM_VERSIONING = OFF )

        SET @sql = 'DELETE FROM [dbo].[OrderHistory] WITH (TABLOCKX) 
        WHERE [Id] = ''' + CAST(@Id AS VARCHAR(40)) + ''''
        EXEC (@sql)

        ALTER TABLE [dbo].[Order] SET ( SYSTEM_VERSIONING = ON (HISTORY_TABLE = [dbo].[OrderHistory]))

    COMMIT TRANSACTION
END

```

"
"Which is the best way to store images for expressjs, mongodb site?
I'm new express, nodejs, mongodb. I'm building a website for a clothing shop. They want to upload images frequently, maybe every 2 days they will upload new images, and those images need to be shown on the website quickly, thumbnails. Click on the image will show detail the product, etc.


My question is what are the best ways to store images? 


1. Store images into hard disk and just store the absolute path into database mongodb?
2. Stored images as binary/base64 into database mongodb?


How to load images into the website quickly?


Thanks


","**Where do we store images?**


**Solution # 1 in MongoDB**


So the first solution is to store the images inside of MongoDB. It can accommodate image files or any file type. So you can take a file and tie it to a record inside of MongoDB and save it directly to your database.


With this approach tying a description page of a particular article of clothing with its corresponding image becomes easy because you can embed this image directly into your the development of that page and your client would be happy with that approach because when the user retrieves a detailed description page of that article of clothing it comes with the image.


So that's one possible solution, just take the image and store directly into MongoDB.


However, I am going to suggest that this is a bad approach. The reason for that, you can tell your client if there is pushback is that they will usually pay for their Mongo instance in terms of the amount of storage that their Mongo copy uses.


So all the more storage they use up, they pay more money per month.


For example, the last time I checked the use of MLab, they were charging $15 per GB. So that's $15 out of your customer's pocket for hosting 1GB worth of images.


For yet another e-commerce website, we are talking about 3GB easily, which translates into 330 images more or less, equally $15 a month.


So if one of their project managers is uploading a new article of clothing once a day, we are talking about a tremendous cost very quickly.


So, personally I think that storing any kind of file type directly inside of MongoDB is really not an option because it will get very costly.


So that's just one possible solution.


**Solution # 2 in HD attached to server**


So let's look at a second solution that might be available to you. You might use a hard drive that's tied your Express server. So when this application gets deployed to some cloud environment like Heroku, Digital Ocean, Linode or AWS you usually get a hard drive associated with your application.


So maybe you take the images and place them inside the local hard drive. This approach is what the vast majority of online posts and articles will advocate for:


[How to upload, display and save images using node.js and express](https://stackoverflow.com/questions/15772394/how-to-upload-display-and-save-images-using-node-js-and-express)


<https://appdividend.com/2019/02/14/node-express-image-upload-and-resize-tutorial-example/>


<https://medium.com/@nitinpatel_20236/image-upload-via-nodejs-server-3fe7d3faa642>


Just with the three I gathered above you have a pretty robust blueprint to get started with.


Every single one says take the file and save it to your local hard drive. In this particular article:


<https://alligator.io/nodejs/uploading-files-multer-express/>


They are showing this code:



```
const storage = multer.diskStorage({
  destination: 'some-destination',
  filename: function (req, file, callback) {
    //..
  }
});

```

They are using the image upload library called `multer` which provides the `diskStorage()` engine for uploading images to disk.


So this is one approach that the development community at-large agrees with.


This is a good approach in the context of a one-to-one mapping.


The problems with this approach starts to arise when we have multiple machines.


An example of this is if you have multiple machines hosted on Digital Ocean or Linode where each environment is a separate instance.


If you have all your images stored in the accompanying hard drive and then you start to scale your server up, they would each have their own separate hard drive.


So you may have a request that comes in through a load balancer and the load balancer decides where to send the request to like the diagram below:


[![enter image description here](https://i.stack.imgur.com/L05al.png)](https://i.stack.imgur.com/L05al.png)


So the issue with the above architecture is if the image gets saved to one of the two hard drives and then later on a request comes in to access that same image, but imagine the request gets routed to the other Express server with a different hard drive where the image does not exist.


This is a problem that comes up when you start to use a service provider like Linode or Digital Ocean where you have a one-to-one mapping between server and hard drive.


It's a short term solution if that is all you need for now, but once that application starts to scale its going to be a problem.


**Solution # 3 outside data store**


This third solution is one I have used in the past with React with Node applications and Ruby on Rails applications as well. In fact, my Ruby on Rails portfolio website uses this solution and it sits on the Heroku platform.


So when the image gets uploaded, rather than the Express API trying to store the file locally like on its own hard drive, it will take the image and use an outside data store to hold all the different images from the app.


The one I use for my portfolio website and what I have used for Node with React applications has been Amazon S3, but there also exists Azure File Storage and Google Cloud Storage. These systems are made to hold a tremendous amount of data and they can be any type of file that you can possibly imagine. Not just images such as in your case, but video files, audio files, etcetera.


There is no cap to the amount of storage you can have with S3, but you don't have to use S3, but it is seen as an industry standard right now, but you can easily just as well use Azure and Google Cloud.


The benefit of this solution that I think your customer will appreciate, Amazon S3 charges you like two pennies per gigabyte per month for storage.


"
"Why the unbound type variables in OCAML object do not appear when class type is used?
I'd like to understand the reason for this behavior of OCAML objects. Suppose I have a class `A` that calls methods of an object of another class `B`. Schematically, A#f calls B#g and B#h. The normal practice in OOP is that I would like to avoid using B as a fixed concrete class, but instead declare only an interface for B. What is the best way to do this in OCAML? I tried several options, and I do not quite understand the reason why some of them work while others don't. Here are the code samples.


Version 1:



```
 # class classA = object
    method f b = b#g + b#h 
   end ;;
 Error: Some type variables are unbound in this type:
     class a : object method f : < g : int; h : int; .. > -> int end
   The method f has type (< g : int; h : int; .. > as 'a) -> int where 'a
   is unbound

```

This behavior is well-known: OCAML correctly infers that `b` has the open object type `<g:int;h:int;..>` but then complains that my class does not declare any type variables. So it seems that `classA` is *required* to have type variables; I then introduced a type variable explicitly.


Version 2:



```
 # class ['a] classA2 = object
   method f (b:'a) = b#g + b#h
   end ;;
 class ['a] classA2 :
   object constraint 'a = < g : int; h : int; .. > method f : 'a -> int end

```

This works, but the class is now explicitly polymorphic with a type constraint, as OCAML shows. It is also confusing that the class type contains a type variable `'a`, and yet I can still say `let x = new classA2` without specifying a type value for `'a`. Why is that possible?


Another drawback of `classA2` is that an explicit type constraint `(b:'a)` contains a type variable. After all, I know that `b` must conform to a fixed interface rather than to an unknown type `'a`. I want OCAML to verify that this interface is indeed correct.


So in version 3 I first declared an interface `classB` as a class type and then declared that `b` must be of this type:



```
 # class type classB = object method g:int method h:int end;;
 class type classB = object method g : int method h : int end
 # class classA3 = object method f (b:classB) = b#g + b#h end;;
 class classA3 : object method f : classB -> int end

```

This works too, but my puzzlement remains: why doesn't `classA3` require explicit polymorphism any more?


Summary of questions:


- Why is it possible to use `new classA2` without specifying a type for `'a` even though `classA2` is declared with a type variable `'a`?
- Why does `classA3` accept a type constraint `(b:classB)` and does not require a bound type variable any more?
- Is the functionality of `classA2` and `classA3` different in some subtle way, and if yes, how?


","This is going to be a bit complex, so hold tight. First, let me add a `classA4` variant, which happens to be what you actually need. 



```
class classA4 = object
  method f : 'a. (#classB as 'a) -> int = fun b -> b#g + b#h
end

```

Classes `classA2`, `classA3` and `classA4` are all subtly different, and the difference lies in how OCaml treats type polymorphism and object polymorphism. Let's assume that two classes, `b1` and `b2`, implement the `classB` type. 


In terms of *object* polymorphism, this means that an expression of type `b1` can be coerced to type `classB` using the coercion syntax `(new b1 :> classB)`. This type coercion discards type information (you no longer know the object is of type `b1`), so it must be made explicit. 


In terms of *type* polymorphism, this means that type `b1` can be used in place of any type variable that has constraint `#classB` (or `< g : int ; h : int ; .. >`). This does not discard any type information (as the type variable is replaced by the actual type), so it is performed by the type inference algorithm.


Method `f` of `classA3` expects a parameter of type `classB`, which means a type coercion is mandatory: 



```
let b = new b1 
let a = new classA3
a # f b (* Type error, expected classB, found b1 *)
a # f (b :> classB) (* Ok ! *)

```

This also means that (as long as you coerce), any class implementing `classB` can be used. 


Method `f` of `classA2` expects a parameter of a type that matches constraint `#classB`, but OCaml mandates that such a type should not be unbound, so it is bound at the class level. This means that every instance of `classA2` will accept parameters of a **single** arbitrary type that implements `classB` (and that type will be type-inferred). 



```
let b1 = new b1 and b2 = new b2
let a  = new classA2 
a # f b1 (* 'a' is type-inferred to be 'b1 classA2' *)
a # f b2 (* Type error, because b1 != b2  *)

```

It is important to note that `classA3` is equivalent to `classB classA2`, which is why it does not require a bound type variable, and also why it is strictly less expressive than `classA2`. 


Method `f` of `classA4` has been given an explicit type using the `'a.` syntax, which binds the type variable at the method level instead of the class level. It's actually an universal quantifier, which means « this method can be called for any type `'a` which implements `#classB` » :



```
let b1 = new b1 and b2 = new b2
let a  = new classA4
a # f b1 (* 'a is chosen to be b1 for this call *)
a # f b2 (* 'a is chosen to be b2 for this call *)

```

"
"ResultSet reset when using recursion
I have a question regarding `ResultSet` objects in Java and recursion. 


I was working on some code for university and whenever I found a descendant I recursed on with that new node but when I came out of the recursion and tried to rs.next() the pointer had gone from pointing to row 1 back to row 0 and when it hit row 0 the rs.next() failed and it returned! I knew there was one thing in there that it hadn't read yet! What is it that causes this? 


The only way I got round that problem was to go through the resultset and get every element and add it into an array list, then loop through the arraylist doing the recursion on each element in the array! Surely this must be a better way around this? 


This is the new code I'm using



```
private Vector<String> getDescendents(String dogname, Vector<String> anc) {
    if (anc == null) anc = new Vector<String>();
    ArrayList<String> tempList = new ArrayList<String>(2);
    try {
        System.out.println(""Inside "");
        childStmt.setString(1,dogname);
        childStmt.setString(2,dogname);
        ResultSet rs = childStmt.executeQuery();
        System.out.println(""Before while ""+rs.getRow());
        while (rs.next()){
            String col1 = rs.getString(1);
            tempList.add(col1);
            anc.add(col1);
        } 
        for (String s:tempList){
            getDescendents(s,anc);
        }

    }
    catch(Exception e) {
        doError(e, ""Failed to execute ancestor query in getBreeding"");
    }
    return anc;
}

```

However before this, I had the `getDescendents` call inside the `while` loop and thus no for loop and no arraylist either, but whenever it actually recursed it would loose track of the resultset when it returned out of the recursion.


Further details :
When I used the debugger (nearly said gdb there lol far too much C) the ID of the result set was the same but the row pointer had returned to 0 and the rs.next call failed!


Once again any explanation is appreciated!


p.s it previously looked like



```
private Vector<String> getDescendents(String dogname, Vector<String> anc) {
    if (anc == null) anc = new Vector<String>();
    ArrayList<String> tempList = new ArrayList<String>(2);
    try {
        System.out.println(""Inside "");
        childStmt.setString(1,dogname);
        childStmt.setString(2,dogname);
        ResultSet rs = childStmt.executeQuery();
        System.out.println(""Before while ""+rs.getRow());
        while (rs.next()){
            String col1 = rs.getString(1);
            anc.add(col1);
            getDescendendts(col1,anc);
        } 

    }
    catch(Exception e) {
        doError(e, ""Failed to execute ancestor query in getBreeding"");
    }
    return anc;
}

```

","It looks like you're re-using `childStmt`; don't do this. From the [`Statement` javadoc](http://docs.oracle.com/javase/6//docs/api/java/sql/Statement.html):



> 
> By default, only one ResultSet object per Statement object can be open
>  at the same time. Therefore, if the reading of one ResultSet object is
>  interleaved with the reading of another, each must have been generated
>  by different Statement objects. All execution methods in the Statement
>  interface implicitly close a statment's current ResultSet object if an
>  open one exists.
> 
> 
> 


You'll have to either save all the rows first, then do the recursive query, or create a new `Statement` for each `ResultSet` you want to fetch.


"
"Remove spurious small islands of noise in an image - Python OpenCV
I am trying to get rid of background noise from some of my images. This is the unfiltered image.


![](https://i.stack.imgur.com/CiBlj.png)


To filter, I used this code to generate a mask of what should remain in the image:



```
 element = cv2.getStructuringElement(cv2.MORPH_RECT, (2,2))
 mask = cv2.erode(mask, element, iterations = 1)
 mask = cv2.dilate(mask, element, iterations = 1)
 mask = cv2.erode(mask, element)

```

With this code and when I mask out the unwanted pixels from the original image, what I get is:
![](https://i.stack.imgur.com/0WDlX.png)


As you can see, all the tiny dots in the middle area are gone, but a lot of those coming from the denser area are also gone. To reduce the filtering, I tried changing the second parameter of `getStructuringElement()` to be (1,1) but doing this gives me the first image as if nothing has been filtered. 


Is there any way where I can apply some filter that is between these 2 extremes?


In addition, can anyone explain to me what exactly does `getStructuringElement()` do? What is a ""structuring element""? What does it do and how does its size (the second parameter) affect the level of filtering?


","A lot of your questions stem from the fact that you're not sure how morphological image processing works, but we can put your doubts to rest. You can interpret the structuring element as the ""base shape"" to compare to. 1 in the structuring element corresponds to a pixel that you want to look at in this shape and 0 is one you want to ignore. There are different shapes, such as rectangular (as you have figured out with `MORPH_RECT`), ellipse, circular, etc.


As such, `cv2.getStructuringElement` returns a structuring element for you. The first parameter specifies the type you want and the second parameter specifies the size you want. In your case, you want a 2 x 2 ""rectangle""... which is really a square, but that's fine.


In a more bastardized sense, you use the structuring element and scan from left to right and top to bottom of your image and you grab pixel neighbourhoods. Each pixel neighbourhood has its centre exactly at the pixel of interest that you're looking at. The size of each pixel neighbourhood is the same size as the structuring element.


# Erosion


For an erosion, you examine all of the pixels in a pixel neighbourhood that are touching the structuring element. If **every non-zero pixel** is touching a structuring element pixel that is 1, then the output pixel in the corresponding centre position with respect to the input is 1. If there is at least one non-zero pixel that **does not** touch a structuring pixel that is 1, then the output is 0.


In terms of the rectangular structuring element, you need to make sure that every pixel in the structuring element is touching a non-zero pixel in your image for a pixel neighbourhood. If it isn't, then the output is 0, else 1. This effectively eliminates small spurious areas of noise and also decreases the area of objects slightly.


The size factors in where the larger the rectangle, the more shrinking is performed. The size of the structuring element is a baseline where any objects that are smaller than this rectangular structuring element, you can consider them as being filtered and not appearing in the output. Basically, choosing a 1 x 1 rectangular structuring element is the same as the input image itself because that structuring element fits all pixels inside it as the pixel is the smallest representation of information possible in an image.


# Dilation


Dilation is the opposite of erosion. If there is at least one non-zero pixel that touches a pixel in the structuring element that is 1, then the output is 1, else the output is 0. You can think of this as slightly enlarging object areas and making small islands bigger.


The implications with size here is that the larger the structuring element, the larger the areas of the objects will be and the larger the isolated islands become.




---


What you're doing is an erosion first followed by a dilation. This is what is known as an **opening** operation. The purpose of this operation is to remove small islands of noise while (trying to) maintain the areas of the larger objects in your image. The erosion removes those islands while the dilation grows back the larger objects to their original sizes.


You follow this with an erosion again for some reason, which I can't quite understand, but that's ok.




---


What I would personally do is perform a **closing** operation first which is a dilation followed by an erosion. Closing helps group areas that are close together into a single object. As such, you see that there are some larger areas that are close to each other that should probably be joined before we do anything else. As such, I would do a closing first, then do an **opening** after so that we can remove the isolated noisy areas. Take note that I'm going to make the closing structuring element size **larger** as I want to make sure I get nearby pixels and the opening structuring element size **smaller** so that I don't want to mistakenly remove any of the larger areas.


Once you do this, I would mask out any extra information with the original image so that you leave the larger areas intact while the small islands go away.


Instead of chaining an erosion followed by a dilation, or a dilation followed by an erosion, use [`cv2.morphologyEx`](http://docs.opencv.org/modules/imgproc/doc/filtering.html#morphologyex), where you can specify `MORPH_OPEN` and `MORPH_CLOSE` as the flags.


As such, I would personally do this, assuming your image is called `spots.png`:



```
import cv2
import numpy as np

img = cv2.imread('spots.png')
img_bw = 255*(cv2.cvtColor(img, cv2.COLOR_BGR2GRAY) > 5).astype('uint8')

se1 = cv2.getStructuringElement(cv2.MORPH_RECT, (5,5))
se2 = cv2.getStructuringElement(cv2.MORPH_RECT, (2,2))
mask = cv2.morphologyEx(img_bw, cv2.MORPH_CLOSE, se1)
mask = cv2.morphologyEx(mask, cv2.MORPH_OPEN, se2)

mask = np.dstack([mask, mask, mask]) / 255
out = img * mask

cv2.imshow('Output', out)
cv2.waitKey(0)
cv2.destroyAllWindows()
cv2.imwrite('output.png', out)

```

The above code is pretty self-explanatory. First, I read in the image and then I convert the image to grayscale and threshold with an intensity of 5 to create a mask of what is considered object pixels. This is a rather clean image and so anything larger than 5 seems to have worked. For the morphology routines, I need to convert the image to `uint8` and scale the mask to 255. Next, we create two structuring elements - one that is a 5 x 5 rectangle for the closing operation and another that is 2 x 2 for the opening operation. I run `cv2.morphologyEx` twice for the opening and closing operations respectively on the thresholded image.


Once I do that, I stack the mask so that it becomes a 3D matrix and divide by 255 so that it becomes a mask of `[0,1]` and then we multiply this mask with the original image so that we can grab the original pixels of the image back and maintaining what is considered a true object from the mask output.


The rest is just for illustration. I show the image in a window, and I also save the image to a file called `output.png`, and its purpose is to show you what the image looks like in this post.


I get this:


![enter image description here](https://i.stack.imgur.com/1JXvU.png)


Bear in mind that it isn't perfect, but it's much better than how you had it before. You'll have to play around with the structuring element sizes to get something that you consider as a good output, but this is certainly enough to get you started.




---


## C++ version


There have been some requests to translate the code I wrote above into the C++ version using OpenCV. I have finally gotten around to writing a C++ version of the code and this has been tested on OpenCV 3.1.0. The code for this is below. As you can see, the code is very similar to that seen in the Python version. However, I used [`cv::Mat::setTo`](http://docs.opencv.org/3.2.0/d3/d63/classcv_1_1Mat.html#a0440e2a164c0b0d8462fb1e487be9876) on a copy of the original image and set whatever was not part of the final mask to 0. This is the same thing as performing an element-wise multiplication in Python.



```
#include <opencv2/opencv.hpp>

using namespace cv;

int main(int argc, char *argv[])
{
    // Read in the image
    Mat img = imread(""spots.png"", CV_LOAD_IMAGE_COLOR);

    // Convert to black and white
    Mat img_bw;
    cvtColor(img, img_bw, COLOR_BGR2GRAY);
    img_bw = img_bw > 5;

    // Define the structuring elements
    Mat se1 = getStructuringElement(MORPH_RECT, Size(5, 5));
    Mat se2 = getStructuringElement(MORPH_RECT, Size(2, 2));

    // Perform closing then opening
    Mat mask;
    morphologyEx(img_bw, mask, MORPH_CLOSE, se1);
    morphologyEx(mask, mask, MORPH_OPEN, se2);

    // Filter the output
    Mat out = img.clone();
    out.setTo(Scalar(0), mask == 0);

    // Show image and save
    namedWindow(""Output"", WINDOW_NORMAL);
    imshow(""Output"", out);
    waitKey(0);
    destroyWindow(""Output"");
    imwrite(""output.png"", out);
}

```

The results should be the same as what you get in the Python version.


"
"converting to list comprehension
I have this code:



```
result = []
for x in [10, 20, 30]:
    for y in [2, 3, 4]:
        if y > 0:
            result.append(x ** y)

```

result



```
[100, 1000, 10000, 400, 8000, 160000, 900, 27000, 810000]

```

I'm trying to convert it to list comprehension with no luck(new in python)


This is my attempt:



```
print [ x ** y if y > 0 for x in [10, 20, 30] for y in [2, 3, 4]]

```

But there is a problem in the statement, any help will be most appropriated.


error:



```
  File ""<stdin>"", line 1
    print [ x ** y if y > 0 for x in [10, 20, 30] for y in [2, 3, 4]]
                              ^
SyntaxError: invalid syntax

```

","The filtering condition has to be at the end, like this



```
print [x ** y for x in [10, 20, 30] for y in [2, 3, 4] if y > 0]

```

because the [grammar for list comprehension](https://docs.python.org/2/reference/expressions.html#list-displays) is defined like this



```
list_display        ::=  ""["" [expression_list | list_comprehension] ""]""
list_comprehension  ::=  expression list_for
list_for            ::=  ""for"" target_list ""in"" old_expression_list [list_iter]
list_iter           ::=  list_for | list_if
list_if             ::=  ""if"" old_expression [list_iter]

```

So only expressions can come before the `for..in` and `if` statement can come only after that.


In your case, `expression` is satisfied by `x ** y` and then `list_for` is satisfied by `for x in [10, 20, 30]` and then another `list_for` is satisfied by `for x in [10, 20, 30]` and finally the `list_if` is satisfied by `if y > 0`. It is of the form



```
[ expression list_for list_for list_if ]

```

BTW, you can do the same with [`itertools.product`](https://docs.python.org/2/library/itertools.html#itertools.product), like this



```
from itertools import product
print [num**power for num,power in product([10, 20, 30], [2, 3, 4]) if power > 0]

```

"
"How to sort ps output to find processes realtime priority and identify processess currently occupied running queue
How to properly identify real-time processes currently occupied CPU queue and count them using ps? I know there is a bunch of fileds like prio,rtprio,pri,nice but do not know correct to use. It seems I need use something like `ps -eo rtprio,prio,cpu,cmd --sort=+rtprio` to get full list, but it do not seems for me right since a lot of processes got with `-` sign at RTPRIO column. For example, I have 48 cores system running Oracle Linux and try to identify following questions:


1. What processes occupied run queue? What's a count of them?
2. How to identify processes that run in Real Time mode or with increased priority?


","A list of non-zero CPU % processes:



```
ps -eo pid,tid,class,rtprio,ni,pri,psr,pcpu,stat,wchan:14,comm --sort=+pcpu | awk '$8!=0.0 {print}' | awk 'NR>1'

```

To count them



```
ps -eo pid,tid,class,rtprio,ni,pri,psr,pcpu,stat,wchan:14,comm --sort=+pcpu | awk '$8!=0.0 {print}' | awk 'NR>1' | wc -l

```

To see this continuously updated, but them in a file called `processes.sh`:



```
#!/bin/bash
ps -eo pid,tid,class,rtprio,ni,pri,psr,pcpu,stat,wchan:14,comm --sort=+pcpu | awk '$8!=0.0 {print}' | awk 'NR>1'

```

and make it executable with `chmod +x processes.sh`. Now run it with watch for live updating:



```
watch ./processes.sh

```

"
"Full screen iOS share Extension
I know how to create an iOS share extension, but what I can't figure out is if there is a way to make it fullscreen such as an Action extension?


Pinterest seem to do this but I'm not sure how.


The docs for an Action extension say to use:



```
          <key>NSExtensionActionWantsFullScreenPresentation</key>
    <true/>

```

In the plist file for the extension but this doesn't seem to have an effect in the Share extension?


Is there anyway to accomplish this?


","You can get ideas from there [iOS Full Screen Share Extension](https://martinnormark.com/present-ios-8-share-extension-as-modal-view/) and you can find the updated syntax for the code snippets below which is compatible with **Swift 3** and **Swift 4**


**EntryViewController**



```
import UIKit

@objc(EntryViewController)

class EntryViewController : UINavigationController {

    override init(rootViewController: UIViewController) {
        super.init(rootViewController: ShareViewController())
    }

    required init(coder aDecoder: NSCoder) {
        super.init(coder: aDecoder)!
    }

    override init(nibName nibNameOrNil: String?, bundle nibBundleOrNil: Bundle?) {
        super.init(nibName: nibNameOrNil, bundle: nibBundleOrNil)
    }

    override func viewWillAppear(_ animated: Bool) {
        super.viewWillAppear(animated)
        self.view.transform = CGAffineTransform(translationX: 0, y: self.view.frame.size.height)
        UIView.animate(withDuration: 0.3, animations: { () -> Void in
            self.view.transform = .identity
        })
    }

}

```

**ShareViewController**



```
import UIKit
import Social

class ShareViewController: SLComposeServiceViewController {

    override func viewDidLoad() {
        super.viewDidLoad()

        self.view.backgroundColor = UIColor.white
        self.navigationItem.title = ""Share""

        self.navigationItem.leftBarButtonItem = UIBarButtonItem(barButtonSystemItem: .cancel, target: self, action: Selector((""cancelButtonTapped:"")))
        self.navigationItem.rightBarButtonItem = UIBarButtonItem(barButtonSystemItem: .save, target: self, action: Selector((""saveButtonTapped:"")))
    }

    func saveButtonTapped(sender: UIBarButtonItem) {
        self.hideExtensionWithCompletionHandler(completion: { (Bool) -> Void in
        self.extensionContext!.completeRequest(returningItems: nil, completionHandler: nil)
        })
    }

    func cancelButtonTapped(sender: UIBarButtonItem) {
        self.hideExtensionWithCompletionHandler(completion: { (Bool) -> Void in
        self.extensionContext!.cancelRequest(withError: NSError())
        })
    }

    func hideExtensionWithCompletionHandler(completion: @escaping (Bool) -> Void) {
        UIView.animate(withDuration: 0.3, animations: { 
        self.navigationController!.view.transform = CGAffineTransform(translationX: 0, y: self.navigationController!.view.frame.size.height)
        }, completion: completion)
    }

}

```

"
".getTime() Alternative without milliseconds
Is there a way to use the Date().getTime() function without getting the time in milliseconds? If not is there a substitute for .getTime() which will only give me the precision in minutes? 


Also I'm not sure how to just strip the milliseconds out of the date object. 



```
var time = new Date().getTime()

Output: 1426515375925

```

","Simple arithmetic. If you want the value in seconds, divide the milliseconds result by 1000:



```
var seconds = new Date().getTime() / 1000;

```

You might want to call `Math.floor()` on that to remove any decimals though:



```
var seconds = Math.floor(new Date().getTime() / 1000);

```


> 
> Is there a different function I can use that will give me only the minutes since 1/1/1970, I just dont want the number to be as precise.
> 
> 
> 


Certainly, divide the seconds by 60, or divide the milliseconds by 60000:



```
var minutes = Math.floor(new Date().getTime() / 60000);

```



---



```
var milliseconds = 1426515375925,
    seconds = Math.floor(milliseconds / 1000),  // 1426515375
    minutes = Math.floor(milliseconds / 60000); // 23775256

```

"
"Best practices for making header file?
#### What things should absolutely always/never be included in a header file?


- Functions: what must go in the header file and what mustn't
- Constants: Is it good practice to define lot of constants in a header file?
- Any other good practices for C header files?


*Example*


I'm writing a parser for a *documented industry standard format* that has a **lots of constants**, is it a good practice to define them in a header file?


","**What to put in headers:**


- The minimal set of `#include` directives that are needed to make the header compilable when the header is included in some source file.
- Preprocessor symbol definitions of things that need to be shared and that can only accomplished via the preprocessor. Even in C, preprocessor symbols are best kept to a minimum.
- Forward declarations of structures that are needed to make the structure definitions, function prototypes, and global variable declarations in the body of the header compilable.
- Definitions of data structures and enumerations that are shared amongst multiple source files.
- Declarations for functions and variables whose definitions will be visible to the linker.
- Inline function definitions, but take care here.


**What doesn't belong in a header:**


- Gratuitous `#include` directives. Those gratuitous includes cause recompilation of things that don't need to be recompiled, and can at times make it so a system can't compile. Don't `#include` a file in a header if the header itself doesn't need that other header file.
- Preprocessor symbols whose intent could be accomplished by some mechanism, any mechanism, other than the preprocessor.
- Lots and lots of structure definitions. Split those up into separate headers.
- Inline definitions of functions that require an additional `#include`, that are subject to change, or that are too big. Those inline functions should have little if any fan out, and if they do have fan out, it should be localized to stuff defined in the header.


**What constitutes the minimal set of `#include` statements?**


This turns out to be a nontrivial question. A TL;DR definition: A header file must include the header files that directly define each of the types directly used in or that directly declare each of the functions used in the header file in question, but must not include anything else. A pointer or C++ reference type does not qualify as direct use; forward references are preferred.


There is a place for a gratuitous `#include` directive, and this is in an automated test. For every header file in a software package, I automatically generate and then compile the following:



```
#include ""path/to/random/header_under_test""
int main () { return 0; }

```

The compilation should be clean (i.e., free of any warnings or errors). Warnings or errors regarding incomplete types or unknown types mean that the header file under test has some missing `#include` directives and/or missing forward declarations. Note well: Just because the test passes does not mean that the set of `#include` directives is sufficient, let alone minimal.




---


### Related


- [Naming Include Guards](https://stackoverflow.com/a/4871964/13903942)


"
"Pandas: excel days360 equivalent
I am working on converting some excel models to python and some of the calculations are using the days360 function in excel, which calc the number of days between two days based on a 360-day year.


I have been looking over pandas.tseries.offsets and I am not seeing anything.


For example in excel,



```
d1 = 2/2/2015
d2 = 1/1/2018

=days(d1, d2) # 1064
=days360(d1, d2) # 1049

```

","There's [a Reddit thread](https://www.reddit.com/r/learnpython/comments/3942zm/python_implementation_of_excels_days360_function/) on `days360` for Pandas - here's the function that's posted there (seems to work with your example input):



```
def days360(start_date, end_date, method_eu=False):
    start_day = start_date.day
    start_month = start_date.month
    start_year = start_date.year
    end_day = end_date.day
    end_month = end_date.month
    end_year = end_date.year

    if (
        start_day == 31 or
        (
            method_eu is False and
            start_month == 2 and (
                start_day == 29 or (
                    start_day == 28 and
                    start_date.is_leap_year is False
                )
            )
        )
    ):
        start_day = 30

    if end_day == 31:
        if method_eu is False and start_day != 30:
            end_day = 1

            if end_month == 12:
                end_year += 1
                end_month = 1
            else:
                end_month += 1
        else:
            end_day = 30

    return (
        end_day + end_month * 30 + end_year * 360 -
        start_day - start_month * 30 - start_year * 360)

```

Testing on OP data:



```
d1 = pd.to_datetime(""2/2/2015"")
d2 = pd.to_datetime(""1/1/2018"")

days360(d1, d2) # 1049

```

"
"WCF - is it bad practice to leave a channel open for a long time?
I'm just learning the ropes around WCF. What I was planning to do was have a duplex channel open between a client and server using NetTcpBinding, and keep that open indefinitely so that the server can initiate requests to the client.


Then I stumbled upon this [blog by Jesse Ezell](http://old.iserviceoriented.com/blog/post/Indisposable+-+WCF+Gotcha+1.aspx), which seems to indicate that it's a Bad Thing to keep a channel open indefinitely, because you can't catch faults, and that causes all manner of instabilities.


Is that correct? If I use NetTcpBinding and keep a reference to an open channel on either side of the relationship, what happens if there's a communication failure? How do I catch the failure event? What other gotchas are there? Is there any difference which .NET framework you're using? (I'm on 4.0.)


","I don't agree with Jesse (*as a side note: he also recommends you use WCF service classes as singletons by default, which is the worst idea ever in my opinion*)..... 


As long as you take good care of catching exceptions on the server (e.g. by implementing the `IErrorHandler` interface in your service class), there's no point here to keep shutting down your channel... especially not in a corporate LAN environment using netTcpBinding.


Contrary to e.g. a database connection which often incurs licensing costs, keeping a network connection to your service machine open shouldn't cause any issues. It's also usually not a limited resource, so constantly opening and closing it seems pointless.


If you do keep your service channel open for a longer period of time, you need to be capable on the client side to handle faults - e.g. you need to be able to recover from a situation where the channel has become faulted, when an exception has occured after all (e.g. the network is down or something like that).


But if you do that, then I don't see any benefit in constantly closing your channel after every call, and reopening for the next...


"
"How to create zipcode boundaries in R
I am trying to create a map that has the name of the 'community' showing the boundaries of multiple zip codes. Data that I have is similar to below. Where the variable is the name of the community and the numbers are the corresponding zipcodes.



```
Tooele       <- c('84074','84029')
NEUtahCo     <- c('84003', '84004', '84042', '84062')
NWUtahCounty <- c('84005','84013','84043','84045')

```

I was able to make a map of the entire area I want using



```
ggmap(get_map(location = c(lon=-111.9, lat= 40.7), zoom = 9))

```

Attached is a picture of what I want.


[![Map of Wasatch Front Communities](https://i.stack.imgur.com/VFFdn.jpg)](https://i.stack.imgur.com/VFFdn.jpg)


","You have a decent foundation for this already by having figured out the shapefile and how it matches the zips you want to show. Simple features (`sf`) make this pretty easy, as does the brand new `ggplot2` v3.0.0 which has the `geom_sf` to plot `sf` objects. 


I wasn't sure if the names of the different areas (counties?) that you have are important, so I just threw them all into little tibbles and bound that into one tibble, `utah_zips`. `tigris` also added `sf` support, so if you set `class = ""sf""`, you get an `sf` object. To keep it simple, I'm just pulling out the columns you need and simplifying one of the names.



```
library(tidyverse)
library(tigris)
library(ggmap)

Tooele       <- c('84074','84029')
NEUtahCo     <- c('84003', '84004', '84042', '84062')
NWUtahCounty <- c('84005','84013','84043','84045')
utah_zips <- bind_rows(
  tibble(area = ""Tooele"", zip = Tooele),
  tibble(area = ""NEUtahCo"", zip = NEUtahCo),
  tibble(area = ""NWUtahCounty"", zip = NWUtahCounty)
)

zips_sf <- zctas(cb = T, starts_with = ""84"", class = ""sf"") %>%
  select(zip = ZCTA5CE10, geometry)

head(zips_sf)
#> Simple feature collection with 6 features and 1 field
#> geometry type:  MULTIPOLYGON
#> dimension:      XY
#> bbox:           xmin: -114.0504 ymin: 37.60461 xmax: -109.0485 ymax: 41.79228
#> epsg (SRID):    4269
#> proj4string:    +proj=longlat +datum=NAD83 +no_defs
#>       zip                       geometry
#> 37  84023 MULTIPOLYGON (((-109.5799 4...
#> 270 84631 MULTIPOLYGON (((-112.5315 3...
#> 271 84334 MULTIPOLYGON (((-112.1608 4...
#> 272 84714 MULTIPOLYGON (((-113.93 37....
#> 705 84728 MULTIPOLYGON (((-114.0495 3...
#> 706 84083 MULTIPOLYGON (((-114.0437 4...

```

Then you can filter the `sf` for just the zips you need—since there's other information (the county names), you can use a join to get everything in one `sf` data frame:



```
utah_sf <- zips_sf %>%
  inner_join(utah_zips, by = ""zip"")
head(utah_sf)
#> Simple feature collection with 6 features and 2 fields
#> geometry type:  MULTIPOLYGON
#> dimension:      XY
#> bbox:           xmin: -113.1234 ymin: 40.21758 xmax: -111.5677 ymax: 40.87196
#> epsg (SRID):    4269
#> proj4string:    +proj=longlat +datum=NAD83 +no_defs
#>     zip         area                       geometry
#> 1 84029       Tooele MULTIPOLYGON (((-112.6292 4...
#> 2 84003     NEUtahCo MULTIPOLYGON (((-111.8497 4...
#> 3 84074       Tooele MULTIPOLYGON (((-112.4191 4...
#> 4 84004     NEUtahCo MULTIPOLYGON (((-111.8223 4...
#> 5 84062     NEUtahCo MULTIPOLYGON (((-111.7734 4...
#> 6 84013 NWUtahCounty MULTIPOLYGON (((-112.1564 4...

```

You already have your basemap figured out, and since `ggmap` makes `ggplot` objects, you can just add on a `geom_sf` layer. The tricks are just to make sure you declare the data you're using, set it to *not* inherit the aes from `ggmap`, and turn off the graticules in `coord_sf`.



```
basemap <- get_map(location = c(lon=-111.9, lat= 40.7), zoom = 9)

ggmap(basemap) +
  geom_sf(aes(fill = zip), data = utah_sf, inherit.aes = F, size = 0, alpha = 0.6) +
  coord_sf(ndiscr = F) +
  theme(legend.position = ""none"")

```

![](https://i.stack.imgur.com/XTOnf.png)


You might want to adjust the position of the basemap, since it cuts off one of the zips. One way is to use `st_bbox` to get the bounding box of `utah_sf`, then use that to get the basemap.


"
"Black line showing on map fragment in android
I am developing a fragment in Android. In my tab there are four fragments, of which one is a map fragment. 


Using this map fragment:



```
<fragment xmlns:android=""http://schemas.android.com/apk/res/android""
          android:id=""@+id/map""
          android:layout_width=""match_parent""
          android:layout_height=""match_parent""
          class=""com.google.android.gms.maps.SupportMapFragment"" />

```

the task is done perfectly. 


The problem is that when the user swipes, a black line is shown on the map fragment for 1 second. I've researched this online but couldn't find a way to prevent this? Why is this occurring and how can I fix it?


","Was also facing this strange issue.Fixed it as follows by applying a transparent View over the mapview fragment seems to resolve the issue. Its a kind of hack.See if it works for you



```
<RelativeLayout
            android:id=""@+id/relativeLayout1""
            android:layout_width=""match_parent""
            android:layout_height=""300dp"" >

            <fragment
                android:id=""@+id/map""
                android:name=""com.google.android.gms.maps.MapFragment""
                android:layout_width=""fill_parent""
                android:layout_height=""fill_parent""/>

            <View
                android:id=""@+id/imageView123""
                android:layout_width=""match_parent""
                android:layout_height=""match_parent""
               android:background=""@android:color/transparent"" />         
    </RelativeLayout>

```

"
"How to get flag arguments in other package
I write program in go, and I have a problem with get variable form command line in other own package using package **flag** . 



> 
> flag.Parse()
> 
> 
> 


Main problem is in **Configuration package**, because I use her many times in many places, so I want to avoid pass pionter to her, and I decided create as an independent module. Now I have problem with get flag to pathFile with configuration. My code looks like below


I run my program with parameter



> 
> program -config=/my/path/config.cfg
> 
> 
> 


and main function



```
func main() {
        flag.Parse()
      // some next operation but not with configuration Type
      // the type of configuration is use in other object 
      //and in other place in my program
      //  here I only want to parse plag from
      // cmd lines
}

```

**package configuration** and file configuration.go so **configuration/configuration.go**



```
var (
    configPath string
    C          *configuration
)

func init() {
    flag.StringVar(&configPath,
        ""config"",
        ""/defaultpath/config.gcfg"",
        ""Path to configuration file"")
    C = &configuration{}
    if err := gcfg.ReadFileInto(C, configPath); err != nil {
        fmt.Printf(""Cannot read configuration file. Program was Terminated error:%s"", err)
        os.Exit(1)
    }
}

type configuration struct {
    Queue struct {
        User                    string
        Host                    string
        Port                    string
}

```


> 
> After compilation I get error: Cannot open log file error:open
>  /defaultpath/config.gcfg: no such file or directory. Program was
>  terminated.
> 
> 
> 


As look above Parsed flags in main are not visible in package configuration, because configuration module try start with default path nor form arguments cmd line. that's why this must be finish error, because this path dosen't exist in my local machine


Question on end is:
How export flag variable from main package to other own package ?


","The `init()` in `configuration` is run before *anything* in `main` is executed, so there's no way the effect of `flag.Parse()` could be seen. You also need to declare the flag variables before you call `flag.Parse()`


You can't export anything from `main`, both because `main` can't be imported, and because you can't have a circular dependency between the packages. Have the `main` package call a function from the configuration package to get the behavior you want. 


You could add an exported `Init()` function in `configuration` to handle the flag options:



```
func Init() {
    flag.StringVar(&configPath,
        ""config"",
        ""/defaultpath/config.gcfg"",
        ""Path to configuration file"")
    flag.Parse()

    C = &configuration{}
    if err := gcfg.ReadFileInto(C, configPath); err != nil {
        fmt.Printf(""Cannot read configuration file. Program was Terminated error:%s"", err)
        os.Exit(1)
    }
}

```

and have the `main` function start with 



```
func main() {
        configuration.Init()
}

```

"
"Has someone made the Parula colormap in Matplotlib?
I know that there's been some [discussion of Matlab copyrighting their new default colormap](http://sourceforge.net/p/matplotlib/mailman/message/32953371/), but I'm wondering if any intrepid user has created the colormap in Matplotlib.


[Viridis](http://bids.github.io/colormap/) is great, but it's a bit dark for what I'm trying to do.


","It looks like Matlab has changed the values slightly since this was answered in 2016. Here are the Parula values from the latest version of Matlab (R2019b Update 3):



```
cm_data = [[0.2422, 0.1504, 0.6603],
[0.2444, 0.1534, 0.6728],
[0.2464, 0.1569, 0.6847],
[0.2484, 0.1607, 0.6961],
[0.2503, 0.1648, 0.7071],
[0.2522, 0.1689, 0.7179],
[0.254, 0.1732, 0.7286],
[0.2558, 0.1773, 0.7393],
[0.2576, 0.1814, 0.7501],
[0.2594, 0.1854, 0.761],
[0.2611, 0.1893, 0.7719],
[0.2628, 0.1932, 0.7828],
[0.2645, 0.1972, 0.7937],
[0.2661, 0.2011, 0.8043],
[0.2676, 0.2052, 0.8148],
[0.2691, 0.2094, 0.8249],
[0.2704, 0.2138, 0.8346],
[0.2717, 0.2184, 0.8439],
[0.2729, 0.2231, 0.8528],
[0.274, 0.228, 0.8612],
[0.2749, 0.233, 0.8692],
[0.2758, 0.2382, 0.8767],
[0.2766, 0.2435, 0.884],
[0.2774, 0.2489, 0.8908],
[0.2781, 0.2543, 0.8973],
[0.2788, 0.2598, 0.9035],
[0.2794, 0.2653, 0.9094],
[0.2798, 0.2708, 0.915],
[0.2802, 0.2764, 0.9204],
[0.2806, 0.2819, 0.9255],
[0.2809, 0.2875, 0.9305],
[0.2811, 0.293, 0.9352],
[0.2813, 0.2985, 0.9397],
[0.2814, 0.304, 0.9441],
[0.2814, 0.3095, 0.9483],
[0.2813, 0.315, 0.9524],
[0.2811, 0.3204, 0.9563],
[0.2809, 0.3259, 0.96],
[0.2807, 0.3313, 0.9636],
[0.2803, 0.3367, 0.967],
[0.2798, 0.3421, 0.9702],
[0.2791, 0.3475, 0.9733],
[0.2784, 0.3529, 0.9763],
[0.2776, 0.3583, 0.9791],
[0.2766, 0.3638, 0.9817],
[0.2754, 0.3693, 0.984],
[0.2741, 0.3748, 0.9862],
[0.2726, 0.3804, 0.9881],
[0.271, 0.386, 0.9898],
[0.2691, 0.3916, 0.9912],
[0.267, 0.3973, 0.9924],
[0.2647, 0.403, 0.9935],
[0.2621, 0.4088, 0.9946],
[0.2591, 0.4145, 0.9955],
[0.2556, 0.4203, 0.9965],
[0.2517, 0.4261, 0.9974],
[0.2473, 0.4319, 0.9983],
[0.2424, 0.4378, 0.9991],
[0.2369, 0.4437, 0.9996],
[0.2311, 0.4497, 0.9995],
[0.225, 0.4559, 0.9985],
[0.2189, 0.462, 0.9968],
[0.2128, 0.4682, 0.9948],
[0.2066, 0.4743, 0.9926],
[0.2006, 0.4803, 0.9906],
[0.195, 0.4861, 0.9887],
[0.1903, 0.4919, 0.9867],
[0.1869, 0.4975, 0.9844],
[0.1847, 0.503, 0.9819],
[0.1831, 0.5084, 0.9793],
[0.1818, 0.5138, 0.9766],
[0.1806, 0.5191, 0.9738],
[0.1795, 0.5244, 0.9709],
[0.1785, 0.5296, 0.9677],
[0.1778, 0.5349, 0.9641],
[0.1773, 0.5401, 0.9602],
[0.1768, 0.5452, 0.956],
[0.1764, 0.5504, 0.9516],
[0.1755, 0.5554, 0.9473],
[0.174, 0.5605, 0.9432],
[0.1716, 0.5655, 0.9393],
[0.1686, 0.5705, 0.9357],
[0.1649, 0.5755, 0.9323],
[0.161, 0.5805, 0.9289],
[0.1573, 0.5854, 0.9254],
[0.154, 0.5902, 0.9218],
[0.1513, 0.595, 0.9182],
[0.1492, 0.5997, 0.9147],
[0.1475, 0.6043, 0.9113],
[0.1461, 0.6089, 0.908],
[0.1446, 0.6135, 0.905],
[0.1429, 0.618, 0.9022],
[0.1408, 0.6226, 0.8998],
[0.1383, 0.6272, 0.8975],
[0.1354, 0.6317, 0.8953],
[0.1321, 0.6363, 0.8932],
[0.1288, 0.6408, 0.891],
[0.1253, 0.6453, 0.8887],
[0.1219, 0.6497, 0.8862],
[0.1185, 0.6541, 0.8834],
[0.1152, 0.6584, 0.8804],
[0.1119, 0.6627, 0.877],
[0.1085, 0.6669, 0.8734],
[0.1048, 0.671, 0.8695],
[0.1009, 0.675, 0.8653],
[0.0964, 0.6789, 0.8609],
[0.0914, 0.6828, 0.8562],
[0.0855, 0.6865, 0.8513],
[0.0789, 0.6902, 0.8462],
[0.0713, 0.6938, 0.8409],
[0.0628, 0.6972, 0.8355],
[0.0535, 0.7006, 0.8299],
[0.0433, 0.7039, 0.8242],
[0.0328, 0.7071, 0.8183],
[0.0234, 0.7103, 0.8124],
[0.0155, 0.7133, 0.8064],
[0.0091, 0.7163, 0.8003],
[0.0046, 0.7192, 0.7941],
[0.0019, 0.722, 0.7878],
[0.0009, 0.7248, 0.7815],
[0.0018, 0.7275, 0.7752],
[0.0046, 0.7301, 0.7688],
[0.0094, 0.7327, 0.7623],
[0.0162, 0.7352, 0.7558],
[0.0253, 0.7376, 0.7492],
[0.0369, 0.74, 0.7426],
[0.0504, 0.7423, 0.7359],
[0.0638, 0.7446, 0.7292],
[0.077, 0.7468, 0.7224],
[0.0899, 0.7489, 0.7156],
[0.1023, 0.751, 0.7088],
[0.1141, 0.7531, 0.7019],
[0.1252, 0.7552, 0.695],
[0.1354, 0.7572, 0.6881],
[0.1448, 0.7593, 0.6812],
[0.1532, 0.7614, 0.6741],
[0.1609, 0.7635, 0.6671],
[0.1678, 0.7656, 0.6599],
[0.1741, 0.7678, 0.6527],
[0.1799, 0.7699, 0.6454],
[0.1853, 0.7721, 0.6379],
[0.1905, 0.7743, 0.6303],
[0.1954, 0.7765, 0.6225],
[0.2003, 0.7787, 0.6146],
[0.2061, 0.7808, 0.6065],
[0.2118, 0.7828, 0.5983],
[0.2178, 0.7849, 0.5899],
[0.2244, 0.7869, 0.5813],
[0.2318, 0.7887, 0.5725],
[0.2401, 0.7905, 0.5636],
[0.2491, 0.7922, 0.5546],
[0.2589, 0.7937, 0.5454],
[0.2695, 0.7951, 0.536],
[0.2809, 0.7964, 0.5266],
[0.2929, 0.7975, 0.517],
[0.3052, 0.7985, 0.5074],
[0.3176, 0.7994, 0.4975],
[0.3301, 0.8002, 0.4876],
[0.3424, 0.8009, 0.4774],
[0.3548, 0.8016, 0.4669],
[0.3671, 0.8021, 0.4563],
[0.3795, 0.8026, 0.4454],
[0.3921, 0.8029, 0.4344],
[0.405, 0.8031, 0.4233],
[0.4184, 0.803, 0.4122],
[0.4322, 0.8028, 0.4013],
[0.4463, 0.8024, 0.3904],
[0.4608, 0.8018, 0.3797],
[0.4753, 0.8011, 0.3691],
[0.4899, 0.8002, 0.3586],
[0.5044, 0.7993, 0.348],
[0.5187, 0.7982, 0.3374],
[0.5329, 0.797, 0.3267],
[0.547, 0.7957, 0.3159],
[0.5609, 0.7943, 0.305],
[0.5748, 0.7929, 0.2941],
[0.5886, 0.7913, 0.2833],
[0.6024, 0.7896, 0.2726],
[0.6161, 0.7878, 0.2622],
[0.6297, 0.7859, 0.2521],
[0.6433, 0.7839, 0.2423],
[0.6567, 0.7818, 0.2329],
[0.6701, 0.7796, 0.2239],
[0.6833, 0.7773, 0.2155],
[0.6963, 0.775, 0.2075],
[0.7091, 0.7727, 0.1998],
[0.7218, 0.7703, 0.1924],
[0.7344, 0.7679, 0.1852],
[0.7468, 0.7654, 0.1782],
[0.759, 0.7629, 0.1717],
[0.771, 0.7604, 0.1658],
[0.7829, 0.7579, 0.1608],
[0.7945, 0.7554, 0.157],
[0.806, 0.7529, 0.1546],
[0.8172, 0.7505, 0.1535],
[0.8281, 0.7481, 0.1536],
[0.8389, 0.7457, 0.1546],
[0.8495, 0.7435, 0.1564],
[0.86, 0.7413, 0.1587],
[0.8703, 0.7392, 0.1615],
[0.8804, 0.7372, 0.165],
[0.8903, 0.7353, 0.1695],
[0.9, 0.7336, 0.1749],
[0.9093, 0.7321, 0.1815],
[0.9184, 0.7308, 0.189],
[0.9272, 0.7298, 0.1973],
[0.9357, 0.729, 0.2061],
[0.944, 0.7285, 0.2151],
[0.9523, 0.7284, 0.2237],
[0.9606, 0.7285, 0.2312],
[0.9689, 0.7292, 0.2373],
[0.977, 0.7304, 0.2418],
[0.9842, 0.733, 0.2446],
[0.99, 0.7365, 0.2429],
[0.9946, 0.7407, 0.2394],
[0.9966, 0.7458, 0.2351],
[0.9971, 0.7513, 0.2309],
[0.9972, 0.7569, 0.2267],
[0.9971, 0.7626, 0.2224],
[0.9969, 0.7683, 0.2181],
[0.9966, 0.774, 0.2138],
[0.9962, 0.7798, 0.2095],
[0.9957, 0.7856, 0.2053],
[0.9949, 0.7915, 0.2012],
[0.9938, 0.7974, 0.1974],
[0.9923, 0.8034, 0.1939],
[0.9906, 0.8095, 0.1906],
[0.9885, 0.8156, 0.1875],
[0.9861, 0.8218, 0.1846],
[0.9835, 0.828, 0.1817],
[0.9807, 0.8342, 0.1787],
[0.9778, 0.8404, 0.1757],
[0.9748, 0.8467, 0.1726],
[0.972, 0.8529, 0.1695],
[0.9694, 0.8591, 0.1665],
[0.9671, 0.8654, 0.1636],
[0.9651, 0.8716, 0.1608],
[0.9634, 0.8778, 0.1582],
[0.9619, 0.884, 0.1557],
[0.9608, 0.8902, 0.1532],
[0.9601, 0.8963, 0.1507],
[0.9596, 0.9023, 0.148],
[0.9595, 0.9084, 0.145],
[0.9597, 0.9143, 0.1418],
[0.9601, 0.9203, 0.1382],
[0.9608, 0.9262, 0.1344],
[0.9618, 0.932, 0.1304],
[0.9629, 0.9379, 0.1261],
[0.9642, 0.9437, 0.1216],
[0.9657, 0.9494, 0.1168],
[0.9674, 0.9552, 0.1116],
[0.9692, 0.9609, 0.1061],
[0.9711, 0.9667, 0.1001],
[0.973, 0.9724, 0.0938],
[0.9749, 0.9782, 0.0872],
[0.9769, 0.9839, 0.0805]]

```

NOTE: This is a copy of ""parula"" from MATLAB. I don't claim any rights to this data, but The Mathworks does. Consult them and/or a lawyer if you want to use it.


"
"CSS: How to make left float div to adjust height dynamically?
How can I make my floated left div with a static content to adjust automatically its height to the same height of the right floated div with a dynamic content?


So what I am trying to accomplish is that the both left and right div will have the same height (left div adjusting automatically to the right div's height)


Below is the sample code.


Thanks in advance :)


Cheers,
Mark 



```
<html>
<head>
    <style type=""text/css"">
        body {
            font-family:verdana;
            font-size:12px;
        }
        .parent {
            border:1px solid red;
            width:530px;
            /*min-height:100%;*/
            padding:5px;
        }
        .left {
            border:1px solid blue;
            width:200px;
            float:left;
            position:relative;
            padding:3px;
        }
        .right {
            border:1px solid green;
            width:300px;
            float:right;
            position: relative;
            padding:3px;
        }
        .clr { clear:both; }
        .footer {
            border:1px solid orange;
            position: relative;
            padding:3px;
            margin-top:5px;
        }
    </style>
</head>
<body>
    <div class=""parent"">
        <div class=""left"">float left div here only static content</div>
        <div class=""right"">
            float right div dynamic content here<br />
            float right div dynamic content here<br />
            float right div dynamic content here<br />
            float right div dynamic content here<br />
            float right div dynamic content here<br />
            float right div dynamic content here<br />
            float right div dynamic content here<br />
            float right div dynamic content here<br />
        </div>
        <div class=""clr""></div>
        <div class=""footer"">Footer here</div>
    </div>
</body>
</html>

```

","Here is the working CSS solution (thanks to pdr for the [link](http://www.ejeliot.com/samples/equal-height-columns/example-6.html)):



```
<html>
<head>
    <style type=""text/css"">
        html, body {
            font-family:verdana;
            font-size:12px;
        }
        .parent {
            border:1px solid red;
            width:530px;
            padding:5px;
            overflow:hidden;
        }
        .left {
            border:1px solid blue;
            width:200px;
            float:left;
            position:relative;
            padding:3px;
        }
        .right {
            border:1px solid green;
            width:300px;
            float:right;
            position: relative;
            padding:3px;
        }

        /* a solution but doesn't work in IE */
        /*
        .left, .right {
            min-height: 100px;
            height: auto;
        }
        */

        .left, .right {
            padding-bottom:8000px;
            margin-bottom:-8000px;
            background:none repeat scroll 0 0 lightblue;
        }

        .clr { clear:both; }
        .footer {
            border:1px solid orange;
            position: relative;
            padding:3px;
            margin-top:5px;
        }
    </style>
</head>
<body>
    <div class=""parent"">
        <div class=""left"">float left div here only static content</div>
        <div class=""right"">
            float right div dynamic content here<br />
            dynamic row<br />
            dynamic row<br />
            dynamic row<br />
            dynamic row<br />
            dynamic row<br />
            dynamic row<br />
            dynamic row<br />
            dynamic row<br />
            dynamic row<br />
            dynamic row<br />
        </div>
        <div class=""clr""></div>
        <div class=""footer"">Footer here</div>
    </div>
</body>
</html>

```

"
"Object oriented programming , inheritance , copy constructors
Suppose I have a base class `Person` and I publicly inherit a class `Teacher` from base class `Person`. 
Now in the main function I write something like this



```
// name will be passed to the base class constructor and 17
// is for derived class constructor.
Teacher object(“name”,17) ;
Teacher object1=object; //call to copy constructor

```

Now I have not written the copy constructor for both the classes, off course the default copy constructors will be called. The Person class’s default copy constructor will first call the base class’s copy constructor.


Now the problem is suppose I write the copy constructor for the base class only, what happens is, the default copy constructor of the derived class will call my written copy constructor.   

Now suppose I write the copy constructor for both the classes . now the copy constructor of the derived class (i.e Teacher) will call the default constructor of the base class but not the copy constructor why?
  

**Is only default copy constructor of the derived class can call the copy constructor of the base class automatically?**


","You have to call the base copy constructor explicitly:



```
Teacher(const Teacher& other) 
    : Person(other) // <--- call Person's copy constructor.
    , num_(other.num_)
{
}

```

Otherwise `Person`'s default constructor will be called.




---


I seem to not fully understand the question so I'll just say everything I think is relevant and hopefully this will help the OP.


All user defined constructors call their base's default constructor by default (unless they explicitly call a different constructor), it doesn't matter if the base's default constructor is user defined or compiler generated. 


When a copy constructor is generated by the compiler it will call the base class's copy constructor.


Compiler defined constructors are not special, they can be called explicitly:



```
class Base {
    int num_
public:
    Base(int n) : num_(n) { }
    // copy constructor defined by compiler
};

class Derived : public Base {
    float flt_;
public:
    Derived(float f, int n) : Base(n), flt_(f) { }
    // Copy constructor
    Derived(const Derived& other)
        : Base(other) // OK to explicitly call compiler generated copy constructor
        , flt_(other.flt_)
    {
    }
};

```

For more details see this [Wikipedia article](http://en.wikipedia.org/wiki/Special_member_functions).


"
"calculating the number of “inversions” in a permutation
Let A be an array of size `N`.
we call a couple of indexes `(i,j)` an ""inverse"" if `i < j` and `A[i] > A[j]`


I need to find an algorithm that receives an array of size `N` (with unique numbers) and return the number of inverses in time of `O(n*log(n))`.


","You can use the [merge sort](http://en.wikipedia.org/wiki/Merge_sort) algorithm.


In the merge algorithm's loop, the left and right halves are both sorted ascendingly, and we want to merge them into a single sorted array. Note that all the elements in the right side have higher indexes than those in the left side.


Assume **array[leftIndex] > array[rightIndex]**. This means that all elements in the left part following the element with index **leftIndex** are also larger than the current one in the right side (because the left side is sorted ascendingly). So the current element in the right side generates **numberOfElementsInTheLeftSide - leftIndex + 1** inversions, so add this to your global inversion count.


Once the algorithm finishes executing you have your answer, and merge sort is **O(n log n)** in the worst case.


"
"Set default value of validation
I am using Joi to validate a payload of a service in my node.js server using hapijs framework. It used to look like this (in my typescript code as well as after compiling to javascript):



```
payload: {
    para1: Joi.number().required(),
    para2: Joi.string()
}

```

Now I want to set default value of the two parameters. If the code is written in javascript, I can do this:



```
payload: {
    para1: Joi.number().required().default(1),
    para2: Joi.string().default(""defaultstring"")
}

```

I tested it in swagger and the default values actually became the values I set.


However, my project is written in typescript. I did the same thing and compiled typescript code. The result javascript looks like this: 



```
 payload: {
    para1: Joi.number().required()[""default""](1),
    para2: Joi.string()[""default""](""defaultstring"")
 }

```

In swagger, the default values are not applied.


Here are my questions:


1. why the code becomes different after compiling?
2. what does `[""default""](""defaultstring"")` mean and what does it do?
3. how can I write typescript code to make sure it can compiled as `Joi.string().default(""defaultstring"")`


**Update**


According to @rsp's post, the format in question 2 is just different way to access object's property. I also get reference from [here](https://www.w3schools.com/js/js_properties.asp). But it doesn't explain if they have any difference. Does anyone have any idea?


**Update2**


[Here](https://stackoverflow.com/questions/4968406/javascript-property-access-dot-notation-vs-brackets) is the difference between the two ways accessing JS property. It seems there is no negative effect using brackets way. However, in my case, the default values are not reflected on swagger. Will be doing research on it.


","In JavaScript this:



```
required().default(1)

```

is the same as this:



```
required()[""default""](1)

```

because you can access object properties either as:



```
object[""propertyName""]

```

or:



```
object.propertyName

```

(with certain restrictions in the second case).


So it's strange that TypeScript would output the longer style if it doesn't have to, but it's also strange that the longer style doesn't work exactly the same as the shorter one.


I would try to manually change the compiled JavaScript to the shorter version and see if that helps. If it doesn't then the problem is somewhere else. My suspicion is that it will not help.


The `.default()` should work in TypeScript because it is defined in [@types/joi](https://www.npmjs.com/package/@types/joi) - see: 


- <https://github.com/DefinitelyTyped/DefinitelyTyped/blob/master/joi/index.d.ts#L272-L273>


But on the other hand there is this comment:



```
// TODO express type of Schema in a type-parameter (.default, .valid, .example etc)

```

Which may suggest that `.default()` implementation is not ready yet - see:


- <https://github.com/DefinitelyTyped/DefinitelyTyped/blob/master/joi/index.d.ts#L6>


and also there's this issue: `joi.d.ts out of date, missing types`


- <https://github.com/DefinitelyTyped/DefinitelyTyped/issues/9332>


"
"Any algorithm for ""Flip all"" (Light Out) game?
in this game: <http://www.mathsisfun.com/games/allout.html>
The solve function can solve any case, no matter how you ""abuse"" the original board. Please tell me the algorithm for solving this game. I have tried to think for days but still found no clue to solve all cases.


**OK, after read some answers and comments (and have a quick look at Light out game), I expand my question:**


Will the game different if I expand the size of the grid (like to 25x25)? Still any possible algorithm to solve *any* case, *in acceptable* time (< 2s)?


","This game is more commonly known as Lights Out, and has a number of elegant solutions, all based in some standard but somewhat advanced mathematics. I won't describe them all here but if you Google a bit you can find all kinds of explanations varying from straightforward procedures to transformations into linear algebra or group theory. A few links:


<http://www.hamusutaa.com/pilot/solution.html>


<http://www.ripon.edu/academics/macs/summation/2010/articles/M.%20Madsen%20-%20Lights%20Out.pdf>


<http://people.math.sfu.ca/~jtmulhol/math302/notes/24-Lights-Out.pdf>


**Edit**: Re: your second question. The algorithm presented in the second link I posted can solve an n x n board in O(n^6) time, meaning you should be able to quickly solve a 25 x 25 board.


"
"Matlab: How to change the linewidth in a figure before actually plotting some?
This question maybe a bit like the link below, but this didn't work for me... <http://nl.mathworks.com/matlabcentral/answers/102530-how-can-i-change-the-default-settings-for-the-linewidth-property-before-i-plot-a-figure-in-matlab>


I'm working on a matlab function that automatically opens your figure in full screen mode and on a second monitor if present. So far, everything works fine. I already achieved to set the fontsize inside the function, so whitout plotting anything and without making xlabel(..) etc.:



```
% Fontsize used at the figure
if ~exist('fontsize_manual','var')|| isempty(fontsize_manual)
    set(gca,'FontSize',16)
else
    set(gca,'Fontsize',fontsize_manual)
end

```

Now is my question: Can I change in a same way the linewidth of the lines that will by plotted in the figure? So also here, predefining the linewidth inside the function and later on in your script plotting some lines etc. I do prefer this works only for the figure you're working on, so that you can change this 'default' for each figure and save them all with different linewidth and fontsizes if needed.


I tried the line below, but that only changed the linewidth of the axis.



```
set(gca,'LineWidth',2)

```

Is there anyone who can help me solving this problem?


%------------------------------------------------------------------------------------------------------------------------------
The answer below is nice, but I detected a new problem.
The code below in found accidentally by solving the previous problem:



```
set(gca,'LineWidth',3)

```

It turned out this changes the width of the axes. But now the problem... Also here this works only on the first figure. (see figure)
![enter image description here](https://i.stack.imgur.com/1kZ13.jpg)


If I also put this code in my session after the plotting in the second figure, the width in the second figure changes. Looks like the right handle isn't reached, or something like that, inside the function, when making the second figure. Do you have any idea what could be wrong here?


","I think what you are after is the `DefaultLineLineWidth` property, to which you can assign a value for a particular figure (or the root). 


Here is a simple code illustrating; basically I create a figure, set its 'visible' property to 'off' and assign a default line linewidth (that sounds weird...). The line plotted has a linewidth of 4, whereas another plot created after has the default width:



```
clear
clc

hFig1 = figure('Visible','off'); %// Create figure, set it to not visible.

set(gcf,'DefaultLineLineWidth',4); %// Assign default linewidth.

x = 1:10;
plot(x,x.^2-5);

set(hFig1,'Visible','on')
title('Figure 1','FontSize',16);

hFig2 = figure;
plot(x,2.*x+rand(1,10));
title('Figure 2','FontSize',16);

```

Plots:


![enter image description here](https://i.stack.imgur.com/grr67.jpg)


![enter image description here](https://i.stack.imgur.com/vwhsu.jpg)


Hope that help!


"
"After pushing to a review repository, ""abort: can't rebase immutable changeset"" on rebase
We have a code review repository where people `hg push -f` all sorts of stuff. After reviews are complete, we pull from the project's central repository, rebase, and push. I recently upgraded to mercurial 2.1 and get this message:



```
abort: can't rebase immutable changeset 43ab8134e7af
(see hg help phases for details)

```

when I try to `hg pull --rebase` from the central repository. How do I fix it?


","In the review repository's .hg/hgrc file, add these lines:



```
[phases]
publish = False

```

The problem is due to a new feature in mercurial 2.1 called [phases](https://www.mercurial-scm.org/wiki/Phases). It's great. [Here](http://www.logilab.org/blogentry/88203) is a nice introduction to its use.


To make the changesets in question mutable right now, use `hg phase -f -d REV` to force REV to be mutable again. Once the hgrc file has been changed, you shouldn't have to do that any more.


As a side note, `hg push -f` is lame. Make an alias `hg review` that pushes with `-f` to that repository.


"
"How can I access my Javascript Variables in PHP?
I have a file called `lightstatuspage.php` and within it, I have HTML, JavaScript and PHP code. I have used some variables within the JavaScript part of the code and I am trying to send these variables to the server by passing them to the PHP part of the code. However, this is not working. 


I am using `$.post(""lightstatuspage.php"", slider_val);` and then in the PHP part, I am calling the variable by doing `$_GET['rangeslider_val'];`. 


What am I doing wrong and what can I do differently to get the variable from JavaScript and send it to the server?



```
    function show_value(x)
    {
       document.getElementById(""slider_value"").innerHTML=x;

       event.preventDefault();
       var slider_val = x;
       var query = new Parse.Query(Post);
       query.first({
          success: function(objects){
            objects.set(""slider_val"", slider_val);
            objects.setACL(new Parse.ACL(Parse.User.current()));

            return objects.save();

            window.location.href = ""lightstatuspage.php?rangeslider_val="" + slider_val;
          }
       })
    }

```

The PHP code is: 



```
<?php
    $_GET['rangeslider_val'];

    print($rangeslider_val);
?>

```

","First Add Jquery



```
<script src='https://code.jquery.com/jquery-1.11.3.min.js'></script>

```

to the end of page before closing body tag.
To send Javascript variable to PHP. the best way is to use Ajax. and add the following code to your javascript.
**Do not forget that the below code should be on an event. For example on a button click or something like that**



```
$( document ).ready(function() {
    var x = $('#input1').val();
    //or var x= 15;
    $.post(""lightstatuspage.php"",
    {
      'rangeslider_val': x
    },
    function(data, status){
    //alert(""Data: "" + data + ""\nStatus: "" + status);
   // you can show alert or not
    });
});

```

and in php, you can use:



```
$value = $_POST['field1'];

```

now your variable is in $value in php.


P.S:
Backend Page and HTML page should be on same domain or you have to check **Cross-Origin Resource Sharing**




---


**Second Solution**
as the User accepted this solution here would be the code:



```
$.get(""lightstatuspage.php?rangeslider_val="" + slider_val,function(res) {
      console.log(res);
});

```

the second way is only the difference between POST and GET method




---


**Third Solution**
if you don't want to use Jquery in your project and you need pure javascript you can use below code



```
    var str = ""Send me to PHP"";
    var xmlhttp = new XMLHttpRequest();
    xmlhttp.onreadystatechange = function() {
        if (xmlhttp.readyState == 4 && xmlhttp.status == 200) {
            console.log(xmlhttp.responseText);
        }
    };
    xmlhttp.open(""GET"", ""lightstatuspage.php?rangeslider_val="" + str, true);
    xmlhttp.send();

```

"
"ssh key passphrase works in windows but not in linux
I'm working to a project in git.
In Windows, I'm using git extensions to manage this project, and to access to the public repository they gave me a .ppk key. I load it into git extension, with the passphrase that they gave me, and it works.


Now I set a linux (ubuntu-32bit) virtual machine, and I want to access also from this machine to the repository.


From another thread that I've seen in this site, I use, to clone the repository, the following command:



```
ssh-agent bash -c 'ssh-add /home/myHome/mykey.ppk; git clone git@serveraddress:project.git'

```

Then, the shell tells me to insert the passphrase



```
Enter passphrase for /home/myHome/mykey.ppk:

```

But when I insert it, it tells me that's a bad passphrase. I've checked it a lot of times, and I'm sure that I use the same passphrase that I use in windows. So how can I use correctly the key in Linux?


Thanks in advance for your replies.


","The Linux SSH client (typically OpenSSH) can't read the PPK format used by the Windows SSH client Putty. You need to convert the ""PPK"" key given to you into an OpenSSH key first. Install ""putty"" on Linux and use the `puttygen` command line tool:



```
$ sudo aptitude install putty
$ mkdir -p ~/.ssh
$ puttygen ~/mykey.ppk -o ~/.ssh/id_rsa -O private-openssh

```

Enter your passphrase, and you'll get an OpenSSH-compatible key in the standard location `~/.ssh/id_rsa`. Afterwards you can just use `ssh-add`(without any arguments!) to add this key to the SSH agent.


Alternatively you can use the PUTTYgen program provided by putty on Windows.


"
"Concatenate constraints in phpunit
My question is, how I concatenate constrains in the clausule with of phpunit?
In the dummy example:



```
$test->expects ($this->once())
     ->method ('increaseValue')
     ->with ($this->greaterThan (0)
     ->will ($this->returnValue (null));

```

The parameter of the method **increaseValue** must be greater than 0, but If I need evaluate that this parameter must be less than 10.


How I concatenate `$this->lessThan(10)`?


","You can use the `logicalAnd` expression:



```
$test->expects ($this->once())
     ->method ('increaseValue')
     ->with ($this->logicalAnd($this->greaterThan(0), $this->lessThan(10)))
     ->will ($this->returnValue (null));

```

For a list of possible functions check the functions in: [`PHPUnit/Framework/Assert.php`](https://github.com/sebastianbergmann/phpunit/blob/3.5/PHPUnit/Framework/Assert.php#L2097) that don't start with ""assert""


## Complete Example



```
<?php

class mockMe {
    public function increaseValue($x) {
    }
}


class fooTest extends PHPUnit_Framework_TestCase {

    public function testMock() {
        $test = $this->getMock('mockMe');
        $test->expects($this->once())
             ->method('increaseValue')
             ->with($this->logicalAnd($this->greaterThan(0), $this->lessThan(10)))
             ->will($this->returnValue(null));
        $test->increaseValue(6);
    }

    public function testMockFails() {
        $test = $this->getMock('mockMe');
        $test->expects($this->once())
             ->method('increaseValue')
             ->with($this->logicalAnd($this->greaterThan(0), $this->lessThan(10)))
             ->will($this->returnValue(null));
        $test->increaseValue(12);
    }

}

```

## Result



```
 phpunit blub.php
PHPUnit 3.5.13 by Sebastian Bergmann.

.F

Time: 0 seconds, Memory: 4.25Mb

There was 1 failure:

1) fooTest::testMockFails
Expectation failed for method name is equal to <string:increaseValue> when invoked 1 time(s)
Parameter 0 for invocation mockMe::increaseValue(<integer:12>) does not match expected value.
Failed asserting that <integer:12> is less than <integer:10>.

/home/.../blub.php:26

```

"
"WCF behind a public reverse proxy which is used for traffic encryption
I have a Silverlight application that connects to a WCF service. Under the basic configuration I am used to, there's no problem connecting this application to its corresponding WCF service. 


However, recently, one of my clients started using an Apache reverse proxy. This proxy is the public server and it's only used to encrypt HTTP traffic via SSL (HTTPS) going between the client and it. This proxy passes all traffic from it to the actual web server that hosts my application. The traffic between the public proxy and the IIS server is just plain HTTP.


**So the traffic flows like this:**
End-User Browser ---HTTPS----> Public Reverse Proxy -----HTTP----> IIS server that hosts the WCF service.


The reverse proxy and IIS are on two separate servers.


I cannot get the Silverlight application to function properly. I am not sure how to configure the endpoints? I get problems whenever I use the public proxy's address as my endpoint address.


The Silverlight application usually has this configuration: 



```
<configuration>
    <system.serviceModel>
        <bindings>
            <basicHttpBinding>
                <binding name=""BasicHttpBinding_IPOTemplateEditorSrv"" maxBufferSize=""2147483647""
                    maxReceivedMessageSize=""2147483647"">
                    <security mode=""TransportWithMessageCredential"" />
                </binding>
            </basicHttpBinding>
        </bindings>
        <client>
            <endpoint address=""https://public-reverse-proxy-url/POTemplateEditorSrv.svc""
                binding=""basicHttpBinding"" bindingConfiguration=""BasicHttpBinding_IPOTemplateEditorSrv""
                contract=""POEditorSrvRef.IPOTemplateEditorSrv"" name=""BasicHttpBinding_IPOTemplateEditorSrv"" />
        </client>
    </system.serviceModel>
</configuration>

```

Note that I am using and I have my endpoint address pointing to the public HTTPS address of the reverse proxy.


Am I missing anything? Is there any additional information to configure the proxy perhaps? Any workarounds that would get my Silverlight client connect to the service?


","Perhaps this answer is a little too obvious, but it simply sounds like the WSDL is advertising an internal host-name as the WCF address - when that address is not the actual public one.
Because IIS is generating the WSDL, it will simply use it's host name in the endpoint addresses - which is not what you want, you want the proxy's address.


Try creating a static copy of your WSDL file, and publish that on your web server. Make sure you replace ALL REFERENCES to the internal host name, with the public proxy host name.
Then modify your WCF client config to point to the static WSDL.
You can find a short explanation here: [Supply a different endpoint address in the WSDL of a WCF web service](https://stackoverflow.com/questions/5073936/supply-a-different-endpoint-address-in-the-wsdl-of-a-wcf-web-service)


If that doesnt work - try using a sniffer (wireshark) to capture what is being sent back and forth - disabling HTTPS might be a piece you need to remove from the equation.
Your web service request appears to be SENT to the proxy, but the proxy is not able to handle the request properly - the perfect scenario to try our your sniffing tools.


When you make a direct request to your SVC using a web browser, the request will look something like this



```
GET /POTemplateEditorSrv.svc HTTP/1.1
Host: public-reverse-proxy-url

```

But when sent via Silverlight, it may look like this



```
GET /POTemplateEditorSrv.svc HTTP/1.1
Host: private-server-address

```

This could be a subtle enough difference to upset the proxy.


"
"Simple Injector - Constructor Inject the current Web API request's HttpRequestMessage
In my Web API project, I have a dependency that requires the current request. 


The code is below:



```
public interface IResourceLinker {
  Uri Link(string routeName, object routeValues);
}

public class ResourceLinker : IResourceLinker {
  private readonly HttpRequestMessage _request;

  public ResourceLinker(HttpRequestMessage request) {
    _request = request;
  }

  public Uri Link(string routeName, object routeValues) {
    return new Uri(_request.GetUrlHelper()
        .Link(routeName, routeValues));
  }
}

public class TestController : ApiController {
  private IResourceLinker _resourceLinker;

  public TestController(IResourceLinker resourceLinker) {
    _resourceLinker = resourceLinker
  }

  public Test Get() {
    var url = _resourceLinker.Link(""routeName"", routeValues);

    // etc.
  }
}

```

Using Simple Injector, is it possible to inject the current request into the container at runtime ?


I tried the following :



```
public class InjectRequestHandler : DelegatingHandler
{
  protected override Task<HttpResponseMessage> SendAsync(
    HttpRequestMessage request, CancellationToken cancellationToken)
  {
    InjectRequest(request);
    return base.SendAsync(request, cancellationToken);
  }

  public static void InjectCurrentRequestIntoContainer(
    HttpRequestMessage request)
  {
    var resolver = (SimpleInjectorDependencyResolver)
      request.GetDependencyScope();
    resolver.Container.Register(() => request);
  }
}

```

but recieved the following error



> 
> The container can't be changed after the first call to GetInstance, GetAllInstances and Verify.
> 
> 
> 


Is there any way to inject the current request into the container at runtime?


","The container blocks any registrations after the registration phase. This splits the usage of the container into two phases: registration and resolve. Simple Injector is not the only container that does this. Autofac for instance, does this even more explicitly by allowing users to make registrations using the `ContainerBuilder` that builds the container using the `Build` method as last step in the configuration phase.


Simple Injector disallows this mainly because making registrations later on during the application lifetime can easily lead to all sorts of problematic behavior such as race conditions. It also makes the DI configuration much harder to understand, since registrations are scattered throughout the application, while you should try to centralize registration when applying DI. As a side effect, this design allows the container have a linear performance characteristic in a multi-threading scenario, because the container's happy path is free of locks.


Simple Injector allows just-in-time registration using the `ResolveUnregisteredType` event, but this is not the way to go in your case.


The problem you are having is that you want to inject an object that is only known at runtime into the object graph. This might not be the best thing to do, since in general you should pass runtime dependencies through method arguments, while compile-time/configuration-time dependencies should be passed through the constructor.


But if you are certain that passing the `HttpRequestMessage` as a constructor argument is the right thing to do, what you need to do is to cache this message during the lifetime of the request and make a registration that allows to return that cached instance.


This is how this would look like:



```
// using SimpleInjector.Advanced; // for IsVerifying()

container.Register<HttpRequestMessage>(() =>
{
    var context = HttpContext.Current;

    if (context == null && container.IsVerifying())
        return new HttpRequestMessage();

    object message = context.Items[""__message""];
    return (HttpRequestMessage)message;
});

```

This registration will retrieve a `HttpRequestMessage` that is cached in the `HttpContext.Items` dictionary. There's an extra check to allow this registration to work during verification (when calling `container.Verify()`), since at that point in time there is no `HttpContext.Current`. But if you're not interest in verifying the container (which you usually [really should](https://simpleinjector.readthedocs.org/en/latest/howto.html#verify-the-containers-configuration) btw), you can minimize it to this:



```
container.Register<HttpRequestMessage>(() =>
    (HttpRequestMessage)HttpContext.Current.Items[""__message""]);

```

With this registration, the only thing you have to do is cache an `HttpRequestMessage` instance *before* `container.GetInstance` is called to get the controller:



```
public static void InjectCurrentRequestIntoContainer(
    HttpRequestMessage request)
{
    HttpContext.Current.Items[""__message""] = request;
}

```

**UPDATE**


The [Web API Integration package](https://www.nuget.org/packages/SimpleInjector.Integration.WebApi/) contains a `GetCurrentHttpRequestMessage` extension method that allows you retrieving the `HttpRequestMessage` of the current request. The Web API Integration Wiki page [describes](https://simpleinjector.readthedocs.org/en/latest/webapiintegration.html#getting-the-current-request-s-httprequestmessage) how to use this extension method.


"
"1-click software release
I am rewriting a vb6 installer into NSIS. One of my priorities is to compile-to-release in the least number of steps possible; ideally, a one click process, in which all needed files are included, registered, and put in the right folders.


At the moment, I am creating several folders. One for app\_path files, one for system\_files, etc. So, every time that a file is modified or a new file added, I just drop it in the right folder and recompile. However, if a completely new file is included, steps like registering a DLL still need to be implemented in code. 


Any suggestion on how to improve this approach?, or on how to design a one-click compile-to-release process? 


EDIT:


The original installer was created using the Package and Deployment wizard from Visual Basic 6.0. Every time that I need to release my software, I found my self updating and adding files, updating the respective lst file, and repackaging everything (CAB files, dlls, etc) into one exe in order to release it from our website. It is a cumbersome process.


Additionally, due to pure design. The original software has been copy pasted in several projects, creating several application and its respective VB6 installers (repeating a lot of common files). Ideally, there would be one single installer, which allows the installation of the different executable and manages all the shared files among them. 


","Time for some serious ""Continous Integration"" for you.


These are the basic ingredients:


- a source code repository ([Subversion](http://subversion.tigris.org/) / [Git](http://git-scm.com/) / [Mercurial](http://mercurial.selenic.com/))
- an automated build tool ([Cruisecontrol.net](http://confluence.public.thoughtworks.org/display/CCNET/Welcome+to+CruiseControl.NET) / [Hudson](http://hudson-ci.org/) or [Jenkins](http://jenkins-ci.org/))
- a package-creator to create an installation-package (I recommend [Advanced Installer](http://www.advancedinstaller.com/))


First you setup your repository. Put all the source files of the application into it. You can use it to manage source code versions, features, branches... but that's a whole topic by itself.


Then get a build tool like [Cruisecontrol.net](http://confluence.public.thoughtworks.org/display/CCNET/Welcome+to+CruiseControl.NET) or [Jenkins](http://jenkins-ci.org/). Make it fetch and compile your code from the repository. The build tool can then take further actions like executing tests or building an installation package. External tools are often integrated as ant/nant tasks. You can integrate almost anything.


That's about everything you need to a one-click to installation-package build cycle.


Of course, there've been written whole books on these subjects. Read some articles on the web and get yourself the software that suits your needs the most. Then read some tutorials. You'll find everything via google.


"
"How can I connect to GNUnet 0.8.1 on 11.10?
As the SOPA/PIPA storm was brewing, I thought I should give some of the new tools people were talking about a try. After reading around and checking what was available in the repos, I picked GNUnet. I want to be able to download Linux ISOs on P2P without anyone being able to tell if I have defected to Sabayon :)


([You shouldn't use BitTorrent over Tor.](https://blog.torproject.org/blog/bittorrent-over-tor-isnt-good-idea) It's too heavy for that network, and your IP address gets revealed at the end point.)


I have installed GNUnet Secured P2P 0.8.1 (package name `gnunet-gtk`) and `gnunet-server`.


I ran the configuration wizard, picking my network interface and all the default options. I got a prompt saying `Unable to change startup process: Success`. I ran it again and this time also ticked ""Start the GNUnet background process"". It then completed with no prompt. (My network interface defaults to `lo` rather than `wlan0` each time. I leave IP/hostname blank because I have a dynamic IP address - should I enter `computername.local`?)


I ran the advanced configuration tool, and changed the hostlists from `http://gnunet.org/hostlist.php http://gnunet.mine.nu:8081/hostlist http://vserver1236.vserver-on.de/hostlist-074 http://mosquito.dyndns.tv/gnunet-hostlist/hostlist http://de.gnunet.org/hostli` to `http://mosquito.dyndns.tv/gnunet-hostlist/hostlist http://de.gnunet.org/hostlist-080 http://vserver1236.vserver-on.de/hostlist-074`. (Per <https://gnunet.org/node/1190> and <http://lists.gnu.org/archive/html/help-gnunet/2010-10/msg00000.html> )


When the GNUnet GUI starts the message bar reads ""Daemon running"". After a while it reads ""Disconnected"".


What do I need to do to connect?


","Apparently, GNUnet version 0.8 is depreciated and all the effort is being invested in developing version 0.9. ([Source](https://bugs.launchpad.net/getdeb.net/+bug/836348/comments/1))


I have used the steps from the [installation instructions for Ubuntu from GNUnet site](https://gnunet.org/book/export/html/445) for installing GNUnet version 0.9. 


First, make sure Subversion is installed on your system:



```
sudo apt-get install subversion

```

Then, install libextractor dependencies:



```
sudo apt-get install automake autopoint autoconf libtool make gcc g++ gettext texinfo
sudo apt-get install zlib1g-dev libgsf-1-dev libmpeg2-4-dev libpoppler-dev

```

Build libextractor:



```
svn checkout https://gnunet.org/svn/Extractor
cd Extractor
./bootstrap
./configure
make; sudo make install; make check
cd ..

```

*Note that 'make check' is optional here.*


Install libmicrohttpd dependencies:



```
sudo apt-get install gnutls-dev libcurl4-gnutls-dev

```

Build libmicrohttpd:



```
svn co https://gnunet.org/svn/libmicrohttpd
cd libmicrohttpd/
./bootstrap
./configure --with-gnutls --enable-messages --enable-curl
make; sudo make install; make check
cd ..

```

*Again, 'make check' is optional (and may fail if you are running services already on certain ports, like 1080).*


Install GNUnet dependencies:



```
sudo apt-get install libltdl7-dev automake autoconf libtool make gcc texinfo
sudo apt-get install libgcrypt11-dev libgmp3-dev libcurl4-gnutls-dev cvs libunistring-dev
sudo apt-get install libmysqlclient15-dev libsqlite3-dev libpq-dev
sudo apt-get install libesmtp-dev libpcap-dev

```

Build GNUnet:


You can specify the location of the GNUnet installation by setting the prefix when calling the configure script: `--prefix=$HOME/gnbuild`


To build GNUnet execute:



```
svn checkout https://gnunet.org/svn/gnunet/
cd gnunet/
./bootstrap
./configure --prefix=$HOME/gnbuild --with-extractor=/usr/local
make; sudo make install

```

You have to set the GNUNET\_PREFIX environment variable according to the prefix you set for the configure script and you have to add your GNUnet installation to your search path:



```
export GNUNET_PREFIX=$HOME/gnbuild
export PATH=$PATH:$HOME/gnbuild/bin

```

Now you can test your installation by running:



```
make check

```

"
"Exporting data from php to excel
I need to export data from php (data retrieved from mysql database) to excel. I'm using Zend Framework. I need to do some changes to my data before exporting to excel. Actually, what I really need is to generate a monthly cash book. 


I have read a lot of documents, but ended up in a mess. I really can't understand how I should begin. Do I really need PEAR? Do I need to download any class library?Isn't there a simple way to do this?


Thanks in advance


","I would suggest you to use great `PHPExcel` library. It is really powerful, supports variety of formats, can do visual formatting and is easy to use.


You can find more about it at their webpage:
~~<http://phpexcel.codeplex.com/>~~. ~~(`PHPExcel` has already moved at <https://github.com/PHPOffice/PHPExcel>)~~


## 2019:


`PHPExcel` has now been superceded by [PhpSpreadsheet GitHub](https://github.com/PHPOffice/PhpSpreadsheet)
.


Example of usage:



```
    $objPHPExcel = new PHPExcel();
    /** You can set many properties */
    $objPHPExcel->getProperties()->setCreator(""My company"")
                 ->setLastModifiedBy(""John Doe"")
                 ->setTitle(""Annual report"")
                 ->setSubject(""Sales"")
                 ->setDescription(""Annual sales report by John Doe"")
                 ->setCategory(""Finance"");

    /** Choose one of sheets */
    $activeSheet = $objPHPExcel->setActiveSheetIndex(0);
    /** Simply set value of cell */
    $activeSheet->setCellValue(""A1"", 'plural');

```

You can do a lot more of course, reading excel files, setting visual styles, creating plots, expressions and lot more.


"
"Type check for iterator?
I have the following line of code in my project:



```
Iterator<Node> i = graph.nodes();

```

graph.nodes() is a method defined in prefuse.


Now I want to add a type check to the iterator, but how can I do that?


","If you use [Guava](http://code.google.com/p/guava-libraries/), you could do



```
Iterator<Node> iterator = Iterators.filter(theOriginalIterator, Node.class);

```

The returned iterator is guaranteed to continue only elements that correspond to the supplied type. ( A similar method is also available for [`Iterable`](http://docs.oracle.com/javase/6/docs/api/java/lang/Iterable.html) btw )


See 


- [`Iterators.filter(iterator, class)`](http://docs.guava-libraries.googlecode.com/git/javadoc/com/google/common/collect/Iterators.html#filter%28java.util.Iterator,%20java.lang.Class%29)
- [`Iterables.filter(iterable, class)`](http://docs.guava-libraries.googlecode.com/git/javadoc/com/google/common/collect/Iterables.html#filter%28java.lang.Iterable,%20java.lang.Class%29)


"
"Encode categorical features with multiple categories per example - sklearn
I'm working on a movie dataset which contains *genre* as a feature. The examples in the dataset may belong to multiple genres at the same time. So, they contain a list of genre labels.


The data looks like this-



```
    movieId                                         genres
0        1  [Adventure, Animation, Children, Comedy, Fantasy]
1        2                     [Adventure, Children, Fantasy]
2        3                                  [Comedy, Romance]
3        4                           [Comedy, Drama, Romance]
4        5                                           [Comedy]

```

I want to vectorize this feature. I have tried *LabelEncoder* and *OneHotEncoder*, but they can't seem to handle these lists directly.


I could vectorize this manually, but I have other similar features that contain too many categories. For those I'd prefer some way to use the *FeatureHasher* class directly.


Is there some way to get these encoder classes to work on such a feature? Or is there a better way to represent such a feature that will make encoding easier? I'd gladly welcome any suggestions.


","[This SO question](https://stackoverflow.com/questions/29034928/pandas-convert-a-column-of-list-to-dummies) has some impressive answers. On your example data, the last answer by Teoretic (using `sklearn.preprocessing.MultiLabelBinarizer`) is 14 times faster than the solution by Paulo Alves (and both are faster than the accepted answer!):



```
from sklearn.preprocessing import MultiLabelBinarizer
mlb = MultiLabelBinarizer()
encoded = pd.DataFrame(mlb.fit_transform(df['genres']), columns=mlb.classes_, index=df.index)
result = pd.concat([df['movieId'], encoded], axis=1)

# Increase max columns to print the entire resulting DataFrame
pd.options.display.max_columns = 50
result
   movieId  Adventure  Animation  Children  Comedy  Drama  Fantasy  Romance
0        1          1          1         1       1      0        1        0
1        2          1          0         1       0      0        1        0
2        3          0          0         0       1      0        0        1
3        4          0          0         0       1      1        0        1
4        5          0          0         0       1      0        0        0

```

"
"how to split dart class in flutter?
I did the following test, but it doesn't work:



```
//main.dart
class Test
{
  static const   a = 10;
  final b = 20;
  final c = a+1;

}

//part.dart
part of 'main.dart';
class Test
{
  final d = a +1;   //<---undefined name 'a'
} 

```

I would like to split the class in [flutter tutorial](https://flutter.io/get-started/codelab/) into multiple files. For example: \_buildSuggestions in a separate file, \_buildRow in a separate file, etc.


**update:**


my solution:


before:



```
//main.dart
class RandomWordsState extends State<RandomWords> {
{
    final _var1;
    final _var2;
    @override
    Widget build(BuildContext context) {
      ...
      body: _buildList(),
    );

    Widget _buildList() { ... }
    Widget _buildRow() { ... }
}

```

after:



```
//main.dart
import 'buildlist.dart';
class RandomWordsState extends State<RandomWords> {
{
    final var1;
    final var2;
    @override
    Widget build(BuildContext context) {
      ...
      body: buildList(this),
    );

}

//buildlist.dart
import 'main.dart';

  Widget buildList(RandomWordsState obj) {
     ... obj.var1 ...
  }

```

","Dart doesn't support partial classes. `part` and `part of` are to split a **library** into multiple files, not a class.


Private (identifiers starting with `_`) in Dart is per library which is usually a `*.dart` file. 


**main.dart**



```
part 'part.dart';

class Test {
  /// When someone tries to create an instance of this class
  /// Create an instance of _Test instead
  factory Test() = _Test;

  /// private constructor that can only be accessed within the same library
  Test._(); 

  static const   a = 10;
  final b = 20;
  final c = a+1;
}

```

**part.dart**



```
part of 'main.dart';
class _Test extends Test {
  /// private constructor can only be called from within the same library
  /// Call the private constructor of the super class
  _Test() : super._();

  /// static members of other classes need to be prefixed with
  /// the class name, even when it is the super class
  final d = Test.a +1;   //<---undefined name 'a'
} 

```

A similar pattern is used in many code-generation scenarios in Dart like in 


- <https://pub.dartlang.org/packages/built_value>
- <https://pub.dartlang.org/packages/built_redux>
- <https://pub.dartlang.org/packages/json_serializable>


and many others.


"
"How to keep $.ajax() async inside $.each() loop but react to result on each iteration
I'm having trouble with some dumb architecture because I'm dumb. I'm trying to loop over YouTube videos posted to Reddit, extract URLs and process them into an .m3u playlist.


Full code at [**Subreddit to YouTube Source Bookmarklet** - Play YouTube music from subreddits in Foobar with foo\_youtube](https://gist.github.com/akaleeroy/b3de63c46ee43d413913)


At some point I get the idea that I could check each URL to see if the video is dead or not, and offer an alternative if they're removed.


So I do AJAX requests to YouTube API and if there's an error I'm supposed to react to it and change that item's URL. 


But the problem is it only works if the AJAX is NOT async - this takes many seconds, during which the page is jammed.


I'd like to let the AJAX be async but I don't know how I should structure my code.


Here is **PSEUDOCODE** of how it is now:



```
var listing = // data from reddit API
$.each(listing, function(key, value) {
    var url = // post URL from reddit posts listing 
    // ( ""http://youtu.be/XmQR4_hDtpY&amp;hd=1"" )
    var aRegex = // to parse YouTube URLs 
    // ( (?:youtube(?:-nocookie)?.com/....bla...bla )
    var videoID = // YouTube video ID, extracted with regex 
    // ( ""XmQR4_hDtpY"" )
    var itemArtist = // parsed from reddit posts listing 
    // ( ""Awesome Artist"" )
    var itemTitle = // parsed from reddit posts listing 
    // ( ""Cool Song (Original Mix)"" )
    var itemURL = // url, further processed 
    // ( ""3dydfy://www.youtube.com/watch?v=XmQR4_hDtpY&hd=1"" )

    $.ajax({
            type: ""HEAD"",
            url: ""https://gdata.youtube.com/feeds/api/videos/"" + videoID,
            error: function() { 
                // If it's no longer available 
                // (removed, deleted account, made private)
                deadVid++; // chalk up another dead one, for showing progress
                itemURL = // dead videos should get a different URL 
           // ( ""3dydfy-search://q=Awesome%20Artist%20Cool%20Song....."" )
            }
        });

    // further process itemURL!
    // at this point I want itemURL changed by the .ajax()'s error callback
    // but I'm trying to keep the requests async 
    // to not jam the page while a hundred HTTP requests happen!
    if (condition){
        itemURL += // append various strings
    }
    // Write the item to the .m3u8 playlist
    playlist += itemURL + '\n';
}// end .each()

```

","Basically you want to know 


- What the errors were and
- When every ajax request is finished


If you push the errors into a list, the results will be ready at the end (order not guaranteed of course).


For the second part, if you keep an array of the ajax promises returned from each $.ajax you can use `$.when` and wait for them all to complete using `always()`.


As a basic example (other details removed):



```
var listing = {}; // data from reddit API
var promises = [];
var results = [];
$.each(listing, function(key, value) {
    // [snip]
    promises.push($.ajax({
            type: ""HEAD"",
            url: ""https://gdata.youtube.com/feeds/api/videos/"" + videoID,
            error: function() { 
                //[snip]
                results.push({
                   itemArtist: itemArtist,
                   videoID: videoID,
                   url: itemURL});
            }
        });
    );
}
// Wait for all promises to complete (pass or fail) 
$.when.apply($, promises).always(function(){
    // process list of failed URLs
});

```

*Apologies for any typos. This was coded straight into the answer, but you get the idea.*


I note you mention 100s of requests, but the browser will only allow a handful through at a time, so no need for additional processing.


If `always` is not working you can add your own deferred objects that resolve on `success` *or* `fail`:



```
var listing = {}; // data from reddit API
var promises = [];
var results = [];
$.each(listing, function(key, value) {
    var deferred = $.Deferred();
    promises.push(deferred.promise());
    // [snip]
    $.ajax({
            type: ""HEAD"",
            url: ""https://gdata.youtube.com/feeds/api/videos/"" + videoID,
            complete: function(){
                // resolve on success or fail
                deferred.resolve();
            },
            error: function() { 
                //[snip]
                results.push({
                   itemArtist: itemArtist,
                   videoID: videoID,
                   url: itemURL});
            }
        });
    );
}
// Wait for all promises to complete (pass or fail) 
$.when.apply($, promises).always(function(){
    // process list of failed URLs
});

```

## Update Dec 2015


Here is another cool way to chain parallel promises together, without using an array (so long as you do not need the data values passed through to the callbacks):


Simplified code:



```
   var promise;   // Undefined is also a resolved promise when passed to $.when()
   $.each(listing, function(key, value) {
       // Each new Ajax promise is chained, in parallel, with the previous promise
       promise = $.when(promise, $.ajax({...}));
   });
   // When they are all finished, fire a final callback
   $.when(promise).always(function(){
       // All done!
   });

```

This has had some criticism, mainly from people that feel it is ""unclean"", but for the simplifying of parallel promise code the trade-off is minimal.


I figured out this was possible when I saw someone use `promise = promise.then(newpromise)` to chain events in sequence. After some experimenting I found I could do the same in parallel using `promise = $.when(promise, newpromise)`


"
"Closing a file after File.Create
I check to see if a file exists with 



```
if(!File.Exists(myPath))
{
    File.Create(myPath);
}

```

However, when I go to create a `StreamReader` with this newly created file, I get an error saying that



> 
> The process cannot access the file '[my file path here]' because it is being used by another process.
> 
> 
> 


There isn't a `File.Close(myPath)` that I can call so that it is closed after being created, so how do I free this resource so that I can open it later in my program?


","[`File.Create(string)`](http://msdn.microsoft.com/en-us/library/d62kzs03.aspx) returns an instance of the [`FileStream` class](http://msdn.microsoft.com/en-us/library/system.io.filestream.aspx). You can call the [`Stream.Close()`](http://msdn.microsoft.com/en-us/library/system.io.stream.close.aspx) method on this object in order to close it and release resources that it's using: 



```
var myFile = File.Create(myPath);
myFile.Close();

```

However, since `FileStream` implements [`IDisposable`](http://msdn.microsoft.com/en-us/library/system.idisposable.aspx), you can take advantage of the [`using` statement](http://msdn.microsoft.com/en-us/library/yh598w02.aspx) (generally the preferred way of handling a situation like this). This will ensure that the stream is closed and disposed of properly when you're done with it:



```
using (var myFile = File.Create(myPath))
{
   // interact with myFile here, it will be disposed automatically
}

```

"
"Create a ""config"" or ""options"" page for a Greasemonkey script
I've written a simple Greasemonkey script, and I'm trying to create a ""config"" page for this script (like the one that is used for Google Chrome extensions. ) Would there be any way to create a config page for a userscript, like the ""options"" pages for Google Chrome extensions? There isn't any way to include an .html page as part of a Greasemonkey script (as far as I know), so I'm looking for other options.



```
// ==UserScript==
// @name       Redirector
// @namespace  http://use.i.E.your.homepage/
// @version    0.1
// @description  enter something useful
// @match      http://*/*
// @copyright  2012+, You
// @run-at document-start
// ==/UserScript==

redirectToPage(""http://www.youtube.com/"", ""http://www.google.com"");

function redirectToPage(page1, page2){
if(window.location.href.indexOf(page1) != -1){
    window.location.href = page2;
}
}

```

","There are a few libraries which provide config pages for userscripts:


### 1) [GM\_config](https://github.com/sizzlemctwizzle/GM_config/)


[![Example GM_config dialog](https://i.stack.imgur.com/TEkAT.png)](https://i.stack.imgur.com/TEkAT.png)


### 2) [MonkeyConfig](https://github.com/odyniec/MonkeyConfig)


[![Example MonkeyConfig dialog](https://i.stack.imgur.com/nLtcz.png)](https://i.stack.imgur.com/nLtcz.png)


### 3) [GM\_registerMenuCommand Submenu JS Module](https://github.com/SloaneFox/code)




---


The usage varies per library, but typically you grant the permissions they need such as `GM_getValue` and `GM_setValue`, and require the library via the `@require` directive, e.g.:



```
// ==UserScript==
// @name          My Userscript
// @description   An example userscript with a config page
// @version       0.0.1
// @require       https://www.example.com/lib/myconfig.js
// @grant         GM_getValue
// @grant         GM_setValue
// @grant         GM_addStyle
// @grant         GM_registerMenuCommand
// ==/UserScript==

const config = new MyConfig({ ... })

```

You then register a menu command which opens the config page/dialog, e.g.:



```
GM_registerMenuCommand('Configure My Userscript!', () => {
    config.open()
})

```

In the case of MonkeyConfig, it can register the command for you:



```
const config = new MonkeyConfig({
    title: 'Configure My Userscript!',
    menuCommand: true,
    // ...
})

```

For advanced uses, the configurator may allow [listeners](https://github.com/sizzlemctwizzle/GM_config/wiki/Event-Callbacks) to be registered for the close/cancel/save events, as well as providing control over the CSS, and other options. Detailed instructions can be found on the [GM\_config wiki](https://github.com/sizzlemctwizzle/GM_config/wiki) and the [MonkeyConfig homepage](https://web.archive.org/web/20190924061647/http://odyniec.net/projects/monkeyconfig/usage.html).


"
"Sublimelinter (PC): How to trigger this shortcut (""Lint This View""): Ctrl K, l?
I am trying to figure out how you use this shortcut, but I am having trouble.
What does it mean by ""Ctrl K, l""? 


I have tried various combinations but in vain.


![enter image description here](https://i.stack.imgur.com/qqvmW.jpg)


","Here is the default config file: [Default (Windows).sublime-keymap](https://github.com/SublimeLinter/SublimeLinter3/blob/master/Default%20%28Windows%29.sublime-keymap)


And here is the string which creates this shortcut:



```
{ ""keys"": [""ctrl+k"", ""l""], ""command"": ""sublimelinter_lint"" },

```

That means you press `ctrl`+`k` and then press `l`. It's a sequence key bind that you're allowed to make in your custom keymap file as well.


Make sure to have `ctrl` released while pressing `l`.




---


P.S. it's a small `L` and not capital `i`. How do I know? I copy pasted it to sublime text, selected and pressed `ctrl`+`k`, `ctrl`+`u`.


This is a default keymap: `{ ""keys"": [""ctrl+k"", ""ctrl+u""], ""command"": ""upper_case"" },`


"
"Forcibly Uninstalling Eclipse 3.5 Plugins
How do you forcibly uninstall Eclipse 3.5 plugins **without using Eclipse**? I made the mistake of installing Google's Android SDK, and now when I try to check for updates via the ""Check For Updates"" or ""Install New Software"", some Android process runs, consumes all my CPU, takes up 3 gigs of memory, and crashes Eclipse. I can't even get to the uninstall dialog before Google's plugin kills everything.


","One thing that you can try, is to go to your Eclipse directory, then ""plugins"" directory. There, remove the files and folders which are related to your SDK.


To find the concerned files, you should download the plugin file at [this link](http://developer.android.com/sdk/eclipse-adt.html#troubleshooting). By checking the content of this archive, you will know exactly what files got installed, and what you should delete.


It's a bit extreme, and you can possibly prevent other things from working if you remove too much, so I recommend you to back this folder prior to doing so.


There might be some configurations remaining, and some unstable state of the plugin, instead of its complete removal, but it could be enough for you to be able to open Eclipse, and remove it properly.


Note that some plugins have configuration files installed in the workspace you use, in the .metadata/.plugins folder.


"
"Subdivisions in statistics
I was having a discussion with a statistics professor a while ago about the different 'flavours' of statistics (frequentist, Bayesian, ...). He posed that he would subdivide statistics in four categories: non-parametric-, robust-, frequentist- and Bayesian statistics. The subdivision is characterized by the amount of assumption the methods make about underlying distributions (non-parametric statistics makes none, while Bayesian makes those assumptions very explicit). 


I was going to to ask if CrossValidated agrees with this subdivision, but since that is a subjective question I'll ask:


1) Is this subdivision widely recognized in statistics;


2) do 'real world' problem usually require one particular method? Ie, given some problem, is there a method most suitable for solving that problem or can multiple methods work for a given problem?


Thanks in advance.


","I wouldn't consider non-parametric or robust as being sub-categories of statistics in the way that frequentist and Bayesian are, simply because there are both frequentist and Bayesian methods for non-parametric and robust statistics. Frequentist and Bayesian are genuine sub-categories as they are based on fundamentally different definitions of a probability. Frequentists and Bayesians will both vary the strength of assumptions made depending on the requirements of the application.


So I would say that particular subdivision into four categories is not widely recognised in statistics. In my opinion, both Bayesian and frequentist methods can be used for most statistical problems, however they are not always equally useful, for example whether a frequentist confidence interval or a Bayesian credible interval is more appropriate depends on whether you want to ask a question about what to expect if the experiment were replicated, or what we can conclude about the statistics as a result of the particular experiment that we have actually performed (I would suggest in most cases it is the latter, but scientists generally use frequentist methods anyway).


"
"What happens when all IP addresses run out?
I recently read an article in which they wrote that that there will be a time when there will be no more IP addresses available. Is that right? Do I need to register a website now just in case?


","No, you don't need to worry.


Registering a website (domain name) is totally separate to having an IP address. Registering a domain name just gives you the name, this name will then be entered into DNS servers, which will allow the name to be resolved into an IP address. This will be the IP address of the web server that holds your site. The company running the web server or their ISP will be responsible for ensuring that they have valid IP address, you don't need to worry about it. It can also change, so you don't need to reserve a specific address.


Something called [Network address translation (NAT)](http://en.wikipedia.org/wiki/Network_address_translation) has relived a lot of the pressure on [IPv4](http://en.wikipedia.org/wiki/IPv4) addresses. It allows a single external IP address to be mapped to a large number of IP addresses internally within a smaller network. (Such as a home or company network).


We are moving towards support for [IPv6](http://en.wikipedia.org/wiki/IPv6), which will expand the number of available addresses significantly. Most modern operating systems now support it, and a a lot of network hardware is starting to support IPv6 now.


Unless you are a registrar or ISP you do not need to significantly worry about IpV4 addresses running out. Just consider it when buying hardware and try to buy IPv6 compatible hardware where possible.


You can read more about it [here](http://en.wikipedia.org/wiki/IPv4_address_exhaustion).


"
"Area2D Not Triggering Object in Godot
Here is my code: (just so you know I am a beginner and I just started this week, though I do have knowledge with other languages and game engines)



```
func _on_Area2D_area_entered(area):
    get_parent().get_node(""Level 1/Area2D/Flag"").rotation_degrees += 1

```

What I was trying to accomplish was that the Player GameObject would see if its in the area of the Flag, and if it is, the flag would rotate.


I am not sure where the issue is. I think it is probably in the second line. I provided a screenshot below if I did the setup wrong. I have looked at the other close questions asked on the same topic, but they did not answer my question.


[![enter image description here](https://i.stack.imgur.com/77iYc.png)](https://i.stack.imgur.com/77iYc.png)


The ""Player"" GameObject is the one with the script that contains the detection if its in the area2D.


","If you want to check if the `Area2D` is positioned correctly during runtime enable Debug -> Visible Collision Shapes.


If you want to check if `_on_Area2D_area_entered` is running, add breakpoints (or use print).




---


Did you get an error?


If there isn't a `Node` there, this expression will cause an error in runtime:



```
get_parent().get_node(""Level 1/Area2D/Flag"")

```

If you want to be able to check, you can use [`get_node_or_null`](https://docs.godotengine.org/en/stable/classes/class_node.html#class-node-method-get-node-or-null) and [`is_instance_valid`](https://docs.godotengine.org/en/stable/classes/class_@gdscript.html#class-gdscript-method-is-instance-valid).


*Since you didn't mention any error, I'm going to guess the method is not running.*




---


If the method is not running, the most likely culprit is that - I'm guessing given then name of the method - you connected the `""area_entered""` signal but intended to connect the `""body_entered""` signal.


The `""area_entered""` signal will trigger when another `Area2D` enters the `Area2D`. But I only see one `Area2D` in your scene tree. On the other hand the `""body_entered""` will trigger when a `PhysicsBody2D` (e.g `StaticBody2D`, `KinematicBody2D`, `RigidBody2D`) enters the `Area2D`. In either case you get what entered as a parameter of the method.


Other reasons why the `Area2D` might not be detecting what you want include no intersection of `collision_layer` and `collision_mask` and `monitoring` being disabled.




---


And to dismiss a couple possible misconceptions:


- The `""area_entered""` and `""body_entered""` trigger when the `Area2D` or `PhysicsBody2D` respectively enter the `Area2D`, not every frame they are inside. *So `rotation_degrees += 1` is not a rotation animation.*
- You will get notifications of anything that trigger the signals, not just the object to which you connected it. You may have to further filter, e.g. `if body == self:`.




---


*For people arriving here from search, I want to link a similar case: [Enemy is not affected by bullets](https://stackoverflow.com/a/72774189/402022). And also my [full explanation of how to set up physic nodes](https://stackoverflow.com/a/69731270/402022).*


"
"Get two keys at the same time - Assembly 8086
I am building a game in assembly, with two players.
So I'm using the following code to get the current key pressed:



```
mov ah, 01h
int 16h

```

This code gives me the Scan code and the Ascii code of the pressed key
The problem is that two players might press 2 different keys on the same time.
**How do I receive two keys at the same time?**


I have tried few ways to achieve that. I tried getting the input directly by using in `al,060h`. I also tried `int 9h` but still it didn't work.


","If memory serves, with a touch of Googling to refresh myself on the numbers...


The BIOS doesn't offer what you want. It's managing an input stream of text characters. Which is strongly related to catching key down events but will not allow you to tell which keys are depressed now, only which keys at some point weren't depressed and then became depressed in the past.


You need to use `int 21h`, function `25h` to install your own handler for `int 9h`. That way you'll get the key up and key down events directly from the hardware.


Within your handler, read port `60h` to determine why the interrupt has occurred. The low seven bits are a key code. If the top bit is clear then the key has transitioned to pressed. If the top bit is set then the key has transitioned to unpressed.


You then need to out `20h` to port `20h` to acknowledge the interrupt.


So you might keep a table of 128 bytes, initialised as all `80h`. In your handler, just store `value` to `value&7h`.


In your actual game code, if you want to know whether key `n` is pressed at that moment, read the nth value from your table and branch on sign. Positive = pressed, negative = unpressed.


(Addendum: you should also get the existing vector when you launch and restore it before you exit, or else you've just accidentally written the hooking part of a TSR, probably without doing the SR part)


"
"what is a 'layer' in a neural network
Below I've drawn a typical feed forward neural network:


[![Typical Feed Forward Neural Network](https://i.stack.imgur.com/KBY0x.png)](https://i.stack.imgur.com/KBY0x.png)


Now my question is, as far as lingo goes, what is a layer?


Could each individual process (rectangle) be considered a layer? or is a layer the combination a single row of the flow diagram? I sometimes see the Multiply + Add as a single layer, and the nonlinear function (relu) as a separate layer. But I would really appreciate a definitive answer.


I often find on online videos teaching people about Neural Networks, the instructors themselves mix up the number of layers within a single example. 


","With your diagram, each row is essentially a layer. But as @beaker states it is not the best way to visualize a neural network.


Taking an image from [here](http://cs231n.github.io/neural-networks-1/) will help make this clear.


[![enter image description here](https://i.stack.imgur.com/Kc50L.jpg)](https://i.stack.imgur.com/Kc50L.jpg)


Layer is a general term that applies to a collection of 'nodes' operating together at a specific depth within a neural network. 


The **input layer** is contains your raw data (you can think of each variable as a 'node'). 


The **hidden layer(s)** are where the *black magic* happens in neural networks. Each layer is trying to *learn* different aspects about the data by minimizing an error/cost function. The most intuitive way to understand these layers is in the context of 'image recognition' such as a face. The first layer may learn edge detection, the second may detect eyes, third a nose, etc. This is not exactly what is happening but the idea is to break the problem up in to components that different levels of abstraction can piece together much like our own brains work (hence the name 'neural networks').


The **output layer** is the simplest, usually consisting of a single output for classification problems. Although it is a single 'node' it is still considered a *layer* in a neural network as it could contain multiple nodes.


"
"Alias template specialisation
Can alias templates (14.5.7) be explicitly specialised (14.7.3)?


My standard-fu fails me, and I can't find a compiler to test on.


The text ""when a template-id refers to the specialization of an alias template"" implies *yes*, but then the example appears to refer to something else, implying *no*.


**NB.** I'm working from n3242, one behind the FDIS, in which the title of this section is ""Aliase templates"". Lol.


","What the Standard means by ""specialization"" is the transformation of a generic template to a *more specialized* entity. For example, instantiating a non-member class template yields a class that's not a template anymore. The term ""specialization"" is two fold, and can refer to a generated specialization (which is a specialization that was instantiated, possibly from a partial specialization) and to an explicit specialization (which is what you referred to). 


Alias templates aren't instantiated and there aren't specializations of them. There is nothing they could instantiate to. Instead, whenever their name is followed by a template argument list, the type denoted is the type you get by replacing the name and argument list by the alias'ed type, replacing all template parameter references with the arguments given in the argument list. That is, rather than the specialization of it being an alias, the alias template itself serves as an alias, without the need to instantiate anything. This replacement is done very early. Consider:



```
template<typename T> using ref = T&;
template<typename T> void f(ref<T> x) { x = 10; }
int main() { int a; f(a); return a; /* 10 */ }

```

The replacement is done at the time `ref<T>` is named (such a names are used to refer to class or function template specializations; hence the spec describes such names to ""refer to the specialization of an alias template""). That is, the parameter of `f` has type `T&`, and thus, `T` can be deduced. This property is preventing explicit or partial specializations of alias templates. Because in order to pick the correct specialization of `ref`, it needs to know `T`. But to know it, it needs to compare `ref<T>` against the argument type to deduce `T`. It's summarized in the paper [N1406, ""Proposed addition to C++: Typedef Templates""](http://www.open-std.org/JTC1/SC22/WG21/docs/papers/2002/n1406.pdf), section 2.2



> 
> **2.2 The Main Choice: Specialization vs. Everything Else**
> 
> 
> After discussion on the reflectors and in the Evolution WG, it turns out that we have to choose between two mutually exclusive models:
> 
> 
> 1. A typedef template is not itself an alias; only the (possibly-specialized) instantiations of the typedef template are aliases. This choice allows us to have specialization of typedef templates.
> 2. A typedef template is itself an alias; it cannot be specialized. This choice would allow:
> 
> 
> 	- deduction on typedef template function parameters (see 2.4)
> 	- a declaration expressed using typedef templates be the same as the declaration without
> 	 typedef templates (see 2.5)
> 	- typedef templates to match template template parameters (see 2.6)
> 
> 
> 


It should be noted that the quoted paper, which favors option 1, did not make it into C++0x.




---


EDIT: Because you desperately want to have a spec quote saying so explicitly. 14.5p3 is one that does



> 
> Because an alias-declaration cannot declare a template-id, it is not possible to partially or explicitly specialize an alias template.
> 
> 
> 


"
"checking to see if same value in two lists
I have 2 lists one :


`[(12,23),(12,45),(12,23),(2,5),(1,2),(2,4),(7,34)]` which goes up to around 1000 elements
and another:


`[(12,23),(12,45),(12,23),(2,5),(1,2),(2,66),(34,7)]` which goes up to around 241 elements.


What I want is to check to see if the lists contain any of the same elements and then put them in a new list.


so the new list becomes 


`[(12,23),(12,45),(12,23),(2,5),(1,2)]`


","This doesn't include duplicates



```
>>> A = [(12,23),(12,45),(12,23),(2,5),(1,2),(2,4),(7,34)]
>>> B = [(12,23),(12,45),(12,23),(2,5),(1,2),(2,66),(34,7)]
>>> set(B).intersection(A) # note: making the smaller list to a set is faster
set([(12, 45), (1, 2), (12, 23), (2, 5)])

```

Or



```
>>> A = [(12,23),(12,45),(12,23),(2,5),(1,2),(2,4),(7,34)]
>>> B = [(12,23),(12,45),(12,23),(2,5),(1,2),(2,66),(34,7)]
>>> filter(set(B).__contains__, A)
[(12, 23), (12, 45), (12, 23), (2, 5), (1, 2)]

```

This returns every item in `B` if it occured in `A`, which produces the result you give in the example, however the `set` is *probably* what you want.


Since I don't know exactly what you are using this for, I'll suggest one more solution which returns a list, containing the items that occur in both lists, the **minimum** amount of times they occurred in either list (unordered). This differs from the set solution above which only returns each item the number of times it occurred in the **other** and doesn't care how many times it occurred in the first. This uses [`Counter`](http://docs.python.org/2/library/collections.html#counter-objects) for the intersection of multisets.



```
>>> from collections import Counter
>>> A = [(12,23),(12,45),(12,23),(2,5),(1,2),(2,4),(7,34)]
>>> B = [(12,23),(12,45),(12,23),(2,5),(1,2),(2,66),(34,7)]
>>> list((Counter(A) & Counter(B)).elements())
[(1, 2), (12, 45), (12, 23), (12, 23), (2, 5)]

```

"
"Sending binary data over Bluetooth/RFCOMM/SPP converts 0x0A to 0x0D 0x0A
I have a Bluetooth RFCOMM service running on Linux via the commands:



```
sdptool add --channel 1 SP
rfcomm watch hci0 1 ""$COMMAND"" {}
# ^ here

```

The `$COMMAND` will write binary data to the file passed as a parameter. I have tested that it behaves correctly by doing:



```
FIFO=$(tempfile)
mkfifo ""$FIFO""
""$COMMAND"" ""$FIFO"" &
cat ""$FIFO"" | hexdump -C # <- output is correct

```

However, when connecting to the service from a different device via SPP/RFCOMM discovery (UUID `00001101-0000-1000-8000-00805F9B34FB`), I see that every instance of `0x0A` (`LF`) is replaced by `0x0D` `0x0A` (`CR` `LF`) in the stream. The issue isn't on the receiving end, because I have tried connecting to a hardware serial device that also sends binary data, and there the conversion doesn't happen. It must be the command in the first snippet (above the `# ^ here` line) that performs the replacement.


Why does the `rfcomm` tool do this replacement and how can I disable it?


","It looks like you are being bitten by the line discipline of the TTY (remember that `rfcomm` does not create a fifo, but a tty).


You may try changing the TTY to *raw mode*, that is, without any magic. The easiest way is with `stty --file <tty> raw`. I don't know if `rfcomm` will admit several commands in its command line but you can do easily with a script: 


### command\_raw



```
#!/bin/bash
stty --file ""$1"" raw
""$COMMAND"" ""$1""

```

And then run:



```
sdptool add --channel 1 SP
rfcomm watch hci0 1 ./command_raw {}

```

If you have the source of the command to be run, you can also change it easily in C:



```
 #include <termios.h>
 #include <unistd.h>

 //WARNING: error checking left as an exercise to the reader!
 void make_raw(int fd)
 {
     struct termios ios;

     //Not a TTY: nothing to do
     if (!isatty(fd))
         return; 

     tcgetattr(fd, &ios);
     cfmakeraw(&ios);
     tcsetattr(fd, TCSANOW, &ios);
 }

```

"
"How to check if a secret variable is empty in if conditional Github Actions
## Context


I want to check in my workflow if a secret is present or not before executing a job.


Something like this:



```
publish:
    runs-on: ubuntu-latest
    if: secrets.AWS_ACCESS_KEY_ID != ''
    steps:
      [ ... ]

```

However, I've got an error like this when using this expression:



```
The workflow is not valid. .github/workflows/release.yml (Line: 11, Col: 9): Unrecognized named-value: 'secrets'...

```

## What I tried


I tried to wrote the expression another way:



```
if: ${{ secrets.AWS_ACCESS_KEY_ID != '' }}

```


```
if: ${{ secrets.AWS_ACCESS_KEY_ID }} != ''

```

## Question


**How to achieve what I want in a Github Actions workflow?**


","The Github Action interpreter currently doesn't identify the `secrets` key word when used in an `if` conditional expression. Therefore, you can't use the `secrets.VARIABLE` syntax there.


Instead, use the environment to carry a result of a secret check and then use an `if` conditional upon the non-secret result.


**job.step Example:**



```
job:
  runs-on: ubuntu-latest
  steps:
    - name: Check for Secret availability
      id: secret-check
      # perform secret check & put boolean result as an output
      shell: bash
      run: |
        if [ ""${{ secrets.MY_KEY }}"" != '' ]; then
          echo ""available=true"" >> $GITHUB_OUTPUT;
        else
          echo ""available=false"" >> $GITHUB_OUTPUT;
        fi

    - name: Check Inadequate Permissions
      if: ${{ steps.secret-check.outputs.available != 'true' }}
      # provide feedback for likely problem, note dependabot cannot access
      # secrets by default. Secondly, this step forces job failure due to
      # missing secret via `exit 1`
      shell: bash
      run: |
        if [ ""${{ github.actor }}"" == ""dependabot[bot]"" ]; then
          echo >&2 ""Unable to access secrets as unprivileged dependabot."";
        else
          echo >&2 ""Inadequate Permissions or missing secret value"";
        fi
        exit 1

    - name: Execute Step requiring secret
      # If you didn't abort step above, then use this conditional
      # if: ${{ steps.secret-check.outputs.available == 'true' }}
      shell: bash
      # Key will be blocked out in log output but will be not empty
      run: |
        echo ""This command is executed with non-empty key: \
          ${{ secrets.MY_KEY }}""

```

If you need to do this at the job level, create a separate `check-secret` job to validate the secrets and then share the result as a defined output.


**Workflow Context Level Example:**



```
jobs:

  check-secret:
    runs-on: ubuntu-latest
    outputs:
      my-key-exists: ${{ steps.my-key-check.outputs.defined }}
    steps:
      - name: Check for Secret availability
        id: my-key-check
        # perform secret check & put boolean result as an output
        shell: bash
        run: |
          if [ ""${{ secrets.AWS_ACCESS_KEY_ID }}"" != '' ]; then
            echo ""defined=true"" >> $GITHUB_OUTPUT;
          else
            echo ""defined=false"" >> $GITHUB_OUTPUT;
          fi

  job1:
    runs-on: ubuntu-latest
    needs: [check-secret]
    if: needs.check-secret.outputs.my-key-exists == 'true'
    steps:
      - run: echo ""This command is executed if AWS_ACCESS_KEY_ID secret IS NOT empty""

  job2:
    runs-on: ubuntu-latest
    needs: [check-secret]
    if: needs.check-secret.outputs.my-key-exists != 'true'
    steps:
      - run: echo ""This command is executed if AWS_ACCESS_KEY_ID secret IS empty""

```

"
"calling ExecutorService.shutDown() in java
i am starting to learn the ExecutorService class. The documentation (and tutorials online) say to always call ExecutorService.shutDown() to reclaim resources. however, the documentation also says that after you call shutDown(), no new tasks will be accepted. so, my question is, do i have always have to instantiate a new ExecutorService whenever i need to parallelize data processing? 


right now i have a List of Callable objects, and i do the following.



```
public void someMethod() {
 List<OuterCallable> outerCallables = getOuterCallables();
 ExecutorService executor = Executor.newFixedThreadPool(NUM_CPUS);
 executor.invokeAll(tasks);
 executor.shutDown();
}

```

however, my OuterCallable also splits data or executes data processing in parallel using InnerCallable.



```
public class OuterCallable implements Callable<Long> {
 public Long call() throws Exception {
  long result = 0L;

  List<InnerCallable> innerCallables = getInnerCallables();
  ExecutorServices executor = Executor.newFixedThreadPool(NUM_CPUS);
  executor.invokeAll(tasks);
  executor.shutDown();

  return result;
 }
}    

```

i can't remember if it was for ExecutorService or the Fork/Join approach, but i remember the documentations and tutorials saying that the actual parallel procedure to manipulate data should not involve I/O operations and everything should be done in memory. however, in my InnerCallable, i am actually making JDBC calls (not shown here).


ultimately, the way i am using ExecutorService works, but i still have lingering concerns.


1. is my approach above good programming practice using ExecutorService?
2. should i be using a singleton instance of ExecutorService?
3. should i be not only avoiding I/O operations inside my parallel methods but also JDBC calls as well?


as a last concern, i was trying to research a little bit on Fork/Join vs ExecutorService. i came across an article that completely blasted the Fork/Join API/classes. is it worth it to learn Fork/Join? i saw a few articles on stackoverflow and elsewhere, where tests are used to compare Fork/Join vs ExecutorService, and there are graphs showing better CPU usage of Fork/Join vs ExecutorService (via Windows Task Manager). however, when i use ExecutorService (JDK 1.7.x), my CPU usage is max. has ExecutorService improved with the latest JDK?


any help/guidance is appreciated.


","You should add `awaitTermination` calls, because `shutDown` returns without waiting for Callables to finish. Other than that,


1. Do you `OuterCallable`s have sequential dependencies? If so, your approach is fine, but using ForkJoinPool would be preferable because it will keep the number of worker threads low. If not, it would be better to submit a big flattened collection of `Callable`s to a single `Executor`.
2. Only if you want to use the ExecutorService in several different routines and want to avoid passing it around. If it's only used in `someMethod`, might as well instantiate it there as you are doing.
3. You should avoid I/O routines that will take a long time to complete. If your ExecutorService has 4 worker threads and all of them are blocking on I/O, the JVM won't be using the CPU at all, even though other Callables may be waiting to do CPU-intensive work. A few JDBC calls is okay as long as the queries do not take long to complete.


"
"ComboBox Cue Banner not italic when DropDownStyle is DropDown
We have a WinForms control that is an extended version of `ComboBox` that supports ""cue banners"" (aka watermarks) when there is no selection or text. Our control is similar to this [implementation making use of CB\_SETCUEBANNER](http://www.aaronlerch.com/blog/2007/12/01/watermarked-edit-controls/).


However, when we set `DropDownStyle` for the control to `ComboBoxStyle.DropDown` (that is, also allows free text entry) the cue banner is showing, just not in italics (which is how it usually shows).


Does anyone know how to draw the cue banner in italics for a combo box in `ComboBoxStyle.DropDown` mode???


","By design. When the Style = DropDown, the text portion of the combobox is a TextBox. Which displays the cue banner in non-italic style. You can verify with [this code](https://stackoverflow.com/questions/4902565/watermark-textbox-in-winforms/4902969#). It is otherwise important to make the distinction between the banner and the actual selection visible when the Style = DropDownList, no doubt the reason they chose to display it italic. TextBox does it differently, it hides the banner when it gets the focus.


Throwing in a non exhausting version:



```
using System;
using System.Windows.Forms;
using System.Runtime.InteropServices;

class CueComboBox : ComboBox {
    private string mCue;
    public string Cue {
        get { return mCue; }
        set {
            mCue = value;
            updateCue();
        }
    }
    private void updateCue() {
        if (this.IsHandleCreated && mCue != null) {
            SendMessage(this.Handle, 0x1703, (IntPtr)0, mCue);
        }
    }
    protected override void OnHandleCreated(EventArgs e) {
        base.OnHandleCreated(e);
        updateCue();
    }
    // P/Invoke
    [DllImport(""user32.dll"", CharSet = CharSet.Unicode)]
    private static extern IntPtr SendMessage(IntPtr hWnd, int msg, IntPtr wp, string lp);
}

```

"
"Verify whether a method parameter is used in the method body
I have an interface which looks like the following



```
interface Evaluator {
    boolean requiresP2();
    EvalResult evaluate(Param1 p1, Param2 p2, Param3 p3);
    // some more methods
}

```

This interface is implemented by several classes. The parameter `p2` of the evaluate method is used by some and not used by others. The method `requiresP2` basically returns a boolean telling whether the evaluate method uses `p2` or not.


Now, this questions may appear a little weird out of context but believe me, it makes sense in our use case. Plus, it would require a lot of time to refactor all the code to eliminate the need for the `requiresP2` method so I would appreciate if we discuss solutions other than a top-to-bottom refactoring of the codebase.


The problem is that the return value of method `requiresP2` is based on how the `evaluate` method is implemented. Therefore everyone must ensure that they update the `requiresP2` method when they change the `evaluate` method.


I am looking for ways so that this can be enforced by the compiler/unit-tests/linters rather than leaving it to the developer's memory.


**EDIT**: I am still exploring the applicability of mocking frameworks to this problem.


I thought that I could reflection in unit tests to inspect `evaluate`'s body in the unit test to check if it refers to `p2` or not and then making sure it matches with the value returned by `requiresP2` method but it seems that it is not possible to inspect method body using reflection.


I am looking for suggestions on how to do this. Any input is appreciated.


","You can use [ASM](http://asm.ow2.io/) to check whether the parameter is used.


To add it to your project using e.g. Apache Ivy, you would add this to `ivy.xml`:



```
<dependency org=""org.ow2.asm"" name=""asm"" rev=""6.1.1"" />

```

Or do the equivalent for Maven, Gradle, etc. Then you can check on the parameter by:



```
import java.io.IOException;
import java.io.InputStream;
import java.io.UncheckedIOException;
import java.util.concurrent.atomic.AtomicBoolean;

import org.objectweb.asm.ClassReader;
import org.objectweb.asm.ClassVisitor;
import org.objectweb.asm.MethodVisitor;
import org.objectweb.asm.Opcodes;

// . . .

public static boolean usesP2(Evaluator evaluator) {
    AtomicBoolean usesP2 = new AtomicBoolean(false);
    String internalName = evaluator.getClass().getName().replace('.', '/');
    String classFileResource = ""/"" + internalName + "".class"";

    ClassVisitor visitor = new ClassVisitor(Opcodes.ASM6) {
        @Override
        public MethodVisitor visitMethod(int access, String name,
                String desc, String signature, String[] exceptions) {
            if (""evaluate"".equals(name)) {
                return new MethodVisitor(Opcodes.ASM6) {
                    @Override
                    public void visitVarInsn(final int insn, final int slot) {
                        if (slot == 2) usesP2.set(true);
                    }
                };
            }
            return super.visitMethod(access, name, desc, signature, exceptions);
        }
    };
    try (InputStream is = Evaluator.class.getResourceAsStream(classFileResource)) {
        ClassReader reader = new ClassReader(is);
        reader.accept(visitor, 0);
    } catch (IOException e) {
        throw new UncheckedIOException(e);
    }
    return usesP2.get();
}

public static void assertCorrectlyDocumentsP2(Evaluator evaluator) {
    boolean usesP2 = usesP2(evaluator);
    if (usesP2 && !evaluator.requiresP2()) {
        throw new AssertionError(evaluator.getClass().getName() +
                "" uses P2 without documenting it"");
    }
    if (!usesP2 && evaluator.requiresP2()) {
        throw new AssertionError(evaluator.getClass().getName() +
                "" says it uses P2 but does not"");
    }
}

```

Unit tests:



```
@Test
public void testFalsePositive() {
    assertCorrectlyDocumentsP2(new FalsePositive());
}

@Test
public static void testFalseNegative() {
    assertCorrectlyDocumentsP2(new FalseNegative());
}

```

(This supposes there are two bad `Evaluator`s, `FalsePositive` and `FalseNegative`, one of which documents that it uses P2 but doesn't, and the other which doesn't document that it uses P2 even though it does, respectively.)


Note: In `usesP2` we check for a variable instruction (an instruction which accesses a local variable) in slot 2 of the stack frame. The slots are numbered from 0, and the first one is `this`. P2 is in slots 2 only because `Evaluator::evaluate` is an *instance method*. If it were a **static method**, we would have to check if **slot 1** were used in order to detect if parameter P2 were used. *Caveat lector*.


"
"Moving files between different WSL2 instances?
If it is possible, how can I move files from one WSL instance to another directly? This would be useful when trying out new distributions, e.g. for copying /home from Ubuntu 18.04 to 20.04.


I am weary of doing this through explorer by accessing [`\\wsl$` in the windows host](https://stackoverflow.com/questions/64185560/where-are-the-files-inside-wsl2-physically-stored), it doesn't seem to reliably transfer all the files and feels wrong overall.


","# How to share data between different WSL instances


## 1. Create shared directory which is not included in any WSL instances



```
mkdir /mnt/wsl/share

```

This will create `/mnt/wsl/share` directory shared across all WSL instances. You can share data via this shared directory.


**WARNING: the shared directory's filesystem lives in memory and thus a `wsl.exe --shutdown` would wipe them from existence.**


However, it is still impossible to directly copy/paste data between WSL instances.


## 2. Create shared directory included in a WSL instance


You can make existing directory in WSL instance shared across other WSL instances by using bind mount.


For example, you have two WSL instances, *Ubuntu-A* and *Ubuntu-B*.


**1. Open `~/.profile` in *Ubuntu-A* and add below code.**



```
# bind mount shared directory
if [ ! -d /mnt/wsl/share-a ]; then
  mkdir /mnt/wsl/share-a
  wsl.exe -d Ubuntu-A -u root mount --bind / /mnt/wsl/share-a/
fi

```

**2. Open `~/.profile` in *Ubuntu-B* and add below code.**



```
# bind mount shared directory
if [ ! -d /mnt/wsl/share-b ]; then
  mkdir /mnt/wsl/share-b
  wsl.exe -d Ubuntu-B -u root mount --bind / /mnt/wsl/share-b/
fi

```

This will automatically mount the root directory (`/`) of *Ubuntu-A* to `/mnt/wsl/shared-a/` and do the same thing for *Ubuntu-B* when you launch WSL.


"
"Building a connection URL for mssql+pyodbc with sqlalchemy.engine.url.URL
# The problem...


I am trying to connect to a MSSql server via SQLAlchemy. Here is my code with fake credentials (*not my real credentials obviously*).


# The code...



```
credentials = {
'username'  : 'SPOTTER_xyz_ACC',
'password'  : '123Goodbye2016!@#',
'host'      : 'MARYLQLT01',
'database'  : 'LRS_DUS',
'port'      : '1560'}

connect_url = sqlalchemy.engine.url.URL(
    'mssql+pyodbc',
    username=credentials['username'],
    password=credentials['password'],
    host=credentials['host'],
    port=credentials['port'],
    query=dict(service_name=credentials['database']))

engine = create_engine(connect_url)
connection=engine.connect()

```

And this is the .pyodbc error that I am getting.



```
(pyodbc.Error) ('IM002', '[IM002] [Microsoft][ODBC Driver Manager] Data source
name not found and no default driver specified (0) (SQLDriverConnect)')

```

# Additional Details


But, here is what is weird... I if make a pyodbc connection and use Pandas.read\_sql, then I can get data without an error. So I do not understand why I am getting a 'pyodbc' error????



```
connection=pyodbc.connect(\
                          'Driver={SQL Server}; \
                          Server=MARYLQLT01; \
                          Database=LRS_DUS; \
                          UID=SPOTTER_xyz_ACC; \
                          PWD=123Goodbye2016!@#')

stmt='select * from PD_SC_All'
df=pd.read_sql(stmt,connection)

```

And then I can see the dataframe.


So, why am I getting the `(pyodbc.Error)` when I try to connect with SQLAlchemy?


## System Information



```
Windows 7
SQLAlchemy: 1.1.11
pyodbc: 4.0.16

Python 3.6.1 |Anaconda custom (64-bit)| (default, May 11 2017, 13:25:24) [MSC v.1900 64 bit (AMD64)]
Type ""copyright"", ""credits"" or ""license"" for more information.

IPython 5.3.0 -- An enhanced Interactive Python.
?         -> Introduction and overview of IPython's features.
%quickref -> Quick reference.
help      -> Python's own help system.
object?   -> Details about 'object', use 'object??' for extra details.

```

","As stated in the [SQLAlchemy documentation](http://docs.sqlalchemy.org/en/latest/dialects/mssql.html#module-sqlalchemy.dialects.mssql.pyodbc), for `mssql+pyodbc` connections



> 
> When using a hostname connection, the driver name must also be specified in the query parameters of the URL.
> 
> 
> 
> ```
> engine = create_engine(""mssql+pyodbc://scott:tiger@myhost:port/databasename?driver=ODBC+Driver+17+for+SQL+Server"")
> 
> ```
> 
> 


For



```
credentials = {
    'username': 'scott',
    'password': 'tiger',
    'host': 'myhost',
    'database': 'databasename',
    'port': '1560'}

```

your statement produces the `connect_url`



```
mssql+pyodbc://scott:tiger@myhost:1560?service_name=databasename

```

What you really need is



```
from sqlalchemy.engine import URL

connect_url = URL.create(
    'mssql+pyodbc',
    username=credentials['username'],
    password=credentials['password'],
    host=credentials['host'],
    port=credentials['port'],
    database=credentials['database'],
    query=dict(driver='ODBC Driver 17 for SQL Server'))

```

which produces



```
mssql+pyodbc://scott:tiger@myhost:1560/databasename?driver=ODBC+Driver+17+for+SQL+Server

```

"
"Maven POM dependencies to ivy.xml file
Is there a maven plugin which i can use to convert the maven pom dependencies including transitive dependencies to an ivy.xml file? 


","Here's an Ant script



```
<project name=""convertPomToIvy"" basedir=""."" default=""all""
  xmlns:ivy=""antlib:fr.jayasoft.ivy.ant""
  xmlns:ac=""antlib:net.sf.antcontrib"">

    <path id=""antlib.classpath"">
        <fileset dir=""/path/to/ivy/libs"" includes=""*.jar""/>
    </path>

    <taskdef uri=""antlib:fr.jayasoft.ivy.ant""
        resource=""fr/jayasoft/ivy/ant/antlib.xml""
        classpathref=""antlib.classpath""
        loaderref=""antlib.classpath.loader""/>

    <target name=""convert"">
        <ivy:convertpom pomFile=""pom.xml"" ivyFile=""ivy.xml"" />
    </target>

</project>

```

From [here](http://vafer.org/blog/20081030002755) or [here](http://stefan222devel.blogspot.com/2010/01/how-to-convert-maven-pomxml-into-ivy.html) (& probably elsewhere)


"
"Nested if's - what's more Pythonic?
Both functions do the same thing. 



```
def function1(self):
    a = self.get_a()
    b = self.get_b()
    c = self.get_c()
    r = None

    if a:
        r = a
        if b:
            r = b
            if c:
                r = c
            else:
                print(""c not set."")
        else:
            print(""b not set."")
    else:
        print(""a not set."")

    return r



def function2(self):
    a = self.get_a()
    b = self.get_b()
    c = self.get_c()
    r = None

    if not a:
        print(""a not set."")
        return r

    r = a
    if not b:
        print(""b not set."")
        return r

    r = b
    if not c:
        print(""c not set."")

    r = c
    return r

```

function1() creates very long lines the more if's are nested which conflicts with PEP8's line-length limit of 78. 


function2() might be harder to read/understand and has more return statements. Line length is no problem here.


Which one is more pythonic?


","One of the principals of Pythonic code is ""flat is better than nested"". On this basis, I'll say `function2()` is objectively more Pythonic. This can be seen in [PEP-20: The Zen of Python](https://www.python.org/dev/peps/pep-0020/):



> 
> **The Zen of Python**
> 
> 
> 
> 
> ```
> Beautiful is better than ugly.
> Explicit is better than implicit.
> Simple is better than complex.
> Complex is better than complicated.
> Flat is better than nested.
> Sparse is better than dense.
> Readability counts.
> Special cases aren't special enough to break the rules.
> Although practicality beats purity.
> Errors should never pass silently.
> Unless explicitly silenced.
> In the face of ambiguity, refuse the temptation to guess.
> There should be one-- and preferably only one --obvious way to do it.
> Although that way may not be obvious at first unless you're Dutch.
> Now is better than never.
> Although never is often better than *right* now.
> If the implementation is hard to explain, it's a bad idea.
> If the implementation is easy to explain, it may be a good idea.
> Namespaces are one honking great idea -- let's do more of those!
> 
> ```
> 
> 


This can be seen by typing `import this` inside the Python interpreter. 


"
"Go func closure in loop
When executing the following code I get what I expect when the first loop is done (sequence from 0 to 9). But when the second loop finishes, the result is not what I expected (I expected the same result as in the first loop, but it prints only '10's):



```
package main

import (
    ""fmt""
    ""sync""
)

func main() {
    var wg sync.WaitGroup

    for i := 0; i < 10; i++ {
        wg.Add(1)

        go func(j int) {
            defer wg.Done()

            fmt.Println(j)
        }(i)
    }

    wg.Wait()

    fmt.Println(""done first"")

    for i := 0; i < 10; i++ {
        wg.Add(1)

        go func() {
            defer wg.Done()

            fmt.Println(i)
        }()
    }

    wg.Wait()

    fmt.Println(""done second"")
}

```

Output:



```
0
1
2
3
4
5
6
7
8
9
done first
10
10
10
10
10
10
10
10
10
10
done second

```

Why doesn't the second loop print a sequence?


","Because the first one gets a copy of the loop counter each time. Whereas the second gets the variable captured as part of a closure.


In the first, you're passing it in here in *every iteration of the loop*:



```
go func(j int) {
    defer wg.Done()

    fmt.Println(j)
}(i) // <------------ its passed in here as part of each loop iteration

```

The second one receives nothing.. so, the loop counter `i` is captured as part of a closure. By the time the first go routine executes, the `for` loop has finished. The loop finishing has set the `i` variable (that is now part of a closure) to 10. Go routine #1 executes and prints the value of `i`.. which is now already 10 and the rest follow suit.


TLDR: The problem here is that the loop is finishing before any go routines are scheduled to be run - its just that quick. Therefore, `i == 10` when the go routines run.


"
"Is there a way to control spatial/memory locality when using Parallel.For
I have a large array of unscaled floats - array length is 40,000,000. In order to scale this array, I thought it would be more performant to use `Parallel.For()`. Here's the sequential version of the for loop that scales the data:



```
for (i = 0; i < rawData.Length; i++)
{
    scaledData[i] = rawData[i] * scale + offset;
}

```

Here is that converted to use `Parallel.For()` e.g.:



```
Parallel.For(0, rawData.Length, i => {
    scaledData[i] = rawData[i] * scale + offset;
});

```

But the performance is worse! My guess, based on observing index/thread combinations, is that `Parallel.For()` is accessing memory in such a way as to cause excess paging. To test this theory, I tried using `Parallel.Invoke()` like so:



```
Parallel.Invoke(
    () => { for (int i =        0; i < 10000000; i++) { dst[i] = src[i] * scale + offset; } },
    () => { for (int i = 10000000; i < 20000000; i++) { dst[i] = src[i] * scale + offset; } },
    () => { for (int i = 20000000; i < 30000000; i++) { dst[i] = src[i] * scale + offset; } },
    () => { for (int i = 30000000; i < 40000000; i++) { dst[i] = src[i] * scale + offset; } },
);

```

And that performs noticeably better but I hate the hard-coded nature of this code. I have 4 processors which is why there are 4 actions passed to Invoke(). 


Is there way to cajole `Parallel.For()` into distributing indices to threads in a fashion that doesn't thrash memory?


","You can use a custom partitioner to get the behavior you want without having to resort to using `Parallel.Invoke`. The `RangPartitioner` is what you want to start with.



```
    var rangePartitioner = Partitioner.Create(0, rawData.Length);

    double[] results = new double[rawData.Length];

    Parallel.ForEach(rangePartitioner, (range, loopState) =>
    {
        for (int i = range.Item1; i < range.Item2; i++)
        {
            scaledData[i] = rawData[i] * scale * offset;
        }
    });

```

You can create a custom partitioner and overload the [`GetPartition()` method](http://msdn.microsoft.com/en-us/library/dd381767%28v=vs.110%29.aspx) to adjust the block size to something that suites your needs.


See [Custom Partitioners for PLINQ and TPL](http://msdn.microsoft.com/en-us/library/dd997411%28v=vs.110%29.aspx) for an in depth discussion.


Will this improve data locality, yes. Provided that your array contains value types. In which case they will be allocated as blocks of contiguous memory. For reference types this is not the case. FWIW I've tried to improve memory locality like this with OK but not astounding improvements. I came to the conclusion that there is potentially a lot of other memory access from the CLR and that this can make it hard to understand the resulting memory access patterns. 


"
"How can I get the name / details of an Action delegate invocation?
I'm testing the performance of some similar method calls that I'm wrapping in some timing and logging statements. I'm passing these methods in via an [Action delegate](http://msdn.microsoft.com/en-us/library/system.action.aspx) parameter.


Is there any way to print details about the call?


For example:



```
var httpResult = TestService(() => serviceHttp.Search(criteria));
var tcpResult = TestService(() => serviceTcp.Search(criteria));
var localResult = TestService(() => servicelocal.Search(criteria));

...

private static double TestService(Action serviceOperation)
{
    const int iterations = 15;
    ...

    for (var i = 0; i < iterations; i++)
    {
        var watch = Stopwatch.StartNew();

        ...

        Console.WriteLine(string.Format(""{0} ElapsedMilliseconds={1}"", ????, watch.ElapsedMilliseconds));
         // Ideally this would print something like ""serviceTcp.DoStuff(...) ElapsedMilliseconds=313""
    }

    ...
}

```

","Change your testing method declaration to



```
private static double TestService(Expression<Action> expression)

```

Call `Compile` method of expression object to get a method for testing:



```
var serviceOperation = expression.Compile();

```

`Expression` object can provide a lot of information about method call, you can start with something like this:



```
private static string GetMethodCallDescription(Expression<Action> expression)
{
    var mce = (MethodCallExpression)expression.Body;
    var method = mce.Method;
    var sb = new StringBuilder();
    sb.Append(method.DeclaringType.Name);
    sb.Append(""."");
    sb.Append(method.Name);
    sb.Append(""("");
    bool firstarg = true;
    foreach(var arg in mce.Arguments)
    {
        if(!firstarg)
        {
            sb.Append("", "");
        }
        else
        {
            firstarg = false;
        }
        sb.Append(arg.ToString());
    }
    sb.Append("")"");
    return sb.ToString();
}

```

"
"import in python 3, explain the output please
Contents :



```
tutorials/maindir/
├── dir1
│   ├── file11.py
│   ├── file12.py
│   ├── __init__.py
│   └── __pycache__
│       ├── file11.cpython-36.pyc
│       └── __init__.cpython-36.pyc
├── dir2
│   ├── file21.py
│   ├── file22.py
│   └── __init__.py
├── file1.py
├── file2.py
├── __init__.py
└── __pycache__
    ├── file1.cpython-36.pyc
    └── __init__.cpython-36.pyc

```

now I have a code giving me following output which I don't get.


Code ::



```
print(""--main--"", dir())
from tutorials.maindir import *
print(""--main--"", dir())

```

OutPut ::



```
--main-- ['__annotations__', '__builtins__', '__cached__', '__doc__', '__file__', '__loader__', '__name__', '__package__', '__spec__']
tutorials
tutorials.maindir
tutorials.maindir.dir1
--main-- ['__annotations__', '__builtins__', '__cached__', '__doc__', '__file__', '__loader__', '__name__', '__package__', '__spec__', 'dir1', 'tutorials']

```

I don't get why `dir1` and `tutorials` are there in the `dir()`'s output.


`maindir.__init__.py` 's Code : 



```
print(__name__)
import tutorials.maindir.dir1.file11

```

EDIT 1 :: 


So if the code is :



```
print(""--main--"", dir())
import tutorials.maindir.dir1.file11
print(""--main--"", dir())

```

The output is :



```
--main-- ['__annotations__', '__builtins__', '__cached__', '__doc__', '__file__', '__loader__', '__name__', '__package__', '__spec__']
tutorials
tutorials.maindir
tutorials.maindir.dir1
--main-- ['__annotations__', '__builtins__', '__cached__', '__doc__', '__file__', '__loader__', '__name__', '__package__', '__spec__', 'tutorials']

```

","When you run an `import` you are causing `__init__.py` to run. [[source](https://docs.python.org/3/tutorial/modules.html#importing-from-a-package)]



> 
> If \_\_all\_\_ is not defined, the statement from sound.effects import \* does not import all submodules from the package sound.effects into the current namespace; it only ensures that the package sound.effects has been imported (possibly running any initialization code in \_\_init\_\_.py) and then **imports whatever names are defined in the package. This includes any names defined (and submodules explicitly loaded) by \_\_init\_\_.py**. It also includes any submodules of the package that were explicitly loaded by previous import statements.
> 
> 
> 


In your case, this includes another `import`:



```
import tutorials.maindir.dir1.file11

```

That causes certain things to be bound to the local namespace: [[source](https://docs.python.org/3/reference/simple_stmts.html#the-import-statement)]



> 
> define a name or names in the local namespace for the scope where the import statement occurs.
> 
> 
> 


And more specifically:



> 
> If the module being imported is not a top level module, then **the name of the top level package** that contains the module is bound in the local namespace as a reference to the top level package. The imported module must be accessed using its full qualified name rather than directly
> 
> 
> 


So in your case, `tutorials` is the top level module which was already imported, and `dir1` is the imported package (chosen by being the parent name).


And then finally `dir()` is going to: [[source](https://docs.python.org/3.7/library/functions.html#dir)]



> 
> Without arguments, return the list of names in the current local scope. With an argument, attempt to return a list of valid attributes for that object.
> 
> 
> 


So to your comment, the top level package and the individual names defined in the import statements are bound to the local namespace which is why they show up, but maindir is not a top level package or an imported package so it does not show up.




---


**EDIT:** (*and additional information*)


First, I'd recommend doing some technical reading on a few things if all of this is super interesting to you or relevant:


- [PEP 420](https://www.python.org/dev/peps/pep-0420/) -- Implicit Namespace Packages
- [PEP 302](https://www.python.org/dev/peps/pep-0302/) -- New Import Hooks
- [PEP 328](https://www.python.org/dev/peps/pep-0328/) -- Imports: Multi-Line and Absolute/Relative


Those PEPs discuss a lot of how modern import works and some of the special and weird parts of it, as well as how to form it to do what you need it to do in your use case.




---


On to your edit question: 



> 
> When the *name variable* is of the form package.module, normally, the **top-level package** (the name up till the first dot) is returned, not the module named by name. However, when a *non-empty fromlist* argument is given, the **module named by name is returned**. [[source](https://docs.python.org/3/library/functions.html#__import__)]
> 
> 
> 


So for your basic long import statement, it will put the top level module into the namespace.



> 
> If the list of identifiers is replaced by a star ('\*'), all public names defined in the module are bound in the local namespace for the scope where the import statement occurs. [[source](https://docs.python.org/3/reference/simple_stmts.html#the-import-statement)]
> 
> 
> 


`import name` gives you *name*  

`import name.one` gives you *name*  

`import name.one.two` gives you *name*  

`from name import one` gives you *one*  

`from name import one as two` gives you *two* 


Now back to your example:  

`from tutorials.maindir import *` will first import the top level module. Then `__init__.py` will run which does another import. That import will skip the top level module since that is already imported, and will move down to the package and then import as the name of the parent, or essentially the right most `.` it will move left one term and import as that name. [[code](https://github.com/python/cpython/blob/50ecedc10309373809cc45595a1027076e808607/Lib/importlib/_bootstrap.py#L948x)].


"
"How to boot load the kernel using EFI stub (efistub) loader?
I have Ubuntu 14.04 running in UEFI mode as only operating system, no dual-boot here. The kernel version is 3.13.0-24-generic. There is an EFI partition. In this case the EFI partition is not at the default `/dev/sda1` but at `/dev/sda3` because I did actually [convert BIOS mode to EFI mode](https://askubuntu.com/questions/509423/which-commands-to-convert-a-ubuntu-bios-install-to-efi-uefi-without-boot-repair). I have used the `grub-efi-amd64` package, though that actually loads GRUB boot menu from UEFI firmware boot menu (UEFI boot loads `\EFI\ubuntu\grubx64.efi`). 


I want to skip that double boot menu loading step, and boot faster, directly from UEFI into the kernel. The [Ubuntu kernels since 12.10 have ""Kernel EFI stub loader""](https://askubuntu.com/questions/510009/does-the-ubuntu-kernel-have-efi-stub-loader-boot-support) feature.


I know I do need to copy the Ubuntu kernel to the EFI partition (possibly rename) and create an entry in UEFI boot menu (for instance using `efibootmgr`). Which exact terminal commands are necessary to do this?


","The commands below are more generic then for kernel version 3.13.0-35 only. 


### 1. Mount the efi partition and copy the kernel files there



> 
> `$ mount /dev/sda3 /boot/efi`
> 
> 
> `$ mkdir -pv /boot/efi/EFI/ubuntu/`
> 
> 
> 
> ```
> $ cp -uv /boot/vmlinuz-* /boot/initrd.img-* /boot/efi/EFI/ubuntu/
> '/boot/vmlinuz-3.13.0-35-generic' -> '/boot/efi/EFI/ubuntu/vmlinuz-3.13.0-35-generic'
> '/boot/initrd.img-3.13.0-35-generic' -> '/boot/efi/EFI/ubuntu/initrd.img-3.13.0-35-generic'
> 
> ```
> 
> 


### 2. Change the kernel file name


Shorten the kernel file name by removing `-generic` because [there seems to be a 39 character length path limit](https://www.kubuntuforums.net/showthread.php?60193-Going-GRUB-less-with-UEFI&p=309923&viewfull=1#post309923) and Rename kernel file(s) to end in `.efi`, this ensures compatibility with most systems



> 
> 
> ```
> $ for f in /boot/efi/EFI/ubuntu/vmlinuz-*-generic; do mv -uv -- ""$f"" ""${f//-generic/}.efi""; done
> '/boot/efi/EFI/ubuntu/vmlinuz-3.13.0-35-generic' -> '/boot/efi/EFI/ubuntu/vmlinuz-3.13.0-35-generic.efi'`
> 
> ```
> 
> 


The above name kernel file name shortening is not enough for a `dpkg` installed **mainline kernel**, because for example `/EFI/ubuntu/vmlinuz-3.16.0-031600rc6.efi` without `-generic` is still 40 characters long.


### 3. Add new entry to EFI boot menu


Replace `3.13.0-35` in this example with your specific kernel version



> 
> `$ kv=3.13.0-35;efibootmgr -c -p 3 -L $kv -l \EFI\ubuntu\vmlinuz-$kv.efi -u root=/dev/sda1 initrd=\\EFI\\ubuntu\\initrd.img-$kv-generic ro rootfstype=ext4 debug ignore_loglevel libata.force=dump_id crashkernel=384M-:128M`
> 
> 
> 


This new boot menu entry will become your default new boot choice. 


You might not need the extra debugging parameters `debug`, `ignore_loglevel`, `libata.force=dump_id` and `crashkernel=384M-:128M`. `Initrd` must be present, otherwise boot hangs at ""*Switched to clocksource tsc.*"" because the root device sda1 cannot be opened.


"
"What does each of these network services do
I have the services listed in the `/etc/init.d` folder starting with 'network':



```
cyrex@cyrex:/etc/init.d$ ls network*
networking                  network-interface-security
network-interface           network-manager

```

What does each of these do. For example which one is related to the gui version of network manager, which one will drop network support for all network cards.


","- What does each of these do?


networking: 



```

#!/bin/sh -e
### BEGIN INIT INFO
# Provides:          networking
# Required-Start:
# Required-Stop:     $local_fs
# Should-Start:      ifupdown
# Should-Stop:       ifupdown
# Default-Start:
# Default-Stop:      0 6
# Short-Description: Raise network interfaces.
### END INIT INFO

```

Regarding the package ifupdown mentioned here: includes commands`ifup` and `ifdown` which may be used to (de)configure network interfaces (`/etc/network/interfaces`).


The other 3 are upstart deamons. From their respective `.conf` files in `/etc/init/`:


network-interface-security:



```

# network-interface-security - configure network device security
#
# This is a one-time start-up script to load AppArmor profiles needed
# before the network comes up.

```

network-interface:



```
# network-interface - configure network device
#
# This service causes network devices to be brought up or down as a result
# of hardware being added or removed, including that which isn't ordinarily
# removable.

```

network-manager:



```
# network-manager - network connection manager
#
# The Network Manager daemon manages the system's network connections,
# automatically switching between the best available.

```

- For example which one is related to the gui version of network manager?


None of these. See the next answer.


- Which one will drop network support for all network cards?


Initiating `networking` will do that but the command used will mostlikely be `ifup --all` and `ifdown --all`. From `man ifup`:



```
-a, --all

If given to ifup, affect all interfaces marked auto. 
Interfaces are brought up in the order in which they are defined in
/etc/network/interfaces. 

If given to ifdown, affect all defined interfaces. 
Interfaces are brought down in the order in which they are currently 
listed in the state file. 
Only interfaces defined in /etc/network/interfaces will be brought down.

```

"
"how to change a selections options based on another select option selected?
here is my html.



```
<select id=""type"">
<option value=""item1"">item1</option>
<option value=""item2"">item2</option>
<option value=""item3"">item3</option>
</select>

<select id=""size"">
<option value="""">-- select one -- </option>
</select>

```

here is the jquery i tried but was unsuccessful.



```
$(document).ready(function() {

if( $(""#type"").val(""item1""))
{
  $(""#size"").html(""<option value='test'>test</option><option value=""test2"">test2</option>);
}
elseif( $(""#type"").val(""item2""))
{
   $(""#size"").html(""<option value='anothertest1'>anothertest1</option>"");
}

});

```

basically what i'm trying to do is if an option is selected in #type then the size select is populated with options associated to it. how can i do this?


thanks


","Here is an example of what you are trying to do => [fiddle](http://jsfiddle.net/YPsqQ/)





```
$(document).ready(function () {
    $(""#type"").change(function () {
        var val = $(this).val();
        if (val == ""item1"") {
            $(""#size"").html(""<option value='test'>item1: test 1</option><option value='test2'>item1: test 2</option>"");
        } else if (val == ""item2"") {
            $(""#size"").html(""<option value='test'>item2: test 1</option><option value='test2'>item2: test 2</option>"");
        } else if (val == ""item3"") {
            $(""#size"").html(""<option value='test'>item3: test 1</option><option value='test2'>item3: test 2</option>"");
        } else if (val == ""item0"") {
            $(""#size"").html(""<option value=''>--select one--</option>"");
        }
    });
});
```


```
<script src=""https://ajax.googleapis.com/ajax/libs/jquery/1.11.1/jquery.min.js""></script>
<select id=""type"">
    <option value=""item0"">--Select an Item--</option>
    <option value=""item1"">item1</option>
    <option value=""item2"">item2</option>
    <option value=""item3"">item3</option>
</select>

<select id=""size"">
    <option value="""">-- select one -- </option>
</select>
```






"
"Line break between inline elements
In the below code, I need to place `a` tag in new line and maintain the center alignment. I cannot change the html. `display:block` makes the more clickable area outside the text.


Below is the screenshot for required result


[![enter image description here](https://i.stack.imgur.com/8XGXZ.jpg)](https://i.stack.imgur.com/8XGXZ.jpg) 





```
div{
  background: #333;
  color: #fff;
  margin: 0 auto;
  padding: 20px;
  text-align: center;
  width: 400px;
}
a{
  color: red;
  display: inline-block;
  clear: both;
}
```


```
<div>
  Lorem ipsum dolor sit amet, consectetur adipisicing elit. <a href="""">This is link in new line</a> Cumque mollitia repellat soluta voluptates molestias veniam reiciendis.
</div>
```






<https://jsfiddle.net/afelixj/nq7ow9eq/>


","Add `display:block` to `a`





```
div{
  background: #333;
  color: #fff;
  margin: 0 auto;
  padding: 20px;
  text-align: center;
  width: 400px;
}
a{
  color: red;
  display: block;
}
```


```
<div>
  Lorem ipsum dolor sit amet, consectetur adipisicing elit. <a href="""">This is link in new line</a> Cumque mollitia repellat soluta voluptates molestias veniam reiciendis.
</div>
```






To remove the extra clickable area





```
div {
  background: #333;
  color: #fff;
  margin: 0 auto;
  padding: 20px;
  text-align: center;
  width: 400px;
}
a {
  color: red;
  display: inline;
}
a:after {
  content: ""\a"";
 white-space: pre;
}
a:before {
  content: ""\a"";
 white-space: pre;
}
```


```
<div>
  Lorem ipsum dolor sit amet, consectetur adipisicing elit. <a href="""">This is link in new line</a> Cumque mollitia repellat soluta voluptates molestias veniam reiciendis.
</div>
```






"
"Replacing items in a dictionary of lists in Python
I am new to python and programming and need some help replacing items in a dictionary of lists. I would like to replace `None` with `'None'` in the dictionary below:



```
dict = {'Chester100': ['Caesar, Augustus', '05/10/2012', '09/09/2012', None],
        'Rochester102': ['Henrich, Norton', '08/18/2014', '12/17/2014', None],
        'Rochester100': ['Caeser, Julius', '08/18/2014', '12/17/2014', None],
        'Rochester101': [None, None, None, '08/18/2012']}

```

My code is as follows:



```
new_dict = {}

for i in dict: #This accesses each dictionary key.
    temp = []
    for j in dict[i]: #This iterates through the inner lists
        if j is None:
            temp.append('None')
        else:
            temp.append(j)
        temp2 = {str(i):temp}
        new_dict.update(temp2)

    print(new_dict)

```

yields



```
{'Chester100': ['Caesar, Augustus', '05/10/2012', '09/09/2012', 'None'], 
'Rochester102': ['Henrich, Norton', '08/18/2014', '12/17/2014', 'None'], 
'Rochester100': ['Caeser, Julius', '08/18/2014', '12/17/2014', 'None'], 
'Rochester101': ['None', 'None', 'None', '08/18/2012']}

```

Is there a way to do this in a fewer lines of code or more efficiently using list comprehension or other methods? Should nested for loop (as I have it in my code) be avoided? Thanks.


**Using Python 3.4.1**


","Use a dictionary comprehension:



```
>>> {k:[e if e is not None else 'None' for e in v] for k,v in di.items()}
{'Rochester102': ['Henrich, Norton', '08/18/2014', '12/17/2014', 'None'], 'Rochester100': ['Caeser, Julius', '08/18/2014', '12/17/2014', 'None'], 'Rochester101': ['None', 'None', 'None', '08/18/2012'], 'Chester100': ['Caesar, Augustus', '05/10/2012', '09/09/2012', 'None']}

```

And don't name a dict `dict` since it will mask the built in function by that name.




---


If you have *huge* dicts or lists, you may want to modify your data in place. If so, this may be the most efficient:



```
for key, value in di.items():
    for i, e in enumerate(value):
        if e is None: di[key][i]='None'    

```

"
"Signal-Desktop: how to export messages?
How can `signal-desktop` messages be exported?


I want to backup my correspondence.


Is it possible at all?


","Yes, it is possible.


Just save this in a file `<yourFilename>`:



```
sigBase=""${HOME}/.config/Signal/"";
key=$( /usr/bin/jq -r '.""key""' ${sigBase}config.json );
db=""${HOME}/.config/Signal/sql/db.sqlite"";
clearTextMsgs=""${sigBase}clearTextMsgs.csv"";

/usr/bin/sqlcipher -list -noheader ""$db"" ""PRAGMA key = \""x'""$key""'\"";select json from messages;"" > ""$clearTextMsgs"";

```

and call it via `bash <yourFilename>`. Or render it executable with `chmod 700 <yourFilename>` and call it directly: `./<yourFilename>`


This script uses `sqlcipher` and `jq` with signal-desktop's database key to open, decrypt and extract all messages in `JSON` format into `clearTextMsgs.csv` inside your signal-desktop folder `~/.config/Signal`.


Besides the key extraction by filtering JSON with `jq` (from `~/.config/Signal/config.json`), the crucial bit happens here:



```
sqlcipher -list -noheader <DB> <SQL>

```

where `<SQL>` contains the `PRAGMA key` definition and the actual SQL statement (`SELECT json FROM messages;`).


One can then use `jq` to access any key/value from the messages backup.


You have to install `sqlcipher` and `jq` for that:



```
sudo apt install sqlcipher jq

```

**Note:** 


While this does extract all messages, we need to specify that ""`all`"" in signal-desktop has the meaning of ""`all messages actually loaded`"". So, in order to extract every single message, the slider of the active contact has to be slid way up, then signal-desktop will load previously not availalble messages (lather rinse repeat until satisfied). Do so as far in the past you would want your messages loaded. This gets tedious quite quickly. Remember to do so for all of your contacts' histories. 


Having that said, it is technically feasable to backup your message history, in practice it is a manual job. A way around this might be a cron job backing up all recent messages, maybe once a day. Then this is likely to contain duplicates and might miss messages in case signal-desktop has been restarted.


In any case, this method is working fine if the (not too far -- read: a couple of months maybe) history is to be searched programmatically once in a while.


"
"strstr on huge mmapped file
I open huge (11Gb) file, mmap it to memmory, and fail to search the string in the file


my code is



```
 if ( (fd  = open(""l"", O_RDONLY)) < 0 )     err_sys(""Cant open file"");
 if ( fstat(fd, &statbuf) < 0 )             err_sys(""Cant get file size"");
 printf(""size is %ld\n"", statbuf.st_size);

 if ( (src = mmap(0, statbuf.st_size, PROT_READ, MAP_SHARED, fd, 0)) == MAP_FAILED )  err_sys(""Cant mmap"");
 printf(""src pointer is at %ld\n"", src);

 char * index = strstr(src, ""bin/bash"");
 printf(""needle is at %ld\n"", index);

```

It works on small files, but on huge sources returns 0. What function should I use to search in huge mmapped files?


The output is:



```
size is 11111745740
src pointer is at 140357526544384
needle is at 0

```

","You should not use `strstr()` to search for text in a memory mapped file:


- If the file is binary, it most likely contains null bytes that will stop the search too soon. This is probably what you observe.
- If the file is pure text, but does not contain a match, `strstr` will keep scanning beyond the end of the file, invoking undefined behavior by attempting to read unmapped memory.


You could instead use a function with equivalent semantics but applied to raw memory instead of C strings, `memmem()`, available on Linux and BSD systems:



```
void *memmem(const void *p1, size_t size1, const void *p2, size_t size2);

```

Note that you also use the wrong `printf` formats: it should be `%p` for `src` and `index` and you might prefer to print the offset as a `ptrdiff_t` or an `unsigned long long`:



```
    if ((fd  = open(""l"", O_RDONLY)) < 0)
        err_sys(""Cannot open file"");
    if (fstat(fd, &statbuf) < 0)
        err_sys(""Cannot get file size"");

    printf(""size is %llu\n"", (unsigned long long)statbuf.st_size);

    if ((src = mmap(0, statbuf.st_size, PROT_READ, MAP_SHARED, fd, 0)) == MAP_FAILED)
        err_sys(""Cannot mmap"");

    printf(""src pointer is at %p\n"", (void*)src);

    char *index = memmem(src, statbuf.st_size, ""bin/bash"", strlen(""bin/bash""));
    printf(""needle is at %p\n"", (void*)index);

    if (index != NULL)
        printf(""needle is at offset %llu\n"", (unsigned long long)(index - src));

```

If `memmem` is not available on your platform, here is a simple implementation:



```
#include <string.h>

void *memmem(const void *haystack, size_t n1, const void *needle, size_t n2) {
    const unsigned char *p1 = haystack;
    const unsigned char *p2 = needle;

    if (n2 == 0)
        return (void*)p1;
    if (n2 > n1)
        return NULL;

    const unsigned char *p3 = p1 + n1 - n2 + 1;
    for (const unsigned char *p = p1; (p = memchr(p, *p2, p3 - p)) != NULL; p++) {
        if (!memcmp(p, p2, n2))
            return (void*)p;
    }

    return NULL;
}    

```

"
"Changing  src based on screen size with Javascript
I'm testing on building a site locally on my machine using bootstrap. 


I have a `<video>` sort of as the header of the site. 


I would like this video to show the full width and height on mobile, and show a cropped/wide version of the video on desktop. I tried using inline media queries in the `<source>` tags, so that the `src` would change but nothing would work.


So I switched gears and used some javascript to change it that way. 


So the crazy thing is, it seems my script works. When I look in chrome dev tools, the `src`does in fact change when I resize my browser screen, however, it does not reflect on the site itself, it keeps whatever `src` I set it to in the html, as if it is ignoring my script. 


I have tried everything I could think of, and I'm just stuck, not sure how to go about it any further. My code is below:


HTML



```
<video class=""col-12"" loop muted autoplay >
  <source id=""hvid"" src=""media/test.mp4"" type=""video/mp4"">
</video>

```

JS



```
let homeVideo = document.getElementById('hvid')
console.log(homeVideo.src)


function myFunction(x) {
  if (x.matches) { // If media query matches
    homeVideo.src = ""media/test.mp4"";
  } else {
   homeVideo.src = ""media/test-3.mp4"";
  }
}

var x = window.matchMedia(""(max-width: 700px)"")
myFunction(x) // Call listener function at run time
x.addListener(myFunction) // Attach listener function on state changes
;
console.log(homeVideo.src)

```

***-Edits-***


JS



```
var w = window.matchMedia(""(max-width: 700px)"");
 var vid = document.getElementById(""vid"");
 var source = document.getElementById(""hvid"");


window.addEventListener(""resize"", function screenres(){
  if (w.matches) {
    vid.pause();
    source.src = ""media/test.mp4"";
    vid.load();
    vid.play();
  } else {
    vid.pause();
    source.src = ""media/test-3.mp4"";
    vid.load();
    vid.play();
  };

});

```

HTML



```
  <div class=""container"">
      <div class=""row"">
          <video id=""vid"" class=""col-12"" loop muted autoplay>
            <source id=""hvid"" src=""media/test.mp4"" type=""video/mp4"">
          </video>
        </div>
    </div>

```

","Just get the viewport size, and based on that value, pause the video, change the `src` link, load the new video and play the new video.


But do note that you will need to refresh the page after changing the browser size to see the video change.


If you want the video to change whenever the screen resizes as well as on page refresh, you will first need to move the above JavaScript to a function and run it when a `resize` event is fired. Then, for the page load, you need to remove the video element from your HTML and add it on page load using the [createElement()](https://developer.mozilla.org/en-US/docs/Web/API/Document/createElement) method with the `src` attribute value also added depending on the viewport width.




---


Check this **[JSFiddle](https://jsfiddle.net/AndrewL64/6pLua510/46/)** or run the following **Code Snippet** for a practical example of what I have described above:





```
/* JavaScript */

  var w = window.matchMedia(""(max-width: 700px)"");
  var vid = document.getElementById(""vid"");
  var source = document.createElement(""source"");
  source.id = ""hvid"";
  source.setAttribute(""type"", ""video/mp4"");
  vid.appendChild(source);
  
  if (w.matches) {
    vid.pause();
    source.removeAttribute(""src"");
    source.setAttribute(""src"", ""https://storage.googleapis.com/coverr-main/mp4/Love-Boat.mp4"");
    vid.load();
    vid.play();
  } else {
    vid.pause();
    source.removeAttribute(""src"");
    source.setAttribute(""src"", ""https://sample-videos.com/video123/mp4/720/big_buck_bunny_720p_1mb.mp4"");
    vid.load();
    vid.play();
  }

window.addEventListener(""resize"", function(){
  var w = window.matchMedia(""(max-width: 700px)"");
  var vid = document.getElementById(""vid"");
  var source = document.getElementById(""hvid"");
  
  if (w.matches) {
    vid.pause();
    source.src = ""https://storage.googleapis.com/coverr-main/mp4/Love-Boat.mp4"";
    vid.load();
    vid.play();
  } else {
    vid.pause();
    source.src = ""https://sample-videos.com/video123/mp4/720/big_buck_bunny_720p_1mb.mp4"";
    vid.load();
    vid.play();
  }
});
```


```
/* CSS */

html, body {margin: 0; padding:0; width: 100%; height: 100%;}.row{display: block !important;}
```


```
<!-- CDN Links -->

<script src=""https://cdnjs.cloudflare.com/ajax/libs/jquery/3.3.1/jquery.min.js""></script><script src=""https://cdnjs.cloudflare.com/ajax/libs/popper.js/1.14.0/umd/popper.min.js""></script><script src=""https://stackpath.bootstrapcdn.com/bootstrap/4.1.0/js/bootstrap.min.js""></script>
<link href=""https://stackpath.bootstrapcdn.com/font-awesome/4.7.0/css/font-awesome.min.css"" rel=""stylesheet""/><link href=""https://stackpath.bootstrapcdn.com/bootstrap/4.1.0/css/bootstrap.min.css"" rel=""stylesheet""/>

<!-- HTML -->

<div class=""container"">
  <div class=""row"">
    <video id=""vid"" class=""col-12"" loop muted autoplay></video>
  </div>  
</div>
```






"
"DirectoryStream.Filter example for listing files that based on certain date/time
I am trying to investigate a `DirecoryStream.Filter` example for `newDirectoryStream` where I can achieve to list all files under a directory(*and all its sub directories*) that are older than 60 days, as an example. 



```
DirectoryStream<Path> dirS = Files.newDirectoryStream(Paths.get(""C:/myRootDirectory""), <DirectoryStream.filter>);

for (Path entry: dirS) {
    System.out.println(entry.toString());
}

```

In the code above, what should be the `DirectoryStream.filter`?   
It will be a great help as I am in a project where I am trying to delete files older than a certain timestamp and pre-java 1.7 `File.listFiles()` just hangs.


Could `Files.walkFileTree()` provide an option?


","If I get this right, you have 2 situations:


1. Create a custom `filter` to select files older than 60 days
2. Traverse through subdirectories (the entire `FileTree`) and gather your information


The custom filter is easier to implement with conditions of 60 days implemented using `Calendar` class:



```
DirectoryStream.Filter<Path> filter = new DirectoryStream.Filter<Path>() {
    @Override
    public boolean accept(Path entry) throws IOException {BasicFileAttributes attr = Files.readAttributes(entry,BasicFileAttributes.class);
    FileTime creationTime = attr.creationTime();
    Calendar cal = Calendar.getInstance();
    int days = cal.fieldDifference(new Date(creationTime.toMillis()),Calendar.DAY_OF_YEAR);
        return (Math.abs(days) > 60);
        }
  };

```

The normal execution would only look for the files in the root directory. To look for subdirectory, your guess of using `walkFileTree()` is right.


However, this requires an implementation of the `FileVisitor` interface, a simple implementation of which luckily is bundled with 7 - `SimpleFileVisitor`.


To traverse through the subdirectories, you can choose to override a directory specific method - I have used `preVisitDirectory` of `SimpleFileVisitor` here:



```
Files.walkFileTree(dirs, new SimpleFileVisitor<Path>() {
            @Override
            public FileVisitResult preVisitDirectory(Path file,
                    BasicFileAttributes attrs) {

```

Since `preVisitDirectory` will be custom made to `return FileVisitResult.CONTINUE;` in case you don't have any additional restrictions, we would leverage `preVisitDirectory` method to iterate through our directory 1 at a time while applying the filter.



```
Files.walkFileTree(dirs, new SimpleFileVisitor<Path>() {
            @Override
            public FileVisitResult preVisitDirectory(Path file,
                    BasicFileAttributes attrs) {

                DirectoryStream.Filter<Path> filter = new DirectoryStream.Filter<Path>() {
                    @Override
                    public boolean accept(Path entry) throws IOException {
                        BasicFileAttributes attr = Files.readAttributes(entry,
                                BasicFileAttributes.class);
                        FileTime creationTime = attr.creationTime();
                        Calendar cal = Calendar.getInstance();
                        int days = cal.fieldDifference(
                                new Date(creationTime.toMillis()),
                                Calendar.DAY_OF_YEAR);
                        return (Math.abs(days) > 60);
                    }
                };
                try (DirectoryStream<Path> stream = Files.newDirectoryStream(
                        file, filter)) {
                    for (Path path : stream) {
                        System.out.println(path.toString());
                    }
                } catch (IOException ex) {
                    ex.printStackTrace();
                }
                return FileVisitResult.CONTINUE;

            }
        }); 

```

This would give you the files from the entire directory and subdirectory structures for the required filter criteria, complete main method below:



```
public static void main(String[] args) throws IOException {
        Path dirs = Paths.get(""C:/"");

        Files.walkFileTree(dirs, new SimpleFileVisitor<Path>() {
            @Override
            public FileVisitResult preVisitDirectory(Path file,
                    BasicFileAttributes attrs) {

                DirectoryStream.Filter<Path> filter = new DirectoryStream.Filter<Path>() {
                    @Override
                    public boolean accept(Path entry) throws IOException {
                        BasicFileAttributes attr = Files.readAttributes(entry,
                                BasicFileAttributes.class);
                        FileTime creationTime = attr.creationTime();
                        Calendar cal = Calendar.getInstance();
                        int days = cal.fieldDifference(
                                new Date(creationTime.toMillis()),
                                Calendar.DAY_OF_YEAR);
                        return (Math.abs(days) > 60);
                    }
                };
                try (DirectoryStream<Path> stream = Files.newDirectoryStream(
                        file, filter)) {
                    for (Path path : stream) {
                        System.out.println(path.toString());
                    }
                } catch (IOException ex) {
                    ex.printStackTrace();
                }
                return FileVisitResult.CONTINUE;

            }
        });

    }

```

"
"ORACLE PL/SQL check whether string is NULL
Query 1:



```
create or replace procedure toUp(code in number)
is sname staff_master.staff_name%type;
recnotfound exception;
begin
select staff_name into sname from staff_master where staff_code=code;
if sname is NULL then
raise recnotfound;
else
update staff_master set staff_name=upper(staff_name) where staff_code=code;
end if;
exception
when recnotfound then dbms_output.put_line('Record not found');
end;

```

Query 2:



```
declare
commsn emp.comm%type;
no_comm exception;
begin
select comm into commsn from emp where empno=7369;
if commsn is NULL then
raise no_comm;
else
dbms_output.put_line('Comm is '||commsn);
end if;
exception
when no_comm then dbms_output.put_line('Commsn for emp doesnt exist');
end;

```

Here in Query 1 I'm checking whether sname is null.. However, when I pass an invalid code as a parameter to the procedure.. sname should be NULL and hence the exception 'recnotfound' must get raised.. but it is showing the following error:



```
SQL> exec toUp(7369);
BEGIN toUp(7369); END;

*
ERROR at line 1:
ORA-01403: no data found
ORA-06512: at ""LAB06TRG15.TOUP"", line 6
ORA-06512: at line 1

```

But when I do the same with Query 2 it is working as expected..
I guess it has something to do with how varchar2 is checked for null.. Am I doing it correctly?


I modified the code as follows :



```
create or replace procedure toUp(code in number)
is
sname staff_master.staff_name%type;
recnotfound exception;
begin
select staff_name into sname from staff_master where staff_code=code;
if sname is NULL then
dbms_output.put_line('a');
raise recnotfound;
else
dbms_output.put_line('b');
--update staff_master set staff_name=upper(staff_name) where staff_code=code;
end if;
exception
when recnotfound then dbms_output.put_line('Record not found');
when no_data_found then raise recnotfound;
end;

```

I get :



```
BEGIN toUp(7369); END;

*
ERROR at line 1:
ORA-06510: PL/SQL: unhandled user-defined exception
ORA-06512: at ""LAB06TRG15.TOUP"", line 16
ORA-01403: no data found
ORA-06512: at line 1

```

How do I solve this?


P.S. I want to do this using Exception only.. Its part of an assignment .. 


","If a query returns no rows then an ""ORA-01403: no data found"" error is raised. Your expectation, I think, is that execution will continue but no value will have been assigned to the variable -- that's not the case.


If what you want to do is check for the existence of a record then use:



```
select count(*)
into   row_found
from   ...
where  ...
   and rownum = 1;

```

this is guaranteed to return a single row with a value of 0 or 1 into the row\_found variable.


With regard to your edit, you are not handling the raising of the user defined exception in the exception handling block. Wrap the SELECT with a BEGIN-END-EXCEPTION.



```
begin
  begin
    select ..
  exception when NO_DATA_FOUND then raise recnotfound;
  end;
  if sname is NULL then
    dbms_output.put_line('a');
    raise recnotfound;
  end if;
exception
  when recnotfound then dbms_output.put_line('Record not found');
end;

```

I'm not clear what you're trying to do here though. Is the sname ever going to be returned as null from the query?


"
"Flink window state size and state management
After reading flink's documentation and searching around, i couldn't entirely understand how flink's handles state in its windows.
Lets say i have an hourly tumbling window with an aggregation function that accumulate msgs into some java pojo or scala case class. 
Will The size of that window be tied to the number of events entering that window in a single hour, or will it just be tied to the pojo/case class, as im accumalting the events into that object. (e.g if counting 10000 msgs into an integer, will the size be close to 10000 \* msg size or size of an int?)
Also, if im using pojos or case classes, does flink handle the state for me (spills to disk if memory exhausted/saves state at check points etc) or must i use flink's state objects for that? 


Thanks for your help! 


","The state size of a window depends on the type of function that you apply. If you apply a `ReduceFunction` or `AggregateFunction`, arriving data is immediately aggregated and the window only holds the aggregated value. If you apply a `ProcessWindowFunction` or `WindowFunction`, Flink collects all input records and applies the function when time (event or processing time depending on the window type) passes the window's end time. 


You can also combine both types of functions, i.e., have an `AggregateFunction` followed by a `ProcessWindowFunction`. In that case, arriving records are immediately aggregated and when the window is closed, the aggregation result is passed as single value to the `ProcessWindowFunction`. This is useful because you have incremental aggregation (due to `ReduceFunction` / `AggregateFunction`) but also access to the window metadata like begin and end timestamp (due to `ProcessWindowFunction`).


How the state is managed depends on the chosen state backend. If you configure the `FsStateBackend` all local state is kept on the heap of the TaskManager and the JVM process is killed with an `OutOfMemoryError` if the state grows too large. If you configure the `RocksDBStateBackend` state is spilled to disk. This comes with de/serialization costs for every state access but gives much more storage for state.


"
"Comparison of a Dataframe column values with a list
Consider this Dataframe:



```
df = pd.DataFrame({'A': [1, 1, 2, 2, 3, 3],
                   'B': [10, 15, 20, 25, 30,35],
                   'C': [100, 150, 200, 250, 300, 350]})

```

This is the code to get values of column C, where it is the first row of each group (Column A):



```
firsts = df.groupby('A').first()['C']

```

So first will be: `(100, 200, 300)`.


Now I want to add new column which it will be `1` if value of column C for row is in `firsts` otherwise it will be `0`.





| A | B | C | D |
| --- | --- | --- | --- |
| 1 | 10 | 100 | 1 |
| 1 | 15 | 150 | 0 |
| 2 | 20 | 200 | 1 |
| 2 | 25 | 250 | 0 |
| 3 | 30 | 300 | 1 |
| 3 | 35 | 350 | 0 |



I used this:



```
df['D'] = df['C'].apply(lambda x: 1 if x in firsts else 0)

```

But the output is:





| A | B | C | D |
| --- | --- | --- | --- |
| 1 | 10 | 100 | 0 |
| 1 | 15 | 150 | 0 |
| 2 | 20 | 200 | 0 |
| 2 | 25 | 250 | 0 |
| 3 | 30 | 300 | 0 |
| 3 | 35 | 350 | 0 |



I appreciate if anyone explain why my solution is wrong and what is actual solution to this problem?


","You can use `isin` method:



```
df['D'] = df.C.isin(firsts).astype(int)

df
#   A   B   C   D
#0  1   10  100 1
#1  1   15  150 0
#2  2   20  200 1
#3  2   25  250 0
#4  3   30  300 1
#5  3   35  350 0

```



---


The reason your approach fails is that python `in` operator check the index of a Series instead of the values, the same as how a dictionary works:



```
firsts
#A
#1    100
#2    200
#3    300
#Name: C, dtype: int64

1 in firsts
# True

100 in firsts
# False

2 in firsts
# True

200 in firsts
# False

```

Modifying your method as follows works:



```
firstSet = set(firsts)
df['C'].apply(lambda x: 1 if x in firstSet else 0)

#0    1
#1    0
#2    1
#3    0
#4    1
#5    0
#Name: C, dtype: int64

```

"
"How do I enable document map in Visual Studio Code?
In the **Sublime text** as well as in **Notepad++** I can enable Document Map.


[Here is a screenshot of the document map of Notepad++.](https://i.stack.imgur.com/wuW3w.png)


","Minimap was added in Visual Studio Code 1.10. The [release notes](https://code.visualstudio.com/updates/v1_10#_preview-minimap) say:



> 
> To enable VS Code's Minimap, set ""editor.minimap.enabled"": true to
>  turn on the rendering of a Minimap for the current file.
> 
> 
> 


To set the preference open VS User Settings (Preferences > Settings). This will open two side-by-side documents. Add the entry and set the value to true. In the side-by-side view you choose between changing the ""User Settings"" (global) or ""Workspace Settings"" (current workspace only).



```
{
    ""typescript.check.tscVersion"": false,
    ""git.confirmSync"": false,
    ""editor.minimap.enabled"": true
}

```

"
"SQL Server Indexes - Initial slow performance after creation
Using SQL Server 2005. This is something I've noticed while doing some performance analysis.


I have a large table with about 100 million rows. I'm comparing the performance of different indexes on the table, to see what the most optimal is for my test scenario which is doing about 10,000 inserts on that table, among other things on other tables. While my test is running, I'm capturing an SQL Profiler trace which I load in to an SQL table when the test has finished so I can analyse the stats.


The first test run after recreating a different set of indexes on the table is very noticeably slower than subsequent runs - typically about 10-15 times slower for the inserts on this table on the first run after the index creation.


Each time, I clear the data and execution plan cache before the test.


What I want to know, is the reason for this initial poorer performance with a newly created set of indexes?
Is there a way I can monitor what is happening to cause this for the first run?


","One possibility is that the default [fill factor](http://msdn.microsoft.com/en-us/library/aa933139(SQL.80).aspx) of zero is coming in to play.


This means that there's 'no room' in the index to accommodate your inserts. When you insert, a page split in the index is needed, which adds some empty space to store the new index information. As you carry out more inserts, more space is created in the index. After a while the rate of splitting will go down, because your inserts are hitting pages that are not fully filled, so splits are not needed. An insert requiring page splits is more expensive than one that doesn't.


You can set the fill factor when you create the index. Its a classic trade off between space used and performance of different operations.


I'm going go include a link to some [Sybase ASE docs](http://manuals.sybase.com/onlinebooks/group-as/asg1250e/ptallbk/@Generic__BookTextView/19288;pt=19235), 'cos they are nicely written and mostly applicable to SQL Server too.


"
"how to map latitude/longitude to a specific US county?
I have a dataframe containing U.S. `latitude` and `longitude` coordinates. 


I would like to create a variable that shows the respective U.S. county that includes these coordinates. 


How can I do that in R/Python?


Thanks!


","You could use [geopy](https://pypi.python.org/pypi/geopy).


Example from the documentation using Nominatim (Open Street Map):



```
>>> from geopy.geocoders import Nominatim
>>> geolocator = Nominatim(user_agent=""http"")
>>> location = geolocator.reverse(""52.509669, 13.376294"")
>>> print(location.address)
Potsdamer Platz, Mitte, Berlin, 10117, Deutschland, European Union
>>> print((location.latitude, location.longitude))
(52.5094982, 13.3765983)
>>> print(location.raw)
{'place_id': '654513', 'osm_type': 'node', ...}

```

The raw output has a dict key `country`, if one can be found.


"
"React Native webview get url
Webview component allowing me to specify a destination url, e.g. facebook.com



```
render() {

    return (
        <WebView
          source={{uri: https://www.facebook.com}}
          style={{marginTop: 20}}
        />
    );
  }

```

However, If i click the link in facebook, how can I get the url being clicked or the url being landed?


","Here is a Webview I used for my previous project.



```
<WebView
       ref=""webview""
       source={{uri:this.state.url}}
       onNavigationStateChange={this._onNavigationStateChange.bind(this)}
       javaScriptEnabled = {true}
       domStorageEnabled = {true}
       injectedJavaScript = {this.state.cookie}
       startInLoadingState={false}
     />

```

for you essential is this line:



```
           onNavigationStateChange={this._onNavigationStateChange.bind(this)}

```

and you can read the URL like this:



```
_onNavigationStateChange(webViewState){
  console.log(webViewState.url)
}

```

If I remember correctly this fires 3 times when loading a url.


webviewState object prototype:



```
{
  canGoBack: bool,
  canGoForward: bool,
  loading: bool,
  target: number,
  title: string,
  url: string,
}

```

The 3rd (final) time this event is called, the loading attribute is false


let me know if it helps.


"
"Create data set from clicks in Shiny ggplot
I'm a Shiny novice, but I'm trying to use it in a project that I'm working on. I'd like to be able to do two things from clicking on a point on a ggplot graph: add a plot character at the specified point (conditioned on info from the sidebar), and add the coordinates (with info from the sidebar) to a data frame. Here's what I've got in terms of code so far:



```
library(shiny)
library(ggplot2)

df = data.frame()


ui = pageWithSidebar(
  headerPanel(""Test""),

  sidebarPanel(
    radioButtons(""orientation"", ""Pick"", c(""L"", ""P"", ""H"")),

    selectInput(
      ""select1"",
      ""Select Here:"",
      c(""Option 1"", ""Option 2"")
    ),

    selectInput(
      ""select2"",
      ""Select Here:"",
      c(""Option 3"", ""Option 4""),
    ),

    radioButtons(""type"", ""Type:"", c(""P"", ""S"")),

    radioButtons(""creator"", ""Creator?"", c(""H"", ""A""))
  ),

  mainPanel(
    plotOutput(""plot1"", click = ""plot_click""),
    verbatimTextOutput(""info""),
    actionButton(""update"", ""Add Event"")
  )
)

server = function(input, output){

  output$plot1 = renderPlot({
    ggplot(df) + geom_rect(xmin = 0, xmax = 100, ymin = 0, ymax = 50, fill = ""red"")
  })

  output$info = renderText({
    paste0(""x = "", input$plot_click$x, ""\ny = "", input$plot_click$y)
  })
}

shinyApp(ui, server)

```

I'm confused on how to add the clicked x and y points from `plot_click` to `df` so that I can add the data to a bigger database. Any help would be appreciated, and I'd be happy to give more info about the project if needed!


","Here's a general framework that you could use:


1. Use `reactiveValues()` to set up a reactive data.frame with columns for `x`, `y`, `inputs`
2. Create a plot using the reactive data.frame with plotting characteristics based on `input`
3. Upon clicking plot, add a new row to the reactive data.frame using `observeEvent`
4. (Optional) Add an `actionButton` to remove the last added point


A simplified example based on your code is below. The table is based on [this answer](https://stackoverflow.com/a/36346950/8099834).


[![enter image description here](https://i.stack.imgur.com/GpsWZ.gif)](https://i.stack.imgur.com/GpsWZ.gif)



```
library(shiny)
library(ggplot2)

ui <- pageWithSidebar(
    headerPanel(""Example""),
    sidebarPanel(
        radioButtons(""color"", ""Pick Color"", c(""Pink"", ""Green"", ""Blue"")),
        selectInput(""shape"", ""Select Shape:"", c(""Circle"", ""Triangle""))
    ),
    mainPanel(
        fluidRow(column(width = 6,
                        h4(""Click plot to add points""),
                        actionButton(""rem_point"", ""Remove Last Point""),
                        plotOutput(""plot1"", click = ""plot_click"")),
                 column(width = 6,
                        h4(""Table of points on plot""),
                        tableOutput(""table"")))
    )
)

server = function(input, output){

    ## 1. set up reactive dataframe ##
    values <- reactiveValues()
    values$DT <- data.frame(x = numeric(),
                            y = numeric(),
                            color = factor(),
                            shape = factor())

    ## 2. Create a plot ##
    output$plot1 = renderPlot({
       ggplot(values$DT, aes(x = x, y = y)) +
            geom_point(aes(color = color,
                           shape = shape), size = 5) +
            lims(x = c(0, 100), y = c(0, 100)) +
            theme(legend.position = ""bottom"") +
            # include so that colors don't change as more color/shape chosen
            scale_color_discrete(drop = FALSE) +
            scale_shape_discrete(drop = FALSE)
    })

    ## 3. add new row to reactive dataframe upon clicking plot ##
    observeEvent(input$plot_click, {
        # each input is a factor so levels are consistent for plotting characteristics
        add_row <- data.frame(x = input$plot_click$x,
                              y = input$plot_click$y,
                              color = factor(input$color, levels = c(""Pink"", ""Green"", ""Blue"")),
                              shape = factor(input$shape, levels = c(""Circle"", ""Triangle"")))
        # add row to the data.frame
        values$DT <- rbind(values$DT, add_row)
    })

    ## 4. remove row on actionButton click ##
    observeEvent(input$rem_point, {
        rem_row <- values$DT[-nrow(values$DT), ]
        values$DT <- rem_row
    })

    ## 5. render a table of the growing dataframe ##
    output$table <- renderTable({
        values$DT
    })
}

shinyApp(ui, server)

```

"
"C# !Conditional attribute?
Does C# have a *not* `Conditional` (`!Conditional`, `NotConditional`, `Conditional(!)`) attribute?




---


i know C# has a [`Conditional` attribute](https://msdn.microsoft.com/en-us/library/system.diagnostics.conditionalattribute):



```
[Conditional(""ShowDebugString"")]
public static void ShowDebugString(string s)
{
   ...
}

```

which is equivalent1 to:



```
public static void ShowDebugString(string s)
{
#if ShowDebugString
   ...
#endif
}

```

But in this case i want the *inverse* behavior (you have to specifically *opt out*):



```
public static void ShowDebugString(string s)
{
#if !RemoveSDS
   ...
#endif
}

```

Which leads me to try:



```
[!Conditional(""RemoveSDS"")]
public static void ShowDebugString(string s)
{
   ...
}

```

which doesn't compile. And:



```
[Conditional(""!RemoveSDS"")]
public static void ShowDebugString(string s)
{
   ...
}

```

which doesn't compile. And:



```
[NotConditional(""RemoveSDS"")]
public static void ShowDebugString(string s)
{
   ...
}

```

which doesn't compile because it's only wishful thinking.


1 [Not true, but true enough. Don't make me bring back the Nitpicker's Corner.](http://blogs.msdn.com/b/oldnewthing/archive/2011/11/07/10234436.aspx) 


","First, having the `Conditional` attribute is *not* equivalent to having `#if` inside the method. Consider:



```
ShowDebugString(MethodThatTakesAges());

```

With the real behaviour of `ConditionalAttribute`, `MethodThatTakesAges` doesn't get called - the entire call *including argument evaluation* is removed from the compiler.


Of course the other point is that it depends on the compile-time preprocessor symbols at the compile time of the *caller*, not of the *method* :)


But no, I don't believe there's anything which does what you want here. I've just checked [the C# spec section which deals with conditional methods and conditional attribute classes](https://learn.microsoft.com/en-us/dotnet/csharp/language-reference/language-specification/attributes#the-conditional-attribute), and there's nothing in there suggesting there's any such mechanism.


"
"How can I route traffic from my main wired router to the wireless computers on the other side of a connected Wi-Fi router?
I have a router that takes my Verizon FiOS as input and has five output ports. The main internal IP address of the router is *192.168.0.1*.


From one of the ports on the wired router, I have my Wi-Fi router connected, and it has an IP address of *192.168.0.10*. The other ports on the wired router all have *192.168.0.x* IP addresses.


On the other side of the Wi-Fi router, the wireless IP addresses are of the form *192.168.1.x*.


From the Wi-Fi network I **can** get to the *192.168.0.x* wired IP addresses, but I **cannot** get to the *192.168.1.x* wireless IP addresses from the wired network.


So, I believe I need to make a change in the Wi-Fi router to allow traffic from the *192.168.0.x* wired IP addresses into the *192.168.1.x* wireless IP addresses. Is this correct?


What do I need to look for in order to make this change (e.g., wireless settings, LAN settings, WAN settings, etc.)? Once there, what type of change should I make?


","You need to set up the device you're calling your ""Wi-Fi router"" to act as a simple Wi-Fi AP instead of acting as a router.


Some wireless router devices allow you to disable their NAT gateway and DHCP server features, and will automatically make their WAN Ethernet port act like a LAN Ethernet port. Others don't have this option, but you can still accomplish the same thing by not using the WAN port (connect the Ethernet cable from the LAN port of the upstream router to a LAN port of the downstream router), and by disabling the DHCP server feature if possible, or setting its DHCP IP address lease pool to be zero length (so it has no addresses it can lease out via DHCP).


By hooking things up this way, your home network becomes one single LAN with a single IP subnet.


"
"where does downcasted object point to?

```
public class Animal{

    int n = 5;

    public static void main(String[] args) {

        Animal a = new Animal();
        Animal ah = new Horse();
        Horse h = new Horse(); 

        System.out.println(h.n); // prints 7
        System.out.println(ah.n); // prints 5
        h = (Horse) ah;
        System.out.println(h.n); // prints 7

    }
}

class Horse extends Animal{

    int n = 7;

}

```

### My question:


Why does `h.n` still print 7 after `h = (Horse) ah`? After the assignment it should point to the same object that `ah` points and the n field points to 5? 


","First, let's call the field `n` of the class `Animal` ""`Animal.n`"" to avoid confusion.


Fields, unlike methods, are not subject to overriding. So in your Horse class, where you may think you are overriding the value of `Animal.n` with 7, you are actually declaring a new variable called `n` (let's call it `Horse.n` to avoid confusion).


So really, you have a class called `Horse` with two fields: `Animal.n` and `Horse.n`. Which field you get when you say ""`n`"" depends upon the *static* type of the variable at the time.


When you have an object whose type is `Horse`, but upcast to an `Animal`, the `n` field refers to `Animal.n`, and has a value of ""5"". Hence `ah.n` is ""5"".


When you have the same object, downcast again to a `Horse`, the `n` field refers to `Horse.n`, and has a value of ""7"". Hence `h.n` is ""7"".


**To clarify**: Indeed, `h` does point to the same object that `ah` points to -- downcasting does not change what object is being pointed at. However, the static type does affect which field of the object is being requested.


"
"What is the traditional name of this Lisp function?
I've seen this function elsewhere, and as I recall, it's got a standard name. I can't remember it though, and perusing Alexandria's `list.lisp` library didn't reveal any secrets to me.



```
(defun familiar-function (list func)
  (loop for e in list
        collect (if (listp e)
                    (familiar-function e func)
                    (funcall func e))))

```

","It looks like [`tree-map`](https://common-lisp.net/project/docudown/documentation/metabang.utilities-package/function-tree--map.html), also called [`maptree`](http://www.informatimago.com/develop/lisp/doc/com.informatimago.common-lisp.cesarum.list.html#MAPTREE):



```
(familiar-function '(() (3 2 3) (2) (2) (55 22 33))
                   #'1+)
=> (NIL (4 3 4) (3) (3) (56 23 34))

```

As mentioned in the comment, your approach assume trees are proper lists and do not apply the function to non-nil `cdr` cells, contrary to the linked functions.


"
"Import Polygon Into Google My Maps
## Background


I have a Google Map, creating with the JavaScript API.. I have a set of polygons which are overlays on the map.


These polygons started out as KML files, but have been converted to encoded polygons. Encoding Algorithm information: **<https://developers.google.com/maps/documentation/utilities/polylinealgorithm>**


**Example:**



```
oosrFrdygOh_sJpchBoog@y{lPyylHtcA_t\praM

```

This is a 4 point polygon in Ohio.


## Question


My question is, is there a way to import these into Google My Maps? I'd like to import these so that I can edit them.


","The answer is no, an encoded polygon cannot be imported. For a shape to be imported to My Maps, it must first be converted to a KML file.


BlueCollar laid out the first step, which is to use Google's encoding API to decode the encoded shapes into Lat/Lng pairs. 



```
var encodedPath = 'oosrFrdygOh_sJpchBoog@y{lPyylHtcA_t\praM';
var path = google.maps.geometry.encoding.decodePath(encodedPath);

```

For the example encoded value, which would return:



```
(39.91816, -85.33082),(38.013470000000005, -85.86915),(38.22091, -83.01270000000001),(39.76632, -83.02369),(39.91816, -85.33082)

```

Those pairs then need to be inserted into a KML file:



```
<?xml version=""1.0"" encoding=""UTF-8""?>
<kml xmlns=""http://www.opengis.net/kml/2.2"">
  <Placemark>
    <name>Test Shape</name>
    <Polygon>
      <extrude>1</extrude>
      <altitudeMode>relativeToGround</altitudeMode>
      <outerBoundaryIs>
        <LinearRing>
          <coordinates>
            -85.33082,39.91816
            -85.86915,38.01347
            -83.01270000000001,38.22091
            -83.02369,39.76632
            -85.33082,39.91816
          </coordinates>
        </LinearRing>
      </outerBoundaryIs>
    </Polygon>
  </Placemark>
</kml>

```

The KML file can then be imported.


"
"Convert auto property to full property
I often need to convert auto properties to full properties with a backing field so that I can implement `INotifyPropertyChanged`. It gets very tedious when a class has 50+ properties.



```
public string MyProperty { get; set;}

```

to



```
private string _myProperty;
public string MyProperty
{
    get
    {
        return _myProperty;
    }
    set
    {
        _myProperty = value;
        OnPropertyChanged(""MyProperty"");
    }
}

```

I was able to create a code snippet that creates a new property in the above format, but I don't know if it is possible to pull in an existing property's name and type and replace it.


I saw [kindofmagic](http://kindofmagic.codeplex.com/) but I really don't want to use arcane magic in my project.


[This question](https://stackoverflow.com/questions/10726918/resharper-convert-auto-property-to-full-property) explains how to do it in Resharper, but I don't have Resharper. I even downloaded the trial and still couldn't figure out how to do it.


Is there some way to do this with code snippets, macros, or even a free extension? It seems like it should be fairly straightforward. 


","If you have notepad++, you could do it via RegEx (quite ugly, but works)


Find what: `(public)\s+([a-zA-z0-9]+)\s+([a-zA-z0-9]+)\s*\{\s*+get;\s*set;\s*\}`


Replace with: `private \2 _\3\; \r\n \1 \2 \3 \r\n \{ \r\n get \{ return _\3\; \} \r\n set \{ _\3=value\; OnPropertyChanged\(\""\3\""\)\; \} \r\n \}`


Make sure ""Regular Expression"" is checked 


This is what the Find/Replace screen looks like:
![FindReplaceImg](https://i.stack.imgur.com/68Civ.png)


And it goes from 


![StartImg](https://i.stack.imgur.com/bc6pe.png)


To:


![EndImg](https://i.stack.imgur.com/onTi8.png)


Edit: Thanks to Britton, here is the Visual Studio equivalent:


Find: `public[^\S\r\n](.+)[^\S\r\n](\b(_\w+|[\w-[0-9_]]\w*)\b)[^\S\r\n]{[^\S\r\n]get;[‌​^\S\r\n]set;[^\S\r\n]}`


Replace: `private $1 _$2;\r\npublic $1 $2 {\r\nget\r\n{\r\nreturn _$2;\r\n}\r\nset\r\n{\r\n_$2 = value; OnPropertyChanged(""$2"");\r\n}\r\n}`


"
"Give access to JSP file/folder only for particular IP address
I made a CMS application using JSP and servlet. I am not using any kind of framework. The CMS has 2 parts:


1. Front end
2. Admin (Back end)


If I hosted it on `www.example.com`, for example, then all my frond end site will show for all users.


But `www.example.com/admin` must be accessible from a couple of IP addresses. Not from all the users.


I found these links:


- [link 1](http://www.jguru.com/faq/view.jsp?EID=445942)
- [link 2](http://middlewaremagic.com/jboss/?p=494)
- [Stack Overflow link](https://stackoverflow.com/questions/3381531/tomcat-restrict-access-by-ip-address)


Link 1 looks bit confusing and Link 2 looks good, but I am not using JBoss.


In the Stack Overflow link they did not mention how to use it.


","As far as I understand from your question you use the `Apache Tomcat` as a web server. In that case use [Remote Address Filter](http://tomcat.apache.org/tomcat-7.0-doc/config/filter.html#Remote_Address_Filter) to restrict access by IP address. It allows comparing IP of the requesting client with regular expressions to either allow or prevent the request based on the results of the comparison.


If you use [Tomcat 7](http://tomcat.apache.org/tomcat-7.0-doc/) you need use class [RemoteAddrFilter](http://tomcat.apache.org/tomcat-7.0-doc/api/org/apache/catalina/filters/RemoteAddrFilter.html) and define regular expressions for necessary IP in the application's configuration file `web.xml`:



```
<filter>
   <filter-name>Remote Address Filter</filter-name>
   <filter-class>org.apache.catalina.filters.RemoteAddrFilter</filter-class>
   <init-param>
      <param-name>allow</param-name> <!-- or deny -->
      <param-value>10\.10\.1[12]\..*</param-value> <!-- regexp for your ip adress -->
   </init-param>
</filter>
<filter-mapping>
   <filter-name>Remote Address Filter</filter-name>
   <url-pattern>*/admin</url-pattern> <!-- the url of your admin page -->
</filter-mapping>

```

You can use hardcoded specific IP address or regular expression patterns. But in some cases regular expressions gives you a lot of flexibility in validation of addresses.


And if you use [6](http://tomcat.apache.org/tomcat-6.0-doc/) or [5](http://tomcat.apache.org/tomcat-5.5-doc/) version of Tomcat you need use class [RemoteAddrValve](http://tomcat.apache.org/tomcat-6.0-doc/api/org/apache/catalina/valves/RemoteAddrValve.html) and define following line in Tomcat's configuration file `server.xml`:



```
<Valve className=”org.apache.catalina.valves.RemoteAddrValve” allow=”10\.10\.1[12]\..*”/> 

```

or



```
 <Valve className=""org.apache.catalina.valves.RemoteAddrValve""
    deny=""86.57.158.37, 213.117.195.*, 124.86.42.*"" /> 

```

More information about [using request filter valves](http://books.google.by/books?id=bgMKmsXVbTAC&pg=PA116&lpg=PA116&dq=tomcat+Remote+Address+Filter+example&source=bl&ots=tpGJMIhsWZ&sig=-he76l1CJEGwXPhayLzvpnMvqMc&sa=X&ei=KeowUMm7J43ItAbbkYDQBg&ved=0CCoQ6AEwBQ#v=onepage&q&f=false).


And interesting article about [securing the administrative web apps with Tomcat](http://thelowedown.wordpress.com/2010/08/07/tomcat6-securing-the-administrative-webapps/).


And by the way there's convenient to not deny requests from `localhost` for testing. So it makes sense to add `127\.0\.0\.1` to your allowable range of IP adresses. 


But don't forget that in some cases a proxy server can be used to get around the IP block. Apply also login authentication for better security.


"
"Nested query object mapping with pg-promise
I'm going through an example from [pg-promise](https://github.com/vitaly-t/pg-promise) for method [map](http://vitaly-t.github.io/pg-promise/Database.html#.map):



```
// Build a list of active users, each with the list of user events:
db.task(t => {
    return t.map('SELECT id FROM Users WHERE status = $1', ['active'], user => {
        return t.any('SELECT * FROM Events WHERE userId = $1', user.id)
            .then(events=> {
                user.events = events;
                return user;
            });
    }).then(t.batch);
})
    .then(data => {
        // success
    })
    .catch(error => {
        // error
    });

```

Let's say the `Event` entity has a one to many relationship with e.g. `Cars`, and I want to list all the `cars` connected to each `event`, how can I use the [map](http://vitaly-t.github.io/pg-promise/Database.html#.map) function when the object I want is more than one level deep?


The result I want could look something like this:



```
[{
    //This is a user
    id: 2,
    first_name: ""John"",
    last_name: ""Doe"",
    events: [{
        id: 4,
        type: 'type',
        cars: [{
            id: 4,
            brand: 'bmw'
        }]
    }]
}]

```

","I'm the author of [pg-promise](https://github.com/vitaly-t/pg-promise).




---



```
function getUsers(t) {
    return t.map('SELECT * FROM Users WHERE status = $1', ['active'], user => {
        return t.map('SELECT * FROM Events WHERE userId = $1', user.id, event => {
            return t.any('SELECT * FROM Cars WHERE eventId = $1', event.id)
                .then(cars => {
                    event.cars = cars;
                    return event;
                });
        })
            .then(t.batch) // settles array of requests for Cars (for each event)
            .then(events => {
                user.events = events;
                return user;
            });
    }).then(t.batch); // settles array of requests for Events (for each user)
}

```

And then using it:



```
db.task(getUsers)
    .then(users => {
        // users = an object tree of users->events->cars
    })
    .catch(error => {
        // error
    });

```

Method [map](http://vitaly-t.github.io/pg-promise/Database.html#map) simplifies mapping retrieved rows into something else, and since we map them into promises, those need to be settled, for which we use method [batch](http://vitaly-t.github.io/pg-promise/Task.html#batch). We do it for each internal array of requests for `cars`, and then on the top level - to settle the array of requests for `events`.


**UPDATE**


It may be easier to read and maintain, if you do the tree logic upside down:



```
function getUsers(t) {
    const getCars = eventId => t.any('SELECT * FROM Cars WHERE eventId = $1', eventId);

    const getEvents = userId => t.map('SELECT * FROM Events WHERE userId = $1', userId, event => {
        return getCars(event.id)
            .then(cars => {
                event.cars = cars;
                return event;
            });
    }).then(t.batch);

    return t.map('SELECT * FROM Users WHERE status = $1', ['active'], user => {
        return getEvents(user.id)
            .then(events => {
                user.events = events;
                return user;
            });
    }).then(t.batch);
}

```



---


There is also a faster, single-query approach that can be found here:


- [get JOIN table as array of results with PostgreSQL/NodeJS](https://stackoverflow.com/questions/39805736/get-join-table-as-array-of-results-with-postgresql-nodejs)


"
"Change IdentityServer4 Entity Framework table names
I am trying to change the default table names created by the PersistedGrantDb and ConfigurationDb for IdentityServer4 and have Entity Framework generate the correct SQL. For example; instead of the using the entity `IdentityServer4.EntityFramework.Entities.ApiResource` using the table `ApiResources`, I want the data to be mapped into a table named `mytesttable`


According to the [documentation](https://learn.microsoft.com/en-us/ef/core/modeling/relational/tables) this should be as simple as adding `ToTable` invocations for each entity that I want to remap in the `DBContext's` `OnModelCreating` method to override the default behaviour of TableName = EntityName. The problem is that this does indeed create a table `mytesttable` but the SQL created by Entity Framework at runtime still uses `ApiResources` in the query and consequently fails. 


The steps I've taken are I've created a `DBContext` that derives from IdentityServer's `ConfigurationDbContext` in order to be able to override `OnModelCreating` and customize the table names:



```
public class MyTestDbContext : ConfigurationDbContext
{
    public MyTestDbContext(DbContextOptions<ConfigurationDbContext> options, ConfigurationStoreOptions storeOptions) : base(options, storeOptions)
    { }


    protected override void OnModelCreating(ModelBuilder modelBuilder)
    {

        Console.WriteLine(""OnModelCreating invoking..."");

        base.OnModelCreating(modelBuilder);

        modelBuilder.Entity<IdentityServer4.EntityFramework.Entities.ApiResource>().ToTable(""mytesttable"");

        base.OnModelCreating(modelBuilder);

        Console.WriteLine(""...OnModelCreating invoked"");
    }
}

```

I've also implemented a `DesignTimeDbContextFactoryBase<MyTestDBContext>` class to manufacture the `MyTestDbContext` instance when invoked at design time via the `dotnet ef migrations` command line syntax. 


This works and an invocation of `dotnet ef migrations add InitialIdentityServerConfigurationDbMigration -c MyTestDbContext -o Data/Migrations/IdentityServer/MyTestContext` creates the initial migrations in my assembly. 


I then start the IdentityServer instance, invoking a test method from `Startup` that contains the following logic: 



```
private static void InitalizeDatabase(IApplicationBuilder app)
{
        using (var serviceScope = app.ApplicationServices.GetService<IServiceScopeFactory>().CreateScope())
         {

            serviceScope.ServiceProvider.GetRequiredService<PersistedGrantDbContext>().Database.Migrate();

            var context = serviceScope.ServiceProvider.GetRequiredService<MyTestDbContext>();
            context.Database.Migrate();

            /* Add some test data here... */
        }
}

```

and this happily wanders through and creates the necessary tables in my PostGRES database using `NpgSQL` provider, including the table named `mytesttable` in place of `ApiResources` for the entity `IdentityServer4.EntityFramework.Entities.ApiResource`. However, when I invoke a command from the IdentityServer instance, the SQL that is generated is still referencing `ApiResources` instead of `mytesttable`:



```
  Failed executing DbCommand (2ms) [Parameters=[], CommandType='Text', CommandTimeout='30']
  SELECT x.""Id"", x.""Description"", x.""DisplayName"", x.""Enabled"", x.""Name""
  FROM ""ApiResources"" AS x
  ORDER BY x.""Id""
  Npgsql.PostgresException (0x80004005): 42P01: relation ""ApiResources"" does not exist

```

Any assistance appreciated.


","There is two parts to this answer; firstly the table names need to be adjusted in IdentityServer's configuration so that it generates queries using the new table names. Secondly; the schema generated by entity framework needs to be amended so that it knows to create the differently named tables for the Identity Framework entities. Read on...


So, first up; the ability to change the table names used in the Entity Framework queries is exposed on `AddOperationalStore` and `AddConfigurationStore` methods that hang off the `AddIdentityServer` middleware method. The `options` argument of the delegate supplied to the configuration methods exposes the table names, for example: `options.{EntityName}.Name = {WhateverTableNameYouWantToUse}` - or `options.ApiResource.Name = mytesttable`. You can also override the schema on a per table basis as well by adjusting the `Schema` property.


The example below uses reflection to update all the entities to use table names prefixed with `idn_`, so `idn_ApiResources`, `idn_ApiScopes` etc:



```
services.AddIdentityServer()
.AddConfigurationStore(options => {
                // Loop through and rename each table to 'idn_{tablename}' - E.g. `idn_ApiResources`
                foreach(var p in options.GetType().GetProperties()) {
                if (p.PropertyType == typeof(IdentityServer4.EntityFramework.Options.TableConfiguration))
                {
                    object o = p.GetGetMethod().Invoke(options, null);
                    PropertyInfo q = o.GetType().GetProperty(""Name"");

                    string tableName = q.GetMethod.Invoke(o, null) as string;
                    o.GetType().GetProperty(""Name"").SetMethod.Invoke(o, new object[] { $""idn_{tableName}"" });
                    
                }
            }

         // Configure DB Context connection string and migrations assembly where migrations are stored  
            options.ConfigureDbContext = builder => builder.UseNpgsql(_configuration.GetConnectionString(""IDPDataDBConnectionString""),
                sql => sql.MigrationsAssembly(typeof(IdentityServer.Data.DbContexts.MyTestDbContext).GetTypeInfo().Assembly.GetName().Name));
}
.AddOperationalStore(options => { 

 // Copy and paste from AddConfigurationStore logic above.
}

```

The second part is to amend the schema generated by entity framework from the IdentityServer entities. To accomplish this you've got two choices; you can either derive from one of IdentityServer's supplied DBContexts; `ConfigurationDbContext` or `PeristedGrantDbContext` and then override the `OnModelCreating` method to remap each IdentityServer entity to the modified table name and then create your initial migration or updated migration as [documented here](https://learn.microsoft.com/en-us/ef/core/modeling/relational/tables) (Fluent Api syntax), **or** you can create the initial migration from the supplied IdentityServer DBContext's `ConfigurationDbContext` and `PersistedGrantDbContext` as per the tutorial [Adding Migrations](https://identityserver4.readthedocs.io/en/release/quickstarts/8_entity_framework.html) section, and then just do a find and replace with a text editor on all the table names and references to those table names in the created migration files.


Whichever method you choose you will still need to use the `dotnet ef migrations ...` command line syntax to create either the initial migration files as shown in [Adding Migrations](https://identityserver4.readthedocs.io/en/release/quickstarts/8_entity_framework.html) or a modified set with table changes and once you've done this, run your IdentityServer project and the schema will be created in the target database.


Note; `OnModelCreating` is invoked via the `dotnet ef migrations` syntax (aka at Design Time) and also at runtime if you call `Database.Migrate()` on your DBContext - E.g. `MyDbContextInstance.Database.Migrate()` (or the async equivalent method).


If you want to use a custom DBContext so you can customise `OnModelCreating`, you need to add a few design time classes which are used when you call `dotnet ef` from the command line and add the new context to `Startup`.


For the sake of completeness below is a hacky rough example where the context target is a PostGres database (use `UseSQLServer` in place of `UseNpgsql` or whatever your backing store is if it differs) and the connection string name is `IDPDataDBConnectionString` in the appsettings.json file and the custom DB context in this case is `MyTestDbContext` which derives from IdentityServer's `ConfigurationDbContext`.


Copy and paste the code, adjust the path to `appsettings.json` (or refactor) and then from the command line execute `dotnet ef migrations add InitialIdentityServerConfigurationDbMigration -c MyTestDbContext -o Data/Migrations/IdentityServer/ConfigurationDbCreatedWithMyTestContext` and you should see the Entity Framework generate the schema migration files using whatever overrides you've placed in `OnModelCreating` on your derived context. The example below also includes some `Console.WriteLine` invocations to make it easier to track what's going on.


Add this to `Startup`:



```
 services.AddDbContext<MyTestDbContext>(options =>
        {
            options.UseNpgsql(_configuration.GetConnectionString(""IDPDataDBConnectionString""));
        }); 

```

**Note** the use of the design time classes also allows you to separate your IdentityServer database migration files into a separate class library if you like. Make sure you target it in `Startup` if you do this (See [here](https://stackoverflow.com/questions/38705694/add-migration-with-different-assembly) for more info).



```
namespace MyIdentityServer.DataClassLibrary.DbContexts
{

public class MyTestDbContext : ConfigurationDbContext
{
    public MyTestDbContext(DbContextOptions<ConfigurationDbContext> options, ConfigurationStoreOptions storeOptions) : base(options, storeOptions)
    { }


    protected override void OnModelCreating(ModelBuilder modelBuilder)
    {

        Console.WriteLine(""OnModelCreating invoking..."");

        base.OnModelCreating(modelBuilder);

        // Map the entities to different tables here
        modelBuilder.Entity<IdentityServer4.EntityFramework.Entities.ApiResource>().ToTable(""mytesttable"");

        Console.WriteLine(""...OnModelCreating invoked"");
    }

}
public class MyTestContextDesignTimeFactory : DesignTimeDbContextFactoryBase<MyTestDbContext>
{

    public MyTestContextDesignTimeFactory()
        : base(""IDPDataDBConnectionString"", typeof(MyTestContextDesignTimeFactory).GetTypeInfo().Assembly.GetName().Name)
    {
    }

    protected override MyTestDbContext CreateNewInstance(DbContextOptions<MyTestDbContext> options)
    {
        var x = new DbContextOptions<ConfigurationDbContext>();

        Console.WriteLine(""Here we go..."");

        var optionsBuilder = newDbContextOptionsBuilder<ConfigurationDbContext>();

        optionsBuilder.UseNpgsql(""IDPDataDBConnectionString"", postGresOptions => postGresOptions.MigrationsAssembly(typeof(MyTestContextDesignTimeFactory).GetTypeInfo().Assembly.GetName().Name));

        DbContextOptions<ConfigurationDbContext> ops = optionsBuilder.Options;

        return new MyTestDbContext(ops, new ConfigurationStoreOptions());
    }
}




/* Enable these if you just want to host your data migrations in a separate assembly and use the IdentityServer supplied DbContexts 

public class ConfigurationContextDesignTimeFactory : DesignTimeDbContextFactoryBase<ConfigurationDbContext>
{

    public ConfigurationContextDesignTimeFactory()
        : base(""IDPDataDBConnectionString"", typeof(ConfigurationContextDesignTimeFactory).GetTypeInfo().Assembly.GetName().Name)
    {
    }

    protected override ConfigurationDbContext CreateNewInstance(DbContextOptions<ConfigurationDbContext> options)
    {
        return new ConfigurationDbContext(options, new ConfigurationStoreOptions());
    }
}

public class PersistedGrantContextDesignTimeFactory : DesignTimeDbContextFactoryBase<PersistedGrantDbContext>
{
    public PersistedGrantContextDesignTimeFactory()
        : base(""IDPDataDBConnectionString"", typeof(PersistedGrantContextDesignTimeFactory).GetTypeInfo().Assembly.GetName().Name)
    {
    }

    protected override PersistedGrantDbContext CreateNewInstance(DbContextOptions<PersistedGrantDbContext> options)
    {
        return new PersistedGrantDbContext(options, new OperationalStoreOptions());
    }
}
*/

public abstract class DesignTimeDbContextFactoryBase<TContext> :
IDesignTimeDbContextFactory<TContext> where TContext : DbContext
{
    protected string ConnectionStringName { get; }
    protected String MigrationsAssemblyName { get; }
    public DesignTimeDbContextFactoryBase(string connectionStringName, string migrationsAssemblyName)
    {
        ConnectionStringName = connectionStringName;
        MigrationsAssemblyName = migrationsAssemblyName;
    }

    public TContext CreateDbContext(string[] args)
    {
        return Create(
            Directory.GetCurrentDirectory(),
            Environment.GetEnvironmentVariable(""ASPNETCORE_ENVIRONMENT""),
            ConnectionStringName, MigrationsAssemblyName);
    }
    protected abstract TContext CreateNewInstance(
        DbContextOptions<TContext> options);

    public TContext CreateWithConnectionStringName(string connectionStringName, string migrationsAssemblyName)
    {
        var environmentName =
            Environment.GetEnvironmentVariable(
                ""ASPNETCORE_ENVIRONMENT"");

        var basePath = AppContext.BaseDirectory;

        return Create(basePath, environmentName, connectionStringName, migrationsAssemblyName);
    }

    private TContext Create(string basePath, string environmentName, string connectionStringName, string migrationsAssemblyName)
    {
        var builder = new ConfigurationBuilder()
            .SetBasePath(basePath)
            .AddJsonFile(@""c:\change\this\path\to\appsettings.json"")
            .AddJsonFile($""appsettings.{environmentName}.json"", true)
            .AddEnvironmentVariables();

        var config = builder.Build();

        var connstr = config.GetConnectionString(connectionStringName);

        if (String.IsNullOrWhiteSpace(connstr) == true)
        {
            throw new InvalidOperationException(
                ""Could not find a connection string named 'default'."");
        }
        else
        {
            return CreateWithConnectionString(connstr, migrationsAssemblyName);
        }
    }

    private TContext CreateWithConnectionString(string connectionString, string migrationsAssemblyName)
    {
        if (string.IsNullOrEmpty(connectionString))
            throw new ArgumentException(
         $""{nameof(connectionString)} is null or empty."",
         nameof(connectionString));

        var optionsBuilder =
             new DbContextOptionsBuilder<TContext>();

        Console.WriteLine(
            ""MyDesignTimeDbContextFactory.Create(string): Connection string: {0}"",
            connectionString);

        optionsBuilder.UseNpgsql(connectionString, postGresOptions => postGresOptions.MigrationsAssembly(migrationsAssemblyName));

        DbContextOptions<TContext> options = optionsBuilder.Options;

        Console.WriteLine(""Instancing...."");

        return CreateNewInstance(options);
    }
}

}

```

Side note; If you've already got a database with the IdentityServer tables in, you can just rename them manually ignoring Entity Framework migrations - the only bit you'll then need is the changes in `Startup` to `AddConfigurationStore` and `AddOperationalStore`.


"
"Keyboard and mouse stuttering on Ubuntu 18.04 with a new laptop
On my LG Gram 17"" (i7 8-th generation), I've got the problem in ubuntu.
Keyboard and mouse sometimes are starting to stutter, repeat once pressed keys and lag.


Did someone else face this issue?


","Found the workaround so far - setting the poll rate for keyboard and mouse to 1000 Hz (originally written he 125Hz by mistake):



```
sudo su
echo 1 > /sys/module/usbhid/parameters/kbpoll
echo 1 > /sys/module/usbhid/parameters/mousepoll

```

and replugged the usb hub.


As a permanent solution - these two parameters `usbhid.kbpoll=1 usbhid.mousepoll=1` should be added to the end of the `GRUB_CMDLINE_LINUX_DEFAULT` variable in the `/etc/default/grub` file.
So the variable should be looking like this:



```
GRUB_CMDLINE_LINUX_DEFAULT=""quiet splash usbhid.kbpoll=1 usbhid.mousepoll=1""

```

...and perform the:



```
update-grub

```

"
"Memory Leakage during removal of an entry in concurrent hashmap
Lets say we remove an entry `<k,v>` from `ConcurrentHashMap`. The remove operation in `ConcurrentHashMap` will clone the nodes preceding the node which has to be deleted. 


Now My question is how Java garbage collect the original nodes preceding the nodes which has to be deleted.


Lets say `1`--->`2`--->`3`--->`4`--->`5`--->`6` is a hashentry list which is maintained by `ConcurrentHashMap`. Now we want to remove `3`.


The following code is the code snippet of the remove method in Java `ConcurrentHashMap`



```
HashEntry newFirst = e.next;
for (HashEntry p = first; p != e; p = p.next) {
    newFirst = new HashEntry(p.key, p.hash, newFirst, p.value);
    tab[index]= newFirst;
}

```

- After 1st iteration 


`1`--->`2`--->`3`--->`4`--->`5`--->`6`


`1A`--->`2A`--->`4`--->`5`--->`6`


A new node `1A` will be created which will point to `4`. 
The Original Node `3` is still pointing to node `4`. 
Thus node `4` is pointed by 2 nodes
- After 2nd Iteration


`1`--->`2`--->`3`--->`4`--->`5`--->`6`


`1A`--->`2A`--->`4`--->`5`--->`6`


The node `4` is a pointer to two list now (`1`--->`2`--->`3`) and (`1A`--->`2A`). The original nodes preceding Node `4` (`1`--->`2`--->`3`) are never removed from list of hashentries.


Ain't this is the case of memory leakage. How the GC will collect (`1`--->`2`--->`3`) as they are still being referenced by `ConcurrentHashMap` ?


","I think you misread that code a bit. First of all, the loop looks like this:



```
HashEntry newFirst = e.next; // the element after the deleted one, in our case: 4
for (HashEntry p = first; p != e; p = p.next) {
    newFirst = new HashEntry(p.key, p.hash, newFirst, p.value);
}
tab[index]= newFirst; // this is outside the loop

```

Secondly, this is a singly-linked list, so the iteration goes like this:



```
Step 0: tab[index] --> 1 --> 2 --> 3 --> 4 --> 5 --> 6
          newFirst ----------------------^
Step 1: tab[index] --> 1 --> 2 --> 3 --> 4 --> 5 --> 6
          newFirst --------------> 1A ---^
Step 2: tab[index] --> 1 --> 2 --> 3 --> 4 --> 5 --> 6
          newFirst -------> 2A --> 1A ---^
Step 3: tab[index] --> 2A--> 1A--> 4 --> 5 --> 6
                 1 --> 2 --> 3 ----^

```

*(Step 0 is the initial state, steps 1 and 2 are the two iterations of the loop and step 3 is `tab[index] = newFirst`)*


**As you can see, after step 3 there's nothing pointing to `1`, so it's eligible for GC and as a result, so are `2` and `3`.**


**P.s.:** Also note how the order of `2A` and `1A` is reversed in the new list. This should have no practical impact unless you have way too many collisions, but in that case the collisions themselves are a far bigger problem than the slight timing inconsistency this introduces.


"
"Implementing polymorphic 'deep' function for traversals and folds
I'm using `lens` together with `xml-lens`. I'd like to make the following function more polymorphic, so that it also works for `Folds` and not only `Traversals`:



```
-- | Traverse a plated structure recursively, trying to match a fold at each level. Don't recurse
-- if the fold matches.
deep :: forall s a. Plated s => Traversal' s a -> Traversal' s a
deep f = let go :: Traversal' s a; go = cloneTraversal $ failing f (plate . go) in go

```

This function works like the [deep](http://hackage.haskell.org/package/hxt-9.3.1.4/docs/Control-Arrow-ArrowTree.html#v%3adeep) function from `hxt`. Is it possible to make it more polymorphic in the way I want?


","This one is rather tricky given the current publicly exposed API.


I took the liberty of expanding the type of `deepOf` to also support indexed folds and indexed traversals along the way as it was easier than not doing so, and makes the implementation be the full one we'd want to export from `lens`, anyways.


Let's import the parts of `lens` that we don't normally show users.



```
{-# LANGUAGE RankNTypes #-}

import Control.Applicative
import Control.Lens
import Control.Lens.Internal.Bazaar
import Control.Lens.Internal.Context
import Control.Lens.Traversal
import Control.Monad.State
import Data.Profunctor.Rep
import Data.Profunctor.Unsafe

```

We'll need a few internal combinators that we don't expose from `Control.Lens.Traversal` that are used to manipulate a `Traversal`/`Fold` as a `BazaarT` and collapse the answer back out.



```
pins :: (Bizarre p w, Corepresentable p) => w a b t -> [Corep p a]
pins = getConst #. bazaar (cotabulate $ \ra -> Const [ra])
{-# INLINE pins #-}

unsafeOuts :: (Bizarre p w, Corepresentable p) => w a b t -> [b] -> t
unsafeOuts = evalState `rmap` bazaar (cotabulate (\_ -> state (unconsWithDefault fakeVal)))
  where fakeVal = error ""unsafePartsOf': not enough elements were supplied""
{-# INLINE unsafeOuts #-}

unconsWithDefault :: a -> [a] -> (a,[a])
unconsWithDefault d []     = (d,[])
unconsWithDefault _ (x:xs) = (x,xs)
{-# INLINE unconsWithDefault #-}

```

Now that we have that, we build a proper version of `deep`.



```
-- |
-- @
-- 'deep' :: 'Plated' s => 'Fold' s a                 -> 'Fold' s a
-- 'deep' :: 'Plated' s => 'Traversal' s s a b        -> 'Traversal' s s a b
-- 'deep' :: 'Plated' s => 'IndexedFold' i s a        -> 'IndexedFold' i s a
-- 'deep' :: 'Plated' s => 'IndexedTraversal' s s a b -> 'Traversal' i s s a b
-- @
deep :: (Plated s, Conjoined p, Applicative f) => Traversing p f s s a b -> Over p f s s a b
deep = deepOf plate

-- |
-- @
-- 'deepOf' :: 'Fold s s'       -> 'Fold' s a                 -> 'Fold' s a
-- 'deepOf' :: 'Traversal' s s' -> 'Traversal' s s a b        -> 'Traversal' s s a b
-- 'deepOf' :: 'Fold s s'       -> 'IndexedFold' i s a        -> 'IndexedFold' i s a
-- 'deepOf' :: 'Traversal' s s' -> 'IndexedTraversal' s s a b -> 'Traversal' i s s a b
-- @
deepOf :: (Plated s, Conjoined p, Applicative f) => LensLike' f s s -> Traversing p f s s a b -> Over p f s s a b
deepOf r l pafb s = case pins b of
  [] -> r (deep l pafb) s
  xs -> unsafeOuts b <$> traverse (corep pafb) xs
  where b = l sell s

```

The guts of `deepOf` are very similar to the existing guts of `failing`, which you were rightly trying to use as your workhorse.



```
failing :: (Conjoined p, Applicative f) => Traversing p f s t a b -> Traversing p f s t a b -> Over p f s t a b
failing l r pafb s = case pins b of
  [] -> runBazaarT (r sell s) pafb
  xs -> unsafeOuts b <$> traverse (corep pafb) xs
  where b = l sell s

```

The only thing different is the [] case, where instead of falling over, what we do is run the whole nested `Traversal`.


I've merely typechecked this and not actually executed it, but it looks right to me.


Feel free to put in an issue on <http://github.com/ekmett/lens/issues> to add these combinators (or some bikeshedded renaming of them), they probably belong in the core `lens` API, lest code like this fall on users, while it is trivial to implement within the library itself.


This is the kind of code we try to write once, so that end users do not have to.


"
"Regex - Without Special Characters
I'm using regex to validate username



```
^[a-zA-Z]+\.[a-zA-Z]{4,10}^'

```

Unfortunately it doesn't affect if the the value contains special characters such as `!@#$%^&*)(':;`


I would glad to get some help for Regex that contains:


- Alphanumeric only (`a-zA-Z0-9`)
- Length between 4 - 10 characters.


","The conditions you specified do not conform to the regexp you posted.


the regexp you posted `^[a-zA-Z]+\.[a-zA-Z]{4,10}^` is erroneous I guess, because of the `^` in the end, **it will never be matched to any expression**, if you want to match with the `^` at the end of the expression, you need to escape it like this `\^`. but `^` alone means ""here is the start of the expression"", while `$` means ""here is the end of the expression"".


Even though, it denotes:


- It starts with alpha (at least 1).
- there must be a '.' period character.
- Now there must be at least 4 alphas.


The regexp you need is really is:



```
^[a-zA-Z0-9]{4,10}$

```

This says:


- It starts with alphanumeric.
- There can be minimum of 4 and maximum of 10 of alphanumeric.
- End of expression.


"
"How to type state and dispatch for useReducer - TypeScript and React
I am currently seeing this error with my current setup.



> 
> Type '({ team: string | null; } | { team: string | null; } | { ...; } | { ...; } | { ...; } | Dispatch<...>)[]' is missing the following properties from type 'State': teamts(2739)
>  index.d.ts(290, 9): The expected type comes from property 'value' which is declared here on type 'IntrinsicAttributes & ProviderProps'
> 
> 
> 


My code is below, if any more detail is required let me know.


actions.ts



```
export const setTeam = (team: string | null) => ({
  type: 'SET_TEAM',
  team,
});

```

reducer.ts



```
export const initialState = {
  team: null,
};

type State = {
  team: string | null;
};

export const GlobalContext = React.createContext<State | null>(null);

export const reducer = (state: State, action: any) => {
  switch (action.type) {
    case actionTypes.SET_TEAM:
      const team = action.team;
      return {
        ...state,
        team,
      };

    default:
      return state;
  }
};

```

App.tsx



```
const App = () => {
  const [state, dispatch] = React.useReducer(reducer, initialState);

  return ( 
    // error with state and dispatch here
    <GlobalContext.Provider value={[state, dispatch]}>
        ...App code
    </GlobalContext.Provider>
  )
}

```

Team.tsx



```
import { GlobalContext } from './reducers';
import { setTeam } from './actions';

const Team = () => {
   const [, dispatch] = React.useContext(GlobalContext);

   return <span onClick={() => dispatch(setTeam('LFC'))}>LFC</span>
}

```

","If you want to pass the state and the dispatch through context you have to type it on the context, you can go with just this line but if you want type safety read further



```
const GlobalContext = React.createContext<[State, React.Dispatch<any>]>([
  { team: null },
  () => {},
])

```

You can change the `<any>` inside `React.Dispatch` to your action types if you want type safety for actions, you would also need to type the action inside the reducer



```
enum TeamTypes {
  SET_TEAM = 'SET_TEAM',
  REMOVE_TEAM = 'REMOVE_TEAM',
}

type State = {
  team: string | null
}

export const initialState: State = {
  team: null,
}

type SetTeamAction = {
  type: typeof TeamTypes.SET_TEAM
  team: string
}

type RemoveTeamAction = {
  type: typeof TeamTypes.REMOVE_TEAM
}

type TeamActionTypes = SetTeamAction | RemoveTeamAction

export const setTeam = (team: string): TeamActionTypes => ({
  type: TeamTypes.SET_TEAM,
  team,
})

export const GlobalContext = React.createContext<
  [State, React.Dispatch<TeamActionTypes>]
>([{ team: null }, () => {}])

export const reducer = (state: State, action: TeamActionTypes): State => {
  switch (action.type) {
    case TeamTypes.SET_TEAM:
      const { team } = action
      return {
        ...state,
        team,
      }
    default:
      return state
  }
}

const App = () => {
  const [state, dispatch] = React.useReducer(reducer, initialState)

  return (
    <GlobalContext.Provider value={[state, dispatch]}></GlobalContext.Provider>
  )
}

```

"
"Java pattern for [j-\*]
Please help me with the pattern matching. I want to build a pattern which will match the word starting with `j-` or `c-` in the following in a string (Say for example)



```
[j-test] is a [c-test]'s name with [foo] and [bar]

```

The pattern needs to find `[j-test]` and `[c-test]` (brackets inclusive).


## What I have tried so far?



```
String template = ""[j-test] is a [c-test]'s name with [foo] and [bar]"";
Pattern patt = Pattern.compile(""\\[[*[j|c]\\-\\w\\-\\+\\d]+\\]"");
Matcher m = patt.matcher(template);
while (m.find()) {
    System.out.println(m.group());
}

```

And its giving output like



```
[j-test]
[c-test]
[foo]
[bar]

```

which is wrong. Please help me, thanks for your time on this thread.


","Inside a character class, you don't need to use alternation to match `j` or `c`. Character class itself means, match any single character from the ones inside it. So, `[jc]` itself will match either `j` or `c`.


Also, you don't need to match the pattern that is after `j-` or `c-`, as you are not bothered about them, as far as they start with `j-` or `c-`.


Simply use this pattern:



```
Pattern patt = Pattern.compile(""\\[[jc]-[^\\]]*\\]"");

```

To explain:



```
Pattern patt = Pattern.compile(""(?x)      ""   // Embedded flag for Pattern.COMMENT
                             + ""\\[       ""   // Match starting `[`
                             + ""    [jc]  ""     // Match j or c
                             + ""    -     ""     // then a hyphen
                             + ""    [^    ""     // A negated character class
                             + ""       \\]""        // Match any character except ] 
                             + ""    ]*    ""     // 0 or more times
                             + ""\\]       ""); // till the closing ]

```

Using `(?x)` flag in the regex, ignores the whitespaces. It is often helpful, to write readable regexes.


"
"How do Delphi XE4 and Xamarin Monotouch differ in the way they target iOS?
I was intrigued to discover that [Delphi XE4](http://www.embarcadero.com/products/delphi) now targets iOS. What are the main differences between the way Monotouch and Delphi XE-4 target the iOS platform?


If I understand correctly, Xamarin Monotouch targets iOS development by providing C# wrappers for the iPhone Cocoa Touch libraries. You still use the XCode Interface Builder to generate the user interface and therefore you need some understanding of iOS concepts like UIViews or Outlets and Actions which are quite different from say, using C# to develop for WinForms or WPF. 


Does the Delphi-XE4 approach work the same way? Does it provide Delphi wrappers for the iPhone libraries?


Also, from an interface building perspective, does its interface builder work similar to the XCode interface builder along with the identical native controls and concepts? How different is it for a Delphi developer to target iOS rather than Windows?


","The Embarcadero approach is quite different from what you describe for Xamarin. 


The Xamarin approach, as you describe it, uses the platform native framework. In contrast, the Delphi iOS solution is based on the FireMonkey cross-platform framework, FMX. The FMX framework does not use native controls. Instead the framework emulates the look and feel of the native controls.


When you develop an FMX mobile app for iOS, it's a straight-forward re-compile to make that app run on Android, the up-coming mobile platform that is coming with XE5. Embarcadero view this as a significant benefit to their approach. It's up to the developer to decide whether or not the common source benefit of the FMX approach outweighs the non-native UI widgets.


All that said, it is perfectly feasible to use the Delphi iOS compiler to generate Cocoa applications. But that would require extra effort and most likely third party components.


The FMX platform is also available on Windows and Mac. So you can develop for those platforms using the same framework as your mobile apps. Obviously the mobile apps will be quite different, but if you master the concepts of FMX on one platform then they transfer easily to other platforms.


"
"How to limit number of attempts someone has to provide username/password credentials for a webpage with htpasswd?
I've enabled htpasswd on a webpage. Is there a way I can limit the number of attempts a user has to enter the correct username/pass credentials before they have to do a hard refresh of the page?


","Not with basic authentication, you'll need to go to a different authentication module in order to do it. I was just reading up on mod\_auth\_pamacea which allows you to use the PAM libraries and all of their goodness with `.htpasswd`. You could setup a separate service for your http authentication and then use pam\_tally.so to limit the number of login attempts. Here are a couple of links for your reading pleasure:


[The Linux-PAM System Administrators' Guide](http://debian.securedservers.com/kernel/pub/linux/libs/pam/Linux-PAM-html/Linux-PAM_SAG.html ""The Linux-PAM System Administrators' Guide"")


[mod\_auth\_pamacea Documentation](http://www.columbia.edu/acis/dev/projects/mod_auth_pamacea/doc.shtml)


I'm sure there are other auth modules that will allow the same functionality, to me this is the closest you'll come to `mod_auth_basic`.


"
"Conditional upsert in MongoDB
I've answered this question on LinkedIn and I thought it's something useful and interesting to share. The question was:


""Suppose we have documents like `{_id: ..., data: ..., timestamp: ...}`.


Is there any way to write update criteria which will satisfy following rules:


1 If there is no documents with following `_id` then insert this document;


2 If there is exists document with following `_id` then


2.1 If new timestamp greater then stored timestamp then update `data`;


2.2 Otherwise do nothing""


","Solution below should do the trick, you just need to ignore dup key errors. Example is given in Mongo shell:



```
> var lastUpdateTime = ISODate(""2013-09-10"")
> var newUpdateTime = ISODate(""2013-09-12"")
>
> lastUpdateTime
ISODate(""2013-09-10T00:00:00Z"")
> newUpdateTime
ISODate(""2013-09-12T00:00:00Z"")
>
> var id = new ObjectId()
> id
ObjectId(""52310502f3bf4823f81e7fc9"")
>
> // collection is empty, first update will do insert:
> db.testcol.update(
... {""_id"" : id, ""ts"" : { $lt : lastUpdateTime } },
... { $set: { ts: lastUpdateTime, data: 123 } },
... { upsert: true, multi: false }
... );
>
> db.testcol.find()
{ ""_id"" : ObjectId(""52310502f3bf4823f81e7fc9""), ""data"" : 123, ""ts"" : ISODate(""2013-09-10T00:00:00Z"") }
>
> // try one more time to check that nothing happens (due to error):
> db.testcol.update(
... {""_id"" : id, ""ts"" : { $lt : lastUpdateTime } },
... { $set: { ts: lastUpdateTime, data: 123 } },
... { upsert: true, multi: false }
... );
E11000 duplicate key error index: test.testcol.$_id_ dup key: { : ObjectId('52310502f3bf4823f81e7fc9') }
>
> var tooOldToUpdate = ISODate(""2013-09-09"")
>
> // update does not happen because query condition does not match
> // and mongo tries to insert with the same id (and fails with dup again):
> db.testcol.update(
... {""_id"" : id, ""ts"" : { $lt : tooOldToUpdate } },
... { $set: { ts: tooOldToUpdate, data: 999 } },
... { upsert: true, multi: false }
... );
E11000 duplicate key error index: test.testcol.$_id_ dup key: { : ObjectId('52310502f3bf4823f81e7fc9') }
>
> // now query cond actually matches, so update rather than insert happens which works
> // as expected:
> db.testcol.update(
... {""_id"" : id, ""ts"" : { $lt : newUpdateTime } },
... { $set: { ts: newUpdateTime, data: 999 } },
... { upsert: true, multi: false }
... );
>
> // check that everything worked:
> db.testcol.find()
{ ""_id"" : ObjectId(""52310502f3bf4823f81e7fc9""), ""data"" : 999, ""ts"" : ISODate(""2013-09-12T00:00:00Z"") }
>

```

The only annoying part are those errors, but they are cheap and safe.


"
"Session in ASP.NET Core MVC
I am migrating an ASP.NET MVC application to ASP.NET Core MVC.


In ASP.NET MVC, I am using some `Session` variables to store or get values like:



```
Session[""UserID""] = user.UserName;
Session[""Role""] = role[0].ROLE_DESCRIPTION;
ViewData[""LEVEL""] = Session[""LEVEL""];

```

But after converting, I get an error:



> 
> The name 'Session' does not exist in the current context
> 
> 
> 


Also, I need to have replacement for this as well in the `.cshtml` file:



```
var UserName = ""@Session[""Name""]"";

```

Is there any other method in ASP.NET Core MVC to store or get values on those variables without changing the behavior?


","You need to add session middleware in your configuration file



```
public void ConfigureServices(IServiceCollection services)  
{  
     //........
    services.AddDistributedMemoryCache();  
    services.AddSession(options => {  
        options.IdleTimeout = TimeSpan.FromMinutes(1);//You can set Time   
    });  
    services.AddMvc();  
} 

public void ConfigureServices(IServiceCollection services)
{
    //......
    app.UseSession();
    //......
}

```

Then in your controller, you can



```
//set session 

HttpContext.Session.SetString(SessionName, ""Jarvik"");  
HttpContext.Session.SetInt32(SessionAge, 24);

//get session

ViewBag.Name = HttpContext.Session.GetString(SessionName);  
ViewBag.Age = HttpContext.Session.GetInt32(SessionAge); 
  

```

"
"MetaPhone Functions (like SoundEx) functions and use in R?
I am wanting to use [MetaPhone](http://en.wikipedia.org/wiki/Metaphone), [Double Metaphone](http://en.wikipedia.org/wiki/Metaphone#Double_Metaphone), [Caverphone](http://en.wikipedia.org/wiki/Caverphone), MetaPhone3, SoundEx, and if anyone has done it yet [NameX](http://namevariants.co.uk/introduction.aspx#section-algorithm) functions within 'R' so I can categorize and summarize like values to minimize data cleansing operations prior to analysis.


I am fully aware that each algorithm has its own strengths and weakness and *would highly **prefer not to use SoundEx*** but it still might work if I cannot find alternatives; as like [mentioned in this post](https://genealogy.stackexchange.com/questions/2026/how-can-i-identify-all-the-possible-alternatives-for-a-surname/2033#2033) *Harper* would match with any of a list of unrelated names under SoundEx but should not in Metaphone for better result matching.


Though I am not sure which would serve my purposes best while still preserving some flexibility so that is the reason I want to take a stab with several of them as well as before looking at the values generate a table like the following.


![enter image description here](https://i.stack.imgur.com/QxcXG.jpg)


[Table Source Link](http://www.sound-ex.com/alternative_double.htm)


Surnames are not the subject of my initial analysis but think it is a good example as I want to effectively consider all like 'sounding' words treated as the same value is really what I am trying to do with a simply call something as values are evaluated. 


Some things I have already looked at:


- I know that a C package could be written and called with [RCpp](https://stackoverflow.com/questions/23488203/call-r-function-from-c-wrapper?rq=1), and there are even C solutions for [SoundEx on SE](https://stackoverflow.com/questions/1626217/soundex-algorithm-in-python-homework-help-request), but I have not written an R package before and looking to avoid re-inventing the wheel if there is a simpler way to do it directly in R or a package exists that has the function available?
- I am aware that the [RecordLinkage](http://journal.r-project.org/archive/2010-2/RJournal_2010-2_Sariyar+Borg.pdf) and now [stringdist](http://www.r-bloggers.com/stringdist-0-8-now-with-soundex/) package have a SoundEx function, but not any form of a MetaPhone function.



> 
> So I am specifically looking for an answer is to how do a MetaPhone / Caverphone function in R and know the ""Value"" so I can group data values by them?
> 
> 
> 


*The additional caveat is I am still consider my self pretty new to R as I am not a daily user of it.*


","The algorithm is pretty straightforward but I, too, could not find an existing R package. If you really need to do this work in R, one short-term option is to install the python module `metaphone` (`pip install metaphone`) then use the `rPython` bridge to use it in R:



```
library(rPython)

python.exec(""from metaphone import doublemetaphone"")
python.call(""doublemetaphone"", ""architect"")
[1] ""ARKTKT"" """"

```

It's not the most elegant solution, but it gets you metaphone operations in R.


The Apache Commons has a [codec library](http://commons.apache.org/proper/commons-codec/) that also implements the metaphone algorithms:



```
library(rJava)

.jinit() # need to have commons-codec-1.10.jar in your CLASSPATH

mp <- .jnew(""org.apache.commons.codec.language.Metaphone"")
.jcall(mp,""S"",""metaphone"", ""architect"")
[1] ""ARXT""

```

You can make the above `.jcall` an R function and use it like any other R function:



```
metaphone <- function(x) {
  .jcall(mp,""S"",""metaphone"", x)  
}

sapply(c(""abridgement"", ""stupendous""), metaphone)

## abridgement  stupendous 
##      ""ABRJ""      ""STPN""

```

The java interface may be more compatible across platforms, too.


Here's a more complete view of using the java interface:



```
library(rJava)

.jinit()

mp <- .jnew(""org.apache.commons.codec.language.Metaphone"")
dmp <- .jnew(""org.apache.commons.codec.language.DoubleMetaphone"")

metaphone <- function(x) {
  .jcall(mp,""S"",""metaphone"", x)  
}

double_metaphone <- function(x) {
  .jcall(dmp,""S"",""doubleMetaphone"", x)  
}

words <- c('Catherine', 'Katherine', 'Katarina', 'Johnathan', 
           'Jonathan', 'John', 'Teresa', 'Theresa', 'Smith', 
           'Smyth', 'Jessica', 'Joshua')

data.frame(metaphone=sapply(words, metaphone),
           double=sapply(words, double_metaphone))

##           metaphone double
## Catherine      K0RN   K0RN
## Katherine      K0RN   K0RN
## Katarina       KTRN   KTRN
## Johnathan      JN0N   JN0N
## Jonathan       JN0N   JN0N
## John             JN     JN
## Teresa          TRS    TRS
## Theresa         0RS    0RS
## Smith           SM0    SM0
## Smyth           SM0    SM0
## Jessica         JSK    JSK
## Joshua           JX     JX

```

"
"How do I transpose music chords using JavaScript?
I was wondering how would one create a javascript function for transposing music chords.


Since I don't expect everyone to be a musician here, I'll try to explain how it works in music theory. I hope I don't forget something. If yes, musicians, please, correct me.


1) The simple chords


The simple chords are almost as simple as an alphabet and it goes like this:



> 
> C, C#, D, D#, E, F, F#, G, G#, A, A# B
> 
> 
> 


From B it loops all over again to C. Therefore, If the original chord is `E` and we want to transpose +1, the resulting chord is `F`. If we transpose +4, the resulting chord is `G#`.


2) Expanded chords


They work almost like the simple chords, but contain a few more characters, which can safely be ignored when transposing. For example:



> 
> Cmi, C#7, Dsus7, Emi, Fsus4, F#mi, G ...
> 
> 
> 


So again, as with the simple chords, if we transpose `Dsus7` + 3 = `Fsus7`


3) Non-root bass tone


A problem arises when the bass plays a different tone than the chord root tone. This is marked by a slash after the chord and also needs to be transposed. Examples:



> 
> C/G, Dmi/A, F#sus7/A#
> 
> 
> 


As with examples 1 and 2, everything is the same, but the part after the slash needs transpose too, therefore:


`C/G` + 5 = `F/C`


`F#sus7/A#` + 1 = `Gsus7/B`


I think this should be all, unless I forgot something.


So basically, imagine you have a javascript variable called `chord` and the transpose value `transpose`. What code would transpose the chord?


Example:



```
var chord = 'F#sus7/C#';
var transpose = 3; // remember this value also may be negative, like ""-4""
... code here ...
var result; // expected result = 'Asus7/E';

```

","How about a little somethin' like this:



```
function transposeChord(chord, amount) {
  var scale = [""C"", ""C#"", ""D"", ""D#"", ""E"", ""F"", ""F#"", ""G"", ""G#"", ""A"", ""A#"", ""B""];
  return chord.replace(/[CDEFGAB]#?/g,
                       function(match) {
                         var i = (scale.indexOf(match) + amount) % scale.length;
                         return scale[ i < 0 ? i + scale.length : i ];
                       });
}

alert(transposeChord(""Dm7/G"", 2)); // gives ""Em7/A""
alert(transposeChord(""Fmaj9#11"", -23)); // gives ""F#maj9#11""

```

Note that I threw in the ""F#maj9#11"" example just to give you more to think about with regard to what makes up a valid chord name: you may find a ""#"" sharp symbol that doesn't follow a letter (in this case it belongs to the ""11"").


And, obviously, my function only understands sharps, not flats, and doesn't understand keys, so, e.g., `transposeChord(""C/E"", 1)` will give ""C#/F"" when it really should be ""C#/E#"".


"
"Azure - difference between service endpoint and private endpoint in simple terms
I am not able to understand the difference between service endpoints and private endpoints clearly.


Need help preferably with an example.


","You might have seen this in the [Private Link FAQ](https://learn.microsoft.com/en-us/azure/private-link/private-link-faq#what-is-the-difference-between-service-endpoints-and-private-endpoints-):



> 
> - Private Endpoints grant network access to specific resources behind a given service providing granular segmentation. Traffic can reach the service resource from on premises without using public endpoints.
> - A Service Endpoint remains a publicly routable IP address. A Private Endpoint is a private IP in the address space of the virtual network where the private endpoint is configured.
> 
> 
> 


For simplicity, let's take the view of a VM in a VNET connecting to a storage account in the same subscription and same Azure region. There are three ways to connect.


[![Connect to PaaS Services](https://i.stack.imgur.com/rHa0q.png)](https://i.stack.imgur.com/rHa0q.png)


1. Default


By default all traffic goes against the public endpoint of the storage account. Source IP of the traffic is the Public IP of the VM.
2. Service Endpoints


Traffic is still directed against the public endpoint of the storage account but the source IP has changed to the private IP of the VM. In fact, the traffic is also using the VNET and Subnet as source in the network dataframe.
3. Private Endpoints


The PaaS service now gets a virtual network interface inside the subnet and traffic from the VM to the storage account is now directed against the private IP address.


By far the best collection of useful information around Private Link that I have seen on the web is in this repository: <https://github.com/dmauser/PrivateLink>


You can also find some examples here: <https://jeffbrown.tech/azure-private-service-endpoint/>


"
"How to use org.apache.felix.scr annotations for a bundle from scratch?
I've started developing a bundle for apache felix and did so with the ops4j pax plugin.


I've created the project structure with `pax-create-project` and in there did the normal `pax-create-bundle`. Then you get the initial project structure for building the whole thing with maven. The important part here is, that your bundle has it's own pom (`bundlename/pom.xml`) and bnd file (`bundlename/osgi.bnd`) but the configuration for `maven-bundle-plugin` is already provided under `poms/compiled/pom.xml`. 
The bundle metadata is configured under `poms/compiled/pom.xml` but the standard activator is configured under the above mentioned `osgi.bnd` file. The default in the bnd file is `Bundle-Activator: ${bundle.namespace}.internal.ExampleActivator`.


Now I wanted to start using annotations from `org.apache.felix.scr` so I included this in the bundles own pom under dependencies:



```
<dependency>
  <groupId>org.apache.felix</groupId>
  <artifactId>org.apache.felix.scr</artifactId>
  <version>1.6.0</version>
</dependency>

```

And I created my interface for the service:



```
package namespace;

public interface Sample {
  void sayHello();
}

```

As well as the implementation with the annotations from `org.apache.felix.scr`:



```
package namespace.internal;

import org.apache.felix.scr.annotations.Activate;
import org.apache.felix.scr.annotations.Component;
import org.apache.felix.scr.annotations.Deactivate;
import namespace.Sample;

@Component
@Service
public class SampleImpl implements Sample{

  @Activate
  void start(){
    System.out.println(""Started SampleImpl."");
  }

  @Deactivate
  void stop(){
    System.out.println(""Stopped SampleImpl."");
  }

  @Override
  public void sayHello() {
    System.out.println(""Hello!"");
  }
}

```

What do I need to place in the `osgi.bnd` file in order to start the components? I had to remove the default entry mentioned above because I`m not using the BundleActivator any more.


Now here is the question: What do I need to do to any of the above mentioned files in order to let the `mvn clean install pax:provision` start up a felix environment that starts my bundle and let me see *any* of the printouts? I have already looked at the log output of the Apache Log Service which I'm starting too and it just tells me:



```
2012.09.28 16:52:44 INFO - Bundle: namespace - BundleEvent STARTED
2012.09.28 16:52:44 INFO - Bundle: namespace - BundleEvent RESOLVED

```

Any hints are very well appreciated. Or links, I've searched for more then hours to find a complete guide to use the maven plugins for scr and bundle together. 


Greetings,
Kjellski 


","This is a modification of my coderwall writeup on this topic: [http://coderwall.com/p/q39uxq](http://coderwall.com/p/q39uxq ""http://coderwall.com/kjellski"") 


First configure the general compilation configuration of the submodules from poms/compiled/pom.xml. Add the plugin right where the maven-bundle-plugin already is:



```
<plugin>
    <groupId>org.apache.felix</groupId>
    <artifactId>maven-scr-plugin</artifactId>
    <version><timeofwriting:1.8.0></version>
    <executions>
        <execution>
            <id>generate-scr-scrdescriptor</id>
            <goals>
                <goal>scr</goal>
            </goals>
        </execution>
    </executions>
</plugin>

```

In the same file, under poms/compiled/pom.xml, look at the instructions for maven-bundle-plugin and add right under the <\_include> statement:



```
<Service-Component>
    OSGI-INF/serviceComponents.xml
</Service-Component>

```

Now depend on the annotations in the subprojects in order to use them:



```
<dependency>
    <groupId>org.apache.felix</groupId>
    <artifactId>org.apache.felix.scr</artifactId>
    <version><timeofwriting:1.6.0></version>
</dependency>
<dependency>
    <groupId>org.apache.felix</groupId>
    <artifactId>
        org.apache.felix.scr.annotations
    </artifactId>
    <version><timeofwriting:1.7.0></version>
</dependency>

```

Now define your **components** with annotations *@Component* and *@Service*. When you want none to be included, like in any api bundle, just insert ""Service-Component: \*"" in the osgi.bnd file. This will cause the bundle plugin to stop complaining about the missing OSGI.INF for the project that doesn't contain any.


Hope this helps somebody ;)


Cheers, 
Kjellski


"
"Best way of returning differences of two json files programatically
I have two json files and I would like to get a json containing the differences. **It is important** that **only** the actual differences of **content** should be shown, regardless of changing the order of some elements. 


What would be the best way to do that? I am searching for a solution as efficient as possible, since `json`s may contain lots of data, and users need jobs to be done as quick as possible.


Note: The `json`s might contain data encoded at different depths. Any programming language is ok, but I would prefer an answer that could easily be implemented in php.


","Basically, what you want is something similar to `array_diff_assoc`, but applied to json objects, and recursive.


The `array_diff` functions are not recursive because of reference issues: it is possible to assign a reference of an array to an entry of that array, making the array infinitely recursive. I don't think it is possible to get the same situation with a json object, thus making a recursive function safe.


Let's suppose that you wish to compute the difference between object A and B, and have the result in object C.
The principle is to loop over each field of A (a `foreach` should do), and when:


- no such field exist in B, copy it over C.
- a similar field exist in B, put in C the result of the difference of A field with B field, which is a recursive call on the diff function with those field as parameter, as well as a fresh object for the result.


The ordering of A should be respected.


"
"Access a subversion repository on a Windows share from Linux
For a long time we put SVN repositories on a Windows share (worked great as only one person was working on one project). 


Now, we are moving dev machines away from Windows to Ubuntu and we need to access those repositories from Ubuntu.


I mounted the svn repositories share like this:



```
//win_serv/svn /mnt/svn cifs credentials=/home/u/.smbcreds,iocharset=utf8,sec=ntlm 0 0

```

I managed to change repository URL to `/mnt/svn/some_project`. However, trying to do `svn up` causes an error:



```
svn: E200033: Another process is blocking the working copy database, or the underlying filesystem does not support file locking; if the working copy is on a network filesystem, make sure file locking has been enabled on the file server
svn: E200033: sqlite[S5]: database is locked
svn: E200033: Additional errors:
svn: E200033: sqlite[S5]: database is locked

```

Am I doing something wrong (besides the 'svn repository on windows share' fact)?


Is it even possible to access an svn repo on a Windows share from Linux?


","I would encourage you not to do this. Either host a SVN server on Windows or on Linux and access the server using svn+ssh protocol. Do not try and develop some kludged method for accessing your SVN repositories.


There are a number of issues you're going to run into that are SVN and SQLite related that you'll run into that require various mounting options such as NOBRL to CIFS.


I'd really encourage you to avoid all this and decide upfront where you want the SVN repositories to reside, and then invest the time in setting up an actual SVN sever on that system vs. proceeding down the path that you're already trying.


Also invest some time in researching the various setup options in the [SVN book](http://svnbook.red-bean.com/), it's free, and spells out the best ways to do what you're doing so you don't destroy or jeopardize losing your data.


"
"data auditing in Cassandra
How to implement auditing for cassandra data?
I am looking for a open source option.


Are there any features of cassandra that help with auditing?
Can I use triggers to log the records into a table? I followed [Triggers](https://github.com/apache/cassandra/tree/trunk/examples/triggers) example and was able to get a record inserted into `triggers_log` table when the updates occur on another table.
But not sure how do I capture the `user/session` details that triggered the update. I have From `CQLSH` terminal, create `users` and `trigger_log table`



```

create table AUDIT_LOG ( 
       transaction_id int,
       entries map<text, text>,  --> to capture the modifications done to the tables
       user varchar,  //authenticated user
       time timestamp, 
       primary key(transaction_id));

```


```

CREATE TABLE users (
  user_id int PRIMARY KEY,
  fname text,
  lname text
);

```

Define the trigger on users table using `CREATE TRIGGER` syntax from `cqlsh`


Below code so far. 



```
public class AuditTrigger implements ITrigger {

    @Override
    public Collection<RowMutation> augment(ByteBuffer key, ColumnFamily update) {

        List<RowMutation> mutations = new ArrayList<RowMutation>();
        for (Column column : update) {
            if (column.value().remaining() > 0) {
                RowMutation mutation = new RowMutation(""mykeyspace"", key);
           //What do I need here to capture the updates to users 
           //table and log the updates into various columns of audit_log
                mutations.add(mutation);
            }
        }
        return mutations;
    }
}

```

If triggers is not the correct approach (any spring AOP approach?), please suggest alternatives. I also tried [Cassandra vs logging activity](https://stackoverflow.com/questions/9604554/cassandra-vs-logging-activity) solution but it does not print the sql executed, authenticated user information.


","Unfortunately at this time, Triggers cannot be used as what you need is the ClientState which contains the user information and is not passed to Triggers.


There are 2 approaches I can think of.(You will need to look at the Cassandra code base for better understanding these approaches)


One approach is AOP i.e to add an agent which would AOP and start Cassandra with the Agent. The class that will need to be pointcut is the QueryProcessor#processStatement method. The call to this method will have the prepared statement and the QueryState as parameters. From the PreparedStatement you can identify the intention of the user. QueryState.getClientState will return the ClientState which is where the user information resides.


The other approach involves custom authenticators and authorizers. Configuring this in Cassandra is described here.


<http://www.datastax.com/documentation/cassandra/2.0/cassandra/security/secure_about_native_authenticate_c.html>


You can have a custom authorizer extending the AllowAllAuthorizer(this will disable permission caching). Whenever you get an authorize request on the Authorizer you can log it. The downside of this approach is that you do not know what the user intends to do with the table, only that he is request some authorization on it. Permission is the one which contains what he wants to do with the table, but it is not passed on to the authorizer.


If you decide on either of these approaches, you are free to post followups if you need more detail.


"
"Constructor to specify zero-initialization of all builtin members?
Is there a simpler way for a class's constructor to specify that all members of built-in type should be zero-initialized? 


This code snippet came up in another post:



```
struct Money
{
    double amountP, amountG, totalChange;
    int twenty, ten, five, one, change;
    int quarter, dime, nickel, penny;

    void foo();
    Money()  {}
};

```

and it turned out that the problem was that the object was instantiated via `Money mc;` and the variables were uninitialized.


The recommended solution was to add the following constructor:



```
Money::Money()
  : amountP(), amountG(), totalChange(),
    twenty(), ten(), five(), one(), change()
    quarter(), dime(), nickel(), penny()
{
}

```

However, this is ugly and not maintenance-friendly. It would be easy to add another member variable and forget to add it to the long list in the constructor, perhaps causing a hard-to-find bug months down the track when the uninitialized variable suddenly stops getting `0` by chance.


","You can use a subobject to initialize en masse. A member works, but then you need to qualify all access. So inheritance is better:



```
struct MoneyData
{
    double amountP, amountG, totalChange;
    int twenty, ten, five, one, change;
    int quarter, dime, nickel, penny;
};

struct Money : MoneyData
{
    void foo();
    Money() : MoneyData() {} /* value initialize the base subobject */
};

```

Demonstration (placement `new` is used to make sure the memory is non-zero before object creation): <http://ideone.com/P1nxN6>


Contrast with a slight variation on the code in the question: <http://ideone.com/n4lOdj>


In both of the above demos, `double` members are removed to avoid possible invalid/NaN encodings.


"
"Proving that ""int \*p = malloc(1); p[0]"" is undefined behavior
I'm trying to convince (citing the specific parts of the C99 standard) a colleague that the following is undefined behavior:



```
int *p = malloc(1);
p[0] = 0;

```

But I cannot find the specific parts in the standard which clearly ensure that this is undefined. **I'm looking specifically for the logical steps in the standard which lead from these lines to the conclusion: *undefined behavior***. Is it the conversion from `void *` to `int *` in the first line? The assignment in the second line?


The only relevant part I can find about `malloc` is that it returns a suitably aligned pointer (7.20.3):



> 
> The pointer returned if the allocation succeeds is suitably aligned so that it may be assigned to a pointer to any type of object and then used to access such an object or an array of such objects in the space allocated (...)
> 
> 
> 


I tried grepping for *space* in the norm, but there's too much noise due to *white space* and other lexical issues.


","Adding from **7.20.3.3 The malloc function** to your quote:



> 
> The malloc function allocates space for an object whose **size** is
>  specified by size and whose value is indeterminate.  
> 
>  The malloc function returns either a **null pointer** or a pointer to the allocated space.
> 
> 
> 


So there are 2 possible sources of undefined behavior, one is overwriting (Size of int is guaranteed to be 16 bits or more, but you are allocating just 1 byte which is 8 bit on almost all systems) the buffer, and second is possible de-referencing of null-pointer.


From **6.5.2.1 Array subscripting**, `p[0] = 0` is equivalent to `*p = 0`. Type of `*p` is an `int` so it would fill `sizeof(*p) * CHAR_BIT` bits with `0` which may not all belong to the allocated buffer causing the UB.


There is no undefined behavior in first line of code (assignation), UB if any would be in second line (de-referencing).


But on machines where `CHAR_BIT` is large and `sizeof(int)` is `1`, this will be well defined behavior for the cases when `malloc` doesn't `return` a null pointer.


"
"Requests-HTML scrape tag image url (Requests-HTML, python)
html to attempt to extract the cpu image from the following [webpage](https://au.pcpartpicker.com/product/jLF48d) i have identified that the image url is in a tag with the class name item: [Chrome inspect tool](https://i.stack.imgur.com/5spll.png)


Here is my code



```
from requests_html import HTMLSession
session = HTMLSession()

r = session.get('https://au.pcpartpicker.com/product/jLF48d')

about = r.html.find('.item')

print(about)

```

This prints 



> 
> Element 'a' class=('item',) onclick='show\_gallery(0, carousel\_images);return false;'
> 
> 
> 


However when I change the print statement to:



```
print(about.absolute_links)

```

I get the following error:



> 
> AttributeError: 'list' object has no attribute 'absolute\_links'
> 
> 
> 


Any idea why this is happening and how i can fix it?


If you require any more information please let me know.


Thanks


","`r.html.find('.item')` returns a list and list has no attribute `absolute_links`. Since there may be not only one node can be found with `.item`, `find()` method gives you a list as expected.


It will be convinient to get a single node with 



```
about = r.html.find('.item')[0]

```

However, this won't give you the img link by `about.absolute_links`
, because the element found here is of `<a>`, not `<img>`



```
about = r.html.find('.item')[0]
img = about.xpath('//img')[0]
img.attrs['src'] # => '//cdn.pcpartpicker.com/static/forever/images/product/55aea2dd64e2e3a3e3b1d678048d8d76.256p.jpg'

```

"
"How to spoof a java package (and thus a part of the API) in OSGi?
Is it possible to spoof a part of the standard Java API, by exporting your own bundle that replaces this API? Of course with the same interfaces.


Would it then be possible for other bundles, not to notice that actually they aren't using the java package, but your own package.


","In principle this is possible. Only catch is that if you specify a class in the 'java' package, the sun classloader will forbid it:



```
Exception in thread ""main"" java.lang.SecurityException: Prohibited package name: java.io
    at java.lang.ClassLoader.preDefineClass(ClassLoader.java:480)
    at java.lang.ClassLoader.defineClass(ClassLoader.java:615)
    at java.security.SecureClassLoader.defineClass(SecureClassLoader.java:124)
    at java.net.URLClassLoader.defineClass(URLClassLoader.java:260)
    at java.net.URLClassLoader.access$000(URLClassLoader.java:56)
    at java.net.URLClassLoader$1.run(URLClassLoader.java:195)
    at java.security.AccessController.doPrivileged(Native Method)
    at java.net.URLClassLoader.findClass(URLClassLoader.java:188)
    at java.lang.ClassLoader.loadClass(ClassLoader.java:307)
    at sun.misc.Launcher$AppClassLoader.loadClass(Launcher.java:301)
    at java.lang.ClassLoader.loadClass(ClassLoader.java:252)
    at java.lang.ClassLoader.loadClassInternal(ClassLoader.java:320)

```

In order to circumvent the problem, you need to provide your extra classes on the boot classpath:



```
$ java -Xbootclasspath:/home/user/Desktop/:/home/user/Desktop/rt.jar java.io.Hack
hacked

```

You can then also override system files, for instance java.io.File:



```
$ java -Xbootclasspath:/home/user/Desktop/:/home/user/Desktop/rt.jar java.io.File
Error occurred during initialization of VM
java.lang.NoSuchFieldError: separatorChar
    at java.lang.Runtime.loadLibrary0(Runtime.java:819)
    at java.lang.System.loadLibrary(System.java:1030)
    at java.lang.System.initializeSystemClass(System.java:1077)

```

(showing that we just overrode java.io.File)


So you can override system classes, the trick is that you must have access to the virtual machine. You can't do it on the fly, which of course is due to security restrictions.


"
"is any difference between std::forward and std::forward?
are these functions equivalent?



```
template <class T>
void foo(T && t)
{
    bar(std::forward<T>(t));
}

template <class T>
void foo2(T && t)
{
    bar(std::forward<decltype(t)>(t));
}

template <class T>
void foo3(T && t)
{
    bar(std::forward(t));
}

```

if they are, can I always use this macro for perfect forwarding?



```
#define MY_FORWARD(var) std::forward<decltype(var)>(var)

```

or just use 



```
bar(std::forward(t));

```

I believe `foo2` and `foo3` are same, but I found people are always use forward like `foo`, is any reason to explicitly write the type?


I understand that `T` and `T&&` are two different types, but I think `std::forward<T>` and `std::forward<T&&>` always give the same result? 




---


Edit:


the reason I want to use macro is I want to save some typing on following C++1y code, I have many similar code in different places



```
#define XLC_FORWARD_CAPTURE(var) var(std::forward<decltype(var)>(var))
#define XLC_MOVE_CAPTURE(var) var(std::move(var))

template <class T, class U>
auto foo(T && func, U && para )
{
    auto val = // some calculation
    return [XLC_FORWARD_CAPTURE(func),
            XLC_FORWARD_CAPTURE(para),
            XLC_MOVE_CAPTURE(val)](){
              // some code use val
              func(std::forward<U>(para)); 
          };
}

```

","
> 
> Are these functions **two** equivalent?
> 
> 
> 


**Yes**, they are equivalent. `decltype(t)` is the same as `T&&`, and when used with `std::forward`, there is no difference between `T` and `T&&`, regardless what `T` is.



> 
> Can I always use this macro for perfect forwarding?
> 
> 
> 


Yes, you can. If you want to make your code unreadable and unmaintainable, then do so. But I strongly advise against it. On the one hand, you gain basically nothing from using this macro. And on the other hand, other developers have to take a look at the definition to understand it, and it can result in subtle errors. For example adding additional parentheses won't work:



```
MY_FORWARD((t))

```

In contrast, the form with `decltype` is perfectly valid. In particular, it is the preferred way of forwarding parameters from generic lambda expressions, because there are no explicit type parameters:



```
[](auto&& t) { foobar(std::forward<decltype(t)>(t)); }

```



---


I ignored the 3rd variant with `std::forward(t)`, because it isn't valid.




---


**Update:** Regarding your example: You can use *call-by-value* instead of *call-by-reference* for the function template `foo`. Then you can use `std::move` instead of `std::forward`. This adds two additional moves to the code, but no additional copy operations. On the other hand, the code becomes much cleaner:



```
template <class T, class U>
auto foo(T func, U para)
{
    auto val = // some calculation
    return [func=std::move(func),para=std::move(para),val=std::move(val)] {
        // some code use val
        func(std::move(para)); 
    };
}

```

"
"HTTP Load Balancer on Google Container Engine using Ingress
I've tried following the following tutorial: <https://cloud.google.com/container-engine/docs/tutorials/http-balancer>


Everything seems to be working up until the end where doing


`kubectl describe ingress basic-ingress`


returns



```
 Name:          basic-ingress
Namespace:      default
Address:        {hidden_external_ip}
Default backend:    nginx:80 ({hidden_internal_ip}:80)
Rules:
  Host  Path    Backends
  ----  ----    --------
  * *   nginx:80 ({hidden_internal_ip}:80)
Annotations:
  backends:     {""k8s-be-00000--0000000000000000"":""Unknown""}
  forwarding-rule:  k8s-fw-default-basic-ingress--0000000000000000
  target-proxy:     k8s-tp-default-basic-ingress--0000000000000000
  url-map:      k8s-um-default-basic-ingress--0000000000000000
Events:
  FirstSeen LastSeen    Count   From                SubobjectPath   Type        Reason  Message
  --------- --------    -----   ----                -------------   --    ------    ------  -------
  1m        1m      1   {loadbalancer-controller }          Normal          ADD default/basic-ingress
  23s       23s     1   {loadbalancer-controller }          Normal          CREATE  ip: {hidden_external_ip}

```

Notice that the backends line ends with ""unknown""


Furthemore, curling the external IP returns this:



```
<html><head>
<meta http-equiv=""content-type"" content=""text/html;charset=utf-8"">
<title>502 Server Error</title>
</head>
<body text=#000000 bgcolor=#ffffff>
<h1>Error: Server Error</h1>
<h2>The server encountered a temporary error and could not complete your request.<p>Please try again in 30 seconds.</h2>
<h2></h2>
</body></html>

```

Any idea what could cause such an issue?


","I figured out two things:


1. The load balancer does a health check on `/` (which is not documented AFAIK), so you have to make sure that your backend returns 200 on that path.
2. There is a significant delay between when the load-balancer is created and when it ""becomes active"" (~10 minutes or so). When it does, the line: `backends: {""k8s-be-00000--0000000000000000"":""Unknown""}` will become either: `backends: {""k8s-be-00000--0000000000000000"":""Healthy""}` or `backends: {""k8s-be-00000--0000000000000000"":""Unhealthy""}`. This delay was the reason the tutorial linked in the question was not working (it really should be documented there...)


"
"jQuery - dynamically loading ASP.NET partial view into modal
**Question background:**


I have an MVC project where I am attempting to create a partial view `modal` content that is then concatenated to the rest of its respective `modal` markup and then finally appending to a 'main' `modal` div.


**The code**:


Main modal div:



```
<div class=""modal fade"" 
     id=""basicModal"" 
     tabindex=""-1"" 
     role=""dialog"" 
     aria-labelledby=""basicModal"" 
     aria-hidden=""true"">
</div>

```

JQuery to trigger the modal popup along with the `AddModal` method to build the modal content:



```
<script>
$(document).ready(function () {
    $("".btn"").click(function () {

        AddModal();

        $(""#basicModal"").modal(""show"");
    });
});
</script>

```

`AddModal` method to build the modal:



```
AddModal = function() {

var partialHtml = $(@Html.Partial(""../Modal/Index"", new { id = 1 }))

    $html = $('<div class=""modal-dialog"">' +
    '<div class=""modal-content"">' +
        '<div class=""modal-header"">' +
            '<button type=""button"" class=""close"" data-dismiss=""modal"" aria-hidden=""true"">X</button>' +
            '<h4 class=""modal-title"" id=""myModalLabel"">Modal title</h4>' +
        '</div>' +
        '<div class=""modal-body"">'+partialHtml+'</div>' +
        '<div class=""modal-footer"">' +
            '<button type=""button"" class=""btn btn-default"" data-dismiss=""modal"">Close</button>' +
            '<button type=""button"" class=""btn btn-primary"">Save changes</button>' +
        '</div>' +
    '</div>' +
    '</div>');

    $(""#basicModal"").append($html);
};

```


Partial View:



```
<h4>Test Partial view</h4>

```

**The issue:**


I am running into an error of `Uncaught SyntaxError: Unexpected token <` which is being caused by the HTML Markup of the partial view as shown:



```
var partialHtml = $(<h4>Test Partial view</h4>)

```

How can I successfully escape this forward slash so that the HTML from my partial view is correctly added to the rest of the dynamically added markup?


","Instead of using `Html.Partial()` and storing that in a JavaScript string, consider using this technique: [Render Partial View Using jQuery in ASP.NET MVC](https://stackoverflow.com/a/1570138/3191599 ""Render Partial View Using jQuery in ASP.NET MVC"")


Basically, in your `AddModal()` method, fire off a GET request that hits an action that returns the partial view as HTML. Then, just replace the contents of `#basicModal` with the returned HTML:



```
AddModal = function() {
    var partialHtml;
    var url = '../Modal/Index?id=1';

    $.get(url, function(data) {
        $(""#basicModal"").html(data);
    });
};

```

I've used this technique to load modals before and it works well. However, one problem with doing it this way is that `$.get()` is an **async** call, so `.modal(""show"")` is probably going to fire before the data has been fetched. We can solve this by having `AddModal` return the promise generated by `$.get()`:



```
AddModal = function() {
    var partialHtml;
    var url = '../Modal/Index?id=1';

    return $.get(url, function(data) {
        $(""#basicModal"").html(data);
    });
};

```

Then you would change your calling line to:



```
AddModal().done(function() {
    $(""#basicModal"").modal(""show"");
});

```

"
"Docker 'latest' Tagging and Pushing
I have a docker image generated from my build process. I am looking to tag the latest build image with tag build id and 'latest'. I see two ways to do this.


First approach - (Add Multiple tags and Push once) 



```
docker tag <id> <user>/<image>:build_id  
docker tag <id> <user>/<image>:latest  
docker push <user>/<image>

```

Second - Tag individually and push 



```
docker tag <id> <user>/<image>:build_id  
docker push <user>/<image>:build_id

docker tag <id> <user>/<image>:latest  
docker push <user>/<image>:latest

```

The docker documentation says if there is an image in the registry with a specific tag already, then docker push with a new image with same tag would overwrite the earlier image. 


1. Are both First and Second Option specified above functionally the same?
2. Is there any preferred way/Best practice?


","(About the original version of the question, which used `docker push` without arguments) `docker push` will not work unless you provide repository name.



```
$ docker push
""docker push"" requires exactly 1 argument.
See 'docker push --help'.

Usage:  docker push [OPTIONS] NAME[:TAG] [flags]

Push an image or a repository to a registry

```

That means, you need to push with repository name. And you can either provide TAG or not.


If you do not provide TAG, you are pushing all images for that repository.




---


In *first approach*, you are pushing all images under `<user>/<image>` repository. 


In *second approach*, you are pushing image one by one.


***Answer of question***



> 
> 1. Are both First and Second Option specified above functionally the same?
> 
> 
> 


Both First and Second Option specified above are functionally the same (in your case).


If you do not provide TAG, you are pushing all images for that repository.


In your case



```
$ docker push <user>/<image>

```

will push both TAG `build_id` and `latest`



> 
> 2. Is there any preferred way/Best practice?
> 
> 
> 


I think, second option is better and preferred


Because, you may not want to push all images. In that case, you can choose which image you want to push following second approach.


"
"How to filter array with NSPredicate for combination of words

```
[filteredArray filterUsingPredicate:
   [NSPredicate predicateWithFormat:@""self BEGINSWITH[cd] %@"", searchText]];

```

`filteredArray` contains simple NSStrings. [hello, my, get, up, seven, etc...];


It will give all strings that begin with `searchText`. 


But if string will be a combination of words like ""my name is"", and `searchText = name`. What would a `NSPredicate` look like to achieve this?


UPDATE:
And how would it have to be if i want to a result with `searchText = name`, but not with `searchText = ame`? Maybe like this: 



```
 [filteredArray filterUsingPredicate:
       [NSPredicate predicateWithFormat:
        @""self BEGINSWITH[cd] %@ or self CONTENTS[cd] %@"", 
        searchText, searchText]];

```

But it should first display the strings that begin with `searchText` and only after those which contain `searchText`.


","
```
[NSPredicate predicateWithFormat:@""self CONTAINS[cd] %@"", searchText];

```

**EDIT after expansion of question**



```
NSArray *beginMatch = [filteredArray filteredArrayUsingPredicate:
  [NSPredicate predicateWithFormat:
    @""self BEGINSWITH[cd] %@"", searchText]];
NSArray *anyMatch = [filteredArray filteredArrayUsingPredicate:
  [NSPredicate predicateWithFormat:
    @""self CONTAINS[cd] %@"", searchText]];

NSMutableArray *allResults = [NSMutableArray arrayWithArray:beginMatch];
for (id obj in anyMatch) {
   if (![allResults containsObject:obj]) {
      [allResults addObject:obj];
   }
}
filteredArray = allResults;

```

This will have the results in the desired order without duplicate entries.


"
"Most efficient method of self referencing tree using Entity Framework
So I have a SQL table which is basically 



```
ID, ParentID, MenuName, [Lineage, Depth]

```

The last two columns are auto-computed to help with searching so we can ignore them for now. 


I'm creating a drop down menu system with multiple categories. 


Unfortunately EF I don't think plays nice with Self referencing tables more than 1 level deep. So I'm left with a few options


1) Create query, order by depth and then create a custom class in C#, populating it one depth at a time. 


2) Find some way to eager load the data in EF, I don't think it is possible for an unlimited amount of levels, only a fixed amount. 


3) Some other way I'm not even sure about. 


Any inputs would be welcomed! 


","I have successfully mapped hierarchical data using EF.


Take for example an `Establishment` entity. This can represent a company, university, or some other unit within a larger organizational structure:



```
public class Establishment : Entity
{
    public string Name { get; set; }
    public virtual Establishment Parent { get; set; }
    public virtual ICollection<Establishment> Children { get; set; }
    ...
}

```

Here is how the Parent / Children properties are mapped. This way, when you set the Parent of 1 entity, the Parent entity's Children collection is automatically updated:



```
// ParentEstablishment 0..1 <---> * ChildEstablishment
HasOptional(d => d.Parent)
    .WithMany(p => p.Children)
    .Map(d => d.MapKey(""ParentId""))
    .WillCascadeOnDelete(false); // do not delete children when parent is deleted

```

Note that so far I haven't included your Lineage or Depth properties. You are right, EF doesn't work well for generating nested hierarchical queries with the above relationships. What I finally settled on was the addition of a new gerund entity, along with 2 new entity properties:



```
public class EstablishmentNode : Entity
{
    public int AncestorId { get; set; }
    public virtual Establishment Ancestor { get; set; }

    public int OffspringId { get; set; }
    public virtual Establishment Offspring { get; set; }

    public int Separation { get; set; }
}

public class Establishment : Entity
{
    ...
    public virtual ICollection<EstablishmentNode> Ancestors { get; set; }
    public virtual ICollection<EstablishmentNode> Offspring { get; set; }

}

```

While writing this up, [hazzik posted an answer that is very similar to this approach](https://stackoverflow.com/a/11565717/304832). I'll continue writing up though, to provide a slightly different alternative. I like to make my Ancestor and Offspring gerund types actual entity types because it helps me get the Separation between the Ancestor and Offspring (what you referred to as Depth). Here is how I mapped these:



```
private class EstablishmentNodeOrm : EntityTypeConfiguration<EstablishmentNode>
{
    internal EstablishmentNodeOrm()
    {
        ToTable(typeof(EstablishmentNode).Name);
        HasKey(p => new { p.AncestorId, p.OffspringId });
    }
}

```

... and finally, the identifying relationships in the Establishment entity:



```
// has many ancestors
HasMany(p => p.Ancestors)
    .WithRequired(d => d.Offspring)
    .HasForeignKey(d => d.OffspringId)
    .WillCascadeOnDelete(false);

// has many offspring
HasMany(p => p.Offspring)
    .WithRequired(d => d.Ancestor)
    .HasForeignKey(d => d.AncestorId)
    .WillCascadeOnDelete(false);

```

Also, I did not use a sproc to update the node mappings. Instead we have a set of internal commands that will derive / compute the Ancestors and Offspring properties based on the Parent & Children properties. However ultimately, you end up being able to do some very similar querying as in hazzik's answer:



```
// load the entity along with all of its offspring
var establishment = dbContext.Establishments
    .Include(x => x.Offspring.Select(y => e.Offspring))
    .SingleOrDefault(x => x.Id == id);

```

The reason for the bridge entity between the main entity and its Ancestors / Offspring is again because this entity lets you get the Separation. Also, by declaring it as an identifying relationship, you can remove nodes from the collection without having to explicitly call DbContext.Delete() on them.



```
// load all entities that are more than 3 levels deep
var establishments = dbContext.Establishments
    .Where(x => x.Ancestors.Any(y => y.Separation > 3));

```

"
"Spread operator vs JSON.parse(JSON.stringify(...)) for immutable objects
I'm working with Redux and recently ran into a problem where I was adding messages to an array and the Redux state was not issueing a re-render on React. I'm using the `react-redux` library for this. Here's an example of the problem I was having:



```
// State structure
structure: { messages: {}, groups: {} }

// ---
newState = { ...prevState };
newState.messages[action.message.group] = action.message;
return newState; 

```

This was updating the state however it wasn't triggering an update to the react component, however replacing `newState = { ...prevState }` with `newState = JSON.parse(JSON.stringify(prevState))` resolved the issue. 


Could anyone explain why this is happening in detail? I was under the impression that the spread operator created a clone of the object and I've never had any problems with it until now. 


","react-redux connected components do a shallow strict equality check to decide if they want to update. see <http://redux.js.org/docs/faq/ImmutableData.html>


The spread operator is like Object.assign and does not deeply clone an object. The reason the JSON thing worked is because you created a whole new object which would pass the strict equality check, however all your components would update needlessly because nothing will pass a strict equality check now.


Object.assign({}, ...prevState, ...newState) would create a new top-level object, but it would not create a new object for any objects nested in prevState or newState. However, you would have to carefully update nested objects so as to avoid needless re-renders. This can get tricky for deeply nested objects and arrays.


I'd recommend checking out the seamless-immutable or immutable packages for managing state. Also, the reselect library can help you extract memoized objects specific to your component's needs.
  
  
  



# UPDATE 08/16/2020


the [immer](https://github.com/immerjs/immer) library is one of the best state management libraries as of today


"
"Firebase Cloud Functions: Error: Unexpected error while acquiring application default credentials: read ECONNRESET
When trying to run my firebase cloud function with http trigger, I kept getting this error but only intermittently:



```
Error: Unexpected error while acquiring application default credentials: read ECONNRESET 

```

There are some unrelated ECONNRESET/firebase SO answers that provided no solution:
[Firebase Storage & Cloud Functions - ECONNRESET](https://stackoverflow.com/q/43224994/2684661)


","Problem: 


To initialize my functions I was using: 



```
const functions = require('firebase-functions');
const admin = require('firebase-admin');    

admin.initializeApp(functions.config().firebase);

```

Solution (for me at least):


I had to use my Google service account credentials:



```
var serviceAccount = require(""./PATH_TO_YOUR_SERVICE_ACCOUNT_FILE_GOES_HERE.json"");

admin.initializeApp({
  credential: admin.credential.cert(serviceAccount),
  databaseURL: ""https://YOUR_FIREBASE_PROJECT_NAME.firebaseio.com""
});

```

To get the firebase project name check the top left corner of your firebase project console (A) .


To Download the service account JSON file click on the (B) gear icon in the top left of the firebase console then click (C) project settings:


[![enter image description here](https://i.stack.imgur.com/lCwva.png)](https://i.stack.imgur.com/lCwva.png)


Then click the service accounts tab:


[![enter image description here](https://i.stack.imgur.com/7QOW2.png)](https://i.stack.imgur.com/7QOW2.png)


Then click the ""Generate New Private Key to download the file. 


You can then move the JSON file into your project functions folder and import it as shown in the code snippet above.


[![enter image description here](https://i.stack.imgur.com/XWtv4.png)](https://i.stack.imgur.com/XWtv4.png)


"
"Python real time varying heat map plotting
I have a 2D grid 50\*50. For each location I have an intensity value(i.e data is like `(x,y,intensity)` for each of those 50\*50 locations). I would like to visualize the data as a heatmap.


The twist is that every second the intensity will change(for most of the locations), which means I will need to re-draw the heatmap every second. I am wondering what is the best library/approach to handle this kind of real-time varing heatmap.


","This really depends on how you get your data, but:



```
import matplotlib.pyplot as plt
import numpy as np
import time

# create the figure
fig = plt.figure()
ax = fig.add_subplot(111)
im = ax.imshow(np.random.random((50,50)))
plt.show(block=False)

# draw some data in loop
for i in range(10):
    # wait for a second
    time.sleep(1)
    # replace the image contents
    im.set_array(np.random.random((50,50)))
    # redraw the figure
    fig.canvas.draw()

```

This should draw 11 random 50x50 images with 1 second intervals. The essential part is `im.set_array` which replaces the image data and `fig.canvas.draw` which redraws the image onto the canvas.




---


If your data is really a list of points in the form `(x, y, intensity)`, you can transform them into a `numpy.array`:



```
import numpy as np

# create an empty array (NaNs will be drawn transparent)
data = np.empty((50,50))
data[:,:] = np.nan

# ptlist is a list of (x, y, intensity) triplets
ptlist = np.array(ptlist)
data[ptlist[:,1].astype('int'), ptlist[:,0].astype('int')] = ptlist[:,2]

```

"
"Do suspended processes still use up CPU/RAM?
I am wondering, if you suspend a process, does it still use the RAM/CPU or does it just free all used RAM and the CPU?


","Suspending all threads in a process will prevent the threads from receiving more CPU time so once any remaining kernel interrupts/operations have completed the process will not consume any CPU power until threads are resumed. The undocumented API to suspend a process simply suspends all the threads in the process while holding a process related lock.


A suspended Win32 process does not release any memory but on a system with little free physical RAM, parts of the application can be moved from RAM to the page file to free up RAM for use by other active applications. [`SetProcessWorkingSetSize`](https://msdn.microsoft.com/en-us/library/windows/desktop/ms686234(v=vs.85).aspx) can be called to force a reduction of pages in the processes working set.


UWP/Store applications have a higher level type of suspension where the application is given a small time window to save its state and then the entire process **might** be terminated. A new process is started when the user switches back to the application if it was terminated.



> 
> The system attempts to keep your app and its data in memory while it's suspended. However, if the system does not have the resources to keep your app in memory, [it will terminate your app](https://learn.microsoft.com/en-us/windows/uwp/launch-resume/app-lifecycle). 
> 
> 
> 


"
"How to detect a background program (in Windows 7) which steals active focus automatically?
A background process/program steals focus from my currently active application. It disturbs me a lot when I type in an app or control an app using keyboard or play online flash games in full-screen.  

I want to remove this program from my PC, but I'm unable to detect it. I've analyzed all processes & scanned for viruses. All things are right from this side.  

I don't want to change focus behavior using registry tweak. Please, help me detecting that app..


","The problem has been fixed. At a time of focus stealing, I pressed `Alt`+`F4` and instantly spotted exiting process from *Process Explorer*.


Note: *Windows Process Explorer* can be [downloaded](http://technet.microsoft.com/pl-pl/sysinternals/bb896653.aspx) as part of *Windows Sysinternals* library.


update:
theres's any nice pogramm showing/logging, what process has focus:
<https://social.technet.microsoft.com/Forums/ie/en-US/dee98b15-0540-4975-b38d-ce3d2ee87c30/figuring-out-whats-stealing-window-focus>


may this helps as well
<http://pcsupport.about.com/od/windowsxp/ht/stealingfocus02.htm>


"
"Entity Framework, Code First and One-to-Many relationship across multiple contexts
I am using VS 2010 and Entity Framework code first (version 6). I have two entities each in its own context and I want to create a one-to-many relationship between them.


Context 1 has the following entity:



```
public class MyTrust
{
    public int MyTrustID { get; set; }
    public string MyTrustName { get; set; }
}

```

and Context 2 has the following entity:



```
public class MyLocation
{
    public int MyLocationID { get; set; }
    public int MyTrustID { get; set; }
    public virtual MyTrust MyTrust { get; set; }
}

```

with the following Fluent API



```
modelBuilder.Entity<MyLocation>()
    .HasRequired(m => m.MyTrust);

```

The migration file for Context 2 contains the correct keys but also creates a new table for `MyTrust` which already exists in the other context.


I know that I can edit the migration file but that is not a solution.


My question is, how to I stop the creation of the second `MyTrust` table.


**UPDATE**


There was a major flaw above in that I pasted the wrong code into Context 2. Now corrected. Apologies.


","You are working with so-called *bounded contexts*. The benefit of such contexts and how to work with them is explained in this [blog by Julie Lerman](http://msdn.microsoft.com/en-us/magazine/jj883952.aspx).


The problem you experience, none of the contexts can be used in migrations, is addressed in this part:



> 
> If you’re doing new development and you want to let Code First create or migrate your database based on your classes, you’ll need to create an “uber-model” using a DbContext that includes all of the classes and relationships needed to build a complete model that represents the database.
> 
> 
> 


Note that you can share the `MyTrust` type between all contexts, if you observe these rules (from Lerman & Miller's book *DbContext*, p 233):



> 
> - An entity can only be attached to one context at a time. This architecture works
>  best with short-lived contexts where the instance to be shared will be completely
>  disassociated from one context before it is attached to another.
> - Entities that are attached to different contexts cannot be attached to one another.
> 
> 
> 


**UPDATE**


In EF6 you can use multiple contexts for one migration path. See [this walkthrough](http://www.dotnet-tricks.com/Tutorial/entityframework/2VOa140214-Entity-Framework-6-Code-First-Migrations-with-Multiple-Data-Contexts.html).


"
"probabilities returned by gensim's get\_document\_topics method doesn't add up to one
Sometimes it returns probabilities for all topics and all is fine, but sometimes it returns probabilities for just a few topics and they don't add up to one, it seems it depends on the document. Generally when it returns few topics, the probabilities add up to more or less 80%, so is it returning just the most relevant topics? Is there a way to force it to return all probabilities?


Maybe I'm missing something but I can't find any documentation of the method's parameters.


","I had the same problem and solved it by including the argument `minimum_probability=0` when calling the `get_document_topics` method of `gensim.models.ldamodel.LdaModel` objects. 



```
    topic_assignments = lda.get_document_topics(corpus,minimum_probability=0)

```

By default, **gensim doesn't output probabilities below 0.01**, so for any document in particular, if there are any topics assigned probabilities under this threshold the sum of topic probabilities for that document will not add up to one.


Here's an example:



```
from gensim.test.utils import common_texts
from gensim.corpora.dictionary import Dictionary
from gensim.models.ldamodel import LdaModel

# Create a corpus from a list of texts
common_dictionary = Dictionary(common_texts)
common_corpus = [common_dictionary.doc2bow(text) for text in common_texts]

# Train the model on the corpus.
lda = LdaModel(common_corpus, num_topics=100)

# Try values of minimum_probability argument of None (default) and 0
for minimum_probability in (None, 0):
    # Get topic probabilites for each document
    topic_assignments = lda.get_document_topics(common_corpus,minimum_probability=minimum_probability)
    probabilities = [ [entry[1] for entry in doc] for doc in topic_assignments ]
    # Print output
    print(f""Calculating topic probabilities with minimum_probability argument = {str(minimum_probability)}"")
    print(f""Sum of probabilites:"")
    for i, P in enumerate(probabilities):
        sum_P = sum(P)
        print(f""\tdoc {i} = {sum_P}"")

```

And the output would be:



```
Calculating topic probabilities with minimum_probability argument = None
Sum of probabilities:
    doc 0 = 0.6733324527740479
    doc 1 = 0.8585712909698486
    doc 2 = 0.7549994885921478
    doc 3 = 0.8019999265670776
    doc 4 = 0.7524996995925903
    doc 5 = 0
    doc 6 = 0
    doc 7 = 0
    doc 8 = 0.5049992203712463
Calculating topic probabilities with minimum_probability argument = 0
Sum of probabilites:
    doc 0 = 1.0000000400468707
    doc 1 = 1.0000000337604433
    doc 2 = 1.0000000079162419
    doc 3 = 1.0000000284053385
    doc 4 = 0.9999999937135726
    doc 5 = 0.9999999776482582
    doc 6 = 0.9999999776482582
    doc 7 = 0.9999999776482582
    doc 8 = 0.9999999930150807

```

This default behaviour is not very clearly stated in the documentation. The default value for `minimum_probability` for the `get_document_topics` method is `None`, however this does not set the probability to zero. Instead the value of `minimum_probability` is set to the value of `minimum_probability` of the `gensim.models.ldamodel.LdaModel` object, which by default is 0.01 as you can see in the [source code](https://github.com/RaRe-Technologies/gensim/blob/996801bb3fb8c4e10a84eefa70f5e2ac738dd47b/gensim/models/ldamodel.py#L347):



```
def __init__(self, corpus=None, num_topics=100, id2word=None,
             distributed=False, chunksize=2000, passes=1, update_every=1,
             alpha='symmetric', eta=None, decay=0.5, offset=1.0, eval_every=10,
             iterations=50, gamma_threshold=0.001, minimum_probability=0.01,
             random_state=None, ns_conf=None, minimum_phi_value=0.01,
             per_word_topics=False, callbacks=None, dtype=np.float32):

```

"
"How to restart PostgreSQL sequences handled by Hibernate?
I've been tasked with restarting a sequence for many different serial columns in a PostgreSQL database. Normally, I would simply use:



```
ALTER SEQUENCE serial RESTART WITH 105;

```

However, it seems that Hibernate is being used to handle the database sequences. I really don't know anything about Hibernate, but my understanding is that hibernate\_sequence is a global sequence for all tables. Is that correct?


I'm assuming that I would need to do the following, then:



```
ALTER SEQUENCE hibernate_sequence RESTART WITH 105;

```

But I am not sure what the consequence will be. Let's assume I have Tables A, B, C, and D. Each of those tables has an ID column of type serial. Will the above SQL statement on hibernate\_sequence restart the ID column for ALL of the tables?


Thanks!


","First, there's an underling problem here: You're treating the value of sequences like they mean something. See the end of the answer.


Now, with Hibernate, it depends on how the user has configured the mappings. A well configured Hibernate instance will use the same sequences everything else does.


Unfortunately, most mappings aren't very good - or rather, I think Hibernate's default behaviour is horrible and should always be overridden. The default sequence used is indeed a shared `hibernate_sequence` for all tables. However, it's entirely possible to have *some* tables using their own sequences, and other tables using the shared `hibernate_sequence`, or to have all tables properly using their own sequences.


To enable a definitive answer you'd need to post your mappings.


If the table(s) you want to change the ID generation of are using the correct sequences for their tables in their Hibernate mappings, you can just go ahead and `ALTER SEQUENCE` that sequence. If your mappings are via JPA, that's done with a `@SequenceGenerator` and `@GeneratedValue` with `GenerationType.SEQUENCE`. If it's done via `orm.xml`, `hbm.xml` or the old-style Hibernate annotations you'll need to check the manual. Or, if Hibernate is the only client that writes to the DB, look at the sequences to see if they're going up using `SELECT * FROM sequence_name`.


If the mappings for the table(s) of interest are using the shared `hibernate_sequence` you probably can't do what you want without fixing the Hibernate mappings. You'd need to change the mappings so the mapping for the table(s) of interest use a different sequence, then``LOCK TABLE` and set the start point for those sequences to the greatest row currently in the table.


**Sequences**: Why do you *need* to change or restart the sequence? It shouldn't matter, since a synthetic primary key is just a meaningless number that you compare for equality with other meaningless numbers. What's the problem you're trying to solve?


"
"Is there a way to extract the type of the props of a JSX Element?
My intent is to **extract the props out of a given JSX element**, is it possible?


This was pretty much my failed attempt...
Thanks in advance for any help ;)



```
function getComponentProps<T extends React.ReactElement>(element: T): ExtractProps<T>;

function Component({ name }: { name: string }) {
  return <h1>{name}</h1>;
}

type ExtractProps<TComponentOrTProps> = TComponentOrTProps extends React.ComponentType<infer TProps>
  ? TProps
  : TComponentOrTProps;

const componentProps = getComponentProps(<Component name=""jon"" />); //type is JSX.Element

```

","For the most part, you can't do this.


In *theory*, the `React.ReactElement` type is generic with a type parameter `P` that depends on the props. So if you were to have a strongly-typed element then you could work backwards.



```
type ElementProps<T extends React.ReactNode> = T extends React.ReactElement<infer P> ? P : never;

```

In *reality*, you will only get a correct props type if you create your element through `React.createElement` rather than JSX.


Any JSX element `<Component name=""John"" />` just gets the type `JSX.Element` which obviously has no information about the props so you cannot work backwards from that to a props type.



```
const e1 = React.createElement(
  Component,
  { name: 'John' }
)

type P1 = ElementProps<typeof e1> // type: {name: string}

console.log(getElementProps(e1)); // will log {name: ""John""}

```


```
const e2 = <Component name=""John"" />

type P2 = ElementProps<typeof e2>  // type: any

console.log(getElementProps(e2)); // will log {name: ""John""}

```

[Playground Link](https://tsplay.dev/NalrPw)


It is much easier to approach the situation from a different angle. You will be able to derive the correct props type if your function takes a component like `Component` or `div` rather than a resolved element. You can use the `ComponentProps` utility type for function and class components and the `JSX.IntrinsicElements` map for built-in ones.


"
"given an %hhu and an input 1025, which part is mapped to an unsigned char?
Is the output of the following programm `1` or `0`, or is it undefined behaviour?



```
int main() {
  unsigned char u = 10;
  sscanf(""1025"",""%hhu"",&u);
  printf(""u, is it 0 or is it 1? u's value is ... %hhu"", u);
}

```

According to `fscanf` [conversion specifier](http://port70.net/~nsz/c/c11/n1570.html#7.21.6.2p12) `%u` with [length modifier](http://port70.net/~nsz/c/c11/n1570.html#7.21.6.2p11) `hh` (i.e. `%hhu`), semantics is defined based on that of `strtoul` function and a mapping to type pointer to unsigned char:



> 
> 12) The conversion specifiers and their meanings are:
> 
> 
> **""u""**
>  Matches an optionally signed decimal integer, whose format is the same
>  as expected for the subject sequence of the strtoul function with the
>  value 10 for the base argument. The corresponding argument shall be a
>  pointer to unsigned integer.
> 
> 
> 11) The length modifiers and their
>  meanings are:
> 
> 
> **""hh""** Specifies that a following d, i, o, u, x,
>  X, or n conversion specifier applies to an argument with type pointer
>  to signed char or unsigned char.
> 
> 
> 


But what happens if an input sequence represents an integral value exceeding 8 bits, which part of the integral value is mapped to the 8 bits of an unsigned char? Is it defined that it has to be the least significant part, does it depend on endianess, is it unspecified, or does it even yield undefined behaviour?


I cannot believe that it is undefined or unspecified behaviour. This would mean that user input might introduce such behaviour in a program using `scanf(""%hhu"",&u)`, and checking user input before every use of `scanf` looks absurd to me.


","Undefined. See [one section up:](http://port70.net/~nsz/c/c11/n1570.html#7.21.6.2p10)



> 
> 10 Except in the case of a % specifier, the input item (or, in the case of a %n directive, the count of input characters) is converted to a type appropriate to the conversion specifier. If the input item is not a matching sequence, the execution of the directive fails: this condition is a matching failure. Unless assignment suppression was indicated by a \*, the result of the conversion is placed in the object pointed to by the first argument following the format argument that has not already received a conversion result. If this object does not have an appropriate type, **or if the result of the conversion cannot be represented in the object, the behavior is undefined**.
> 
> 
> 


"
"ChangeNotifier was used after being disposed
I am new to Flutter and stuck with this
I have a page that uses a class named GoogleMapsNotifier with ChangeNotifier and when I pop the page I want to dispose the Stream inside this class (last function).



```
class GoogleMapsNotifier with ChangeNotifier {
  final geolocatorService = GeolocatorService();
  final placesService = PlacesService();
  final markerService = MarkerService();

  Position? currentLocation;
  List<PlaceSearch> searchResults = [];
  StreamController<Place> selectedLocation = BehaviorSubject<Place>();
  StreamController<LatLngBounds> bounds = BehaviorSubject<LatLngBounds>();
  late Place selectedLocationStatic;
  List<Marker> markers = <Marker>[];

  GoogleMapsNotifier() {
    setCurrentLocation();
  }

  setCurrentLocation() async {
    currentLocation = await geolocatorService.determinePosition();
    selectedLocationStatic = Place(
        geometry: Geometry(
          location: Location(
              lat: currentLocation!.latitude, lng: currentLocation!.longitude),
        ),
        name: '',
        vicinity: '');
    notifyListeners();
  }

  searchPlaces(String searchTerm) async {
    searchResults = await placesService.getAutocomplete(searchTerm);
    notifyListeners();
  }

  setSelectedLocation(String placeId) async {
    var sLocation = await placesService.getPlace(placeId);
    selectedLocation.add(sLocation);
    selectedLocationStatic = sLocation;
    searchResults = [];
    markers = [];
    var newMarker = markerService.createMarkerFromPlace(sLocation);
    markers.add(newMarker);
    var _bounds = markerService.bounds(Set<Marker>.of(markers));
    bounds.add(_bounds as LatLngBounds);
    notifyListeners();
  }

  @override
  void dispose() {
    selectedLocation.close();

    super.dispose();
  }
}

```

and then I have a Go Back button that pops the page and I call this function with Provider before.



```
onTap: () async {
                    Provider.of<GoogleMapsNotifier>(context, listen: false)
                        .dispose();
                    Navigator.pop(context);
                  },

```

It works fine for the first time but when I enter the page for the second time and press Go Back button again, it return an error



> 
> Unhandled Exception: A GoogleMapsNotifier was used after being
> disposed. E/flutter (13173): Once you have called dispose() on a
> GoogleMapsNotifier, it can no longer be used.
> 
> 
> 


How can I fix this?


","The `Provider` should be inside the `Route` you push. If you use a global provider, the instance of `GoogleMapsNotifier` will always be the same. Therefore the second time you go on the page it won't work (since it's the same instance you already disposed the first time)


Here is a concrete example



```
// GOOD
runApp(MaterialApp(...));

...

Navigator.push(
  context,
  MaterialPageRoute(
    builder: (_) => ChangeNotifierProvider<GoogleMapsNotifier>(
      create: (_) => GoogleMapsNotifier(),
      child: ...,
    ),
  ),
);


// BAD
runApp(
  ChangeNotifierProvider<GoogleMapsNotifier>(
    create: (_) => GoogleMapsNotifier(),
    child: MaterialApp(
      home: ...,
    ),
  )
);

...

Navigator.push(
  context,
  MaterialPageRoute(
    builder: (_) => ...,
  ),
);

```

"
"Changing php version in mamp
I am using the latest version of MAMP.
In my MAMP settings, I can find the latest version of PHP as 7.1.1,
but when I run `php -v` i get this:



```
PHP 5.5.14 (cli) (built: Sep  9 2014 19:09:25) 
Copyright (c) 1997-2014 The PHP Group
Zend Engine v2.5.0, Copyright (c) 1998-2014 Zend Technologies

```

I am not able to change it.
There is no folder like `php5.5.14` inside `MAMP/bin/php/`


","That's because it's referring to the pre-installed PHP of your *OSX*. You need to change that to point to your MAMP PHP. To do so you need to edit the `.bash_profile` with a text editor and add the MAMP version of PHP to the `PATH` variable.


Follow these steps ([source](https://gist.github.com/irazasyed/5987693)):



> 
> 1. Within the Terminal, run
> 
> 
> 



```
vim ~/.bash_profile

```


> 
> 2. Type i and then paste the following at the top of the file:
> 
> 
> 



```
export PATH=/Applications/MAMP/bin/php/php7.1.1/bin:$PATH

```


> 
> 3. Hit ESC, Type `:wq`, and hit *Enter*
> 4. In Terminal, run:
> 
> 
> 



```
 source ~/.bash_profile

```


> 
> 5. In Terminal, type in which php again and look for the updated string.
>  If everything was successful, It should output the new path to MAMP
>  PHP install.
> 
> 
> 


In case it doesn't output the correct path, try closing the terminal
window (exit fully) and open again, it should apply the changes.


"
"Java best practices, add to collection before or after object has been modified?
Say you are adding x number of objects to a collection, and after or before adding them to a collection you are modifying the objects attributes. When would you add the element to the collection before or after the object has been modified.


Option A)



```
public static void addToCollection(List<MyObject> objects) {
    MyObject newObject = new MyObject();
    objects.add(newObject);
    newObject.setMyAttr(""ok"");
}

```

Option B)



```
public static void addToCollection(List<MyObject> objects) {
    MyObject newObject = new MyObject();
    newObject.setMyAttr(""ok"");
    objects.add(newObject);
}

```

","To be on the safe side, you should modify before adding, unless there is a specific reason you cannot do this, and you know the collection can handle the modification. The example can reasonably be assumed to be safe, since the general `List` contract does not depend upon object attributes - but that says nothing about specific implementations, which may have additional behavior that depends upon the object's value.


TreeSet, and Maps in general do no tolerate modifying objects after they have been inserted, because the structure of the collection is dependent upon the attributes of the object. For trees, any attributes used by the comparator cannot be changed once the item has been added. For maps, it's the hashCode that must remain constant.


So, in general, modify first, and then add. This becomes even more important with concurrent collections, since adding first can lead to other collection users seeing an object before it been assigned it's final state.


"
"Character class subtraction, converting from Java syntax to RegexBuddy
Which regular expression engine does Java uses?


In a tool like RegexBuddy if I use



```
[a-z&&[^bc]]

```

that expression in Java is good but in RegexBuddy it has not been understood.


In fact it reports:



> 
> Match a single character present in
>  the list below `[a-z&&[^bc]`
> 
> 
> - A character in the range between `a` and `z` : **`a-z`**
> - One of the characters `&[^bc` : **`&&[^bc`**
> - Match the character `]` literally : **`]`**
> 
> 
> 


but i want to match a character between `a` and `z` intersected with a character that is not `b` or `c`


","Like most regex flavors, [`java.util.regex.Pattern`](http://java.sun.com/javase/6/docs/api/java/util/regex/Pattern.html) has its own specific features with syntax that may not be fully compatible with others; this includes character class union, intersection and subtraction:



> 
> - `[a-d[m-p]]` : `a` through `d`, or `m` through `p`: `[a-dm-p]` *(union)*
> - `[a-z&&[def]]` : `d`, `e`, or `f` *(intersection)*
> - `[a-z&&[^bc]]` : `a` through `z`, except for `b` and `c`: `[ad-z]` *(subtraction)*
> 
> 
> 


The most important ""caveat"" of Java regex is that `matches` attempts to match a pattern against the *whole* string. This is atypical of most engines, and can be a source of confusion at times.


### See also


- [regular-expressions.info/Flavor Comparison](http://www.regular-expressions.info/refflavors.html) and [Java Flavor Notes](http://www.regular-expressions.info/java.html)




---


### On character class subtraction


Subtraction allows you to define for example ""all consonants"" in Java as `[a-z&&[^aeiou]]`.


This syntax is specific to Java. In XML Schema, .NET, JGSoft and RegexBuddy, it's `[a-z-[aeiou]]`. Other flavors may not support this feature at all.


### References


- [regular-expressions.info/Character Classes in XML Regular Expressions](http://www.regular-expressions.info/xmlcharclass.html)
- [MSDN - Regular Expression Character Classes - Subtraction](http://msdn.microsoft.com/en-us/library/20bw873z.aspx#CharacterClassSubtraction)


### Related questions


- [What is the point behind character class intersections in Java’s Regex?](https://stackoverflow.com/questions/1106387/what-is-the-point-behind-character-class-intersections-in-javas-regex)


"
"Expected Primary Expression (Solidity)
I am creating a simple smart contract, however, I am getting an error on my last function (""ViewNotes"") stating that the compiler was ""Expected Primary Expression""? Can I not check the value at a mapping (of address => string) against the value *0* ?


My code: 



```
pragma solidity ^0.4.4;

contract Logistics{

address public owner;
mapping(address => string) notes;

modifier onlyOwner() {
    require(msg.sender == owner);
    _;
}

constructor(address genesis) public {
   owner = genesis;
}

function sign(string signedNote) public onlyOwner{
        notes[owner] = signedNote; //gaurenteed that msg.sender == owner
}

function transferOwnership(address nuOwner) onlyOwner {
    owner = nuOwner;
}

function viewNotes(address participant) public returns(string){ // signed note on success nothing on fail
    if(notes[participant] !== 0){
        return (notes(participant));   
    }
}

```

}


","There are a couple issues. The primary issue is that you misspelled `!=`. (You have an extra equals sign. `!==` is an operator in JavaScript, but not in Solidity.)


Once you fix that, you'll find that you can't compare a `string` to the number `0`. You probably want to check the string's length? You'll need to cast to `bytes` to do that:



```
function viewNotes(address participant) public returns (string) {
    if (bytes(notes[participant]).length != 0) {
        return notes[participant];
    }
}

```

That said, I believe this is probably equivalent to just:



```
function viewNotes(address participant) public returns (string) {
    return notes[participant];
}

```

And you could instead just make `notes` `public`:



```
mapping(address => string) public notes;

```

That way, Solidity will generate a getter function for you, and people can just call `notes(addr)`, making `viewNotes` redundant.


Fixing up a couple other warnings, getting rid of the modifier in favor of a direct ownership check, and assigning initial ownership to the deployer, here's my take on the contract:



```
pragma solidity ^0.4.24;

contract Logistics{
    address public owner = msg.sender;
    mapping(address => string) public notes;

    function sign(string note) public {
        require(msg.sender == owner);
        notes[owner] = note;
    }

    function transferOwnership(address newOwner) public {
        require(msg.sender == owner);
        owner = newOwner;
    }
}

```

"
"Cleanup & Arrange practices during integration testing to avoid dirty databases
I'm coding tests in C# and I settled with this structure:



```
try
{
    // ==========
    // ARRANGE
    // ==========

    // Insert into the database all test data I'll need during the test

    // ==========
    // ACT
    // ==========

    // Do what needs to be tested

    // ==========
    // ASSERT
    // ==========

    // Check for correct behavior
}
finally
{
    // ==========
    // CLEANUP
    // ==========

    // Inverse of ARRANGE, delete the test data inserted during this test
}

```

The concept was ""every test cleans up the mess it makes"". However, *some* tests are leaving the database dirty and the failing the tests that come after.


**What's the right way to do this?** (minimize errors, minimize run time)


- `Deletes everything` » `Insert defaults` » `Insert test data` » Run test?
- `Insert defaults` » `Insert test data` » Run test » `Delete everything`?
- **Currently**:


	- (per session) `Deletes everything` » `Insert defaults`
	- (per test) `Insert test data` » Run test » `Delete test data`


","Besides the fact that this is an integration test as opposed to a unit test, the operations you describe typically go in `Setup` and/or `Teardown` methods. Frameworks like *nUnit* allow one to decorate class methods with these attributes to indicate whether the method is a setup method or teardown method.


Then your tests should become cleaner and smaller as the setup and cleanup is done outside of the test itself.


Most likely multiple tests can re-use the same data so that is a plus as well as opposed to insert/remove on every test. Going back to *nUnit*, the `FixtureSetup` and `FixtureTeardown` attributes help to setup data for multiple tests at once.


I would use a test framework over a try/catch as many of these testing features are built into the framework itself. xUnit, nUnit, even the Microsoft built in testing framework are all solid choices and will help with the setup and clean up of database records in a consistent fashion.


"
"Cannot follow Shortcuts in CIFS-mounted windows shares
I have a windows share:



```
\\MYSERVER\SHARE\
                 MyFolder
                 MyShortcut\
                            SubFolder1
                            Subfolder2

```

Using a windows box, I can navigate through the shortcut to reach the subfolders.


But using a linux box with CentOS release 6.5, I can not navigate through the shortcut.


I have the following in my fstab:



```
//MYSERVER/SHARE /mnt/g cifs    uid=ME,gid=ME,noauto,credentials=/home/ME/.smbcredentials,iocharset=utf8,file_mode=0777,dir_mode=0777   0 0

```

I can cd to /mnt/g and /mnt/g/MyFolder
but cannot cd to /mnt/g/MyShortcut


In /mnt/g, if I typed: cd MyShortcut:



```
-bash: cd: MyShortcut: Invalid argument

```

If I typed: ls -l



```
drwx--x--x. 2 ME ME 0    Apr 17 12:16 MyShortcut
drwxrwxrwx. 1 ME ME 4096 Sep 19  2013 MyFolder

```

Update: I also tried to mount \MYSERVER\SHARE\MyShortcut from fstab, but ended with errors.


","Windows implements its shortcuts as `.lnk` files (typically you don't see the `.lnk` extension as it's one of the few extensions that remains hidden even if you tell the Folder Settings control panel to display them). When you use shortcuts in Windows, these files are read by `explorer.exe` to change the current folder when navigating through folders.


Many Linux filesystems feature [symlinks](http://en.wikipedia.org/wiki/Symbolic_link) which do the same thing, but work on the filesystem level and thus independent of any program. Windows-style `.lnk` shortcuts are not supported by the file manager you are using (probably `nautilus`).


The reason why the `cd` command is failing is because you are asking Linux to `cd` to a file and not a directory/folder.


Some workarounds:


- The `.lnk` file is just a text file, so you can open it in `gedit` or `vi` and path manually to its equivalent CIFS destination.
- Windows supports UNIX/Linux-style symlinks with the `mklink` command (on NTFS filesystems) - however I haven't tried accessing them via CIFS share to see if they'd work.


"
"selecting from either of two columns in SELECT
In SQL Server 2008R2 I have two tables, `CONFIG_DATA` and `CNA_LIST`:


`CONFIG_DATA` has two fields: 



```
ID [varchar(20)]    Address [int]
01                  141516  
02                  132132

```

`CNA_LIST` has three fields: 



```
Address [int]  IP1 [varchar(20)]   IP2 [varchar(20)]
141516         1.2.3.4             (null)
132132         (null)              2.3.4.5

```

These are field devices that have one or two IP addresses, in the `IP1` or `IP2` field (can be either, or both).


I would like to list the `CONFIG_DATA` table with `IP1` from the `CNA_LIST` table, and if it's `NULL`, the `IP2`.


The `Address` field links the two tables. 


this would work fine:



```
select a.ID, a.Address, b.IP1
from CONFIG_DATA a, CNA_LIST b where a.Address = b.Address

```

but if `IP1` is `null` I need `IP2`, with this column listed as `""IP""`:


I've tried `IF(b.IP1 <> '')b.IP1 ELSE b.IP2` but the syntax checker complains.


","You can use either the [isnull()](https://msdn.microsoft.com/en-us/library/ms184325.aspx) or [coalesce()](https://msdn.microsoft.com/en-us/library/ms190349.aspx) functions to do this:



```
select a.ID, a.Address, isnull(b.IP1, b.IP2) as ""IP""
from CONFIG_DATA a
inner join CNA_LIST b on a.Address = b.Address

```

I also rewrote your query to use explicit joins instead of the older implicit join syntax. 


`isnull()` takes two arguments (*check, replacement*) and returns the second if the first is `NULL`. 
  
`coalesce()` works almost the same way but can take more than two arguments and returns the first non-`NULL` argument or `NULL` if all arguments are `NULL`.


There's also a third way that doesn't rely on the build-in function, which is ti use the [case](https://msdn.microsoft.com/en-us/library/ms181765.aspx) expression like this: `case when b.IP1 is not null then b.IP1 else b.IP2 end` 


"
"IPtables : Limit number of new ssh connections per minute
I need to ensure on my server that maximum new ssh connections per minute are not more then 5.



```
sudo /sbin/iptables -A INPUT -p tcp --syn --dport 22 -m connlimit --connlimit-above 5 -j REJECT

```

Above IPtables rule work for me, but it will not allow new connections after one minute.


Any pointers how to achieve this?


","
```
# SSH Anti-Bruteforce
# Next create a limiting rule. I limit to 3 hits in 60 seconds for ssh with this
/sbin/iptables -A INPUT -p tcp --dport 22 -m state --state NEW -m recent --set --name SSH
/sbin/iptables -A INPUT -p tcp --dport 22 -m state --state NEW -j ssh-whitelist
/sbin/iptables -A INPUT -p tcp --dport 22 -m state --state NEW -m recent --update --seconds 60 --hitcount 3 --rttl --name SSH -j LOG --log-level info --log-prefix ""Anti SSH-Bruteforce: ""
/sbin/iptables -A INPUT -p tcp --dport 22 -m state --state NEW -m recent --update --seconds 60 --hitcount 3 --rttl --name SSH -j DROP

```

"
"Minimize all windows on the active monitor
I'm looking for a way to use a keyboard shortcut to minimize all the windows on the monitor that currently has my mouse on it. I found this page: <http://vbcity.com/forums/t/163339.aspx> but that appears to be hardcoded to minimize windows on a specific monitor, rather than the active monitor. Plus, I have no idea how to compile that code (not a VB coder). Anybody know a way that I can make this happen? I'm on Windows 7 ultimate, 3 monitors.


","Under Windows 7, here is the list of the keyboard shortcuts that manage multiple windows :


**`Win`+`d`**  

Minimize all windows on all Monitors. Press again to restore previous state 


**`Win`+`m`**  

Minimize all windows on all Monitors.


**`Win`+`Shift`+`m`**  

Restore previously minimized windows on current Monitor 


**`Win`+`Home`**  

Set all windows to Minimized on current Monitor except active 


**`Win`+`Space`**  

Preview Desktop / make windows transparent (May not work with all Settings)


Source : [127 useful Keyboard Shortcuts for Windows 7](http://www.shortcutworld.com/en/win/Windows_7.html).


The OP notes that Actual Multiple Monitors has an option for ""Minimize All command affects only the monitor with mouse"".


![image](https://i.stack.imgur.com/hqRS5.jpg)


"
"Convert SQL Binary to byte array
Given some data stored in a SQL binary field:



```
0x83C8BB02E96F2383870CC1619B6EC...

```

I'd like to convert it into a byte array, but it doesn't seem like I can just cast this directly into a byte like so:



```
byte[] someBytes = (byte) 0x83C8BB02E96F2383870CC1619B6EC...;

```

What am I missing here?


","The data stored in the SQL field is in binary. The '0x83..' string you quoted is just a hexadecimal representation of that binary data.


If you just want to copy/paste the hexadecimal data into your C# code (as you seem to have written), then you'll need to convert it from hexadecimal to binary. .NET provides a (rather obscure) class for this sort of thing:



```
using System.Runtime.Remoting.Metadata.W3cXsd2001.SoapHexBinary

public byte[] ToBytes(string hex)
{
    var shb = SoapHexBinary.Parse(hex);
    return shb.Value;
}

public void TestConvert()
{
    byte[] someBytes = ToBytes(""83C8BB02E96F2383870CC1619B6EC"");
}

```

If your binary data is coming from your database in a different form than a copy/pasted string, you'll need to provide more information about your setup.


"
"Avoid else condition in 'When' in kotlin
As per documentation of [When](https://kotlinlang.org/docs/reference/control-flow.html#when-expression) in Kotlin, else is not mandatory if the compiler knows all the values are covered. This is very in case of emums or sealed class but how to do it in case of arrays for numbers 1 to 5 (startRating).



```
private fun starMapping(startRating: Int): String {

    return when (startRating) {
        1 -> ""Perfect""
        2 -> ""Great""
        3-> ""Okay""
        4-> ""Bad""
        5-> ""Terrible""
        // don't want to add else as I believe it is prone to errors.
    }
}

```

Something similar to this



```
return when (AutoCompleteRowType.values()[viewType]) {
        AutoCompleteRowType.ITEM -> ItemView(
                LayoutInflater.from(parent.context).inflate(R.layout.item_venue_autocomplete_item_info, parent, false))

        AutoCompleteRowType.SECTION -> SectionView(
                LayoutInflater.from(parent.context).inflate(R.layout.item_venue_autocomplete_section, parent, false)
        )
    }

```

","Using `when` statement it is impossible to exclude `else` clause in case of using ints, because compiler doesn't know what to return if `startRating` is not in 1..5 range. You can, for example, throw an `IllegalStateException` if the value is not in the required range:



```
private fun starMapping(startRating: Int): String {
    return when (startRating) {
        1 -> ""Perfect""
        2 -> ""Great""
        3-> ""Okay""
        4-> ""Bad""
        5 -> ""Terrible""
        else -> throw IllegalStateException(""Invalid rating param value"")
    }
}

```

Or you can do something like this:



```
return when {
    startRating <= 1 -> ""Perfect""
    startRating == 2 -> ""Great""
    startRating == 3 -> ""Okay""
    startRating == 4 -> ""Bad""
    else -> ""Terrible""
}

```

But `else` clause is required.


"
"Correct daemon behaviour (from PEP 3143) explained
I have some tasks [for my RPi] in Python that involve a lot of `sleep`ing: do something that takes a second or two or three, then go wait for several minutes or hours. 
I want to pass control back to the OS (Linux) in that sleep time. For this, I should daemonise those tasks. One way is by using Python's Standard daemon process library. 


But daemons aren't so easy to understand. As per the Rationale paragraph of [PEP 3143](http://legacy.python.org/dev/peps/pep-3143/#correct-daemon-behaviour), a well behaved daemon should do the following. 


- Close all open file descriptors.
- Change current working directory.
- Reset the file access creation mask.
- Run in the background.
- Disassociate from process group.
- Ignore terminal I/O signals.
- Disassociate from control terminal.
- Don't reacquire a control terminal.
- Correctly handle the following circumstances:
	- Started by System V init process.
	- Daemon termination by SIGTERM signal.
	- Children generate SIGCLD signal.


For a Linux/Unix novice like me, some of this is hardly an explanation. But I want to know why I do what I do. So what is the rationale behind this rationale? 


","PEP 3142 took these requirements from [Unix Network Programming](http://books.google.nl/books/about/UNIX_Network_Programming.html?id=ptSC4LpwGA0C) ('UNP') by the late W. Richard Stevens. The explanation below is quoted or summarised from that book. It's not so easily found online, and it may be illegal to download. So I borrowed it from the library. Pages referred to are in the second edition, Volume 1 (1998). (The PEP refers to the first edition, 1990.) 


**Close all open file descriptors.** 



> 
> ""We close any open descriptors inherited from the process that executed the daemon (ie the shell). [..] Some daemons open `/dev/null` for reading and writing and duplicate the descriptor to standard input, standard output and standard error."" 
> 
> 
> 


(This ['Howdy World' Python daemon](https://stackoverflow.com/questions/473620/how-do-you-create-a-daemon-in-python/9047339#9047339) demonstrates this.) 



> 
> ""This guarantees that the common descriptors are open, and a read from any of these descriptors returns 0 (End Of File) and the kernel just discards anything written to any of these three descriptors. The reason for opening these descriptors is so that any library function called by the daemon that assumes it can read from standard input or write to standard output or standard error, will not fail. Alternately, some daemons open a log file that they will write to while running and duplicate its descriptor to standard output and standard error"". (UNP p. 337) 
> 
> 
> 


**Change current working directory**



> 
> ""A printer daemon might change to the printer's spool directory, where it does all its work. [...] The daemon could have been started anywhere in the filesystem, and if it remains there, that filesystem cannot be unmounted."" (UNP p 337) 
> 
> 
> 


Why would you want to unmount a filesystem? Two reasons:  

1. You want to separate (and be able mount and unmount) directories that can fill up with user data from directories dedicated to the OS.  

2. If you start a daemon from, say, a USB-stick, you want to be able to unmount that stick without interfering with the daemon. 


**Reset the file access creation mask.**



> 
> ""So that if the daemon creates its own files, permission bits in the inherited file mode creation mask do not affect the permission bits of the new files."" (UNP, p 337) 
> 
> 
> 


**Run in the background.**  

By definition, 



> 
> ""a daemon is a process that runs in the background and is independent of control from all terminals"". (UNP p 331) 
> 
> 
> 


**Disassociate from process group.**  

In order to understand this, you need to understand what a process group is, and that means you need to know what `fork` does. 


*What fork does* 


`fork` is the only way (in Unix) to create a new process. (in Linux, there is also `clone`). Key in understanding `fork` is that it returns *twice* when called (once): once in the calling process (= parent) with the process ID of the newly created process (= child), and once in the child. ""All descriptors known by the parent when forking, are shared with the child when fork returns."" (UNP p 102). 
When a process wants to execute another program, it creates a new process by calling fork, which creates a copy of itself. Then one of them (usually the child) calls the new program. (UNP, p 102) 


*Why disassociate from process group* 


The point is that a session leader may acquire a controlling terminal. A daemon should never do this, it must stay in the background. This is achieved by calling `fork` twice: the parent forks to create a child, the child forks to create a grandchild. Parent and child are terminated, but grandchild remains. But because it's a grandchild, it's not a session leader, and therefor can't acquire a controlling terminal. (Summarised from UNP par 12.4 p 335) 


The double fork is discussed in more detail [here](https://stackoverflow.com/questions/881388/what-is-the-reason-for-performing-a-double-fork-when-creating-a-daemon), and in the comments below. 


**Ignore terminal I/O signals.** 



> 
> ""Signals generated from terminal keys must not affect any daemons started from that terminal earlier"". (UNP p. 331) 
> 
> 
> 


**Disassociate from control terminal and don't reacquire a control terminal.**  

By now, the reasons are obvious: 



> 
> ""If the daemon is started from a terminal, we want to be able to use that terminal for other tasks at a later time. For example, if we start the daemon from a terminal, log off the terminal, and someone else logs in on that terminal, we do not want any daemon error messages appearing during the next user's terminal session."" (UNP p 331) 
> 
> 
> 


**Correctly handle the following circumstances:** 


- *Started by System V init process* 


	- A daemon should be launchable at boot time, obviously.
- *Daemon termination by SIGTERM signal* 


	- SIGTERM means Signal Terminate. At shutdown, the init process normally sends SIGTERM to all processess, waits usually 5 to 20 seconds to give them time to clean up and terminate. (UNP, p 135) Also, a child can send SIGTERM to its parent, when its parent should stop doing what it's doing. (UNP p 408)
- *Children generate SIGCLD signal* 


	- Stevens discusses SIGCHLD, not SIGCLD. The difference between them isn't important for understanding daemon behaviour. If a child terminates, it sends SIGCHLD to it's parent. If a parent doesn't catch it, the child becomes a zombie (UNP p 118). Oh what fun.


On a final note, when I started to find answers to my question in UNP, it soon struck me I really should read more of it. It's 900+ (!) pages, from 1998 (!) but I believe the concepts and the explanations in UNP stand the test of time, gloriously. Stevens not only knew very well what he was talking about, he also understood what was difficult about it, and made it easier to understand. That's really rare. 


"
"Difference between /.../ and m/.../ in Perl
What is difference between `/.../` and `m/.../`?



```
use strict;
use warnings;

my $str = ""This is a testing for modifier"";

if ($str =~ /This/i) { print ""Modifier...\n"";  }
if ($str =~ m/This/i) { print ""W/O Modifier...\n"";  }

```

However, I checked with this site for [Reference](https://perldoc.perl.org/perlre.html#Modifiers) not clearly understand with the theory


","There's no difference. If you just supply `/PATTERN/` then it assumes `m`. However, if you're using an alternative delimiter, you need to supply the `m`. E.g. `m|PATTERN|` won't work as `|PATTERN|`. 


In your example, `i` is the modifier as it's after the pattern. `m` is the operation. (as opposed to `s`, `tr`, `y` etc.)


Perhaps slightly confusingly - you *can* use `m` as a modifier, but only if you put if *after* the match. 


`m/PATTERN/m` will cause `^` and `$` to match differently than in `m/PATTERN/`, but it's the trailing `m` that does this, not the leading one. 


"
"What does `~` mean in Haskell?
I'm studying the `mtl` library and trying to do some MonadTransformers of my own. I was checking the `Control.Monad.State.StateT` declaration, and across all the code, I see this syntax:



```
execStateT :: (Monad m) => StateT s m a -> s -> m s
execStateT m s = do
  ~(_, s') <- runStateT m s
  return s'

```

What does this `~` operand mean?


","This is the notation for a lazy pattern in Haskell. I can't say that I'm familiar with it but [from here](http://www.haskell.org/tutorial/patterns.html):



> 
> It is called a lazy pattern, and has
>  the form ~pat. Lazy patterns are
>  irrefutable: matching a value v
>  against ~pat always succeeds,
>  regardless of pat. Operationally
>  speaking, if an identifier in pat is
>  later ""used"" on the right-hand-side,
>  it will be bound to that portion of
>  the value that would result if v were
>  to successfully match pat, and ⊥
>  otherwise.
> 
> 
> 


Also, [this section](http://en.wikibooks.org/wiki/Haskell/Laziness#When_does_it_make_sense_to_use_lazy_patterns.3F) may be useful.


"
"Locking down Firebase DB access to specific apps
I think from searching the web this is not technically possible but I want to ask again in case I'm missing something.


I have an app that uses Firebase. Reading and writing is locked down through security rules for authorised users only but there's certain information I want unauthorised users to be able to access (so I don't have to put a login wall in front of them, influencing churn).


What I want to know is, is there any way of locking down this read access that only *my* app can call the DB? I know I can lock down domains to prevent someone writing localhost scrapers but what's to stop someone cloning and re-skinning an app and pointing it to the same back end? Is it possible to achieve this using your certificates fingerprint?


","There is no way to limit access to your database to just your app. That just doesn't match with the cloud-based nature of the Firebase APIs. Anyone that knows the URL of your database can in in principle access it, and security rule are **the** way to ensure all access is authorized.


Note that security rules are not an all-or-nothing approach: you can require sign-in for some parts of your database, while leaving other parts publicly readable. But you *can't* make the publicly readable parts only be readable by your own app.


Some previous questions on the same topic:


- [how to make sure only my own website (clientside code) can talk to Firebase backend?](https://stackoverflow.com/questions/18890330/how-to-make-sure-only-my-own-website-clientside-code-can-talk-to-firebase-back) (pretty much my go-to answer for this)
- [How to allow only my app to access firebase without a login?](https://stackoverflow.com/questions/38395025/how-to-allow-only-my-app-to-access-firebase-without-a-login)
- [Restrict Firebase database access to one Android app](https://stackoverflow.com/questions/41818782/restrict-firebase-database-access-to-one-android-app)
- [How to allow only my app to access firebase without a login?](https://stackoverflow.com/questions/38395025/how-to-allow-only-my-app-to-access-firebase-without-a-login/49284538)


**Update**: since May 2021 you **can** actually restrict access to just users of *your* App by implementing [Firebase App Check](https://firebase.google.com/docs/app-check).


"
"Reset checkboxes to false
I have a Google spreadsheet where column A has checkboxes in each row. I have written a script to perform a function on all rows where the checkboxes are checked, but I want to add in at the end a reset function so that all checked boxes are unchecked again after the script is run.


I've tried using a for loop like this:



```
var dataRange = sheet.getRange('A3:A');
var values = dataRange.getValues();

for (var i = 0; i < values.length; i++) {
  for (var j = 0; j < values[i].length; j++) {     
    if (values[i][j] == true) {
      values[i].setValue(false);
    }
  }    
} 

```

But clearly this doesn't work as I get an error.


Does anyone know how this could be done?


","How about this modification? I think that there are several solutions for your situation. So please think of this as one of them.


### Modification points :


- The reason of the issue is `values[i].setValue(false);`. `values[i]` is an array. Please use the range for `setValue()`.
	- But to use `setValue()` in the for loop leads to higher cost. So in this modification, I used `setValues()`.
- Put ""false"" to `values`, if `values[i][j]` is ""true"".
- Put the modified values to the sheet using `setValues()`.


### Modified script :



```
var dataRange = sheet.getRange('A3:A');
var values = dataRange.getValues();
for (var i = 0; i < values.length; i++) {
  for (var j = 0; j < values[i].length; j++) {
    if (values[i][j] == true) {
      values[i][j] = false; // Modified
    }
  }
}
dataRange.setValues(values); // Added

```

### Reference :


- [setValue()](https://developers.google.com/apps-script/reference/spreadsheet/range#setValue(Object))
- [setValues()](https://developers.google.com/apps-script/reference/spreadsheet/range#setvaluesvalues)


If this was not what you want, please tell me. I would like to modify it.


"
"Multiplatform crash reporting?
What would be the best way to implement crash reporting in a cross-platform app in c++ (windows, mac, linux)? Looked into google breakpad, but spent a few weeks trying to make socorro (UI) work to no avail.


edit: @Nim to clear up, I want to send a stack-trace to my server when some exception occurs.


","When we went crossplatform (from being previously Windows-only), we spent a lot of time researching the possible tools for exactly this. Google Breakpad really is the best crossplatform solution out there. 


Your problem is Socorro, which is huge, obfuscated, and confusing. I can suggest two routes:


- Drop the current version of Socorro, which is way too much in love with newfangled OSS toys like Hadoop, and roll back to Socorro 1.x. You'll need to do a bit more work to set up your processors, and your dumps will go into a Postgres table, but you'll have way fewer moving parts and crazy infrastructure to deal with.
- Forget about Socorro altogether and write your own processors for the Breakpad dumps and a UI to view them with. Breakpad has already done the really hard work -- parsing the symbols and catching the exceptions -- so all you need to write is a collector that the dumps can get uploaded to, a farm to run the Breakpad processors on them, someplace to store the processed results, and something to view them.


Knowing what we do now, we would have probably gone with the second option rather than trying to integrate Socorro — the amount of time it's taken us to find a version of Socorro that works for us and adapt it to our workflow has already exceeded what it would have taken for us to just write a system that does exactly what we want.


"