Here's my solution to the issue. Tried to keep it straight and simple.

#include <iostream>
using namespace std;

class Game{
public:
  void init(){
    cout << "Hi" << endl;
  }
}g;

class b : Game{  //class b uses/imports class Game
public:
  void h(){
    init(); //Use function from class Game
  }
}A;

int main()
{
  A.h();
  return 0;
}

You've only got a forward declaration for student in the header file; you need to place the struct declaration in the header file, not the .cpp. The method definitions will be in the .cpp (assuming you have any).

I used postman chrome extension until it became deprecated. Chrome extension also less usable and powerful then native postman application. So, it became not very convenient to use chrome extension. I have found next approach:

1. copy any request in chrome/any other browser as CURL request (image 1)
2. import to postman copied request (image 2)
3. save imported request in postman's list

image 1

image 2

An instance of Date always represents a date and cannot be empty. You can use a null value of date2 to represent "there is no date". You will have to check for null whenever you use date2 to avoid a NullPointerException, like when rendering your page:

if (date2 != null)
  // print date2
else
  // print nothing, or a default value

Let's fit the model:

> library(ISwR)
> fit <- lm(metabolic.rate ~ body.weight, rmr)
> summary(fit)

Call:
lm(formula = metabolic.rate ~ body.weight, data = rmr)

Residuals:
    Min      1Q  Median      3Q     Max 
-245.74 -113.99  -32.05  104.96  484.81 

Coefficients:
            Estimate Std. Error t value Pr(>|t|)    
(Intercept) 811.2267    76.9755  10.539 2.29e-13 ***
body.weight   7.0595     0.9776   7.221 7.03e-09 ***
---
Signif. codes:  0 ‘***’ 0.001 ‘**’ 0.01 ‘*’ 0.05 ‘.’ 0.1 ‘ ’ 1 

Residual standard error: 157.9 on 42 degrees of freedom
Multiple R-squared: 0.5539, Adjusted R-squared: 0.5433 
F-statistic: 52.15 on 1 and 42 DF,  p-value: 7.025e-09 

The 95% confidence interval for the slope is the estimated coefficient (7.0595) ± two standard errors (0.9776).

This can be computed using confint:

> confint(fit, 'body.weight', level=0.95)
               2.5 % 97.5 %
body.weight 5.086656 9.0324

I had this issue when i cloned from a git repository and I had it resolved when I created a new project and re-inserted the src folder from the old project.

The src folder is the only folder needed when deploying your angular application but you must reconfigure the dev environment, using this solution.

Another excellent way of dealing with this is to create your TEXT field without the unique constraint and add a sibling VARCHAR field that is unique and contains a digest (MD5, SHA1, etc.) of the TEXT field. Calculate and store the digest over the entire TEXT field when you insert or update the TEXT field then you have a uniqueness constraint over the entire TEXT field (rather than some leading portion) that can be searched quickly.

[self.navigationController.navigationBar.layer setZPosition:-0.1];
UIView *view = [[UIView alloc]initWithFrame:CGRectMake(10, 20, 35, 35)];
[view setBackgroundColor:[UIColor redColor]];
[self.navigationController.view addSubview:view];
[self.navigationController.view bringSubviewToFront:view];
self.navigationController.view.clipsToBounds = NO;
[self.navigationController.navigationBar.layer setZPosition:0.0];

Sorry for necromancing the thread, but whenever you iterate over files by globbing, it's good practice to avoid the corner case where the glob does not match (which makes the loop variable expand to the (un-matching) glob pattern string itself).

For example:

for filename in Data/*.txt; do
    [ -e "$filename" ] || continue
    # ... rest of the loop body
done

Reference: Bash Pitfalls

Since you have created an image from the Dockerfile, the image currently is not in active state. In order to work you need to run this image inside a container.

The $ docker images command describes how many images are currently available in the local repository. and

docker ps -a

shows how many containers are currently available, i.e. the list of active and exited containers.

There are two ways to run the image in the container:

$ docker run [OPTIONS] IMAGE[:TAG|@DIGEST] [COMMAND] [ARG...]

In detached mode:

-d=false: Detached mode: Run container in the background, print new container id

In interactive mode:

-i :Keep STDIN open even if not attached

Here is the Docker run command

$ docker run image_name:tag_name

For more clarification on Docker run, you can visit Docker run reference.

It's the best material to understand Docker.

Expanding on eumiro's comment, normally tuple(l) will convert a list l into a tuple:

In [1]: l = [4,5,6]

In [2]: tuple
Out[2]: <type 'tuple'>

In [3]: tuple(l)
Out[3]: (4, 5, 6)

However, if you've redefined tuple to be a tuple rather than the type tuple:

In [4]: tuple = tuple(l)

In [5]: tuple
Out[5]: (4, 5, 6)

then you get a TypeError since the tuple itself is not callable:

In [6]: tuple(l)
TypeError: 'tuple' object is not callable

You can recover the original definition for tuple by quitting and restarting your interpreter, or (thanks to @glglgl):

In [6]: del tuple

In [7]: tuple
Out[7]: <type 'tuple'>

You are implementing LocationListener in your activity MainActivity. The call for concurrent location updates will therefor be like this:

mLocationClient.requestLocationUpdates(mLocationRequest, this);

Be sure that the LocationListener you're implementing is from the google api, that is import this:

import com.google.android.gms.location.LocationListener;

and not this:

import android.location.LocationListener;

and it should work just fine.

It's also important that the LocationClient really is connected before you do this. I suggest you don't call it in the onCreate or onStart methods, but in onResume. It is all explained quite well in the tutorial for Google Location Api: https://developer.android.com/training/location/index.html

Change MySQL character:

Client

default-character-set=utf8

mysqld

character_set_server=utf8

We should not write default-character-set=utf8 in mysqld, because that could result in an error like:

start: Job failed to start

At last:

 +--------------------------+----------------------------+
 | Variable_name            | Value                      |
 +--------------------------+----------------------------+
 | character_set_client     | utf8                       |
 | character_set_connection | utf8                       |
 | character_set_database   | utf8                       |
 | character_set_filesystem | binary                     |
 | character_set_results    | utf8                       |
 | character_set_server     | utf8                       |
 | character_set_system     | utf8                       |
 | character_sets_dir       | /usr/share/mysql/charsets/ |
 +--------------------------+----------------------------+

For creating real UML class diagrams:

In Visual Studio 2013 Ultimate you can do this without any external tools.

• In the menu, click on Architecture, New Diagram Select UML Class Diagram
• This will ask you to create a new Modeling Project if you don't have one already.

You will have a empty UMLClassDiagram.classdiagram.

• Again, go to Architecture, Windows, Architecture Explorer.
• A window will pop up with your namespaces, Choose Class View.
• Then a list of sub-namespaces will appear, if any. Choose one, select the classes and drag them to the empty UMLClassDiagram1.classdiagram window.

Reference: Create UML Class Diagrams from Code

When an iFrame loads, it initially contains the #document, so checking the load state might best work by checking what's there now..

if ($('iframe').contents().find('body').children().length > 0) {
    // is loaded
} else {
    // is not loaded
}

I think this does what you are asking for:

webdeveloper.com - decimal to fraction

Width/height gets you a decimal, converted to a fraction with ":" in place of '/' gives you a "ratio".

@AVeryDev

6) To attach the loaded view to your view controller's view:

[self.view addSubview:myViewFromNib];

Presumably, it is necessary to remove it from the view to avoid memory leaks.

To clarify: the view controller has several IBOutlets, some of which are connected to items in the original nib file (as usual), and some are connected to items in the loaded nib. Both nib's have the same owner class. The loaded view overlays the original one.

Hint: set the opacity of the main view in the loaded nib to zero, then it won't obscure the items from the original nib.

Try this to get the month in number form:

Month(DateAdd("m", -3, Now))

It will give you 12 for December.

So in your case you would use Month(DateAdd("m", -1, Now)) to just subract one month.

The file may be existing but may have a different path. Try writing the absolute path for the file.

Try os.listdir() function to check that atleast python sees the file.

Try it like this:

file1 = open(r'Drive:\Dir\recentlyUpdated.yaml')

cv:Mat mat;
int rows = mat.rows;
int cols = mat.cols;

cv::Size s = mat.size();
rows = s.height;
cols = s.width;

Also note that stride >= cols; this means that actual size of the row can be greater than element size x cols. This is different from the issue of continuous Mat and is related to data alignment.

Tried other variants, this can be proper

    Logger logger = Logger.getLogger(MyClass.class.getName());        
    Level level = Level.ALL;
    for(Handler h : java.util.logging.Logger.getLogger("").getHandlers())    
        h.setLevel(level);
    logger.setLevel(level);
// this must be shown
    logger.fine("fine");
    logger.info("info");

There's a very simple way if you got your APK allready on your Smartphone. Just use one of these APPs:

Package Name Viewer Apps

1: No.

2: As a short answer: The 65th character ("=" sign) is used only as a complement in the final process of encoding a message.

You will not have a '=' sign if your string has a multiple of 3 characters number, because Base64 encoding takes each three bytes (8 bits) and represents them as four printable characters in the ASCII standard.

Details:

(a) If you want to encode

ABCDEFG <=> [ABC] [DEF] [G

Base64 will deal with the first block (producing 4 characters) and the second (as they are complete). But for the third it will add a double == in the output in order to complete the 4 needed characters. Thus, the result will be QUJD REVG Rw== (without spaces).

(b) If you want to encode

ABCDEFGH <=> [ABC] [DEF] [GH

similarly, it will add just a single = in the end of the output to get 4 characters.

The result will be QUJD REVG R0g= (without spaces).

if (iv_new_pwd.isSelected()) {
                iv_new_pwd.setSelected(false);
                Log.d("mytag", "in case 1");
                edt_new_pwd.setInputType(InputType.TYPE_CLASS_TEXT);
            } else {
                Log.d("mytag", "in case 1");
                iv_new_pwd.setSelected(true);
                edt_new_pwd.setInputType(InputType.TYPE_CLASS_TEXT | InputType.TYPE_TEXT_VARIATION_PASSWORD);
            }

I've also faced a similar situation where I needed a script which can take care of javac and then java(ing) my java program. So, I came up with this BATCH script.

::  @author Rudhin Menon
::  Created on 09/06/2015
::
::  Auto-Concrete is a build tool, which monitor the file under
::  scrutiny for any changes, and compiles or runs the same once
::  it got changed.
::  
::  ========================================
::  md5sum and gawk programs are prerequisites for this script.
::  Please download them before running auto-concrete.
::  ========================================
::  
::  Happy coding ...

@echo off

:: if filename is missing
if [%1] EQU [] goto usage_message

:: Set cmd window name
title Auto-Concrete v0.2

cd versions

if %errorlevel% NEQ 0 (
    echo creating versions directory
    mkdir versions
    cd versions
)

cd ..

javac "%1"

:loop
    :: Get OLD HASH of file
    md5sum "%1" | gawk '{print $1}' > old
    set /p oldHash=<old
    copy "%1" "versions\%oldHash%.java"

    :inner_loop
    :: Get NEW HASH of the same file
    md5sum "%1" | gawk '{print $1}' > new
    set /p newHash=<new

    :: While OLD HASH and NEW HASH are the same
    :: keep comparing OLD HASH and NEW HASH
    if "%newHash%" EQU "%oldHash%" (
        :: Take rest before proceeding
        ping -w 200 0.0.0.0 >nul
        goto inner_loop
    )

    :: Once they differ, compile the source file
    :: and repeat everything again
    echo.
    echo ========= %1 changed on %DATE% at %TIME% ===========
    echo.       
    javac "%1"
goto loop

:usage_message
echo Usage : auto-concrete FILENAME.java

Above batch script will check the file for any changes and compile if any changes are done, you can tweak it for compiling whenever you want. Happy coding :)

The reason is that the data doesn't match the datatype. I have come across the same issues that I forgot to make the fields match. Though my case is not same as yours, but it shows the similar error message.

The situation is that I copy a table, but accidently I misspell one field, so I change it using the ALTER after creating the database. And the order of fields in both table is not identical. so when I use the INSERT INTO TableName SELECT * FROM TableName, the result showed the similar errors: Operand type clash: datetime is incompatible with uniqueidentifier

This is a simiple example:

use example
go
create table Test1 (
    id int primary key,
    item uniqueidentifier,
    inserted_at datetime
    )
go
create table Test2 (
    id int primary key,
    inserted_at datetime
    )
go
alter table Test2 add item uniqueidentifier;
go

--insert into Test1 (id, item, inserted_at) values (1, newid(), getdate()), (2, newid(), getdate());


insert into Test2 select * from Test1;

select * from Test1;
select * from Test2;

The error message is:

Msg 206, Level 16, State 2, Line 24
Operand type clash: uniqueidentifier is incompatible with datetime

I believe this is controlled by the css color property applied to the element.

This solution is by far the best working one. In your manifest file add

<activity
    android:configChanges="keyboardHidden|orientation|screenSize"
    android:name="your activity name"
    android:label="@string/app_name" 
    android:screenOrientation="landscape">
  </activity

And in your activity class add the following code

@Override
public void onConfigurationChanged(Configuration newConfig) 
{
    super.onConfigurationChanged(newConfig);
    if (newConfig.orientation == Configuration.ORIENTATION_PORTRAIT) {
        //your code
    } else if (newConfig.orientation == Configuration.ORIENTATION_LANDSCAPE) {
  //your code

    }
}

Code coverage tools, such as Emma, Cobertura, and Clover, will instrument your code and record which parts of it gets invoked by running a suite of tests. This is very useful, and should be an integral part of your development process. It will help you identify how well your test suite covers your code.

However, this is not the same as identifying real dead code. It only identifies code that is covered (or not covered) by tests. This can give you false positives (if your tests do not cover all scenarios) as well as false negatives (if your tests access code that is actually never used in a real world scenario).

I imagine the best way to really identify dead code would be to instrument your code with a coverage tool in a live running environment and to analyse code coverage over an extended period of time.

If you are runnning in a load balanced redundant environment (and if not, why not?) then I suppose it would make sense to only instrument one instance of your application and to configure your load balancer such that a random, but small, portion of your users run on your instrumented instance. If you do this over an extended period of time (to make sure that you have covered all real world usage scenarios - such seasonal variations), you should be able to see exactly which areas of your code are accessed under real world usage and which parts are really never accessed and hence dead code.

I have never personally seen this done, and do not know how the aforementioned tools can be used to instrument and analyse code that is not being invoked through a test suite - but I am sure they can be.

An alternative way is to use DateTime.DaysInMonth to get the number of days in the current month as suggested by @Jade

Since we know the first day of the month will always 1 we can use it as default for the first day with the current Month & year as current.year,current.Month,1.

var now = DateTime.Now; // get the current DateTime 

//Get the number of days in the current month
int daysInMonth = DateTime.DaysInMonth (now.Year, now.Month); 
    
//First day of the month is always 1
var firstDay = new DateTime(now.Year,now.Month,1); 
    
//Last day will be similar to the number of days calculated above
var lastDay = new DateTime(now.Year,now.Month,daysInMonth);

//So
rdpStartDate.SelectedDate = firstDay;
rdpEndDate.SelectedDate = lastDay; 

And if you want to create the list with some elements to start with:

var emailList = new List<EmailData>
{
   new EmailData { FirstName = "John", LastName = "Doe", Location = "Moscow" },
   new EmailData {.......}
};

You have this simple command:

$ free -h

You need to double check the PATH environment setting. C:\Program Files\Java\jdk-13 you currently have there is not correct. Please make sure you have the bin subdirectory for the latest JDK version at the top of the PATH list.

java.exe executable is in C:\Program Files\Java\jdk-13\bin directory, so that is what you need to have in PATH.

Use this tool to quickly verify or edit the environment variables on Windows. It allows to reorder PATH entries. It will also highlight invalid paths in red.

If you want your code to run on lower JDK versions as well, change the target bytecode version in the IDE. See this answer for the relevant screenshots.

See also this answer for the Java class file versions. What happens is that you build the code with Java 13 and 13 language level bytecode (target) and try to run it with Java 8 which is the first (default) Java version according to the PATH variable configuration.

The solution is to have Java 13 bin directory in PATH above or instead of Java 8. On Windows you may have C:\Program Files (x86)\Common Files\Oracle\Java\javapath added to PATH automatically which points to Java 8 now:

If it's the case, remove the highlighted part from PATH and then logout/login or reboot for the changes to have effect. You need to Restart as administrator first to be able to edit the System variables (see the button on the top right of the system variables column).

Found a solution not involving Flex, because Flex doesn't work in older Browsers. Example:

.container {
    display:block;
    position:relative;
    height:150px;
    width:1024px;
    margin:0 auto;
    padding:0px;
    border:0px;
    background:#ececec;
    margin-bottom:10px;
    text-align:justify;
    box-sizing:border-box;
    white-space:nowrap;
    font-size:0pt;
    letter-spacing:-1em;
}

.cols {
    display:inline-block;
    position:relative;
    width:32%;
    height:100%;
    margin:0 auto;
    margin-right:2%;
    border:0px;
    background:lightgreen;  
    box-sizing:border-box;
    padding:10px;
    font-size:10pt;
    letter-spacing:normal;
}

.cols:last-child {
    margin-right:0;
}

In your specific case, you can set the containing a element to be:

a {
    display: block;
    text-align: center;
}

JS Bin demo.

Try:

const result = Math.round((data.expense / data.income) * 100)

What worked for me was to import "github.com/gorilla/handlers" and then use it this way:

router := mux.NewRouter()
router.HandleFunc("/config", getConfig).Methods("GET")
router.HandleFunc("/config/emcServer", createEmcServers).Methods("POST")

headersOk := handlers.AllowedHeaders([]string{"X-Requested-With", "Content-Type"})
originsOk := handlers.AllowedOrigins([]string{"*"})
methodsOk := handlers.AllowedMethods([]string{"GET", "HEAD", "POST", "PUT", "OPTIONS"})

log.Fatal(http.ListenAndServe(":" + webServicePort, handlers.CORS(originsOk, headersOk, methodsOk)(router)))

As soon as I executed an Ajax POST request and attaching JSON data to it, Chrome would always add the Content-Type header which was not in my previous AllowedHeaders config.

I have a cell which is linked to online stock database and updated frequently. I want to trigger a macro whenever the cell value is updated.

I believe this is similar to cell value change by a program or any external data update but above examples somehow do not work for me. I think the problem is because excel internal events are not triggered, but thats my guess.

I did the following,

Private Sub Worksheet_Change(ByVal Target As Range) 
  If Not Intersect(Target, Target.Worksheets("Symbols").Range("$C$3")) Is Nothing Then
   'Run Macro
End Sub

Softwares run on the OS on a very simple premise - they require memory. The device OS provides it in the form of RAM. The amount of memory required may vary - some softwares need huge memory, some require paltry memory. Most (if not all) users run multiple applications on the OS simultaneously, and given that memory is expensive (and device size is finite), the amount of memory available is always limited. So given that all softwares require a certain amount of RAM, and all of them can be made to run at the same time, OS has to take care of two things:

1. That the software always runs until user aborts it, i.e. it should not auto-abort because OS has run out of memory.
2. The above activity, while maintaining a respectable performance for the softwares running.

Now the main question boils down to how the memory is being managed. What exactly governs where in the memory will the data belonging to a given software reside?

Possible solution 1: Let individual softwares specify explicitly the memory address they will use in the device. Suppose Photoshop declares that it will always use memory addresses ranging from 0 to 1023 (imagine the memory as a linear array of bytes, so first byte is at location 0, 1024th byte is at location 1023) - i.e. occupying 1 GB memory. Similarly, VLC declares that it will occupy memory range 1244 to 1876, etc.

Advantages:

1. Every application is pre-assigned a memory slot, so when it is installed and executed, it just stores its data in that memory area, and everything works fine.

Disadvantages:

1. This does not scale. Theoretically, an app may require a huge amount of memory when it is doing something really heavy-duty. So to ensure that it never runs out of memory, the memory area allocated to it must always be more than or equal to that amount of memory. What if a software, whose maximal theoretical memory usage is 2 GB (hence requiring 2 GB memory allocation from RAM), is installed in a machine with only 1 GB memory? Should the software just abort on startup, saying that the available RAM is less than 2 GB? Or should it continue, and the moment the memory required exceeds 2 GB, just abort and bail out with the message that not enough memory is available?

2. It is not possible to prevent memory mangling. There are millions of softwares out there, even if each of them was allotted just 1 kB memory, the total memory required would exceed 16 GB, which is more than most devices offer. How can, then, different softwares be allotted memory slots that do not encroach upon each other's areas? Firstly, there is no centralized software market which can regulate that when a new software is being released, it must assign itself this much memory from this yet unoccupied area, and secondly, even if there were, it is not possible to do it because the no. of softwares is practically infinite (thus requiring infinite memory to accommodate all of them), and the total RAM available on any device is not sufficient to accommodate even a fraction of what is required, thus making inevitable the encroaching of the memory bounds of one software upon that of another. So what happens when Photoshop is assigned memory locations 1 to 1023 and VLC is assigned 1000 to 1676? What if Photoshop stores some data at location 1008, then VLC overwrites that with its own data, and later Photoshop accesses it thinking that it is the same data is had stored there previously? As you can imagine, bad things will happen.

So clearly, as you can see, this idea is rather naive.

Possible solution 2: Let's try another scheme - where OS will do majority of the memory management. Softwares, whenever they require any memory, will just request the OS, and the OS will accommodate accordingly. Say OS ensures that whenever a new process is requesting for memory, it will allocate the memory from the lowest byte address possible (as said earlier, RAM can be imagined as a linear array of bytes, so for a 4 GB RAM, the addresses range for a byte from 0 to 2^32-1) if the process is starting, else if it is a running process requesting the memory, it will allocate from the last memory location where that process still resides. Since the softwares will be emitting addresses without considering what the actual memory address is going to be where that data is stored, OS will have to maintain a mapping, per software, of the address emitted by the software to the actual physical address (Note: that is one of the two reasons we call this concept Virtual Memory. Softwares are not caring about the real memory address where their data are getting stored, they just spit out addresses on the fly, and the OS finds the right place to fit it and find it later if required).

Say the device has just been turned on, OS has just launched, right now there is no other process running (ignoring the OS, which is also a process!), and you decide to launch VLC. So VLC is allocated a part of the RAM from the lowest byte addresses. Good. Now while the video is running, you need to start your browser to view some webpage. Then you need to launch Notepad to scribble some text. And then Eclipse to do some coding.. Pretty soon your memory of 4 GB is all used up, and the RAM looks like this:

Problem 1: Now you cannot start any other process, for all RAM is used up. Thus programs have to be written keeping the maximum memory available in mind (practically even less will be available, as other softwares will be running parallelly as well!). In other words, you cannot run a high-memory consuming app in your ramshackle 1 GB PC.

Okay, so now you decide that you no longer need to keep Eclipse and Chrome open, you close them to free up some memory. The space occupied in RAM by those processes is reclaimed by OS, and it looks like this now:

Suppose that closing these two frees up 700 MB space - (400 + 300) MB. Now you need to launch Opera, which will take up 450 MB space. Well, you do have more than 450 MB space available in total, but...it is not contiguous, it is divided into individual chunks, none of which is big enough to fit 450 MB. So you hit upon a brilliant idea, let's move all the processes below to as much above as possible, which will leave the 700 MB empty space in one chunk at the bottom. This is called compaction. Great, except that...all the processes which are there are running. Moving them will mean moving the address of all their contents (remember, OS maintains a mapping of the memory spat out by the software to the actual memory address. Imagine software had spat out an address of 45 with data 123, and OS had stored it in location 2012 and created an entry in the map, mapping 45 to 2012. If the software is now moved in memory, what used to be at location 2012 will no longer be at 2012, but in a new location, and OS has to update the map accordingly to map 45 to the new address, so that the software can get the expected data (123) when it queries for memory location 45. As far as the software is concerned, all it knows is that address 45 contains the data 123!)! Imagine a process that is referencing a local variable i. By the time it is accessed again, its address has changed, and it won't be able to find it any more. The same will hold for all functions, objects, variables, basically everything has an address, and moving a process will mean changing the address of all of them. Which leads us to:

Problem 2: You cannot move a process. The values of all variables, functions and objects within that process have hardcoded values as spat out by the compiler during compilation, the process depends on them being at the same location during its lifetime, and changing them is expensive. As a result, processes leave behind big "holes" when they exit. This is called External Fragmentation.

Fine. Suppose somehow, by some miraculous manner, you do manage to move the processes up. Now there is 700 MB of free space at the bottom:

Opera smoothly fits in at the bottom. Now your RAM looks like this:

Good. Everything is looking fine. However, there is not much space left, and now you need to launch Chrome again, a known memory-hog! It needs lots of memory to start, and you have hardly any left...Except.. you now notice that some of the processes, which were initially occupying large space, now is not needing much space. May be you have stopped your video in VLC, hence it is still occupying some space, but not as much as it required while running a high resolution video. Similarly for Notepad and Photos. Your RAM now looks like this:

Holes, once again! Back to square one! Except, previously, the holes occurred due to processes terminating, now it is due to processes requiring less space than before! And you again have the same problem, the holes combined yield more space than required, but they are scattered around, not much of use in isolation. So you have to move those processes again, an expensive operation, and a very frequent one at that, since processes will frequently reduce in size over their lifetime.

Problem 3: Processes, over their lifetime, may reduce in size, leaving behind unused space, which if needed to be used, will require the expensive operation of moving many processes. This is called Internal Fragmentation.

Fine, so now, your OS does the required thing, moves processes around and start Chrome and after some time, your RAM looks like this:

Cool. Now suppose you again resume watching Avatar in VLC. Its memory requirement will shoot up! But...there is no space left for it to grow, as Notepad is snuggled at its bottom. So, again, all processes has to move below until VLC has found sufficient space!

Problem 4: If processes needs to grow, it will be a very expensive operation

Fine. Now suppose, Photos is being used to load some photos from an external hard disk. Accessing hard-disk takes you from the realm of caches and RAM to that of disk, which is slower by orders of magnitudes. Painfully, irrevocably, transcendentally slower. It is an I/O operation, which means it is not CPU bound (it is rather the exact opposite), which means it does not need to occupy RAM right now. However, it still occupies RAM stubbornly. If you want to launch Firefox in the meantime, you can't, because there is not much memory available, whereas if Photos was taken out of memory for the duration of its I/O bound activity, it would have freed lot of memory, followed by (expensive) compaction, followed by Firefox fitting in.

Problem 5: I/O bound jobs keep on occupying RAM, leading to under-utilization of RAM, which could have been used by CPU bound jobs in the meantime.

So, as we can see, we have so many problems even with the approach of virtual memory.

There are two approaches to tackle these problems - paging and segmentation. Let us discuss paging. In this approach, the virtual address space of a process is mapped to the physical memory in chunks - called pages. A typical page size is 4 kB. The mapping is maintained by something called a page table, given a virtual address, all now we have to do is find out which page the address belong to, then from the page table, find the corresponding location for that page in actual physical memory (known as frame), and given that the offset of the virtual address within the page is same for the page as well as the frame, find out the actual address by adding that offset to the address returned by the page table. For example:

On the left is the virtual address space of a process. Say the virtual address space requires 40 units of memory. If the physical address space (on the right) had 40 units of memory as well, it would have been possible to map all location from the left to a location on the right, and we would have been so happy. But as ill luck would have it, not only does the physical memory have less (24 here) memory units available, it has to be shared between multiple processes as well! Fine, let's see how we make do with it.

When the process starts, say a memory access request for location 35 is made. Here the page size is 8 (each page contains 8 locations, the entire virtual address space of 40 locations thus contains 5 pages). So this location belongs to page no. 4 (35/8). Within this page, this location has an offset of 3 (35%8). So this location can be specified by the tuple (pageIndex, offset) = (4,3). This is just the starting, so no part of the process is stored in the actual physical memory yet. So the page table, which maintains a mapping of the pages on the left to the actual pages on the right (where they are called frames) is currently empty. So OS relinquishes the CPU, lets a device driver access the disk and fetch the page no. 4 for this process (basically a memory chunk from the program on the disk whose addresses range from 32 to 39). When it arrives, OS allocates the page somewhere in the RAM, say first frame itself, and the page table for this process takes note that page 4 maps to frame 0 in the RAM. Now the data is finally there in the physical memory. OS again queries the page table for the tuple (4,3), and this time, page table says that page 4 is already mapped to frame 0 in the RAM. So OS simply goes to the 0th frame in RAM, accesses the data at offset 3 in that frame (Take a moment to understand this. The entire page, which was fetched from disk, is moved to frame. So whatever the offset of an individual memory location in a page was, it will be the same in the frame as well, since within the page/frame, the memory unit still resides at the same place relatively!), and returns the data! Because the data was not found in memory at first query itself, but rather had to be fetched from disk to be loaded into memory, it constitutes a miss.

Fine. Now suppose, a memory access for location 28 is made. It boils down to (3,4). Page table right now has only one entry, mapping page 4 to frame 0. So this is again a miss, the process relinquishes the CPU, device driver fetches the page from disk, process regains control of CPU again, and its page table is updated. Say now the page 3 is mapped to frame 1 in the RAM. So (3,4) becomes (1,4), and the data at that location in RAM is returned. Good. In this way, suppose the next memory access is for location 8, which translates to (1,0). Page 1 is not in memory yet, the same procedure is repeated, and the page is allocated at frame 2 in RAM. Now the RAM-process mapping looks like the picture above. At this point in time, the RAM, which had only 24 units of memory available, is filled up. Suppose the next memory access request for this process is from address 30. It maps to (3,6), and page table says that page 3 is in RAM, and it maps to frame 1. Yay! So the data is fetched from RAM location (1,6), and returned. This constitutes a hit, as data required can be obtained directly from RAM, thus being very fast. Similarly, the next few access requests, say for locations 11, 32, 26, 27 all are hits, i.e. data requested by the process is found directly in the RAM without needing to look elsewhere.

Now suppose a memory access request for location 3 comes. It translates to (0,3), and page table for this process, which currently has 3 entries, for pages 1, 3 and 4 says that this page is not in memory. Like previous cases, it is fetched from disk, however, unlike previous cases, RAM is filled up! So what to do now? Here lies the beauty of virtual memory, a frame from the RAM is evicted! (Various factors govern which frame is to be evicted. It may be LRU based, where the frame which was least recently accessed for a process is to be evicted. It may be first-come-first-evicted basis, where the frame which allocated longest time ago, is evicted, etc.) So some frame is evicted. Say frame 1 (just randomly choosing it). However, that frame is mapped to some page! (Currently, it is mapped by the page table to page 3 of our one and only one process). So that process has to be told this tragic news, that one frame, which unfortunate belongs to you, is to be evicted from RAM to make room for another pages. The process has to ensure that it updates its page table with this information, that is, removing the entry for that page-frame duo, so that the next time a request is made for that page, it right tells the process that this page is no longer in memory, and has to be fetched from disk. Good. So frame 1 is evicted, page 0 is brought in and placed there in the RAM, and the entry for page 3 is removed, and replaced by page 0 mapping to the same frame 1. So now our mapping looks like this (note the colour change in the second frame on the right side):

Saw what just happened? The process had to grow, it needed more space than the available RAM, but unlike our earlier scenario where every process in the RAM had to move to accommodate a growing process, here it happened by just one page replacement! This was made possible by the fact that the memory for a process no longer needs to be contiguous, it can reside at different places in chunks, OS maintains the information as to where they are, and when required, they are appropriately queried. Note: you might be thinking, huh, what if most of the times it is a miss, and the data has to be constantly loaded from disk into memory? Yes, theoretically, it is possible, but most compilers are designed in such a manner that follows locality of reference, i.e. if data from some memory location is used, the next data needed will be located somewhere very close, perhaps from the same page, the page which was just loaded into memory. As a result, the next miss will happen after quite some time, most of the upcoming memory requirements will be met by the page just brought in, or the pages already in memory which were recently used. The exact same principle allows us to evict the least recently used page as well, with the logic that what has not been used in a while, is not likely to be used in a while as well. However, it is not always so, and in exceptional cases, yes, performance may suffer. More about it later.

Solution to Problem 4: Processes can now grow easily, if space problem is faced, all it requires is to do a simple page replacement, without moving any other process.

Solution to Problem 1: A process can access unlimited memory. When more memory than available is needed, the disk is used as backup, the new data required is loaded into memory from the disk, and the least recently used data frame (or page) is moved to disk. This can go on infinitely, and since disk space is cheap and virtually unlimited, it gives an illusion of unlimited memory. Another reason for the name Virtual Memory, it gives you illusion of memory which is not really available!

Cool. Earlier we were facing a problem where even though a process reduces in size, the empty space is difficult to be reclaimed by other processes (because it would require costly compaction). Now it is easy, when a process becomes smaller in size, many of its pages are no longer used, so when other processes need more memory, a simple LRU based eviction automatically evicts those less-used pages from RAM, and replaces them with the new pages from the other processes (and of course updating the page tables of all those processes as well as the original process which now requires less space), all these without any costly compaction operation!

Solution to Problem 3: Whenever processes reduce in size, its frames in RAM will be less used, so a simple LRU based eviction can evict those pages out and replace them with pages required by new processes, thus avoiding Internal Fragmentation without need for compaction.

As for problem 2, take a moment to understand this, the scenario itself is completely removed! There is no need to move a process to accommodate a new process, because now the entire process never needs to fit at once, only certain pages of it need to fit ad hoc, that happens by evicting frames from RAM. Everything happens in units of pages, thus there is no concept of hole now, and hence no question of anything moving! May be 10 pages had to be moved because of this new requirement, there are thousands of pages which are left untouched. Whereas, earlier, all processes (every bit of them) had to be moved!

Solution to Problem 2: To accommodate a new process, data from only less recently used parts of other processes have to be evicted as required, and this happens in fixed size units called pages. Thus there is no possibility of hole or External Fragmentation with this system.

Now when the process needs to do some I/O operation, it can relinquish CPU easily! OS simply evicts all its pages from the RAM (perhaps store it in some cache) while new processes occupy the RAM in the meantime. When the I/O operation is done, OS simply restores those pages to the RAM (of course by replacing the pages from some other processes, may be from the ones which replaced the original process, or may be from some which themselves need to do I/O now, and hence can relinquish the memory!)

Solution to Problem 5: When a process is doing I/O operations, it can easily give up RAM usage, which can be utilized by other processes. This leads to proper utilization of RAM.

And of course, now no process is accessing the RAM directly. Each process is accessing a virtual memory location, which is mapped to a physical RAM address and maintained by the page-table of that process. The mapping is OS-backed, OS lets the process know which frame is empty so that a new page for a process can be fitted there. Since this memory allocation is overseen by the OS itself, it can easily ensure that no process encroaches upon the contents of another process by allocating only empty frames from RAM, or upon encroaching upon the contents of another process in the RAM, communicate to the process to update it page-table.

Solution to Original Problem: There is no possibility of a process accessing the contents of another process, since the entire allocation is managed by the OS itself, and every process runs in its own sandboxed virtual address space.

So paging (among other techniques), in conjunction with virtual memory, is what powers today's softwares running on OS-es! This frees the software developer from worrying about how much memory is available on the user's device, where to store the data, how to prevent other processes from corrupting their software's data, etc. However, it is of course, not full-proof. There are flaws:

1. Paging is, ultimately, giving user the illusion of infinite memory by using disk as secondary backup. Retrieving data from secondary storage to fit into memory (called page swap, and the event of not finding the desired page in RAM is called page fault) is expensive as it is an IO operation. This slows down the process. Several such page swaps happen in succession, and the process becomes painfully slow. Ever seen your software running fine and dandy, and suddenly it becomes so slow that it nearly hangs, or leaves you with no option that to restart it? Possibly too many page swaps were happening, making it slow (called thrashing).

So coming back to OP,

Why do we need the virtual memory for executing a process? - As the answer explains at length, to give softwares the illusion of the device/OS having infinite memory, so that any software, big or small, can be run, without worrying about memory allocation, or other processes corrupting its data, even when running in parallel. It is a concept, implemented in practice through various techniques, one of which, as described here, is Paging. It may also be Segmentation.

Where does this virtual memory stand when the process (program) from the external hard drive is brought to the main memory (physical memory) for the execution? - Virtual memory doesn't stand anywhere per se, it is an abstraction, always present, when the software/process/program is booted, a new page table is created for it, and it contains the mapping from the addresses spat out by that process to the actual physical address in RAM. Since the addresses spat out by the process are not real addresses, in one sense, they are, actually, what you can say, the virtual memory.

Who takes care of the virtual memory and what is the size of the virtual memory? - It is taken care of by, in tandem, the OS and the software. Imagine a function in your code (which eventually compiled and made into the executable that spawned the process) which contains a local variable - an int i. When the code executes, i gets a memory address within the stack of the function. That function is itself stored as an object somewhere else. These addresses are compiler generated (the compiler which compiled your code into the executable) - virtual addresses. When executed, i has to reside somewhere in actual physical address for duration of that function at least (unless it is a static variable!), so OS maps the compiler generated virtual address of i into an actual physical address, so that whenever, within that function, some code requires the value of i, that process can query the OS for that virtual address, and OS in turn can query the physical a

Have you tried Tools > Formula Auditing?

This has been fully answered here.

See Oracle's docs on SELECT to see how subquery factoring works, and Mark's example:

WITH employee AS (SELECT * FROM Employees)
SELECT * FROM employee WHERE ID < 20
UNION ALL
SELECT * FROM employee WHERE Sex = 'M'

You can use console.dir(object) to write that objects properties to the console.

Above answers all assume that the object in question is an object with behaviour. An advanced strategy in OOP is to separate data objects (that do zip, only have fields) and behaviour objects.

With data objects, it is perfectly fine to omit getters and instead have public fields. They usually don't have setters, since they most commonly are immutable - their fields are set via the constructors, and never again. Have a look at Bob Martin's Clean Code or Pryce and Freeman's Growing OO Software... for details.

I’d really suggest that people who encounter this problem go and find a third party database comparison tool.

Reason – these tools save a lot of time and make the process less error prone.

I’ve used comparison tools from ApexSQL (Diff and Data Diff) but you can’t go wrong with other tools marc_s and Marina Nastenko already pointed out.

If you’re absolutely sure that you are only going to compare tables once then SQL is fine but if you’re going to need this from time to time you’ll be better off with some 3rd party tool.

If you don’t have budget to buy it then just use it in trial mode to get the job done.

I hope new readers will find this useful even though it’s a late answer…

IOS 10 Update:

I had a problem with building the fatlib with iphoneos10.0 because the regular expression in the script only expects 9.x and lower and returns 0.0 for ios 10.0

to fix this just replace

SDK_VERSION=$(echo ${SDK_NAME} | grep -o '.\{3\}$')

with

SDK_VERSION=$(echo ${SDK_NAME} | grep -o '[\\.0-9]\{3,4\}$')

You need the secret string which was used to generate encrypt token. This code works for me:

protected string GetName(string token)
    {
        string secret = "this is a string used for encrypt and decrypt token"; 
        var key = Encoding.ASCII.GetBytes(secret);
        var handler = new JwtSecurityTokenHandler();
        var validations = new TokenValidationParameters
        {
            ValidateIssuerSigningKey = true,
            IssuerSigningKey = new SymmetricSecurityKey(key),
            ValidateIssuer = false,
            ValidateAudience = false
        };
        var claims = handler.ValidateToken(token, validations, out var tokenSecure);
        return claims.Identity.Name;
    }

@vhu I did the SHOW VARIABLES LIKE "secure_file_priv"; and it returned C:\ProgramData\MySQL\MySQL Server 8.0\Uploads\ so when I plugged that in, it still didn't work.

When I went to the my.ini file directly I discovered that the path is formatted a bit differently:

C:/ProgramData/MySQL/MySQL Server 8.0/Uploads

Then when I ran it with that, it worked. The only difference was the direction of the slashes.

You know what has worked for me really well on windows.

My Computer > Properties > Advanced System Settings > Environment Variables >

Just add the path as C:\Python27 (or wherever you installed python)

OR

Then under system variables I create a new Variable called PythonPath. In this variable I have C:\Python27\Lib;C:\Python27\DLLs;C:\Python27\Lib\lib-tk;C:\other-folders-on-the-path

This is the best way that has worked for me which I hadn't found in any of the docs offered.

EDIT: For those who are not able to get it, Please add

C:\Python27;

along with it. Else it will never work.

As the other answers said, the function you need is cv2.rectangle(), but keep in mind that the coordinates for the bounding box vertices need to be integers if they are in a tuple, and they need to be in the order of (left, top) and (right, bottom). Or, equivalently, (xmin, ymin) and (xmax, ymax).

$("input[name='mygroup'][value='5']").attr("checked", true);

You can pass you session variables from your php script to JQUERY using JSON such as

JS:

jQuery("#rowed2").jqGrid({
    url:'yourphp.php?q=3', 
    datatype: "json", 
    colNames:['Actions'], 
    colModel:[{
                name:'Actions',
                index:'Actions', 
                width:155,
                sortable:false
              }], 
    rowNum:30, 
    rowList:[50,100,150,200,300,400,500,600], 
    pager: '#prowed2', 
    sortname: 'id',
    height: 660,        
    viewrecords: true, 
    sortorder: 'desc',
    gridview:true,
    editurl: 'yourphp.php', 
    caption: 'Caption', 
    gridComplete: function() { 
        var ids = jQuery("#rowed2").jqGrid('getDataIDs'); 
        for (var i = 0; i < ids.length; i++) { 
            var cl = ids[i]; 
            be = "<input style='height:22px;width:50px;' `enter code here` type='button' value='Edit' onclick=\"jQuery('#rowed2').editRow('"+cl+"');\" />"; 
            se = "<input style='height:22px;width:50px;' type='button' value='Save' onclick=\"jQuery('#rowed2').saveRow('"+cl+"');\" />"; 
            ce = "<input style='height:22px;width:50px;' type='button' value='Cancel' onclick=\"jQuery('#rowed2').restoreRow('"+cl+"');\" />"; 
            jQuery("#rowed2").jqGrid('setRowData', ids[i], {Actions:be+se+ce}); 
        } 
    }
}); 

PHP

// start your session
session_start();

// get session from database or create you own
$session_username = $_SESSION['John'];
$session_email = $_SESSION['[email protected]'];

$response = new stdClass();
$response->session_username = $session_username;
$response->session_email = $session_email;

$i = 0;
while ($row = mysqli_fetch_array($result)) { 
    $response->rows[$i]['id'] = $row['ID']; 
    $response->rows[$i]['cell'] = array("", $row['rowvariable1'], $row['rowvariable2']); 
    $i++; 
} 

echo json_encode($response);
// this response (which contains your Session variables) is sent back to your JQUERY 

My case was kind of more complicated than most of the ones described in the answers. I was the owner of two private repositories repo-A and repo-B in a Github organization and needed to pip install repo-A during the python unittests of repo-B, as a Github action.

Steps I followed to solve this task:

• Created a Personal Access Token for my account. As for its permissions, I only needed to keep the default ones, .i.e. repo - Full control of private repositories.
• Created a repository secret under repo-B, pasted my Personal Access Token in there and named it PERSONAL_ACCESS_TOKEN. This was important because, unlike the solution proposed by Jamie, I didn't need to explicitly expose my precious raw Personal Access Token inside the github action .yml file.
• Finally, I pip installed the package from source via HTTPS as follows export PERSONAL_ACCESS_TOKEN=${{ secrets.PERSONAL_ACCESS_TOKEN }} && pip install git+https://${PERSONAL_ACCESS_TOKEN}@github.com/MY_ORG_NAME/MY_REPO_NAME.git

var list3 = list1.Where(x => !list2.Any(z => z.Id == x.Id)).ToList();

Note: list3 will contain the items or objects that are not in both lists. Note: Its ToList() not toList()

To answer the question of going from an existing python datetime to a pandas Timestamp do the following:

    import time, calendar, pandas as pd
    from datetime import datetime
    
    def to_posix_ts(d: datetime, utc:bool=True) -> float:
        tt=d.timetuple()
        return (calendar.timegm(tt) if utc else time.mktime(tt)) + round(d.microsecond/1000000, 0)
    
    def pd_timestamp_from_datetime(d: datetime) -> pd.Timestamp:
        return pd.to_datetime(to_posix_ts(d), unit='s')
    
    dt = pd_timestamp_from_datetime(datetime.now())
    print('({}) {}'.format(type(dt), dt))

Output:

(<class 'pandas._libs.tslibs.timestamps.Timestamp'>) 2020-09-05 23:38:55

I was hoping for a more elegant way to do this but the to_posix_ts is already in my standard tool chain so I'm moving on.

I had this issue as well. The branch looked like 'username/master' which seemed to confuse git as it looked like a remote address I defined. For me using this

git merge origin/username/master

worked perfectly fine.

There should be three pages here:

1. Initial login page with a form that asks for your username, but not your password.
2. You didn't mention this one, but I'd check whether the client computer is recognized, and if not, then challenge the user with either a CAPTCHA or else a security question. Otherwise the phishing site can simply use the tendered username to query the real site for the security image, which defeats the purpose of having a security image. (A security question is probably better here since with a CAPTCHA the attacker could have humans sitting there answering the CAPTCHAs to get at the security images. Depends how paranoid you want to be.)
3. A page after that that displays the security image and asks for the password.

I don't see this short, linear flow being sufficiently complex to warrant using Spring Web Flow.

I would just use straight Spring Web MVC for steps 1 and 2. I wouldn't use Spring Security for the initial login form, because Spring Security's login form expects a password and a login processing URL. Similarly, Spring Security doesn't provide special support for CAPTCHAs or security questions, so you can just use Spring Web MVC once again.

You can handle step 3 using Spring Security, since now you have a username and a password. The form login page should display the security image, and it should include the user-provided username as a hidden form field to make Spring Security happy when the user submits the login form. The only way to get to step 3 is to have a successful POST submission on step 1 (and 2 if applicable).

You can use the rpad and lpad functions to pad numbers to the right or to the left, respectively. Note that this does not work directly on numbers, so you'll have to use ::char or ::text to cast them:

SELECT RPAD(numcol::text, 3, '0'), -- Zero-pads to the right up to the length of 3
       LPAD(numcol::text, 3, '0'), -- Zero-pads to the left up to the length of 3
FROM   my_table

I did a little experimenting to see if I could keep the database name as part of the url. I prefer the promise syntax but it should still work for the callback syntax. Notice below that client.db() is called without passing any parameters.

MongoClient.connect(
    'mongodb://localhost:27017/mytestingdb', 
    { useNewUrlParser: true}
)
.then(client => {

    // The database name is part of the url.  client.db() seems 
    // to know that and works even without a parameter that 
    // relays the db name.
    let db = client.db(); 

    console.log('the current database is: ' + db.s.databaseName);
    // client.close() if you want to

})
.catch(err => console.log(err));

My package.json lists monbodb ^3.2.5.

The 'useNewUrlParser' option is not required if you're willing to deal with a deprecation warning. But it is wise to use at this point until version 4 comes out where presumably the new driver will be the default and you won't need the option anymore.

If this helps anybody, I encountered this while using purr::map() with a function I wrote which was something like this:

find_nearby_shops <- function(base_account) {
   states_table %>% 
        filter(state == base_account$state) %>% 
        left_join(target_locations, by = c('border_states' = 'state')) %>% 
        mutate(x_latitude = base_account$latitude,
               x_longitude = base_account$longitude) %>% 
        mutate(dist_miles = geosphere::distHaversine(p1 = cbind(longitude, latitude), 
                                                     p2 = cbind(x_longitude, x_latitude))/1609.344)
}

nearby_shop_numbers <- base_locations %>% 
    split(f = base_locations$id) %>% 
    purrr::map_df(find_nearby_shops) 

I would get this error sometimes with samples, but most times I wouldn't. The root of the problem is that some of the states in the base_locations table (PR) did not exist in the states_table, so essentially I had filtered out everything, and passed an empty table on to mutate. The moral of the story is that you may have a data issue and not (just) a code problem (so you may need to clean your data.)

Thanks for agstudy and zx8754's answers above for helping with the debug.

Using these libraries:

• gson-2.8.5.jar
• json-20180813.jar

Using this code:

List<Object[]> testNovedads = crudService.createNativeQuery(
                "SELECT cantidad, id FROM NOVEDADES GROUP BY id ");

        Gson gson = new Gson();
        String json = gson.toJson(new TestNovedad());
        JSONObject jsonObject = new JSONObject(json);
        Collection<TestNovedad> novedads = new ArrayList<>();
        for (Object[] object : testNovedads) {
            Iterator<String> iterator = jsonObject.keys();
            int pos = 0;
            for (Iterator i = iterator; i.hasNext();) {
                jsonObject.put((String) i.next(), object[pos++]);
            }
            novedads.add(gson.fromJson(jsonObject.toString(), TestNovedad.class));
        }
        for (TestNovedad testNovedad : novedads) {
            System.out.println(testNovedad.toString());
        }

/**
 * Autores: Chalo Mejia 
 * Fecha: 01/10/2020
 */
package org.main;

import java.io.Serializable;

public class TestNovedad implements Serializable {

    private static final long serialVersionUID = -6362794385792247263L;

    private int id;

    private int cantidad;

    public TestNovedad() {
        // TODO Auto-generated constructor stub
    }

    public int getId() {
        return id;
    }

    public void setId(int id) {
        this.id = id;
    }

    public int getCantidad() {
        return cantidad;
    }

    public void setCantidad(int cantidad) {
        this.cantidad = cantidad;
    }

    @Override
    public String toString() {
        return "TestNovedad [id=" + id + ", cantidad=" + cantidad + "]";
    }

}

Both algorithms generally works like this

Step 1: take the next unsorted element from the unsorted list then

Step 2: put it in the right place in the sorted list.

One of the steps is easier for one algorithm and vice versa.

Insertion sort: We take the first element of the unsorted list, put it in the sorted list, somewhere. We know where to take the next element (the first position in the unsorted list), but it requires some work to find where to put it (somewhere). Step 1 is easy.

Selection sort: We take the element somewhere from the unsorted list, then put it in the last position of the sorted list. We need to find the next element (it most likely is not in the first position of the unsorted list, but rather, somewhere) then put it right at the end of the sorted list. Step 2 is easy

This is an extension method derived from @Seattle Leonard 's answer:

 public static T GetMax<T,U>(this IEnumerable<T> data, Func<T,U> f) where U:IComparable
 {
     return data.Aggregate((i1, i2) => f(i1).CompareTo(f(i2))>0 ? i1 : i2);
 }

delete from t
where id in (1, 4, 6, 7)

Modify your CSS like this:

.vertical_banner {_x000D_
    border: 1px solid #E9E3DD;_x000D_
    float: left;_x000D_
    height: 210px;_x000D_
    margin: 2px;_x000D_
    padding: 4px 2px 10px 10px;_x000D_
    text-align: left;_x000D_
    width: 117px;_x000D_
    position:relative;_x000D_
}_x000D_
_x000D_
#bottom_link{_x000D_
   position:absolute;                  /* added */_x000D_
   bottom:0;                           /* added */_x000D_
   left:0;                           /* added */_x000D_
}

<div class="vertical_banner">_x000D_
    <div id="bottom_link">_x000D_
         <input type="submit" value="Continue">_x000D_
       </div>_x000D_
</div>

You could use the checked selector to grab only the selected ones (negating the need to know the count or to iterate over them all yourself):

$("input[name='user_group[]']:checked")

With those checked items, you can either create a collection of those values or do something to the collection:

var values = new Array();
$.each($("input[name='user_group[]']:checked"), function() {
  values.push($(this).val());
  // or you can do something to the actual checked checkboxes by working directly with  'this'
  // something like $(this).hide() (only something useful, probably) :P
});

I don't think this is possible with AD. The distinguishedName attribute is the only thing I know of that contains the OU piece on which you're trying to search, so you'd need a wildcard to get results for objects under those OUs. Unfortunately, the wildcard character isn't supported on DNs.

If at all possible, I'd really look at doing this in 2 queries using OU=Staff... and OU=Vendors... as the base DNs.

string.find(character)  
string.index(character)  

Perhaps you'd like to have a look at the documentation to find out what the difference between the two is.

Everyone else's method doesn't account for whitespaces. Obviously nobody really considers a whitespace a special character.

Use this method to detect special characters not including whitespaces:

import re

def detect_special_characer(pass_string): 
  regex= re.compile('[@_!#$%^&*()<>?/\|}{~:]') 
  if(regex.search(pass_string) == None): 
    res = False
  else: 
    res = True
  return(res)

$sql = "SELECT * FROM table_name ORDER BY ID DESC LIMIT 1";
$records = mysql_query($sql);

you can change LIMIT 1 to LIMIT any number you want

This will show you the last INSERTED row first.

Found a nice solution to this problem elsewhere, no messing around with tags on the button:

- (void)buttonPressedAction:(id)sender {

    NSSet *touches = [event allTouches];
    UITouch *touch = [touches anyObject];
    CGPoint currentTouchPosition = [touch locationInView:self.tableView];
    NSIndexPath *indexPath = [self.tableView indexPathForRowAtPoint: currentTouchPosition];

    // do stuff with the indexPath...
}

ndb is an improved debugging experience for Node.js, enabled by Chrome DevTools

https://github.com/GoogleChromeLabs/ndb

Loosely speaking, time complexity is a way of summarising how the number of operations or run-time of an algorithm grows as the input size increases.

Like most things in life, a cocktail party can help us understand.

O(N)

When you arrive at the party, you have to shake everyone's hand (do an operation on every item). As the number of attendees N increases, the time/work it will take you to shake everyone's hand increases as O(N).

Why O(N) and not cN?

There's variation in the amount of time it takes to shake hands with people. You could average this out and capture it in a constant c. But the fundamental operation here --- shaking hands with everyone --- would always be proportional to O(N), no matter what c was. When debating whether we should go to a cocktail party, we're often more interested in the fact that we'll have to meet everyone than in the minute details of what those meetings look like.

O(N^2)

The host of the cocktail party wants you to play a silly game where everyone meets everyone else. Therefore, you must meet N-1 other people and, because the next person has already met you, they must meet N-2 people, and so on. The sum of this series is x^2/2+x/2. As the number of attendees grows, the x^2 term gets big fast, so we just drop everything else.

O(N^3)

You have to meet everyone else and, during each meeting, you must talk about everyone else in the room.

O(1)

The host wants to announce something. They ding a wineglass and speak loudly. Everyone hears them. It turns out it doesn't matter how many attendees there are, this operation always takes the same amount of time.

O(log N)

The host has laid everyone out at the table in alphabetical order. Where is Dan? You reason that he must be somewhere between Adam and Mandy (certainly not between Mandy and Zach!). Given that, is he between George and Mandy? No. He must be between Adam and Fred, and between Cindy and Fred. And so on... we can efficiently locate Dan by looking at half the set and then half of that set. Ultimately, we look at O(log_2 N) individuals.

O(N log N)

You could find where to sit down at the table using the algorithm above. If a large number of people came to the table, one at a time, and all did this, that would take O(N log N) time. This turns out to be how long it takes to sort any collection of items when they must be compared.

Best/Worst Case

You arrive at the party and need to find Inigo - how long will it take? It depends on when you arrive. If everyone is milling around you've hit the worst-case: it will take O(N) time. However, if everyone is sitting down at the table, it will take only O(log N) time. Or maybe you can leverage the host's wineglass-shouting power and it will take only O(1) time.

Assuming the host is unavailable, we can say that the Inigo-finding algorithm has a lower-bound of O(log N) and an upper-bound of O(N), depending on the state of the party when you arrive.

Space & Communication

The same ideas can be applied to understanding how algorithms use space or communication.

Knuth has written a nice paper about the former entitled "The Complexity of Songs".

Theorem 2: There exist arbitrarily long songs of complexity O(1).

PROOF: (due to Casey and the Sunshine Band). Consider the songs Sk defined by (15), but with

V_k = 'That's the way,' U 'I like it, ' U
U   = 'uh huh,' 'uh huh'

for all k.

You can do this with the scp command, which uses the ssh protocol to copy files across machines. It extends the syntax of cp to allow references to other systems:

scp username1@hostname1:/path/to/file username2@hostname2:/path/to/other/file

Copy something from this machine to some other machine:

scp /path/to/local/file username@hostname:/path/to/remote/file

Copy something from another machine to this machine:

scp username@hostname:/path/to/remote/file /path/to/local/file

Copy with a port number specified:

scp -P 1234 username@hostname:/path/to/remote/file /path/to/local/file

The Content-Security-Policy meta-tag allows you to reduce the risk of XSS attacks by allowing you to define where resources can be loaded from, preventing browsers from loading data from any other locations. This makes it harder for an attacker to inject malicious code into your site.

I banged my head against a brick wall trying to figure out why I was getting CSP errors one after another, and there didn't seem to be any concise, clear instructions on just how does it work. So here's my attempt at explaining some points of CSP briefly, mostly concentrating on the things I found hard to solve.

For brevity I won’t write the full tag in each sample. Instead I'll only show the content property, so a sample that says content="default-src 'self'" means this:

<meta http-equiv="Content-Security-Policy" content="default-src 'self'">

1. How can I allow multiple sources?

You can simply list your sources after a directive as a space-separated list:

content="default-src 'self' https://example.com/js/"

Note that there are no quotes around parameters other than the special ones, like 'self'. Also, there's no colon (:) after the directive. Just the directive, then a space-separated list of parameters.

Everything below the specified parameters is implicitly allowed. That means that in the example above these would be valid sources:

https://example.com/js/file.js
https://example.com/js/subdir/anotherfile.js

These, however, would not be valid:

http://example.com/js/file.js
^^^^ wrong protocol

https://example.com/file.js
                   ^^ above the specified path

2. How can I use different directives? What do they each do?

The most common directives are:

• default-src the default policy for loading javascript, images, CSS, fonts, AJAX requests, etc
• script-src defines valid sources for javascript files
• style-src defines valid sources for css files
• img-src defines valid sources for images
• connect-src defines valid targets for to XMLHttpRequest (AJAX), WebSockets or EventSource. If a connection attempt is made to a host that's not allowed here, the browser will emulate a 400 error

There are others, but these are the ones you're most likely to need.

3. How can I use multiple directives?

You define all your directives inside one meta-tag by terminating them with a semicolon (;):

content="default-src 'self' https://example.com/js/; style-src 'self'"

4. How can I handle ports?

Everything but the default ports needs to be allowed explicitly by adding the port number or an asterisk after the allowed domain:

content="default-src 'self' https://ajax.googleapis.com http://example.com:123/free/stuff/"

The above would result in:

https://ajax.googleapis.com:123
                           ^^^^ Not ok, wrong port

https://ajax.googleapis.com - OK

http://example.com/free/stuff/file.js
                 ^^ Not ok, only the port 123 is allowed

http://example.com:123/free/stuff/file.js - OK

As I mentioned, you can also use an asterisk to explicitly allow all ports:

content="default-src example.com:*"

5. How can I handle different protocols?

By default, only standard protocols are allowed. For example to allow WebSockets ws:// you will have to allow it explicitly:

content="default-src 'self'; connect-src ws:; style-src 'self'"
                                         ^^^ web Sockets are now allowed on all domains and ports.

6. How can I allow the file protocol file://?

If you'll try to define it as such it won’t work. Instead, you'll allow it with the filesystem parameter:

content="default-src filesystem"

7. How can I use inline scripts and style definitions?

Unless explicitly allowed, you can't use inline style definitions, code inside <script> tags or in tag properties like onclick. You allow them like so:

content="script-src 'unsafe-inline'; style-src 'unsafe-inline'"

You'll also have to explicitly allow inline, base64 encoded images:

content="img-src data:"

8. How can I allow eval()?

I'm sure many people would say that you don't, since 'eval is evil' and the most likely cause for the impending end of the world. Those people would be wrong. Sure, you can definitely punch major holes into your site's security with eval, but it has perfectly valid use cases. You just have to be smart about using it. You allow it like so:

content="script-src 'unsafe-eval'"

9. What exactly does 'self' mean?

You might take 'self' to mean localhost, local filesystem, or anything on the same host. It doesn't mean any of those. It means sources that have the same scheme (protocol), same host, and same port as the file the content policy is defined in. Serving your site over HTTP? No https for you then, unless you define it explicitly.

I've used 'self' in most examples as it usually makes sense to include it, but it's by no means mandatory. Leave it out if you don't need it.

But hang on a minute! Can't I just use content="default-src *" and be done with it?

No. In addition to the obvious security vulnerabilities, this also won’t work as you'd expect. Even though some docs claim it allows anything, that's not true. It doesn't allow inlining or evals, so to really, really make your site extra vulnerable, you would use this:

content="default-src * 'unsafe-inline' 'unsafe-eval'"

... but I trust you won’t.

Further reading:

http://content-security-policy.com

http://en.wikipedia.org/wiki/Content_Security_Policy

The easiest is Simple there is a tutorial, no WEB-INF not Servlet API no dependencies. Just a simple lightweight HTTP server in a single JAR.

  function randomPassword(length) {
        var chars = "abcdefghijklmnopqrstuvwxyz#$%ABCDEFGHIJKLMNOP1234567890";
        var pass = "";
        for (var x = 0; x < length; x++) {
            var i = Math.floor(Math.random() * chars.length);
            pass += chars.charAt(i);
        }
        var pattern = false;
        var passwordpattern = new RegExp("[^a-zA-Z0-9+]+[0-9+]+[A-Z+]+[a-z+]");
        pattern = passwordpattern.test(pass);
        if (pattern == true)
            { alert(pass); }
        else
         { randomPassword(length); }
    }

try this to create the random password with atleast one special character

I also encountered the same problem then I did following things.

1. I import the library project into my AndroidStudio IDE as a module using menu File -> Import module menus

2. Then I went to my main module in which I want the library project as a dependent project

3. Right click on the main module (in my case its name is app) -> open module setting -> go into dependencies tab -> click on + button (you will get it on right side of window) -> click on module dependency -> select your library project from list

Apply the changes and click the OK button.

It worked for me. I hope it will help others too.

Swift 4

How about this small trick?

func showToast(controller: UIViewController, message : String, seconds: Double) {
    let alert = UIAlertController(title: nil, message: message, preferredStyle: .alert)
    alert.view.backgroundColor = UIColor.black
    alert.view.alpha = 0.6
    alert.view.layer.cornerRadius = 15

    controller.present(alert, animated: true)

    DispatchQueue.main.asyncAfter(deadline: DispatchTime.now() + seconds) {
        alert.dismiss(animated: true)
    }
}

Example of calling:

showToast(controller: self, message : "This is a test", seconds: 2.0)

Output:

the above is not bad, however its kinda different for 11.10

sudo apt-get install ruby1.9 rubygems1.9

that will install ruby 1.9

when linking, you just use ls /usr/bin | grep ruby it should output ruby1.9.1

so then you sudo ln -sf /usr/bin/ruby1.9.1 /usr/bin/ruby and your off to the races.

You solve the issue with a try/ catch. This crash happens when user close the app before the start intent.

try
{
    Intent mIntent = new Intent(getActivity(),MusicHome.class);
    mIntent.putExtra("SigninFragment.user_details", bundle);
    startActivity(mIntent);
}
catch (Exception e) {
    e.printStackTrace();
}

I'm not certain what's wrong with your code specifically, but I use this function I found online (URL in the comments) for checking if a file exists:

Private Function File_Exists(ByVal sPathName As String, Optional Directory As Boolean) As Boolean
    'Code from internet: http://vbadud.blogspot.com/2007/04/vba-function-to-check-file-existence.html
    'Returns True if the passed sPathName exist
    'Otherwise returns False
    On Error Resume Next
    If sPathName <> "" Then

        If IsMissing(Directory) Or Directory = False Then

            File_Exists = (Dir$(sPathName) <> "")
        Else

            File_Exists = (Dir$(sPathName, vbDirectory) <> "")
        End If

    End If
End Function

The -Wno-unused-variable switch usually does the trick. However, that is a very useful warning indeed if you care about these things in your project. It becomes annoying when GCC starts to warn you about things not in your code though.

I would recommend you keeping the warning on, but use -isystem instead of -I for include directories of third-party projects. That flag tells GCC not to warn you about the stuff you have no control over.

For example, instead of -IC:\\boost_1_52_0, say -isystem C:\\boost_1_52_0.

Hope it helps. Good Luck!

Goto View -> Show Symbol Select as per screen-shot, you will get correct

Make sure there is an namespace definition (xmlns) for the namespace your control belong to.

xmlns:myControls="clr-namespace:YourCustomNamespace.Controls;assembly=YourAssemblyName"

<myControls:thecontrol/>

Simplest way I've found that doesn't have inconsistency across platforms:

setVisible(false); setVisible(true);

Since you tagged Java-ee as well - then YES it is possible.

and in core java as well it is possible using static blocks

and check this How can you run a Java program without main method?

Edit:
as already pointed out in other answers - it does not support from Java 7

def my_add_fn():
   print "SUM:%s"%sum(map(int,raw_input("Enter 2 numbers seperated by a space").split()))

def my_quit_fn():
   raise SystemExit

def invalid():
   print "INVALID CHOICE!"

menu = {"1":("Sum",my_add_fn),
        "2":("Quit",my_quit_fn)
       }
for key in sorted(menu.keys()):
     print key+":" + menu[key][0]

ans = raw_input("Make A Choice")
menu.get(ans,[None,invalid])[1]()

Abstract methods means there is no default implementation for it and an implementing class will provide the details.

Essentially, you would have

abstract class AbstractObject {
   public abstract void method();
}

class ImplementingObject extends AbstractObject {
  public void method() {
    doSomething();
  }
}

So, it's exactly as the error states: your abstract method can not have a body.

There's a full tutorial on Oracle's site at: http://download.oracle.com/javase/tutorial/java/IandI/abstract.html

The reason you would do something like this is if multiple objects can share some behavior, but not all behavior.

A very simple example would be shapes:

You can have a generic graphic object, which knows how to reposition itself, but the implementing classes will actually draw themselves.

(This is taken from the site I linked above)

abstract class GraphicObject {
    int x, y;
    ...
    void moveTo(int newX, int newY) {
        ...
    }
    abstract void draw();
    abstract void resize();
}

class Circle extends GraphicObject {
    void draw() {
        ...
    }
    void resize() {
        ...
    }
}
class Rectangle extends GraphicObject {
    void draw() {
        ...
    }
    void resize() {
        ...
    }
}

You can exclude tasks

 gradle build --exclude-task test 

https://docs.gradle.org/current/userguide/command_line_interface.html#sec:command_line_executing_tasks

Example code:

var d = new Date();
var time = d.toISOString().replace(/.*?T(\d+:\d+:\d+).*/, "$1");

Output:

"13:45:20"

This is not a reply (I cant post comments), just few random ideas might be helpful. Unfortunately I've never dealt with citrix, only with regular windows servers.

_0. Ensure you're not a victim of Windows Firewall, or any other personal firewall that selectively blocks processes.

Add 10 minutes Sleep() to the first line of your .NET app, then run both VBScript file and your stand-alone application, run sysinternals process explorer, and compare 2 processes.

_1. Same tab, "command line" and "current directory". Make sure they are the same.

_2. "Environment" tab. Make sure they are the same. Normally child processes inherit the environment, but this behaviour can be easily altered.

The following check is required if by "run my script" you mean anything else then double-clicking the .VBS file:

_3. Image tab, "User". If they differ - it may mean user has no access to the network (like localsystem), or user token restricted to delegation and thus can only access local resources (like in the case of IIS NTLM auth), or user has no access to some local files it wants.

I feel like reinventing the wheel...

public static T[] ConvertToArray<T>(this IEnumerable<T> enumerable)
{
    if (enumerable == null)
        throw new ArgumentNullException("enumerable");

    return enumerable as T[] ?? enumerable.ToArray();
}

Using the "with" statement with shorthand object literal syntax

Nobody has demonstrated this method yet, probably because it's terrible and you shouldn't do it, but I feel like it has to be listed.

var o = {a:1,b:2,c:3,d:4,e:4,f:5}
with(o){
  var output =  {a,b,f}
}
console.log(output)

Pro: You don't have to type the property names twice.

Cons: The "with" statement is not recommended for many reasons.

Conclusion: It works great, but don't use it.

The top 2 answers (from Adam Robinson and Andrejs Cainikovs) are kinda, sorta correct, in that they do technically work, but their explanations are wrong and so could be misleading in many cases. For example, while the SQL_Latin1_General_CP1_CI_AS collation will work in many cases, it should not be assumed to be the appropriate case-insensitive collation. In fact, given that the O.P. is working in a database with a case-sensitive (or possibly binary) collation, we know that the O.P. isn't using the collation that is the default for so many installations (especially any installed on an OS using US English as the language): SQL_Latin1_General_CP1_CI_AS. Sure, the O.P. could be using SQL_Latin1_General_CP1_CS_AS, but when working with VARCHAR data, it is important to not change the code page as it could lead to data loss, and that is controlled by the locale / culture of the collation (i.e. Latin1_General vs French vs Hebrew etc). Please see point # 9 below.

The other four answers are wrong to varying degrees.

I will clarify all of the misunderstandings here so that readers can hopefully make the most appropriate / efficient choices.

1. Do not use UPPER(). That is completely unnecessary extra work. Use a COLLATE clause. A string comparison needs to be done in either case, but using UPPER() also has to check, character by character, to see if there is an upper-case mapping, and then change it. And you need to do this on both sides. Adding COLLATE simply directs the processing to generate the sort keys using a different set of rules than it was going to by default. Using COLLATE is definitely more efficient (or "performant", if you like that word :) than using UPPER(), as proven in this test script (on PasteBin).

   There is also the issue noted by @Ceisc on @Danny's answer:

   In some languages case conversions do not round-trip. i.e. LOWER(x) != LOWER(UPPER(x)).

   The Turkish upper-case "I" is the common example.

2. No, collation is not a database-wide setting, at least not in this context. There is a database-level default collation, and it is used as the default for altered and newly created columns that do not specify the COLLATE clause (which is likely where this common misconception comes from), but it does not impact queries directly unless you are comparing string literals and variables to other string literals and variables, or you are referencing database-level meta-data.

3. No, collation is not per query.

4. Collations are per predicate (i.e. something operand something) or expression, not per query. And this is true for the entire query, not just the WHERE clause. This covers JOINs, GROUP BY, ORDER BY, PARTITION BY, etc.

5. No, do not convert to VARBINARY (e.g.convert(varbinary, myField) = convert(varbinary, 'sOmeVal')) for the following reasons:

   1. that is a binary comparison, which is not case-insensitive (which is what this question is asking for)
   2. if you do want a binary comparison, use a binary collation. Use one that ends with _BIN2 if you are using SQL Server 2008 or newer, else you have no choice but to use one that ends with _BIN. If the data is NVARCHAR then it doesn't matter which locale you use as they are all the same in that case, hence Latin1_General_100_BIN2 always works. If the data is VARCHAR, you must use the same locale that the data is currently in (e.g. Latin1_General, French, Japanese_XJIS, etc) because the locale determines the code page that is used, and changing code pages can alter the data (i.e. data loss).
   3. using a variable-length datatype without specifying the size will rely on the default size, and there are two different defaults depending on the context where the datatype is being used. It is either 1 or 30 for string types. When used with CONVERT() it will use the 30 default value. The danger is, if the string can be over 30 bytes, it will get silently truncated and you will likely get incorrect results from this predicate.
   4. Even if you want a case-sensitive comparison, binary collations are not case-sensitive (another very common misconception).

6. No, LIKE is not always case-sensitive. It uses the collation of the column being referenced, or the collation of the database if a variable is compared to a string literal, or the collation specified via the optional COLLATE clause.

7. LCASE is not a SQL Server function. It appears to be either Oracle or MySQL. Or possibly Visual Basic?

8. Since the context of the question is comparing a column to a string literal, neither the collation of the instance (often referred to as "server") nor the collation of the database have any direct impact here. Collations are stored per each column, and each column can have a different collation, and those collations don't need to be the same as the database's default collation or the instance's collation. Sure, the instance collation is the default for what a newly created database will use as its default collation if the COLLATE clause wasn't specified when creating the database. And likewise, the database's default collation is what an altered or newly created column will use if the COLLATE clause wasn't specified.

9. You should use the case-insensitive collation that is otherwise the same as the collation of the column. Use the following query to find the column's collation (change the table's name and schema name):

   SELECT col.*
   FROM   sys.columns col
   WHERE  col.[object_id] = OBJECT_ID(N'dbo.TableName')
   AND    col.[collation_name] IS NOT NULL;
   

   Then just change the _CS to be _CI. So, Latin1_General_100_CS_AS would become Latin1_General_100_CI_AS.

   If the column is using a binary collation (ending in _BIN or _BIN2), then find a similar collation using the following query:

   SELECT *
   FROM   sys.fn_helpcollations() col
   WHERE  col.[name] LIKE N'{CurrentCollationMinus"_BIN"}[_]CI[_]%';
   

   For example, assuming the column is using Japanese_XJIS_100_BIN2, do this:

   SELECT *
   FROM   sys.fn_helpcollations() col
   WHERE  col.[name] LIKE N'Japanese_XJIS_100[_]CI[_]%';
   

For more info on collations, encodings, etc, please visit: Collations Info

var consolidatedChildren =
    from c in children
    group c by new
    {
        c.School,
        c.Friend,
        c.FavoriteColor,
    } into gcs
    select new ConsolidatedChild()
    {
        School = gcs.Key.School,
        Friend = gcs.Key.Friend,
        FavoriteColor = gcs.Key.FavoriteColor,
        Children = gcs.ToList(),
    };

var consolidatedChildren =
    children
        .GroupBy(c => new
        {
            c.School,
            c.Friend,
            c.FavoriteColor,
        })
        .Select(gcs => new ConsolidatedChild()
        {
            School = gcs.Key.School,
            Friend = gcs.Key.Friend,
            FavoriteColor = gcs.Key.FavoriteColor,
            Children = gcs.ToList(),
        });

Cast the dtype of the col to str and you can perform vectorised slicing calling str:

In [29]:
df['new_col'] = df['First'].astype(str).str[0]
df

Out[29]:
   First  Second new_col
0    123     234       1
1     22    4353       2
2     32     355       3
3    453     453       4
4     45     345       4
5    453     453       4
6     56      56       5

if you need to you can cast the dtype back again calling astype(int) on the column

If you are copy-pasting code into R, it sometimes won't accept some special characters such as "~" and will appear instead as a "?". So if a certain character is giving an error, make sure to use your keyboard to enter the character, or find another website to copy-paste from if that doesn't work.

The cross-platform way to do this is to print('\a'). This will send the ASCII Bell character to stdout, and will hopefully generate a beep (a for 'alert'). Note that many modern terminal emulators provide the option to ignore bell characters.

Since you're on Windows, you'll be happy to hear that Windows has its own (brace yourself) Beep API, which allows you to send beeps of arbitrary length and pitch. Note that this is a Windows-only solution, so you should probably prefer print('\a') unless you really care about Hertz and milliseconds.

The Beep API is accessed through the winsound module: http://docs.python.org/library/winsound.html

To calculate the difference between two dates in Years, Months, Days, Minutes, Seconds, Milliseconds using TypeScript/ JavaScript

dateDifference(actualDate) {
            // Calculate time between two dates:
            const date1 = actualDate; // the date you already commented/ posted
            const date2: any = new Date(); // today

            let r = {}; // object for clarity
            let message: string;

            const diffInSeconds = Math.abs(date2 - date1) / 1000;
            const days = Math.floor(diffInSeconds / 60 / 60 / 24);
            const hours = Math.floor(diffInSeconds / 60 / 60 % 24);
            const minutes = Math.floor(diffInSeconds / 60 % 60);
            const seconds = Math.floor(diffInSeconds % 60);
            const milliseconds = 
           Math.round((diffInSeconds - Math.floor(diffInSeconds)) * 1000);

            const months = Math.floor(days / 31);
            const years = Math.floor(months / 12);

            // the below object is just optional 
            // if you want to return an object instead of a message
            r = {
                years: years,
                months: months,
                days: days,
                hours: hours,
                minutes: minutes,
                seconds: seconds,
                milliseconds: milliseconds
            };

            // check if difference is in years or months
            if (years === 0 && months === 0) {
                // show in days if no years / months
                if (days > 0) {
                    if (days === 1) {
                        message = days + ' day';
                    } else { message = days + ' days'; }
                }  else if (hours > 0) {
                    if (hours === 1) {
                        message = hours + ' hour';
                    } else {
                        message = hours + ' hours';
                    }
                } else {
                    // show in minutes if no years / months / days
                    if (minutes === 1) {
                        message = minutes + ' minute';
                    } else {message = minutes + ' minutes';}  
                }
            } else if (years === 0 && months > 0) {
                // show in months if no years
                if (months === 1) {
                    message = months + ' month';
                } else {message = months + ' months';}
            } else if (years > 0) {
                // show in years if years exist
                if (years === 1) {
                    message = years + ' year';
                } else {message = years + ' years';}
            }

            return 'Posted ' + message + ' ago'; 
     // this is the message a user see in the view
        }

However, you can update the above logic for the message to show seconds and milliseconds too or else use the object 'r' to format the message whatever way you want.

If you want to directly copy the code, you can view my gist with the above code here

Easiest true sync solution (i.e. no yield/async) I could come up with that works in all OS's without any dependencies is to call the node process to eval an in-line setTimeout expression:

const sleep = (ms) => require("child_process")
    .execSync(`"${process.argv[0]}" -e setTimeout(function(){},${ms})`);

Using java.nio.Files along with java.nio.file.StandardOpenOption

    PrintWriter out = null;
    BufferedWriter bufWriter;

    try{
        bufWriter =
            Files.newBufferedWriter(
                Paths.get("log.txt"),
                Charset.forName("UTF8"),
                StandardOpenOption.WRITE, 
                StandardOpenOption.APPEND,
                StandardOpenOption.CREATE);
        out = new PrintWriter(bufWriter, true);
    }catch(IOException e){
        //Oh, no! Failed to create PrintWriter
    }

    //After successful creation of PrintWriter
    out.println("Text to be appended");

    //After done writing, remember to close!
    out.close();

This creates a BufferedWriter using Files, which accepts StandardOpenOption parameters, and an auto-flushing PrintWriter from the resultant BufferedWriter. PrintWriter's println() method, can then be called to write to the file.

The StandardOpenOption parameters used in this code: opens the file for writing, only appends to the file, and creates the file if it does not exist.

Paths.get("path here") can be replaced with new File("path here").toPath(). And Charset.forName("charset name") can be modified to accommodate the desired Charset.

def updateJsonFile():
    jsonFile = open("replayScript.json", "r") # Open the JSON file for reading
    data = json.load(jsonFile) # Read the JSON into the buffer
    jsonFile.close() # Close the JSON file

    ## Working with buffered content
    tmp = data["location"] 
    data["location"] = path
    data["mode"] = "replay"

    ## Save our changes to JSON file
    jsonFile = open("replayScript.json", "w+")
    jsonFile.write(json.dumps(data))
    jsonFile.close()

Try this code working fine check codeigniter version

Just go to 'system' > 'core' > 'CodeIgniter.php' and look for the lines,

/**
 * CodeIgniter Version
 *
 * @var    string
 *
 */
    define('CI_VERSION', '3.0.0');

Alternate method to check codeigniter version, you can echo the constant value 'CI_VERSION' somewhere in codeigniter controller/view file.

<?php
    echo CI_VERSION;
?>

More Information with demo: how to check codeigniter version

You can write Bean Editors (details are in the Spring Docs) if you want to add further value and write to custom types.

Don't delete it: for just one commit git cherry-pick is enough.

But if you had several commits on the wrong branch, that is where git rebase --onto shines:

Suppose you have this:

 x--x--x--x <-- master
           \
            -y--y--m--m <- y branch, with commits which should have been on master

, then you can mark master and move it where you would want to be:

 git checkout master
 git branch tmp
 git checkout y
 git branch -f master

 x--x--x--x <-- tmp
           \
            -y--y--m--m <- y branch, master branch

, reset y branch where it should have been:

 git checkout y
 git reset --hard HEAD~2 # ~1 in your case, 
                         # or ~n, n = number of commits to cancel

 x--x--x--x <-- tmp
           \
            -y--y--m--m <- master branch
                ^
                |
                -- y branch

, and finally move your commits (reapply them, making actually new commits)

 git rebase --onto tmp y master
 git branch -D tmp


 x--x--x--x--m'--m' <-- master
           \
            -y--y <- y branch

You have properly guarded your code block when you call wait() method by using synchronized(this).

But you have not taken same precaution when you call notify() method without using guarded block : synchronized(this) or synchronized(someObject)

If you refer to oracle documentation page on Object class, which contains wait() ,notify(), notifyAll() methods, you can see below precaution in all these three methods

This method should only be called by a thread that is the owner of this object's monitor

Many things have been changed in last 7 years and let's have look into other alternatives to synchronized in below SE questions:

Why use a ReentrantLock if one can use synchronized(this)?

Synchronization vs Lock

Avoid synchronized(this) in Java?

Yet another method:

s = 'hello'

h = ''.join([hex(ord(i)) for i in s]);

# outputs: '0x680x650x6c0x6c0x6f'

This basically splits the string into chars, does the conversion through hex(ord(char)), and joins the chars back together. In case you want the result without the prefix 0x then do:

h = ''.join([str(hex(ord(i)))[2:4] for i in s]);

# outputs: '68656c6c6f'

Tested with Python 3.5.3.

    public static function normalizeUrl(string $url) {
        $parts = parse_url($url);
        return $parts['scheme'] .
            '://' .
            $parts['host'] .
            implode('/', array_map('rawurlencode', explode('/', $parts['path'])));

    }

I got a situation like the code below

 for(id<-0 to 99) {
    try {
      var symbol = ctx.read("$.stocks[" + id + "].symbol").toString
      var name = ctx.read("$.stocks[" + id + "].name").toString
      stocklist(symbol) = name
    }catch {
      case ex: com.jayway.jsonpath.PathNotFoundException=>{break}
    }
  }

I am using a java lib and the mechanism is that ctx.read throw a Exception when it can find nothing. I was trapped in the situation that :I have to break the loop when a Exception was thrown, but scala.util.control.Breaks.break using Exception to break the loop ,and it was in the catch block thus it was caught.

I got ugly way to solve this: do the loop for the first time and get the count of the real length. and use it for the second loop.

take out break from Scala is not that good,when you are using some java libs.

I tried all of stack overflow and all didn't works. But this works for me:

1. Open System Configuration
2. Click Service tab
3. Uncheck all of Hyper-V related

It is a bug in IE8.

Starting with your second question: “margin: 0 auto” centers a block, but only when width of the block is set to be less that width of parent. Usually, they get to be the same. That is why text in the example below is not centered.

<div style="height: 100px; width: 500px; background-color: Yellow;">    
    <b style="display: block; margin: 0 auto; ">text</b>
</div>

Once the display style of the b element is set to block, its width defaults to the parents width. CSS spec 10.3.3 Block-level, non-replaced elements in normal flow describes how: “If 'width' is set to 'auto', any other 'auto' values become '0' and 'width' follows from the resulting equality.” The equality mentioned there is

'margin-left' + 'border-left-width' + 'padding-left' + 'width' + 'padding-right' + 'border-right-width' + 'margin-right' = width of containing block

So, normally all autos result in a block width being equal to the width of containing block.

However, this calculation should not be applied to INPUT, which is a replaced element. Replaced elements are covered by 10.3.4 Block-level, replaced elements in normal flow. Text there says: “The used value of 'width' is determined as for inline replaced elements.” The relevant part of 10.3.2 Inline, replaced elements is: “if 'width' has a computed value of 'auto', and the element has an intrinsic width, then that intrinsic width is the used value of 'width'”.

I guess that the scenario CSS cares about is IMG element. Stackoverflow logo in this example will be centered by all browsers.

<div style="height: 100px; width: 500px; background-color: Yellow;">    
    <img style="display: block; margin: 0 auto; " border="0" src="http://stackoverflow.com/content/img/so/logo.png" alt="">
</div>

INPUT element should behave the same way.

Standardize your module's imports and exports then you won't risk hitting problems with misspelled property names.

module.exports = Component should become export default Component.

CommonJS uses module.exports as a convention, however, this means that you are just working with a regular Javascript object and you are able to set the value of any key you want (whether that's exports, exoprts or exprots). There are no runtime or compile-time checks to tell you that you've messed up.

If you use ES6 (ES2015) syntax instead, then you are working with syntax and keywords. If you accidentally type exoprt default Component then it will give you a compile error to let you know.

In your case, you can simplify the Speaker component.

import React from 'react';

export default React.createClass({
  render() {
    return (
      <h1>Speaker</h1>
    )
  }
});

You could do

if item not in mylist:
     mylist.append(item)

But you should really use a set, like this :

myset = set()
myset.add(item)

EDIT: If order is important but your list is very big, you should probably use both a list and a set, like so:

mylist = []
myset = set()
for item in ...:
    if item not in myset:
        mylist.append(item)
        myset.add(item)

This way, you get fast lookup for element existence, but you keep your ordering. If you use the naive solution, you will get O(n) performance for the lookup, and that can be bad if your list is big

Or, as @larsman pointed out, you can use OrderedDict to the same effect:

from collections import OrderedDict

mydict = OrderedDict()
for item in ...:
    mydict[item] = True

I faced the same error, but only with files cloned from git that were assigned to a proprietary plugin. I realized that even after cloning the files from git, I needed to create a new project or import a project in eclipse and this resolved the error.

The Scanner class is bases on logic implemented in String next(Pattern) method. The additional API method like nextDouble() or nextFloat(). Provide the pattern inside.

Then class description says:

A simple text scanner which can parse primitive types and strings using regular expressions.

A Scanner breaks its input into tokens using a delimiter pattern, which by default matches whitespace. The resulting tokens may then be converted into values of different types using the various next methods.

From the description it can be sad that someone has forgot about char as it is a primitive type for sure.

But the concept of class is to find patterns, a char has no pattern is just next character. And this logic IMHO caused that nextChar has not been implemented.

If you need to read a filed char by char you can used more efficient class.

The state of the checkbox will be reflected on whatever model you have it bound to, in this case, $scope.answers[item.questID]

As you are probably aware, ints are stored internally in binary. Typically an int contains 32 bits, but in some environments might contain 16 or 64 bits (or even a different number, usually but not necessarily a power of two).

But for this example, let's look at 4-bit integers. Tiny, but useful for illustration purposes.

Since there are four bits in such an integer, it can assume one of 16 values; 16 is two to the fourth power, or 2 times 2 times 2 times 2. What are those values? The answer depends on whether this integer is a signed int or an unsigned int. With an unsigned int, the value is never negative; there is no sign associated with the value. Here are the 16 possible values of a four-bit unsigned int:

bits  value
0000    0
0001    1
0010    2
0011    3
0100    4
0101    5
0110    6
0111    7
1000    8
1001    9
1010   10
1011   11
1100   12
1101   13
1110   14
1111   15

... and Here are the 16 possible values of a four-bit signed int:

bits  value
0000    0
0001    1
0010    2
0011    3
0100    4
0101    5
0110    6
0111    7
1000   -8
1001   -7
1010   -6
1011   -5
1100   -4
1101   -3
1110   -2
1111   -1

As you can see, for signed ints the most significant bit is 1 if and only if the number is negative. That is why, for signed ints, this bit is known as the "sign bit".

Yes, this is new in PHP 5.3. It returns either the value of the test expression if it is evaluated as TRUE, or the alternative value if it is evaluated as FALSE.

Well, Izhar solution was OK but, personally, I am trying to avoid from code that looks as this:

 if (Build.VERSION.SDK_INT >= Build.VERSION_CODES.KITKAT) {
//Do what you need for this SDK
};

As well, I don't like to duplicate code either. In your answer I have to add such line of code in all Activities:

setStatusBarColor(findViewById(R.id.statusBarBackground),getResources().getColor(android.R.color.white));

So, I took Izhar solution and used XML to get the same result: Create a layout for the StatusBar status_bar.xml

<View xmlns:android="http://schemas.android.com/apk/res/android"
     android:layout_width="match_parent"
     android:layout_height="@dimen/statusBarHeight"
     android:background="@color/primaryColorDark"
     android:elevation="@dimen/statusBarElevation">

Notice the height and elevation attributes, these will be set in values, values-v19, values-v21 further down.

Add this layout to your activities layout using include, main_activity.xml:

<RelativeLayout xmlns:android="http://schemas.android.com/apk/res/android"
android:layout_width="match_parent"
android:layout_height="match_parent"
android:background="@color/Black" >

<include layout="@layout/status_bar"/>
<include android:id="@+id/app_bar" layout="@layout/app_bar"/>
//The rest of your layout       
</RelativeLayout>

For the Toolbar, add top margin attribute:

<android.support.v7.widget.Toolbar xmlns:android="http://schemas.android.com/apk/res/android"
xmlns:app="http://schemas.android.com/apk/res-auto"
android:layout_width="match_parent"
android:layout_height="?android:attr/actionBarSize"
android:background="@color/primaryColor"
app:theme="@style/MyCustomToolBarTheme"
app:popupTheme="@style/ThemeOverlay.AppCompat.Dark"
android:elevation="@dimen/toolbarElevation"
android:layout_marginTop="@dimen/appBarTopMargin"
android:textDirection="ltr"
android:layoutDirection="ltr">

In your appTheme style-v19.xml and styles-v21.xml, add the windowTranslucent attr:

styles-v19.xml, v21:

<resources>
<item name="android:windowTranslucentStatus">true</item>
</resources>

And finally, on your dimens, dimens-v19, dimens-v21, add the values for the Toolbar topMargin, and the height of the statusBarHeight: dimens.xml for less than KitKat:

<resources>
<dimen name="toolbarElevation">4dp</dimen>
<dimen name="appBarTopMargin">0dp</dimen>
<dimen name="statusBarHeight">0dp</dimen>
</resources>

The status bar height is always 24dp dimens-v19.xml for KitKat and above:

<resources>
<dimen name="statusBarHeight">24dp</dimen>
<dimen name="appBarTopMargin">24dp</dimen>
</resources>

dimens-v21.xml for Lolipop, just add the elevation if needed:

<resources>
<dimen name="statusBarElevation">4dp</dimen>
</resources>

This is the result for Jellybean KitKat and Lollipop:

I know this thread is old, but since a Google search brought me here, it will also do to other people who may find this useful.

Microsoft recenly launched Visual Studio Online, which is free for projects with up to 5 users:

http://www.visualstudio.com/en-us/products/visual-studio-online-overview-vs.aspx

I have been using it for a while, and it integrates completely with Visual Studio 2013. It claims integration with other IDEs too. Apart from TFS, Git can also be used with it.

I know this thread is old, but since a Google search brought me here

Another thing that might not have been mentioned already is that urllib.urlencode() will encode empty values in the dictionary as the string None instead of having that parameter as absent. I don't know if this is typically desired or not, but does not fit my use case, hence I have to use quote_plus.

maybe you can try this way

SELECT * FROM Product P WHERE (CASE WHEN @Status = 'published' THEN (CASE WHEN P.Status IN (1, 3) THEN 'TRUE' ELSE FALSE END) WHEN @Status = 'standby' THEN (CASE WHEN P.Status IN (2, 5, 9, 6) THEN 'TRUE' ELSE 'FALSE' END) WHEN @Status = 'deleted' THEN (CASE WHEN P.Status IN (4, 5, 8, 10) THEN 'TRUE' ELSE 'FALSE' END) ELSE (CASE WHEN P.Status IN (1, 3) THEN 'TRUE' ELSE 'FALSE' END) END) = 'TRUE'

In this way if @Status = 'published', the query will check if P.Status is among 1 or 3, it will return TRUE else 'FALSE'. This will be matched with TRUE at the end

Hope it helps.

You could write something like that :

public static bool HasMethod(this object objectToCheck, string methodName)
{
    var type = objectToCheck.GetType();
    return type.GetMethod(methodName) != null;
} 

Edit : you can even do an extension method and use it like this

myObject.HasMethod("SomeMethod");

null is not an object, it is a primitive value. For example, you cannot add properties to it. Sometimes people wrongly assume that it is an object, because typeof null returns "object". But that is actually a bug (that might even be fixed in ECMAScript 6).

The difference between null and undefined is as follows:

• undefined: used by JavaScript and means “no value”. Uninitialized variables, missing parameters and unknown variables have that value.

  > var noValueYet;
  > console.log(noValueYet);
  undefined
  
  > function foo(x) { console.log(x) }
  > foo()
  undefined
  
  > var obj = {};
  > console.log(obj.unknownProperty)
  undefined
  

  Accessing unknown variables, however, produces an exception:

  > unknownVariable
  ReferenceError: unknownVariable is not defined
  

• null: used by programmers to indicate “no value”, e.g. as a parameter to a function.

Examining a variable:

console.log(typeof unknownVariable === "undefined"); // true

var foo;
console.log(typeof foo === "undefined"); // true
console.log(foo === undefined); // true

var bar = null;
console.log(bar === null); // true

As a general rule, you should always use === and never == in JavaScript (== performs all kinds of conversions that can produce unexpected results). The check x == null is an edge case, because it works for both null and undefined:

> null == null
true
> undefined == null
true

A common way of checking whether a variable has a value is to convert it to boolean and see whether it is true. That conversion is performed by the if statement and the boolean operator ! (“not”).

function foo(param) {
    if (param) {
        // ...
    }
}
function foo(param) {
    if (! param) param = "abc";
}
function foo(param) {
    // || returns first operand that can't be converted to false
    param = param || "abc";
}

Drawback of this approach: All of the following values evaluate to false, so you have to be careful (e.g., the above checks can’t distinguish between undefined and 0).

• undefined, null
• Booleans: false
• Numbers: +0, -0, NaN
• Strings: ""

You can test the conversion to boolean by using Boolean as a function (normally it is a constructor, to be used with new):

> Boolean(null)
false
> Boolean("")
false
> Boolean(3-3)
false
> Boolean({})
true
> Boolean([])
true

Or you simply manage it like full IIS by using Jexus Manager for IIS Express, an open source project I work on

https://jexusmanager.com

Start a site and the process will be launched for you.

You can use the -as operator. If casting succeed you get back a number:

$numberAsString -as [int]

Ken Clark's answer didn't work in my case. It might not work in yours either. If not, try this:

SELECT * 
from table T

INNER JOIN
  (
  select id, MIN(point) MinPoint
  from table T
  group by AccountId
  ) NewT on T.id = NewT.id and T.point = NewT.MinPoint

ORDER BY game desc

Python evaluates expressions from left to right. Notice that while evaluating an assignment, the right-hand side is evaluated before the left-hand side.

Python docs: Evaluation order

That means the following for the expression a,b = b,a :

• The right-hand side b,a is evaluated, that is to say, a tuple of two elements is created in the memory. The two elements are the objects designated by the identifiers b and a, that were existing before the instruction is encountered during the execution of the program.
• Just after the creation of this tuple, no assignment of this tuple object has still been made, but it doesn't matter, Python internally knows where it is.
• Then, the left-hand side is evaluated, that is to say, the tuple is assigned to the left-hand side.
• As the left-hand side is composed of two identifiers, the tuple is unpacked in order that the first identifier a be assigned to the first element of the tuple (which is the object that was formerly b before the swap because it had name b)
  and the second identifier b is assigned to the second element of the tuple (which is the object that was formerly a before the swap because its identifiers was a)

This mechanism has effectively swapped the objects assigned to the identifiers a and b

So, to answer your question: YES, it's the standard way to swap two identifiers on two objects.
By the way, the objects are not variables, they are objects.

I got this in my browser when I changed from using http://localhost to http://145.900.50.20 (where 145.900.50.20 is my local IP address) and then changed back to localhost -- it was necessary to stay with the IP address once I had changed to that once

You need to do it through an ArrayAdapter which will adapt your ArrayList (or any other collection) to your items in your layout (ListView, Spinner etc.).

This is what the Android developer guide says:

A ListAdapter that manages a ListView backed by an array of arbitrary objects. By default this class expects that the provided resource id references a single TextView. If you want to use a more complex layout, use the constructors that also takes a field id. That field id should reference a TextView in the larger layout resource.

However the TextView is referenced, it will be filled with the toString() of each object in the array. You can add lists or arrays of custom objects. Override the toString() method of your objects to determine what text will be displayed for the item in the list.

To use something other than TextViews for the array display, for instance ImageViews, or to have some of data besides toString() results fill the views, override getView(int, View, ViewGroup) to return the type of view you want.

So your code should look like:

public class YourActivity extends Activity {

    private ListView lv;

    public void onCreate(Bundle saveInstanceState) {
         setContentView(R.layout.your_layout);

         lv = (ListView) findViewById(R.id.your_list_view_id);

         // Instanciating an array list (you don't need to do this, 
         // you already have yours).
         List<String> your_array_list = new ArrayList<String>();
         your_array_list.add("foo");
         your_array_list.add("bar");

         // This is the array adapter, it takes the context of the activity as a 
         // first parameter, the type of list view as a second parameter and your 
         // array as a third parameter.
         ArrayAdapter<String> arrayAdapter = new ArrayAdapter<String>(
                 this, 
                 android.R.layout.simple_list_item_1,
                 your_array_list );

         lv.setAdapter(arrayAdapter); 
    }
}

In Activity A

private void startSwitcher() {
    int yourInt = 200;
    Intent myIntent = new Intent(A.this, B.class);
    intent.putExtra("yourIntName", yourInt);
    startActivity(myIntent);
}

in Activity B

int score = getIntent().getIntExtra("yourIntName", 0);

You can pass data from PHP to javascript but the only way to get data from javascript to PHP is via AJAX.

The reason for that is you can build a valid javascript through PHP but to get data to PHP you will need to get PHP running again, and since PHP only runs to process the output, you will need a page reload or an asynchronous query.

A single liner, using find:

find -type f -exec grep -lm1 "PATTERN" {} \; -a -quit

First you need to define the format of date column.

df['Date'] = pd.to_datetime(df.Date, format='%Y-%m-%d %H:%M:%S')

For your case base format can be set to;

df['Date'] = pd.to_datetime(df.Date, format='%Y-%m-%d')

After that you can set/change your desired output as follows;

df['Date'] = df['Date'].dt.strftime('%Y-%m-%d')

I just had this error message happen. When searching for why, I figured out that there's a related issue that can occur if you're not paying attention - the same error occurs if the directory you are trying to move into does not exist.

I had this problem with my PLAY Java application. This is my stack trace for that exception:

javax.persistence.PersistenceException: Error[Incorrect string value: '\xE0\xA6\xAC\xE0\xA6\xBE...' for column 'product_name' at row 1]
  at io.ebean.config.dbplatform.SqlCodeTranslator.translate(SqlCodeTranslator.java:52)
  at io.ebean.config.dbplatform.DatabasePlatform.translate(DatabasePlatform.java:192)
  at io.ebeaninternal.server.persist.dml.DmlBeanPersister.execute(DmlBeanPersister.java:83)
  at io.ebeaninternal.server.persist.dml.DmlBeanPersister.insert(DmlBeanPersister.java:49)
  at io.ebeaninternal.server.core.PersistRequestBean.executeInsert(PersistRequestBean.java:1136)
  at io.ebeaninternal.server.core.PersistRequestBean.executeNow(PersistRequestBean.java:723)
  at io.ebeaninternal.server.core.PersistRequestBean.executeNoBatch(PersistRequestBean.java:778)
  at io.ebeaninternal.server.core.PersistRequestBean.executeOrQueue(PersistRequestBean.java:769)
  at io.ebeaninternal.server.persist.DefaultPersister.insert(DefaultPersister.java:456)
  at io.ebeaninternal.server.persist.DefaultPersister.insert(DefaultPersister.java:406)
  at io.ebeaninternal.server.persist.DefaultPersister.save(DefaultPersister.java:393)
  at io.ebeaninternal.server.core.DefaultServer.save(DefaultServer.java:1602)
  at io.ebeaninternal.server.core.DefaultServer.save(DefaultServer.java:1594)
  at io.ebean.Model.save(Model.java:190)
  at models.Product.create(Product.java:147)
  at controllers.PushData.xlsupload(PushData.java:67)
  at router.Routes$$anonfun$routes$1.$anonfun$applyOrElse$40(Routes.scala:690)
  at play.core.routing.HandlerInvokerFactory$$anon$3.resultCall(HandlerInvoker.scala:134)
  at play.core.routing.HandlerInvokerFactory$$anon$3.resultCall(HandlerInvoker.scala:133)
  at play.core.routing.HandlerInvokerFactory$JavaActionInvokerFactory$$anon$8$$anon$2$$anon$1.invocation(HandlerInvoker.scala:108)
  at play.core.j.JavaAction$$anon$1.call(JavaAction.scala:88)
  at play.http.DefaultActionCreator$1.call(DefaultActionCreator.java:31)
  at play.core.j.JavaAction.$anonfun$apply$8(JavaAction.scala:138)
  at scala.concurrent.Future$.$anonfun$apply$1(Future.scala:655)
  at scala.util.Success.$anonfun$map$1(Try.scala:251)
  at scala.util.Success.map(Try.scala:209)
  at scala.concurrent.Future.$anonfun$map$1(Future.scala:289)
  at scala.concurrent.impl.Promise.liftedTree1$1(Promise.scala:29)
  at scala.concurrent.impl.Promise.$anonfun$transform$1(Promise.scala:29)
  at scala.concurrent.impl.CallbackRunnable.run$$$capture(Promise.scala:60)
  at scala.concurrent.impl.CallbackRunnable.run(Promise.scala)
  at play.core.j.HttpExecutionContext$$anon$2.run(HttpExecutionContext.scala:56)
  at play.api.libs.streams.Execution$trampoline$.execute(Execution.scala:70)
  at play.core.j.HttpExecutionContext.execute(HttpExecutionContext.scala:48)
  at scala.concurrent.impl.CallbackRunnable.executeWithValue(Promise.scala:68)
  at scala.concurrent.impl.Promise$KeptPromise$Kept.onComplete(Promise.scala:368)
  at scala.concurrent.impl.Promise$KeptPromise$Kept.onComplete$(Promise.scala:367)
  at scala.concurrent.impl.Promise$KeptPromise$Successful.onComplete(Promise.scala:375)
  at scala.concurrent.impl.Promise.transform(Promise.scala:29)
  at scala.concurrent.impl.Promise.transform$(Promise.scala:27)
  at scala.concurrent.impl.Promise$KeptPromise$Successful.transform(Promise.scala:375)
  at scala.concurrent.Future.map(Future.scala:289)
  at scala.concurrent.Future.map$(Future.scala:289)
  at scala.concurrent.impl.Promise$KeptPromise$Successful.map(Promise.scala:375)
  at scala.concurrent.Future$.apply(Future.scala:655)
  at play.core.j.JavaAction.apply(JavaAction.scala:138)
  at play.api.mvc.Action.$anonfun$apply$2(Action.scala:96)
  at scala.concurrent.Future.$anonfun$flatMap$1(Future.scala:304)
  at scala.concurrent.impl.Promise.$anonfun$transformWith$1(Promise.scala:37)
  at scala.concurrent.impl.CallbackRunnable.run$$$capture(Promise.scala:60)
  at scala.concurrent.impl.CallbackRunnable.run(Promise.scala)
  at akka.dispatch.BatchingExecutor$AbstractBatch.processBatch(BatchingExecutor.scala:55)
  at akka.dispatch.BatchingExecutor$BlockableBatch.$anonfun$run$1(BatchingExecutor.scala:91)
  at scala.runtime.java8.JFunction0$mcV$sp.apply(JFunction0$mcV$sp.java:12)
  at scala.concurrent.BlockContext$.withBlockContext(BlockContext.scala:81)
  at akka.dispatch.BatchingExecutor$BlockableBatch.run(BatchingExecutor.scala:91)
  at akka.dispatch.TaskInvocation.run(AbstractDispatcher.scala:40)
  at akka.dispatch.ForkJoinExecutorConfigurator$AkkaForkJoinTask.exec(ForkJoinExecutorConfigurator.scala:43)
  at akka.dispatch.forkjoin.ForkJoinTask.doExec(ForkJoinTask.java:260)
  at akka.dispatch.forkjoin.ForkJoinPool$WorkQueue.runTask(ForkJoinPool.java:1339)
  at akka.dispatch.forkjoin.ForkJoinPool.runWorker(ForkJoinPool.java:1979)
  at akka.dispatch.forkjoin.ForkJoinWorkerThread.run(ForkJoinWorkerThread.java:107)
Caused by: java.sql.SQLException: Incorrect string value: '\xE0\xA6\xAC\xE0\xA6\xBE...' for column 'product_name' at row 1
  at com.mysql.jdbc.SQLError.createSQLException(SQLError.java:1074)
  at com.mysql.jdbc.MysqlIO.checkErrorPacket(MysqlIO.java:4096)
  at com.mysql.jdbc.MysqlIO.checkErrorPacket(MysqlIO.java:4028)
  at com.mysql.jdbc.MysqlIO.sendCommand(MysqlIO.java:2490)
  at com.mysql.jdbc.MysqlIO.sqlQueryDirect(MysqlIO.java:2651)
  at com.mysql.jdbc.ConnectionImpl.execSQL(ConnectionImpl.java:2734)
  at com.mysql.jdbc.PreparedStatement.executeInternal(PreparedStatement.java:2155)
  at com.mysql.jdbc.PreparedStatement.executeUpdate(PreparedStatement.java:2458)
  at com.mysql.jdbc.PreparedStatement.executeUpdate(PreparedStatement.java:2375)
  at com.mysql.jdbc.PreparedStatement.executeUpdate(PreparedStatement.java:2359)
  at com.zaxxer.hikari.pool.ProxyPreparedStatement.executeUpdate(ProxyPreparedStatement.java:61)
  at com.zaxxer.hikari.pool.HikariProxyPreparedStatement.executeUpdate(HikariProxyPreparedStatement.java)
  at io.ebeaninternal.server.type.DataBind.executeUpdate(DataBind.java:82)
  at io.ebeaninternal.server.persist.dml.InsertHandler.execute(InsertHandler.java:122)
  at io.ebeaninternal.server.persist.dml.DmlBeanPersister.execute(DmlBeanPersister.java:73)
  ... 59 more

I was trying to save a record using io.Ebean. I fixed it by re creating my database with utf8mb4 collation, and applied play evolution to re create all tables so that all tables should be recreated with utf-8 collation.

CREATE DATABASE inventory CHARACTER SET utf8mb4 COLLATE utf8mb4_unicode_ci;

Given:

declare @hexStr varchar(16), @intVal int

IntToHexStr:

select @hexStr = convert(varbinary, @intVal, 1)

HexStrToInt:

declare
    @query varchar(100),
    @parameters varchar(50)

select
    @query = 'select @result = convert(int,' + @hb + ')',
    @parameters = '@result int output'

exec master.dbo.Sp_executesql @query, @parameters, @intVal output

Use the property card_view:cardBackgroundColor:

<android.support.v7.widget.CardView xmlns:card_view="http://schemas.android.com/apk/res-auto"
    android:id="@+id/card_view"
    android:layout_width="fill_parent"
    android:layout_height="150dp"
    android:layout_gravity="center"
    card_view:cardCornerRadius="4dp"
    android:layout_margin="10dp"
    card_view:cardBackgroundColor="#fff"
    >

Try this:

mongod --fork --logpath /var/log/mongodb.log

You may need to create the db-folder:

mkdir -p /data/db

If you get any 'Permission denied'-error, I'ld recommend changing the permissions of the particular files instead of running mongod as root.

you need to set URL as 10.0.2.2:portNr

portNr = the given port by ASP.NET Development Server my current service is running on localhost:3229/Service.svc

so my url is 10.0.2.2:3229

i'd fixed my problem this way

i hope it helps...

Using Hadley Wickham's purrr package this is quite simple. Use split to split the passed data_frame into groups, then use map to apply the summary function to each group.

library(purrr)

df %>% split(.$group) %>% map(summary)

try this:

ren "File 1-1" "File 1 - %date:/=-% %time::=-%"

1st add font.ttf file on font Folder. Then Add this line in onCreate method

    Typeface typeface = ResourcesCompat.getFont(getApplicationContext(), R.font.myfont);
    mytextView.setTypeface(typeface);

And here is my xml

            <TextView
            android:id="@+id/idtext1"
            android:layout_width="wrap_content"
            android:layout_height="wrap_content"
            android:layout_marginTop="7dp"
            android:gravity="center"
            android:text="My Text"
            android:textColor="#000"
            android:textSize="10sp"
        />

As explained in the previous answer it provides a simple interface to the consuming client. For example: "watch ESPN" is the intended function. But it involves several steps like:

1. Switch on TV if required;
2. Check for satellite/cable functioning;
3. Switch to ESPN if required.

But the facade will simplify this and just provide "watch ESPN" function to the client.

It seems that the only correct answer here so far has been given by romnex: "onDestroy() may not be called at all". Even though in practice, in almost all cases it will, there is no guarantee: The documentation on finish() only promises that the result of the activity is propagated back to the caller, but nothing more. Moreover, the lifecycle documentation clarifies that the activity is killable by the OS as soon as onStop() finishes (or even earlier on older devices), which, even though unlikely and therefore rare to observe in a simple test, might mean that the activity might be killed while or even before onDestroy() is executed.

So if you want to make sure some work is done when you call finish(), you cannot put it in onDestroy(), but will need to do in the same place where you call finish(), right before actually calling it.

You can use formatDate() to format the date in services or component ts. syntax:-

    formatDate(value: string | number | Date, format: string, locale: string, timezone?: string): string

import the formatDate() from common module like this,

    import { formatDate } from '@angular/common';

and just use it in the class like this ,

    formatDate(new Date(), 'MMMM dd yyyy', 'en');

You can also use the predefined format options provided by angular like this ,

    formatDate(new Date(), 'shortDate', 'en');

You can see all other predefined format options here ,

https://angular.io/api/common/DatePipe

Passing it a few ws will ignore the display width.

iOS 9 and OSX 10.11 require TLSv1.2 SSL for all hosts you plan to request data from unless you specify exception domains in your app's Info.plist file.

The syntax for the Info.plist configuration looks like this:

<key>NSAppTransportSecurity</key>
<dict>
  <key>NSExceptionDomains</key>
  <dict>
    <key>yourserver.com</key>
    <dict>
      <!--Include to allow subdomains-->
      <key>NSIncludesSubdomains</key>
      <true/>
      <!--Include to allow insecure HTTP requests-->
      <key>NSExceptionAllowsInsecureHTTPLoads</key>
      <true/>
      <!--Include to specify minimum TLS version-->
      <key>NSExceptionMinimumTLSVersion</key>
      <string>TLSv1.1</string>
    </dict>
  </dict>
</dict>

If your application (a third-party web browser, for instance) needs to connect to arbitrary hosts, you can configure it like this:

<key>NSAppTransportSecurity</key>
<dict>
    <!--Connect to anything (this is probably BAD)-->
    <key>NSAllowsArbitraryLoads</key>
    <true/>
</dict>

If you're having to do this, it's probably best to update your servers to use TLSv1.2 and SSL, if they're not already doing so. This should be considered a temporary workaround.

As of today, the prerelease documentation makes no mention of any of these configuration options in any specific way. Once it does, I'll update the answer to link to the relevant documentation.

The int values are already zero after initialization, as everyone has mentioned. If you have a situation where you actually do need to set array values to zero and want to optimize that, use System.arraycopy:

static private int[] zeros = new float[64];
...
int[] values = ...
if (zeros.length < values.length) zeros = new int[values.length];
System.arraycopy(zeros, 0, values, 0, values.length);

This uses memcpy under the covers in most or all JRE implementations. Note the use of a static like this is safe even with multiple threads, since the worst case is multiple threads reallocate zeros concurrently, which doesn't hurt anything.

You could also use Arrays.fill as some others have mentioned. Arrays.fill could use memcpy in a smart JVM, but is probably just a Java loop and the bounds checking that entails.

Benchmark your optimizations, of course.

As you can see, 01/01/1970 returns 0, which means it is the lowest possible date.

new Date('1970-01-01Z00:00:00:000') //returns Thu Jan 01 1970 01:00:00 GMT+0100 (Central European Standard Time)
new Date('1970-01-01Z00:00:00:000').getTime() //returns 0
new Date('1970-01-01Z00:00:00:001').getTime() //returns 1

XML Schemas describe hierarchial data models and may not map well to a relational data model. Mapping XSD's to database tables is very similar mapping objects to database tables, in fact you could use a framework like Castor that does both, it allows you to take a XML schema and generate classes, database tables, and data access code. I suppose there are now many tools that do the same thing, but there will be a learning curve and the default mappings will most like not be what you want, so you have to spend time customizing whatever tool you use.

XSLT might be the fastest way to generate exactly the code that you want. If it is a small schema hardcoding it might be faster than evaluating and learing a bunch of new technologies.

Late answer, but I figured it's worth keeping this updated. Expanding on Andrea Turri answer to reflect updated jQuery API and .success/.error deprecated methods.

As of jQuery 1.8.* the preferred way of doing this is to use .done() and .fail(). Jquery Docs

e.g.

$('#my_get_related_keywords').click(function() {

    var ajaxRequest = $.ajax({
        type: "POST",
        url: "HERE PUT THE PATH OF YOUR SERVICE OR PAGE",
        data: '{"HERE YOU CAN PUT DATA TO PASS AT THE SERVICE"}',
        contentType: "application/json; charset=utf-8",
        dataType: "json"});

    //When the request successfully finished, execute passed in function
    ajaxRequest.done(function(msg){
           //do something
    });

    //When the request failed, execute the passed in function
    ajaxRequest.fail(function(jqXHR, status){
        //do something else
    });
});

@media queries serve this purpose. Here's an example:

@media only screen and (max-width: 991px) and (min-width: 769px){
 /* CSS that should be displayed if width is equal to or less than 991px and larger 
  than 768px goes here */
}

@media only screen and (max-width: 991px){
 /* CSS that should be displayed if width is equal to or less than 991px goes here */
}

General algorithms for date manipulation convert dates to and from Julian Day Numbers. Here is a link to a description of such algorithms, a description of the best algorithms currently known, and the mathematical proofs of each of them: http://web.archive.org/web/20140910060704/http://mysite.verizon.net/aesir_research/date/date0.htm

The accepted answer already mentions PEP 238. I just want to add a quick look behind the scenes for those interested in what's going on without reading the whole PEP.

Python maps operators like +, -, * and / to special functions, such that e.g. a + b is equivalent to

a.__add__(b)

Regarding division in Python 2, there is by default only / which maps to __div__ and the result is dependent on the input types (e.g. int, float).

Python 2.2 introduced the __future__ feature division, which changed the division semantics the following way (TL;DR of PEP 238):

• / maps to __truediv__ which must "return a reasonable approximation of the mathematical result of the division" (quote from PEP 238)
• // maps to __floordiv__, which should return the floored result of /

With Python 3.0, the changes of PEP 238 became the default behaviour and there is no more special method __div__ in Python's object model.

If you want to use the same code in Python 2 and Python 3 use

from __future__ import division

and stick to the PEP 238 semantics of / and //.

If you have already set labels for your axis like how @andyhasit and @Marcus mentioned, and would like to change it at a later time, then you can try this:

chart.options.scales.yAxes[ 0 ].scaleLabel.labelString = "New Label";

Full config for reference:

var chartConfig = {
    type: 'line',
    data: {
      datasets: [ {
        label: 'DefaultLabel',
        backgroundColor: '#ff0000',
        borderColor: '#ff0000',
        fill: false,
        data: [],
      } ]
    },
    options: {
      responsive: true,
      scales: {
        xAxes: [ {
          type: 'time',
          display: true,
          scaleLabel: {
            display: true,
            labelString: 'Date'
          },
          ticks: {
            major: {
              fontStyle: 'bold',
              fontColor: '#FF0000'
            }
          }
        } ],
        yAxes: [ {
          display: true,
          scaleLabel: {
            display: true,
            labelString: 'value'
          }
        } ]
      }
    }
  };

• In your Xampp folder, open php.ini file inside the PHP folder i.e xampp\php\php.ini (with a text editor).

• Search for extension=mysqli (Ctrl+F), if there are two, look for the one that has been uncommented (without ";" behind)

• Change the mysqli with the correct path address i.e extension=C:\xampp\php\ext\php_mysqli.dll.

• On your Xampp control panel, stop and start apache and MySQL

I came up with the following extension that takes care of removing elements from an Array, assuming the elements in the Array implement Equatable:

extension Array where Element: Equatable {
  
  mutating func removeEqualItems(_ item: Element) {
    self = self.filter { (currentItem: Element) -> Bool in
      return currentItem != item
    }
  }

  mutating func removeFirstEqualItem(_ item: Element) {
    guard var currentItem = self.first else { return }
    var index = 0
    while currentItem != item {
      index += 1
      currentItem = self[index]
    }
    self.remove(at: index)
  }
  
}
  

Usage:

var test1 = [1, 2, 1, 2]
test1.removeEqualItems(2) // [1, 1]

var test2 = [1, 2, 1, 2]
test2.removeFirstEqualItem(2) // [1, 1, 2]

It may be handy when using shared memory, among other uses... For example: http://www.boost.org/doc/libs/1_51_0/doc/html/interprocess/synchronization_mechanisms.html#interprocess.synchronization_mechanisms.conditions.conditions_anonymous_example

The workaround that works in Mobile Safari at this time of writing, is to have the the third argument in addEventListener be { passive: false }, so the full workaround looks like this:

document.addEventListener('touchmove', function (event) {
  if (event.scale !== 1) { event.preventDefault(); }
}, { passive: false });

You may want to check if options are supported to remain backwards compatible.

Strings are always modelled as immutable (atleast in heigher level languages python/java/javascript/Scala/Objective-C).

So any string operations like concatenation, replacements always returns a new string which contains intended value, whereas the original string will still be same.

Debug.Print outputs to the "Immediate" window.

Also, you can simply type ? and then a statement directly into the immediate window (and then press Enter) and have the output appear right below, like this:

This can be very handy to quickly output the property of an object...

? myWidget.name

...to set the property of an object...

myWidget.name = "thingy"

...or to even execute a function or line of code, while in debugging mode:

Sheet1.MyFunction()

It seems to me that all other answers here (so far) fail to address the race-condition that occurs with their proposed solutions.

Any code where you first check for the files existence, and then, a few lines later in your program, you create it, runs the risk of the file being created while you weren't looking and causing you problems (or you causing the owner of "that other file" problems).

If you want to avoid this sort of thing, I would suggest something like the following (untested):

import os

def open_if_not_exists(filename):
    try:
        fd = os.open(filename, os.O_CREAT | os.O_EXCL | os.O_WRONLY)
    except OSError, e:
        if e.errno == 17:
            print e
            return None
        else:
            raise
    else:
        return os.fdopen(fd, 'w')

This should open your file for writing if it doesn't exist already, and return a file-object. If it does exists, it will print "Ooops" and return None (untested, and based solely on reading the python documentation, so might not be 100% correct).

Use the TimeSpan object to capture your initial time element and use the methods such as AddHours or AddMinutes. To substract 3 hours, you will do AddHours(-3). To substract 45 mins, you will do AddMinutes(-45)

<strong> and <em> add extra semantic meaning to your document. It just so happens that they also give a bold and italic style to your text.

You could of course override their styling with CSS.

<b> and <i> on the other hand only apply font styling and should no longer be used. (Because you're supposed to format with CSS, and if the text was actually important then you would probably make it "strong" or "emphasised" anyway!)

Hope that makes sense.

You can also use shorthand for describe as desc for table description.

desc [db_name.]table_name;

or

use db_name;
desc table_name;

You can also use explain for table description.

explain [db_name.]table_name;

See official doc

Will give output like:

+----------+-------------+------+-----+---------+-------+
| Field    | Type        | Null | Key | Default | Extra |
+----------+-------------+------+-----+---------+-------+
| id       | int(10)     | NO   | PRI | NULL    |       |
| name     | varchar(20) | YES  |     | NULL    |       |
| age      | int(10)     | YES  |     | NULL    |       |
| sex      | varchar(10) | YES  |     | NULL    |       |
| sal      | int(10)     | YES  |     | NULL    |       |
| location | varchar(20) | YES  |     | Pune    |       |
+----------+-------------+------+-----+---------+-------+

With the constructor:

// create a vector with 20 integer elements
std::vector<int> arr(20);

for(int x = 0; x < 20; ++x)
   arr[x] = x;

echo test > afile.txt

redirects stdout to afile.txt. This is the same as doing

echo test 1> afile.txt

To redirect stderr, you do:

echo test 2> afile.txt

>& is the syntax to redirect a stream to another file descriptor - 0 is stdin, 1 is stdout, and 2 is stderr.

You can redirect stdout to stderr by doing:

echo test 1>&2 # or echo test >&2

Or vice versa:

echo test 2>&1

So, in short... 2> redirects stderr to an (unspecified) file, appending &1 redirects stderr to stdout.

I use a jQuery plugin called ColorBox, it is

1. Very easy to use
2. lightweight
3. customizable
4. the nicest popup dialog I have seen for jQuery yet

Once you have your key stored in a PEM file, you can read it back easily using PemObject and PemReader classes provided by BouncyCastle, as shown in this this tutorial.

Create a PemFile class that encapsulates file handling:

import java.io.FileInputStream;
import java.io.FileNotFoundException;
import java.io.IOException;
import java.io.InputStreamReader;

import org.bouncycastle.util.io.pem.PemObject;
import org.bouncycastle.util.io.pem.PemReader;

public class PemFile {

    private PemObject pemObject;

    public PemFile(String filename) throws FileNotFoundException, IOException {
        PemReader pemReader = new PemReader(new InputStreamReader(
                new FileInputStream(filename)));
        try {
            this.pemObject = pemReader.readPemObject();
        } finally {
            pemReader.close();
        }
    }

    public PemObject getPemObject() {
        return pemObject;
    }
}

Then instantiate private and public keys as usual:

import java.io.FileNotFoundException;
import java.io.IOException;
import java.security.KeyFactory;
import java.security.NoSuchAlgorithmException;
import java.security.NoSuchProviderException;
import java.security.PrivateKey;
import java.security.PublicKey;
import java.security.Security;
import java.security.spec.InvalidKeySpecException;
import java.security.spec.PKCS8EncodedKeySpec;
import java.security.spec.X509EncodedKeySpec;

import org.apache.log4j.Logger;
import org.bouncycastle.jce.provider.BouncyCastleProvider;

public class Main {

    protected final static Logger LOGGER = Logger.getLogger(Main.class);

    public final static String RESOURCES_DIR = "src/main/resources/rsa-sample/";

    public static void main(String[] args) throws FileNotFoundException,
            IOException, NoSuchAlgorithmException, NoSuchProviderException {
        Security.addProvider(new BouncyCastleProvider());
        LOGGER.info("BouncyCastle provider added.");

        KeyFactory factory = KeyFactory.getInstance("RSA", "BC");
        try {
            PrivateKey priv = generatePrivateKey(factory, RESOURCES_DIR
                    + "id_rsa");
            LOGGER.info(String.format("Instantiated private key: %s", priv));

            PublicKey pub = generatePublicKey(factory, RESOURCES_DIR
                    + "id_rsa.pub");
            LOGGER.info(String.format("Instantiated public key: %s", pub));
        } catch (InvalidKeySpecException e) {
            e.printStackTrace();
        }
    }

    private static PrivateKey generatePrivateKey(KeyFactory factory,
            String filename) throws InvalidKeySpecException,
            FileNotFoundException, IOException {
        PemFile pemFile = new PemFile(filename);
        byte[] content = pemFile.getPemObject().getContent();
        PKCS8EncodedKeySpec privKeySpec = new PKCS8EncodedKeySpec(content);
        return factory.generatePrivate(privKeySpec);
    }

    private static PublicKey generatePublicKey(KeyFactory factory,
            String filename) throws InvalidKeySpecException,
            FileNotFoundException, IOException {
        PemFile pemFile = new PemFile(filename);
        byte[] content = pemFile.getPemObject().getContent();
        X509EncodedKeySpec pubKeySpec = new X509EncodedKeySpec(content);
        return factory.generatePublic(pubKeySpec);
    }
}

Hope this helps.

It's always a good idea, and while it's not always required, the Windows standard is to include both.

\n actually represents a Line Feed, or the number 10, and canonically a Line Feed means just "move down one row" on terminals and teletypes.

\r represents CR, a Carriage Return, or the number 13. On Windows, Unix, and most terminals, a CR moves the cursor to the beginning of the line. (This is not the case for 8-bit computers: most of those do advance to the next line with a CR.)

Anyway, some processes, such as the text console, might add a CR automatically when you send an LF. However, since the CR simply moves to the start of the line, there's no harm in sending the CR twice.

On the other hand, dialog boxes, text boxes, and other display elements require both CR and LF to properly start a new line.

Since there's really no downside to sending both, and both are required in some situations, the simplest policy is to use both, if you're not sure.

Hi you can use Renderer2 and elementRef with just a few lines of code:

constructor(private readonly elementRef: ElementRef,
          private renderer: Renderer2) {
}
ngOnInit() {
 const script = this.renderer.createElement('script');
 script.src = 'http://iknow.com/this/does/not/work/either/file.js';
 script.onload = () => {
   console.log('script loaded');
   initFile();
 };
 this.renderer.appendChild(this.elementRef.nativeElement, script);
}

the onload function can be used to call the script functions after the script is loaded, this is very useful if you have to do the calls in the ngOnInit()

Generally, make an "inverse" commit, using:

git revert 364705c

then send it to the remote as usual:

git push

This won't delete the commit: it makes an additional commit that undoes whatever the first commit did. Anything else, not really safe, especially when the changes have already been propagated.

The printStackTrace() helps the programmer understand where the actual problem occurred. The printStackTrace() method is a member of the class Throwable in the java.lang package.

Apply DataAnnotation like:

[DisplayFormat(DataFormatString = "{0:MMM dd, yyyy}")]

In the very right bottom corner, left to the smiley there was the icon saying "Plain Text". When you click it, the menu with all languages appears where you can choose your desired language.

Personally, I'd like to use range on a channel in a goroutine:

https://play.golang.org/p/qt48vvDu8cd

package main

import (
    "fmt"
    "sync"
)

func main() {
    var wg sync.WaitGroup
    c := make(chan bool)
    wg.Add(1)
    go func() {
        defer wg.Done()
        for b := range c {
            fmt.Printf("Hello %t\n", b)
        }
    }()
    c <- true
    c <- true
    close(c)
    wg.Wait()
}

Dave has written a great post about this: http://dave.cheney.net/2013/04/30/curious-channels.

Your task declaration is incorrectly combining the Copy task type and project.copy method, resulting in a task that has nothing to copy and thus never runs. Besides, Copy isn't the right choice for renaming a directory. There is no Gradle API for renaming, but a bit of Groovy code (leveraging Java's File API) will do. Assuming Project1 is the project directory:

task renABCToXYZ {     doLast {         file("ABC").renameTo(file("XYZ"))     } } 

Looking at the bigger picture, it's probably better to add the renaming logic (i.e. the doLast task action) to the task that produces ABC.

del /s /q directorytobedeleted

Working Solution :

For Resize image with size lower then 100Kb

WriteableBitmap bitmap = new WriteableBitmap(140,140);
bitmap.SetSource(dlg.File.OpenRead());
image1.Source = bitmap;

Image img = new Image();
img.Source = bitmap;
WriteableBitmap i;

do
{
    ScaleTransform st = new ScaleTransform();
    st.ScaleX = 0.3;
    st.ScaleY = 0.3;
    i = new WriteableBitmap(img, st);
    img.Source = i;
} while (i.Pixels.Length / 1024 > 100);

More Reference at http://net4attack.blogspot.com/

For the question

How can i run a jar file in command prompt but with arguments

.

To pass arguments to the jar file at the time of execution

java -jar myjar.jar arg1 arg2

In the main() method of "Main-Class" [mentioned in the manifest.mft file]of your JAR file. you can retrieve them like this:

String arg1 = args[0];
String arg2 = args[1];

The process of creating executables for many platforms can be a little tedious, so I suggest to use a script:

#!/usr/bin/env bash

package=$1
if [[ -z "$package" ]]; then
  echo "usage: $0 <package-name>"
  exit 1
fi
package_name=$package

#the full list of the platforms: https://golang.org/doc/install/source#environment
platforms=(
"darwin/386"
"dragonfly/amd64"
"freebsd/386"
"freebsd/amd64"
"freebsd/arm"
"linux/386"
"linux/amd64"
"linux/arm"
"linux/arm64"
"netbsd/386"
"netbsd/amd64"
"netbsd/arm"
"openbsd/386"
"openbsd/amd64"
"openbsd/arm"
"plan9/386"
"plan9/amd64"
"solaris/amd64"
"windows/amd64"
"windows/386" )

for platform in "${platforms[@]}"
do
    platform_split=(${platform//\// })
    GOOS=${platform_split[0]}
    GOARCH=${platform_split[1]}
    output_name=$package_name'-'$GOOS'-'$GOARCH
    if [ $GOOS = "windows" ]; then
        output_name+='.exe'
    fi

    env GOOS=$GOOS GOARCH=$GOARCH go build -o $output_name $package
    if [ $? -ne 0 ]; then
        echo 'An error has occurred! Aborting the script execution...'
        exit 1
    fi
done

I checked this script on OSX only

gist - go-executable-build.sh

I could not get the mount command to work without specifying the dev block to mount as /system

#cat /proc/mounts returns ( only the system line here )
/dev/stl12 /system rfs ro,relatime,vfat,log_off,check=no,gid/uid/rwx,iocharset=utf8 0 0

so my working command is
mount -o rw,remount -t rfs /dev/stl12 /system

Following two lines fixed my issue

view.heightAnchor.constraintEqualToConstant(50).active = true;
view.widthAnchor.constraintEqualToConstant(350).active = true;

Swift version -

var DynamicView=UIView(frame: CGRectMake(100, 200, 100, 100))
DynamicView.backgroundColor=UIColor.greenColor()
DynamicView.layer.cornerRadius=25
DynamicView.layer.borderWidth=2
self.view.addSubview(DynamicView)
DynamicView.heightAnchor.constraintEqualToConstant(50).active = true;
DynamicView.widthAnchor.constraintEqualToConstant(350).active = true;

var DynamicView2=UIView(frame: CGRectMake(100, 310, 100, 100))
DynamicView2.backgroundColor=UIColor.greenColor()
DynamicView2.layer.cornerRadius=25
DynamicView2.layer.borderWidth=2
self.view.addSubview(DynamicView2)
DynamicView2.heightAnchor.constraintEqualToConstant(50).active = true;
DynamicView2.widthAnchor.constraintEqualToConstant(350).active = true;

var DynamicView3:UIView=UIView(frame: CGRectMake(10, 420, 355, 100))
DynamicView3.backgroundColor=UIColor.greenColor()
DynamicView3.layer.cornerRadius=25
DynamicView3.layer.borderWidth=2
self.view.addSubview(DynamicView3)

let yourLabel:UILabel = UILabel(frame: CGRectMake(110, 10, 200, 20))
yourLabel.textColor = UIColor.whiteColor()
//yourLabel.backgroundColor = UIColor.blackColor()
yourLabel.text = "mylabel text"
DynamicView3.addSubview(yourLabel)
DynamicView3.heightAnchor.constraintEqualToConstant(50).active = true;
DynamicView3.widthAnchor.constraintEqualToConstant(350).active = true;

let stackView   = UIStackView()
stackView.axis  = UILayoutConstraintAxis.Vertical
stackView.distribution  = UIStackViewDistribution.EqualSpacing
stackView.alignment = UIStackViewAlignment.Center
stackView.spacing   = 30

stackView.addArrangedSubview(DynamicView)
stackView.addArrangedSubview(DynamicView2)
stackView.addArrangedSubview(DynamicView3)

stackView.translatesAutoresizingMaskIntoConstraints = false;

self.view.addSubview(stackView)

//Constraints
stackView.centerXAnchor.constraintEqualToAnchor(self.view.centerXAnchor).active = true
stackView.centerYAnchor.constraintEqualToAnchor(self.view.centerYAnchor).active = true

You can achieve this with the display property:

html, body {
    height:100%;
    margin:0;
    padding:0;
}
#section1 {
    width:100%; /*full width*/
    min-height:90%;
    text-align:center;
    display:table; /*acts like a table*/
}
h1{
    margin:0;
    padding:0;
    vertical-align:middle; /*middle centred*/
    display:table-cell; /*acts like a table cell*/
}

Demo: http://jsfiddle.net/a3Kns/

I had another requirement like only few stash have to be removed, below code would be helpful in that case.

#!/bin/sh
for i in `seq 5 8`
do
   git stash drop stash@{$i}
done

/* will delete from 5 to 8 index*/

Without jQuery:

document.getElementById('file').onchange = function(){

  var file = this.files[0];

  var reader = new FileReader();
  reader.onload = function(progressEvent){
    // Entire file
    console.log(this.result);

    // By lines
    var lines = this.result.split('\n');
    for(var line = 0; line < lines.length; line++){
      console.log(lines[line]);
    }
  };
  reader.readAsText(file);
};

HTML:

<input type="file" name="file" id="file">

Remember to put your javascript code after the file field is rendered.

this one also gives the no.of lines in a file.

a=open('filename.txt','r')
l=a.read()
count=l.splitlines()
print(len(count))

100% is relative to the base font size, which, if you haven't set it, would be the browser's user-agent default.

To get the effect you're after, I would use a piece of JavaScript code to adjust the base font size relative to the window dimensions.

Gwerder's solution wont work because hash = hmac.read(); happens before the stream is done being finalized. Thus AngraX's issues. Also the hmac.write statement is un-necessary in this example.

Instead do this:

var crypto    = require('crypto');
var hmac;
var algorithm = 'sha1';
var key       = 'abcdeg';
var text      = 'I love cupcakes';
var hash;

hmac = crypto.createHmac(algorithm, key);

// readout format:
hmac.setEncoding('hex');
//or also commonly: hmac.setEncoding('base64');

// callback is attached as listener to stream's finish event:
hmac.end(text, function () {
    hash = hmac.read();
    //...do something with the hash...
});

More formally, if you wish, the line

hmac.end(text, function () {

could be written

hmac.end(text, 'utf8', function () {

because in this example text is a utf string

For people who are not restricted to parseInt, you can use the bitwise OR operator (which implicitly calls ToInt32 to its operands).

var value = s | 0;

// NaN | 0     ==>> 0
// ''  | 0     ==>> 0
// '5' | 0     ==>> 5
// '33Ab' | 0  ==>> 0
// '0x23' | 0  ==>> 35
// 113 | 0     ==>> 113
// -12 | 0     ==>> -12
// 3.9 | 0     ==>> 3

Note: ToInt32 is different from parseInt. (i.e. parseInt('33Ab') === 33)

Just use in your if (key=="smk")

if (key=="smk") { window.open('http://www.smkproduction.eu5.org','_blank'); }

Here is a way to do this with the Walrus Operator (new in Python 3.8)

f = open("a.txt", "r")

while (c := f.read(n)):
    process(c)

f.close()

Useful Python Docs (3.8):

Walrus operator: https://docs.python.org/3/whatsnew/3.8.html#assignment-expressions

Methods of file objects: https://docs.python.org/3/tutorial/inputoutput.html#methods-of-file-objects

What about one of these?

object.methods.sort
Class.methods.sort

   // Do you want to download an image?
   // But are u denied access?
   // well here is the solution.

    public static void DownloadImage(String search, String path) {

    // This will get input data from the server
    InputStream inputStream = null;

    // This will read the data from the server;
    OutputStream outputStream = null;

    try {
        // This will open a socket from client to server
        URL url = new URL(search);

       // This user agent is for if the server wants real humans to visit
        String USER_AGENT = "Mozilla/5.0 (Windows NT 10.0; Win64; x64) AppleWebKit/537.36 (KHTML, like Gecko) Chrome/56.0.2924.87 Safari/537.36";

       // This socket type will allow to set user_agent
        URLConnection con = url.openConnection();

        // Setting the user agent
        con.setRequestProperty("User-Agent", USER_AGENT);

        // Requesting input data from server
        inputStream = con.getInputStream();

        // Open local file writer
        outputStream = new FileOutputStream(path);

        // Limiting byte written to file per loop
        byte[] buffer = new byte[2048];

        // Increments file size
        int length;

        // Looping until server finishes
        while ((length = inputStream.read(buffer)) != -1) {
            // Writing data
            outputStream.write(buffer, 0, length);
        }
    } catch (Exception ex) {
        Logger.getLogger(WebCrawler.class.getName()).log(Level.SEVERE, null, ex);
     }

     // closing used resources
     // The computer will not be able to use the image
     // This is a must

     outputStream.close();
     inputStream.close();
}

Debugging build step failures is indeed very annoying.

The best solution I have found is to make sure that each step that does real work succeeds, and adding a check after those that fails. That way you get a committed layer that contains the outputs of the failed step that you can inspect.

A Dockerfile, with an example after the # Run DB2 silent installer line:

#
# DB2 10.5 Client Dockerfile (Part 1)
#
# Requires
#   - DB2 10.5 Client for 64bit Linux ibm_data_server_runtime_client_linuxx64_v10.5.tar.gz
#   - Response file for DB2 10.5 Client for 64bit Linux db2rtcl_nr.rsp 
#
#
# Using Ubuntu 14.04 base image as the starting point.
FROM ubuntu:14.04

MAINTAINER David Carew <[email protected]>

# DB2 prereqs (also installing sharutils package as we use the utility uuencode to generate password - all others are required for the DB2 Client) 
RUN dpkg --add-architecture i386 && apt-get update && apt-get install -y sharutils binutils libstdc++6:i386 libpam0g:i386 && ln -s /lib/i386-linux-gnu/libpam.so.0 /lib/libpam.so.0
RUN apt-get install -y libxml2


# Create user db2clnt
# Generate strong random password and allow sudo to root w/o password
#
RUN  \
   adduser --quiet --disabled-password -shell /bin/bash -home /home/db2clnt --gecos "DB2 Client" db2clnt && \
   echo db2clnt:`dd if=/dev/urandom bs=16 count=1 2>/dev/null | uuencode -| head -n 2 | grep -v begin | cut -b 2-10` | chgpasswd && \
   adduser db2clnt sudo && \
   echo '%sudo ALL=(ALL) NOPASSWD:ALL' >> /etc/sudoers

# Install DB2
RUN mkdir /install
# Copy DB2 tarball - ADD command will expand it automatically
ADD v10.5fp9_linuxx64_rtcl.tar.gz /install/
# Copy response file
COPY  db2rtcl_nr.rsp /install/
# Run  DB2 silent installer
RUN mkdir /logs
RUN (/install/rtcl/db2setup -t /logs/trace -l /logs/log -u /install/db2rtcl_nr.rsp && touch /install/done) || /bin/true
RUN test -f /install/done || (echo ERROR-------; echo install failed, see files in container /logs directory of the last container layer; echo run docker run '<last image id>' /bin/cat /logs/trace; echo ----------)
RUN test -f /install/done

# Clean up unwanted files
RUN rm -fr /install/rtcl

# Login as db2clnt user
CMD su - db2clnt

Just to add up my bit:
Remember, you're gonna need to have at least 2 areas in your MVC application to get the routeValues: { area="" } working; otherwise the area value will be used as a query-string parameter and you link will look like this: /?area=

If you don't have at least 2 areas, you can fix this behavior by:
1. editing the default route in RouteConfig.cs like this:

routes.MapRoute(
    name: "Default",
    url: "{controller}/{action}/{id}",
    defaults: new { area = "", controller = "Home", action = "Index", id = UrlParameter.Optional }
);

OR
2. Adding a dummy area to your MVC project.

In this navbar CSS, set to own color:

/* Navbar */_x000D_
.navbar-default {_x000D_
    background-color: #F8F8F8;_x000D_
    border-color: #E7E7E7;_x000D_
}_x000D_
/* Title */_x000D_
.navbar-default .navbar-brand {_x000D_
    color: #777;_x000D_
}_x000D_
.navbar-default .navbar-brand:hover,_x000D_
.navbar-default .navbar-brand:focus {_x000D_
    color: #5E5E5E;_x000D_
}_x000D_
/* Link */_x000D_
.navbar-default .navbar-nav > li > a {_x000D_
    color: #777;_x000D_
}_x000D_
.navbar-default .navbar-nav > li > a:hover,_x000D_
.navbar-default .navbar-nav > li > a:focus {_x000D_
    color: #333;_x000D_
}_x000D_
.navbar-default .navbar-nav > .active > a, _x000D_
.navbar-default .navbar-nav > .active > a:hover, _x000D_
.navbar-default .navbar-nav > .active > a:focus {_x000D_
    color: #555;_x000D_
    background-color: #E7E7E7;_x000D_
}_x000D_
.navbar-default .navbar-nav > .open > a, _x000D_
.navbar-default .navbar-nav > .open > a:hover, _x000D_
.navbar-default .navbar-nav > .open > a:focus {_x000D_
    color: #555;_x000D_
    background-color: #D5D5D5;_x000D_
}_x000D_
/* Caret */_x000D_
.navbar-default .navbar-nav > .dropdown > a .caret {_x000D_
    border-top-color: #777;_x000D_
    border-bottom-color: #777;_x000D_
}_x000D_
.navbar-default .navbar-nav > .dropdown > a:hover .caret,_x000D_
.navbar-default .navbar-nav > .dropdown > a:focus .caret {_x000D_
    border-top-color: #333;_x000D_
    border-bottom-color: #333;_x000D_
}_x000D_
.navbar-default .navbar-nav > .open > a .caret, _x000D_
.navbar-default .navbar-nav > .open > a:hover .caret, _x000D_
.navbar-default .navbar-nav > .open > a:focus .caret {_x000D_
    border-top-color: #555;_x000D_
    border-bottom-color: #555;_x000D_
}

Based on your comments in this answer, I do what you're trying to do like this:

module.exports = function (app, db) {
    var module = {};

    module.auth = function (req, res) {
        // This will be available 'outside'.
        // Authy stuff that can be used outside...
    };

    // Other stuff...
    module.pickle = function(cucumber, herbs, vinegar) {
        // This will be available 'outside'.
        // Pickling stuff...
    };

    function jarThemPickles(pickle, jar) {
        // This will be NOT available 'outside'.
        // Pickling stuff...

        return pickleJar;
    };

    return module;
};

I structure pretty much all my modules like that. Seems to work well for me.