Just be sure you have used double apostrophe ('')

String text = java.text.MessageFormat.format("You''re about to delete {0} rows.", 5);
System.out.println(text);

Edit:

Within a String, a pair of single quotes can be used to quote any arbitrary characters except single quotes. For example, pattern string "'{0}'" represents string "{0}", not a FormatElement. ...

Any unmatched quote is treated as closed at the end of the given pattern. For example, pattern string "'{0}" is treated as pattern "'{0}'".

Source http://docs.oracle.com/javase/7/docs/api/java/text/MessageFormat.html

Check if your image data contains some header information at the beginning:

imageCode = "data:image/png;base64,iVBORw0KGgoAAAANSUhEUgAAAMgAAABkC...

This will cause the above error.

Just remove everything in front of and including the first comma, and you good to go.

imageCode = "iVBORw0KGgoAAAANSUhEUgAAAMgAAABkC...

If you use this:

contentType: "application/json"

AJAX won't sent GET or POST params to the server.... dont know why.

It took me hours to lear it today.

Just Use:

$.ajax(
  { url : 'http://blabla.com/wsGetReport.php',
    data : myFormData, type : 'POST', dataType : 'json', 
    // contentType: "application/json", 
    success : function(wsQuery) { }
  }
)

My Answer is also related to Config section. If you assign values from Config file at static class of C# or Module.VB of VB, you will get this error at run time.

add key="LogPath" value="~/Error_Log/"

Using Forward slash in Web.Config also leads to this error on Run time. I just resolved this issue by putting BackSlash

add key="LogPath" value="~\Error_Log\"

I faced the same problem, and resolved it by changing the BodyStyle attribut value to "WebMessageBodyStyle.Bare" :

[OperationContract]
[WebGet(BodyStyle = WebMessageBodyStyle.Bare, RequestFormat = WebMessageFormat.Json,
        ResponseFormat = WebMessageFormat.Json, UriTemplate = "GetProjectWithGeocodings/{projectId}")]
GeoCod_Project GetProjectWithGeocodings(string projectId);

The returned object will no longer be wrapped.

I have been working on a Java servlet application that needs to construct very dynamic SQL statements for adhoc reporting purposes. The basic function of the app is to feed a bunch of named HTTP request parameters into a pre-coded query, and generate a nicely formatted table of output. I used Spring MVC and the dependency injection framework to store all of my SQL queries in XML files and load them into the reporting application, along with the table formatting information. Eventually, the reporting requirements became more complicated than the capabilities of the existing parameter mapping frameworks and I had to write my own. It was an interesting exercise in development and produced a framework for parameter mapping much more robust than anything else I could find.

The new parameter mappings looked as such:

select app.name as "App", 
       ${optional(" app.owner as "Owner", "):showOwner}
       sv.name as "Server", sum(act.trans_ct) as "Trans"
  from activity_records act, servers sv, applications app
 where act.server_id = sv.id
   and act.app_id = app.id
   and sv.id = ${integer(0,50):serverId}
   and app.id in ${integerList(50):appId}
 group by app.name, ${optional(" app.owner, "):showOwner} sv.name
 order by app.name, sv.name

The beauty of the resulting framework was that it could process HTTP request parameters directly into the query with proper type checking and limit checking. No extra mappings required for input validation. In the example query above, the parameter named serverId would be checked to make sure it could cast to an integer and was in the range of 0-50. The parameter appId would be processed as an array of integers, with a length limit of 50. If the field showOwner is present and set to "true", the bits of SQL in the quotes will be added to the generated query for the optional field mappings. field Several more parameter type mappings are available including optional segments of SQL with further parameter mappings. It allows for as complex of a query mapping as the developer can come up with. It even has controls in the report configuration to determine whether a given query will have the final mappings via a PreparedStatement or simply ran as a pre-built query.

For the sample Http request values:

showOwner: true
serverId: 20
appId: 1,2,3,5,7,11,13

It would produce the following SQL:

select app.name as "App", 
       app.owner as "Owner", 
       sv.name as "Server", sum(act.trans_ct) as "Trans"
  from activity_records act, servers sv, applications app
 where act.server_id = sv.id
   and act.app_id = app.id
   and sv.id = 20
   and app.id in (1,2,3,5,7,11,13)
 group by app.name,  app.owner,  sv.name
 order by app.name, sv.name

I really think that Spring or Hibernate or one of those frameworks should offer a more robust mapping mechanism that verifies types, allows for complex data types like arrays and other such features. I wrote my engine for only my purposes, it isn't quite read for general release. It only works with Oracle queries at the moment and all of the code belongs to a big corporation. Someday I may take my ideas and build a new open source framework, but I'm hoping one of the existing big players will take up the challenge.

This query will use index if you have it for signup_date field

SELECT users.id, DATE_FORMAT(users.signup_date, '%Y-%m-%d') 
    FROM users 
    WHERE signup_date >= CURDATE() && signup_date < (CURDATE() + INTERVAL 1 DAY)

When you print it with this print 'I\nwould\nexpect\nmultiple\nlines' you would get:

I
would
expect
multiple
lines

The \n is a new line character specially used for marking END-OF-TEXT. It signifies the end of the line or text. This characteristics is shared by many languages like C, C++ etc.

To setup you Ruby development environment on Windows:

1. Install Ruby via RubyInstaller: http://rubyinstaller.org/downloads/

2. Check your ruby version: Start - Run - type in cmd to open a windows console

3. Type in ruby -v
4. You will get something like that: ruby 2.0.0p353 (2013-11-22) [i386-mingw32]

For Ruby 2.4 or later, run the extra installation at the end to install the DevelopmentKit. If you forgot to do that, run ridk install in your windows console to install it.

For earlier versions:

5. Download and install DevelopmentKit from the same download page as Ruby Installer. Choose an ?exe file corresponding to your environment (32 bits or 64 bits and working with your version of Ruby).
6. Follow the installation instructions for DevelopmentKit described at: https://github.com/oneclick/rubyinstaller/wiki/Development-Kit#installation-instructions. Adapt it for Windows.
7. After installing DevelopmentKit you can install all needed gems by just running from the command prompt (windows console or terminal): gem install {gem name}. For example, to install rails, just run gem install rails.

Hope this helps.

No in my opinion , you can create a reference variable of an interface but you can not create an instance of an interface just like an abstract class.

You're pretty close:

var myTag = $(':selected', element).attr("myTag");

Coupled with reflection, I found Activator.CreateInstance to be very helpful in mapping stored procedure result to a custom class as described in the following answer.

if (x % 1 == 0)
    // x is an integer

Here x is a numeric primitive: short, int, long, float or double

This is my mix:

overflow-y: scroll;
height: 13em; // Initial height.
resize: vertical; // Allow user to change the vertical size.
max-height: 31em; // If you want to constrain the max size.

From Copy images from url to server, delete all images after

function getimg($url) {         
    $headers[] = 'Accept: image/gif, image/x-bitmap, image/jpeg, image/pjpeg';              
    $headers[] = 'Connection: Keep-Alive';         
    $headers[] = 'Content-type: application/x-www-form-urlencoded;charset=UTF-8';         
    $user_agent = 'php';         
    $process = curl_init($url);         
    curl_setopt($process, CURLOPT_HTTPHEADER, $headers);         
    curl_setopt($process, CURLOPT_HEADER, 0);         
    curl_setopt($process, CURLOPT_USERAGENT, $user_agent); //check here         
    curl_setopt($process, CURLOPT_TIMEOUT, 30);         
    curl_setopt($process, CURLOPT_RETURNTRANSFER, 1);         
    curl_setopt($process, CURLOPT_FOLLOWLOCATION, 1);         
    $return = curl_exec($process);         
    curl_close($process);         
    return $return;     
} 

$imgurl = 'http://www.foodtest.ru/images/big_img/sausage_3.jpg'; 
$imagename= basename($imgurl);
if(file_exists('./tmp/'.$imagename)){continue;} 
$image = getimg($imgurl); 
file_put_contents('tmp/'.$imagename,$image);       

I had the same issue and this piece of code worked out for me.

names(data)[names(data) == "oldVariableName"] <- "newVariableName"

In short, this code does the following:

names(data) looks into all the names in the dataframe (data)

[names(data) == oldVariableName] extracts the variable name (oldVariableName) you want to get renamed and <- "newVariableName" assigns the new variable name.

var iframe = $('iframe'); // or some other selector to get the iframe
$('[tokenid=' + token + ']', iframe.contents()).addClass('border');

Also note that if the src of this iframe is pointing to a different domain, due to security reasons, you will not be able to access the contents of this iframe in javascript.

First give your input type submit a name, like this name='submitform'.

and then put this in your php file

if (isset($_POST['submitform']))
    {   
    ?>
<script type="text/javascript">
window.location = "http://www.google.com/";
</script>      
    <?php
    }

Don't forget to change the url to yours.

Git Graph seems like a decent extension. After installing, you can open the graph view from the bottom status bar.

Besides len you can also use operator.length_hint (requires Python 3.4+). For a normal list both are equivalent, but length_hint makes it possible to get the length of a list-iterator, which could be useful in certain circumstances:

>>> from operator import length_hint
>>> l = ["apple", "orange", "banana"]
>>> len(l)
3
>>> length_hint(l)
3

>>> list_iterator = iter(l)
>>> len(list_iterator)
TypeError: object of type 'list_iterator' has no len()
>>> length_hint(list_iterator)
3

But length_hint is by definition only a "hint", so most of the time len is better.

I've seen several answers suggesting accessing __len__. This is all right when dealing with built-in classes like list, but it could lead to problems with custom classes, because len (and length_hint) implement some safety checks. For example, both do not allow negative lengths or lengths that exceed a certain value (the sys.maxsize value). So it's always safer to use the len function instead of the __len__ method!

This could be a CORS issue. try enabling CORS for you api.

For WebApi

  var cors = new EnableCorsAttribute("*", "*", "*");
            config.EnableCors(cors);

I found that use onclick='this.checked = false;' worked to a certain extent. A radio button that was clicked would not be selected. However, if there was a radio button that was already selected (e.g., a default value), that radio button would become unselected.

<!-- didn't completely work -->
<input type="radio" name="r1" id="r1" value="N" checked="checked" onclick='this.checked = false;'>N</input>
<input type="radio" name="r1" id="r1" value="Y" onclick='this.checked = false;'>Y</input>

For this scenario, leaving the default value alone and disabling the other radio button(s) preserves the already selected radio button and prevents it from being unselected.

<!-- preserves pre-selected value -->
<input type="radio" name="r1" id="r1" value="N" checked="checked">N</input>
<input type="radio" name="r1" id="r1" value="Y" disabled>Y</input>

This solution is not the most elegant way of preventing the default value from being changed, but it will work whether or not javascript is enabled.

For me the cause of this issue under Rails 4 was a missing,

<%= csrf_meta_tags %>

Line in my main application layout. I had accidently deleted it when I rewrote my layout.

If this isn't in the main layout you will need it in any page that you want a CSRF token on.

resize() not only allocates memory, it also creates as many instances as the desired size which you pass to resize() as argument. But reserve() only allocates memory, it doesn't create instances. That is,

std::vector<int> v1;
v1.resize(1000); //allocation + instance creation
cout <<(v1.size() == 1000)<< endl;   //prints 1
cout <<(v1.capacity()==1000)<< endl; //prints 1

std::vector<int> v2;
v2.reserve(1000); //only allocation
cout <<(v2.size() == 1000)<< endl;   //prints 0
cout <<(v2.capacity()==1000)<< endl; //prints 1

Output (online demo):

1
1
0
1

So resize() may not be desirable, if you don't want the default-created objects. It will be slow as well. Besides, if you push_back() new elements to it, the size() of the vector will further increase by allocating new memory (which also means moving the existing elements to the newly allocated memory space). If you have used reserve() at the start to ensure there is already enough allocated memory, the size() of the vector will increase when you push_back() to it, but it will not allocate new memory again until it runs out of the space you reserved for it.

You can use like:

"12-18" -like "*-*"

Or split for contains:

"12-18" -split "" -contains "-"

>>> class name(object):
...     def __init__(self, name):
...         self.name = name
... 
>>> person1 = name("jean")
>>> person2 = name("dean")
>>> person1.name
'jean'
>>> person2.name
'dean'
>>>

???????

yourStringStream.str()

Correct, when you drag a view controller object onto your storyboard in order to create a new scene, it doesn't automatically make the new class for you, too.

Having added a new view controller scene to your storyboard, you then have to:

1. Create a UIViewController subclass. For example, go to your target's folder in the project navigator panel on the left and then control-click and choose "New File...". Choose a "Cocoa Touch Class":

   

   And then select a unique name for the new view controller subclass:

   

2. Specify this new subclass as the base class for the scene you just added to the storyboard.

   

3. Now hook up any IBOutlet and IBAction references for this new scene with the new view controller subclass.

This URL format worked like a charm:

http://maps.google.com/maps?&z={INSERT_MAP_ZOOM}&mrt={INSERT_TYPE_OF_SEARCH}&t={INSERT_MAP_TYPE}&q={INSERT_MAP_LAT_COORDINATES}+{INSERT_MAP_LONG_COORDINATES}

Example for Mount Everest:

http://maps.google.com/maps?&z=15&mrt=yp&t=k&q=27.9879012+86.9253141

Full reference here:

https://moz.com/ugc/everything-you-never-wanted-to-know-about-google-maps-parameters

-- EDIT --

Apparently the zoom parameter stopped working, here's the updated format.

Format

https://www.google.com/maps/@?api=1&map_action=map&basemap=satellite&center={LAT},{LONG}&zoom={ZOOM}

Example

https://www.google.com/maps/@?api=1&map_action=map&basemap=satellite&center=27.9879012,86.9253141&zoom=14

It took me some time to get the conversion working both ways, so here are the two extension methods I came up with:

using System.Drawing;
using System.Drawing.Imaging;
using System.IO;
using System.Windows.Media.Imaging;

public static class BitmapConversion {

    public static Bitmap ToWinFormsBitmap(this BitmapSource bitmapsource) {
        using (MemoryStream stream = new MemoryStream()) {
            BitmapEncoder enc = new BmpBitmapEncoder();
            enc.Frames.Add(BitmapFrame.Create(bitmapsource));
            enc.Save(stream);

            using (var tempBitmap = new Bitmap(stream)) {
                // According to MSDN, one "must keep the stream open for the lifetime of the Bitmap."
                // So we return a copy of the new bitmap, allowing us to dispose both the bitmap and the stream.
                return new Bitmap(tempBitmap);
            }
        }
    }

    public static BitmapSource ToWpfBitmap(this Bitmap bitmap) {
        using (MemoryStream stream = new MemoryStream()) {
            bitmap.Save(stream, ImageFormat.Bmp);

            stream.Position = 0;
            BitmapImage result = new BitmapImage();
            result.BeginInit();
            // According to MSDN, "The default OnDemand cache option retains access to the stream until the image is needed."
            // Force the bitmap to load right now so we can dispose the stream.
            result.CacheOption = BitmapCacheOption.OnLoad;
            result.StreamSource = stream;
            result.EndInit();
            result.Freeze();
            return result;
        }
    }
}

DAO is native to Access and by far the best for general use. ADO has its place, but it is unlikely that this is it.

 Dim rs As DAO.Recordset
 Dim db As Database
 Dim strSQL as String

 Set db=CurrentDB

 strSQL = "select * from table where some condition"

 Set rs = db.OpenRecordset(strSQL)

 Do While Not rs.EOF

    rs.Edit
    rs!SomeField = "Abc"
    rs!OtherField = 2
    rs!ADate = Date()
    rs.Update

    rs.MoveNext
Loop

Using curl:

curl -v \
    -F "r=releases" \
    -F "g=com.acme.widgets" \
    -F "a=widget" \
    -F "v=0.1-1" \
    -F "p=tar.gz" \
    -F "file=@./widget-0.1-1.tar.gz" \
    -u myuser:mypassword \
    http://localhost:8081/nexus/service/local/artifact/maven/content

You can see what the parameters mean here: https://support.sonatype.com/entries/22189106-How-can-I-programatically-upload-an-artifact-into-Nexus-

To make the permissions for this work, I created a new role in the admin GUI and I added two privileges to that role: Artifact Download and Artifact Upload. The standard "Repo: All Maven Repositories (Full Control)"-role is not enough. You won't find this in the REST API documentation that comes bundled with the Nexus server, so these parameters might change in the future.

On a Sonatype JIRA issue, it was mentioned that they "are going to overhaul the REST API (and the way it's documentation is generated) in an upcoming release, most likely later this year".

Another way to do this is to tell Mongoose that all you need is a plain JavaScript version of the returned doc by using lean() in the query chain. That way Mongoose skips the step of creating the full model instance and you directly get a doc you can modify:

MyModel.findOne().lean().exec(function(err, doc) {
    doc.addedProperty = 'foobar';
    res.json(doc);
});

Using "fill_parent" alone for the layout_width will do the trick:

 <ImageView
    android:id="@+id/image"
    android:layout_width="fill_parent"
    android:layout_height="wrap_content"
    android:layout_marginRight="6dip"
    android:background="#0000" 
    android:src="@drawable/icon1" />

**check this out** 
<!DOCTYPE html>
<html>
<head>
<script src="http://ajax.googleapis.com/ajax/libs/jquery/1.11.1/jquery.min.js"></script>
<script>
$(document).ready(function(){
    $(document).keypress(function(e)
        {

            var keynum;
            if(window.event)
            { // IE                 
                keynum = e.keyCode;
            }
                else if(e.which)
                    { 
                    // Netscape/Firefox/Opera                   
                    keynum = e.which;
                    }
                    alert(String.fromCharCode(keynum));
                    var unicode=e.keyCode? e.keyCode : e.charCode;
                    alert(unicode);
        });
});  

</script>
</head>
<body>

<input type="text"></input>
</body>
</html>

You may try this

#headerAgentInfoDetailsPhone
{
    white-space:pre
}
#headerAgentInfoDetailsPhone:after {
    content:"Office: XXXXX \A Mobile: YYYYY ";
}

Js Fiddle

in addition to all answers above:

if you want to use a gif for your github repository README.md and don't want to address it from your root directory, it's not enough if you just copy the url of your browser, for example your browser URL is sth like:

https://github.com/ashkan-nasirzadeh/simpleShell/blob/master/README%20assets/shell-gif.gif

but you should open your gif in your github account and right click on it and click copy image address or sth like that which is sth like this:

https://github.com/ashkan-nasirzadeh/simpleShell/blob/master/README%20assets/shell-gif.gif?raw=true

Old post I know, but using Select(x => x) can be useful to split the EF Core (or even just Linq) expression up into a query builder.

This is handy for adding dynamic conditions.

For example:

public async Task<User> GetUser(Guid userId, string userGroup, bool noTracking = false)
{
    IQueryable<User> queryable = _context.Users.Select(x => x);

    if(!string.IsNullOrEmpty(userGroup))
        queryable = queryable.Where(x => x.UserGroup == userGroup);

    if(noTracking)
        queryable = queryable.AsNoTracking();

    return await queryable.FirstOrDefaultAsync(x => x.userId == userId);
}

The following worked for me

const conString = "postgres://YourUserName:YourPassword@YourHostname:5432/YourDatabaseName";

Difference between forEach() & map()

forEach() just loop through the elements. It's throws away return values and always returns undefined.The result of this method does not give us an output .

map() loop through the elements allocates memory and stores return values by iterating main array

Example:

   var numbers = [2,3,5,7];

   var forEachNum = numbers.forEach(function(number){
      return number
   })
   console.log(forEachNum)
   //output undefined

   var mapNum = numbers.map(function(number){
      return number
   })
   console.log(mapNum)
   //output [2,3,5,7]

map() is faster than forEach()

JSONArray jsonChildArray = (JSONArray) jsonChildArray.get("LanguageLevels");
    JSONObject secObject = (JSONObject) jsonChildArray.get(1);

I think this should work, but i do not have the possibility to test it at the moment..

In simple words

$.ajax("info.txt").done(function(data) {
  alert(data);
}).fail(function(data){
  alert("Try again champ!");
});

if its get the info.text then it will alert and whatever function you add or if any how unable to retrieve info.text from the server then alert or error function.

Use a TreeMap, although having a map "look like that" is a bit nebulous--you could also just sort the keys based on your criteria and iterate over the map, retrieving each object.

You may be looking for an automated tool that will automatically generate a lot of stuff for you. But here's a free, generally powerful diagramming tool useful not only for UML but for all kinds of diagramming tasks. It accepts as input and outputs to a wide variety of commonly used file formats. It's called yEd, and it's worth a look

It works for me.

ALTER USER 'root'@'localhost' IDENTIFIED WITH mysql_native_password BY 'password'

Adding a shell script e.g. run.sh makes it much more easier:

#!/usr/bin/env bash
export JAVA_PROGRAM_ARGS=`echo "$@"`
mvn exec:java -Dexec.mainClass="test.Main" -Dexec.args="$JAVA_PROGRAM_ARGS"

Then you are able to execute:

./run.sh arg1 arg2 arg3

In case you have only one controller and you want to access every action on root you can skip controller name like this

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

Have you tried the SVG text element?

.append("text").text(function(d, i) { return d[whichevernode];})

rect element doesn't permit text element inside of it. It only allows descriptive elements (<desc>, <metadata>, <title>) and animation elements (<animate>, <animatecolor>, <animatemotion>, <animatetransform>, <mpath>, <set>)

Append the text element as a sibling and work on positioning.

UPDATE

Using g grouping, how about something like this? fiddle

You can certainly move the logic to a CSS class you can append to, remove from the group (this.parentNode)

Yes, #id selectors combined with a multiple selector (comma) is perfectly valid in both jQuery and CSS.

However, for your example, since <script> comes before the elements, you need a document.ready handler, so it waits until the elements are in the DOM to go looking for them, like this:

<script>
  $(function() {
    $("#segement1,#segement2,#segement3").hide()
  });
</script>

<div id="segement1"></div>
<div id="segement2"></div>
<div id="segement3"></div>

You can also use .replaceWith(), like this:

$("b").replaceWith(function() { return $(this).contents(); });

Or if you know it's just a string:

$("b").replaceWith(function() { return this.innerHTML; });

This can make a big difference if you're unwrapping a lot of elements since either approach above is significantly faster than the cost of .unwrap().

It seems that IDLE changes its current working dir to location of the script that is executed, while when running the script using cmd doesn't do that and it leaves CWD as it is.

To change current working dir to the one containing your script you can use:

import os
os.chdir(os.path.dirname(__file__))
print(os.getcwd())

The __file__ variable is available only if you execute script from file, and it contains path to the file. More on it here: Python __file__ attribute absolute or relative?

InnoDB supports RENAME TABLE statement to move table from one database to another. To use it programmatically and rename database with large number of tables, I wrote a couple of procedures to get the job done. You can check it out here - SQL script @Gist

To use it simply call the renameDatabase procedure.

CALL renameDatabase('old_name', 'new_name');

Tested on MariaDB and should work ideally on all RDBMS using InnoDB transactional engine.

If you wanted to search some elements based on a regex, you can use the filter function. For example, say you wanted to make sure that in all the input boxes, the user has only entered numbers, so let's find all the inputs which don't match and highlight them.

$("input:text")
    .filter(function() {
        return this.value.match(/[^\d]/);
    })
    .addClass("inputError")
;

Of course if it was just something like this, you could use the form validation plugin, but this method could be applied to any sort of elements you like. Another example to show what I mean: Find all the elements whose id matches /[a-z]+_\d+/

$("[id]").filter(function() {
    return this.id.match(/[a-z]+_\d+/);
});

I had this issue recently,

Problem statement: Mine was a windows service that I run locally by attaching VS debugger. When I stop debugging and try to restart/stop the service (under services.msc) I used to get the mentioned error.

Solution:

1. Open up Task manager.
2. Search for the service (based on the exe name and not service name, for those that are different).
3. Kill the service.

On doing the above the service is stopped.

The trick is pretty simple... The request cycle is still pretty much alive. You can just add a new variable that will create a temporary, calling

app.get('some/url/endpoint', middleware1, middleware2);

Since you can handle your request in the first middleware

(req, res, next) => {
    var yourvalue = anyvalue
}

In middleware 1 you handle your logic and store your value like below:

req.anyvariable = yourvalue

In middleware 2 you can catch this value from middleware 1 doing the following:

(req, res, next) => {
    var storedvalue = req.yourvalue
}

you can use default ng-submitted is set if the form was submitted.

https://docs.angularjs.org/api/ng/directive/form

example: http://jsbin.com/cowufugusu/1/

I am just having hard time to understand concept behind putting wait() in object class For this questions sake consider as if wait() and notifyAll() are in thread class

In the Java language, you wait() on a particular instance of an Object – a monitor assigned to that object to be precise. If you want to send a signal to one thread that is waiting on that specific object instance then you call notify() on that object. If you want to send a signal to all threads that are waiting on that object instance, you use notifyAll() on that object.

If wait() and notify() were on the Thread instead then each thread would have to know the status of every other thread. How would thread1 know that thread2 was waiting for access to a particular resource? If thread1 needed to call thread2.notify() it would have to somehow find out that thread2 was waiting. There would need to be some mechanism for threads to register the resources or actions that they need so others could signal them when stuff was ready or available.

In Java, the object itself is the entity that is shared between threads which allows them to communicate with each other. The threads have no specific knowledge of each other and they can run asynchronously. They run and they lock, wait, and notify on the object that they want to get access to. They have no knowledge of other threads and don't need to know their status. They don't need to know that it is thread2 which is waiting for the resource – they just notify on the resource and whomever it is that is waiting (if anyone) will be notified.

In Java, we then use objects as synchronization, mutex, and communication points between threads. We synchronize on an object to get mutex access to an important code block and to synchronize memory. We wait on an object if we are waiting for some condition to change – some resource to become available. We notify on an object if we want to awaken sleeping threads.

// locks should be final objects so the object instance we are synchronizing on,
// never changes
private final Object lock = new Object();
...
// ensure that the thread has a mutex lock on some key code
synchronized (lock) {
    ...
    // i need to wait for other threads to finish with some resource
    // this releases the lock and waits on the associated monitor
    lock.wait();
    ...
    // i need to signal another thread that some state has changed and they can
    // awake and continue to run
    lock.notify();
}

There can be any number of lock objects in your program – each locking a particular resource or code segment. You might have 100 lock objects and only 4 threads. As the threads run the various parts of the program, they get exclusive access to one of the lock objects. Again, they don't have to know the running status of the other threads.

This allows you to scale up or down the number of threads running in your software as much as you want. You find that the 4 threads is blocking too much on outside resources, then you can increase the number. Pushing your battered server too hard then reduce the number of running threads. The lock objects ensure mutex and communication between the threads independent of how many threads are running.

Another case to use a multi line function would be to circumvent sql server from pushing down the where clause.

For example, I have a table with a table names and some table names are formatted like C05_2019 and C12_2018 and and all tables formatted that way have the same schema. I wanted to merge all that data into one table and parse out 05 and 12 to a CompNo column and 2018,2019 into a year column. However, there are other tables like ACA_StupidTable which I cannot extract CompNo and CompYr and would get a conversion error if I tried. So, my query was in two part, an inner query that returned only tables formatted like 'C_______' then the outer query did a sub-string and int conversion. ie Cast(Substring(2, 2) as int) as CompNo. All looks good except that sql server decided to put my Cast function before the results were filtered and so I get a mind scrambling conversion error. A multi statement table function may prevent that from happening, since it is basically a "new" table.

For InnoDB, the following seems to work: create the new empty database, then rename each table in turn into the new database:

RENAME TABLE old_db.table TO new_db.table;

You will need to adjust the permissions after that.

For scripting in a shell, you can use either of the following:

mysql -u username -ppassword old_db -sNe 'show tables' | while read table; \ 
    do mysql -u username -ppassword -sNe "rename table old_db.$table to new_db.$table"; done

OR

for table in `mysql -u root -ppassword -s -N -e "use old_db;show tables from old_db;"`; do mysql -u root -ppassword -s -N -e "use old_db;rename table old_db.$table to new_db.$table;"; done;

Notes:

• There is no space between the option -p and the password. If your database has no password, remove the -u username -ppassword part.

• If some table has a trigger, it cannot be moved to another database using above method (will result Trigger in wrong schema error). If that is the case, use a traditional way to clone a database and then drop the old one:

  mysqldump old_db | mysql new_db

• If you have stored procedures, you can copy them afterwards:

  mysqldump -R old_db | mysql new_db

Another way to link Angular Paginator with the data table using Slice Pipe.Here data is fetched only once from server.

View:

 <div class="col-md-3" *ngFor="let productObj of productListData | 
     slice: lowValue : highValue">
       //actual data dispaly  
 </div>

<mat-paginator [length]="productListData.length" [pageSize]="pageSize" 
   (page)="pageEvent = getPaginatorData($event)">
</mat-paginator> 

Component

    pageIndex:number = 0;
    pageSize:number = 50;
    lowValue:number = 0;
    highValue:number = 50;       

  getPaginatorData(event){
     console.log(event);
     if(event.pageIndex === this.pageIndex + 1){
        this.lowValue = this.lowValue + this.pageSize;
        this.highValue =  this.highValue + this.pageSize;
       }
    else if(event.pageIndex === this.pageIndex - 1){
       this.lowValue = this.lowValue - this.pageSize;
       this.highValue =  this.highValue - this.pageSize;
      }   
       this.pageIndex = event.pageIndex;
 }

IMO, you should have a look at the Observer Pattern, and this is how most of the listeners work

This would be easy to understand:

#include <iostream>

using namespace std;

void Merge(int *a, int *L, int *R, int p, int q)
{
    int i, j=0, k=0;
    for(i=0; i<p+q; i++)
    {
        if(j==p)                       //When array L is empty
        {
            *(a+i) = *(R+k);
            k++;
        }
        else if(k==q)                  //When array R is empty
        {
            *(a+i) = *(L+j);
            j++;
        }
        else if(*(L+j) < *(R+k))  //When element in L is smaller than element in R
        {
            *(a+i) = *(L+j);
            j++;
        }
        else   //When element in R is smaller or equal to element in L
        {
            *(a+i) = *(R+k);
            k++;
        }
    }
}

void MergeSort(int *a, int len)
{
    int i, j;
    if(len > 1)
    {
        int p = len/2 + len%2;      //length of first array
        int q = len/2;              //length of second array
        int L[p];                   //first array
        int R[q];                   //second array
        for(i=0; i<p; i++)
        {
            L[i] = *(a+i);      //inserting elements in first array
        }
        for(i=0; i<q; i++)
        {
            R[i] = *(a+p+i);    //inserting elements in second array
        }
        MergeSort(&L[0], p);
        MergeSort(&R[0], q);
        Merge(a, &L[0], &R[0], p, q);   //Merge arrays L and R into A
    }
    else
    {
        return;        //if array only have one element just return
    }
}

int main()
{
    int i, n;
    int a[100000];
    cout<<"Enter numbers to sort. When you are done, enter -1\n";
    i=0;
    while(true)
    {
        cin>>n;
        if(n==-1)
        {
            break;
        }
        else
        {
            a[i] = n;
            i++;
        }
    }
    int len = i;
    MergeSort(&a[0], len);
    for(i=0; i<len; i++)
    {
        cout<<a[i]<<" ";
    }

    return 0;
}

If you're inside an iframe that don't have cross domain src, or src is empty:

Then:

function getOriginUrl() {
    var href = document.location.href;
    var referrer = document.referrer;
    // Check if window.frameElement not null
    if(window.frameElement) {
        href = window.frameElement.ownerDocument.location.href;
        // This one will be origin
        if(window.frameElement.ownerDocument.referrer != "") {
            referrer = window.frameElement.ownerDocument.referrer;
        }
    }
    // Compare if href not equal to referrer
    if(href != referrer) {
        // Take referrer as origin
        return referrer;
    } else {
        // Take href
        return href
    }
}

If you're inside an iframe with cross domain src:

Then:

function getOriginUrl() {
    var href = document.location.href;
    var referrer = document.referrer;
    // Detect if you're inside an iframe
    if(window.parent != window) {
        // Take referrer as origin
        return referrer;
    } else {
        // Take href
        return href;
    }
}

Cookies that 'expire at end of the session' expire unpredictably from the user's perspective!

On iOS with Safari they expire whenever you switch apps!

On Android with Chrome they don't expire when you close the browser.

On Windows desktop running Chrome they expire when you close the browser. That's not when you close your website's tab; its when you close all tabs. Nor do they expire if there are any other browser windows open. If users run web apps as windows they might not even know they are browser windows. So your cookie's life depends on what the user is doing with some apparently unrelated app.

I had a similar issue in Ubuntu 16.04 and what help me was that i installed php-mysql for php 7.2. I would recommend you run the following command if you have php 7.2 or install php mysql depending on your version of PHP. Make sure that you have installed the DBAL packege

apt-get install php7.2-mysql

systemctl restart apache2

The accepted answer is good and simple approach but have limitation of information it can hold with tag. As sometime more information needed.

You can create a custom button and add properties as many as you like they will hold info you wanna pass:

class CustomButton: UIButton {
    var orderNo = -1
    var clientCreatedDate = Date(timeIntervalSince1970: 1)
}

Make button of this type in Storyboard or programmatically:

protocol OrderStatusDelegate: class {
    func orderStatusUpdated(orderNo: Int, createdDate: Date)
}

class OrdersCell: UITableViewCell {
    @IBOutlet weak var btnBottom: CustomButton!
    weak var delegate: OrderStatusDelegate?
}

While configuring the cell add values to these properties:

 func configureCell(order: OrderRealm, index: Int) {
     btnBottom.orderNo = Int(order.orderNo)
     btnBottom.clientCreatedDate = order.clientCreatedDate
 }

When tapped access those properties in button's action (within cell's subclass) that can be sent through delegate:

@IBAction func btnBumpTapped(_ sender: Any) {
    if let button = sender as? CustomButton {
        let orderNo = button.orderNo
        let createdDate = button.clientCreatedDate
        delegate?.orderStatusUpdated(orderNo: orderNo, createdDate: createdDate)
    }
}

The latest version of sqlcmd adds the -w option to remove extra space after the field value; however, it does not put quotes around strings, which can be a problem with CSV files when importing a field value that contains a comma.

The shiny package provides the convenient functions validate() and need() for checking that variables are both available and valid. need() evaluates an expression. If the expression is not valid, then an error message is returned. If the expression is valid, NULL is returned. One can use this to check if a variable is valid. See ?need for more information.

I suggest defining a function like this:

is.valid <- function(x) {
  require(shiny)
  is.null(need(x, message = FALSE))  
}

This function is.valid() will return FALSE if x is FALSE, NULL, NA, NaN, an empty string "", an empty atomic vector, a vector containing only missing values, a logical vector containing only FALSE, or an object of class try-error. In all other cases, it returns TRUE.

That means, need() (and is.valid()) covers a really broad range of failure cases. Instead of writing:

if (!is.null(x) && !is.na(x) && !is.nan(x)) {
  ...
}

one can write simply:

if (is.valid(x)) {
  ...
}

With the check for class try-error, it can even be used in conjunction with a try() block to silently catch errors: (see https://csgillespie.github.io/efficientR/programming.html#communicating-with-the-user)

bad = try(1 + "1", silent = TRUE)
if (is.valid(bad)) {
  ...
}

MySQL

COALESCE(field, 'default')

For example:

  SELECT
    t.id,
    COALESCE(d.field, 'default')
  FROM
     test t
  LEFT JOIN
     detail d ON t.id = d.item

Also, you can use multiple columns to check their NULL by COALESCE function. For example:

mysql> SELECT COALESCE(NULL, 1, NULL);
        -> 1
mysql> SELECT COALESCE(0, 1, NULL);
        -> 0
mysql> SELECT COALESCE(NULL, NULL, NULL);
        -> NULL

This will allow file names to have spaces (reproducible example).

# Select files of interest, here, only text files for ex.
find -type f -exec file {} \; > findresult.txt
grep ": ASCII text$" findresult.txt > textfiles.txt
# leave only the path to the file removing suffix and prefix
sed -i -e 's/:.*$//' textfiles.txt
sed -i -e 's/\.\///' textfiles.txt

#write a script that deletes the files in textfiles.txt
IFS_backup=$IFS
IFS=$(echo "\n\b")
for f in $(cat textfiles.txt); 
do 
rm "$f"; 
done
IFS=$IFS_backup

# save script as "some.sh" and run: sh some.sh

You probablly have 2 different versions of hibernate-jpa-api on the classpath. To check that run:

mvn dependency:tree >dep.txt

Then search if there are hibernate-jpa-2.0-api and hibernate-jpa-2.1-api. And exclude the excess one.

I got it working by

• Installing the Google Play Services through the Android SDK Manager
• Using a Galaxy Nexus Device (4.65", 720 x 1280: xhdpi)
• Targeting the Android 4.2.2 Google API Level 17

Here is a good article on the matter - "Use references when you can, and pointers when you have to."

I don't have a php-cgi binary on Ubuntu, so I did this:

% alias php-cgi="php -r '"'parse_str(implode("&", array_slice($argv, 2)), $_GET); include($argv[1]);'"' --"
% php-cgi test1.php foo=123
<html>
You set foo to 123.
</html>

%cat test1.php
<html>You set foo to <?php print $_GET['foo']?>.</html>

Hi I know this was asked a while ago but I've just figured this out and it might help someone else. Not sure if this is exactly what you're looking for but this is how I call a stored proc and view the output using SQL Developer.
In SQL Developer when viewing the proc, right click and choose 'Run' or select Ctrl+F11 to bring up the Run PL/SQL window. This creates a template with the input and output params which you need to modify. My proc returns a sys_refcursor. The tricky part for me was declaring a row type that is exactly equivalent to the select stmt / sys_refcursor being returned by the proc:

DECLARE
  P_CAE_SEC_ID_N NUMBER;
  P_FM_SEC_CODE_C VARCHAR2(200);
  P_PAGE_INDEX NUMBER;
  P_PAGE_SIZE NUMBER;
  v_Return sys_refcursor;
  type t_row is record (CAE_SEC_ID NUMBER,FM_SEC_CODE VARCHAR2(7),rownum number, v_total_count number);
  v_rec t_row;

BEGIN
  P_CAE_SEC_ID_N := NULL;
  P_FM_SEC_CODE_C := NULL;
  P_PAGE_INDEX := 0;
  P_PAGE_SIZE := 25;

  CAE_FOF_SECURITY_PKG.GET_LIST_FOF_SECURITY(
    P_CAE_SEC_ID_N => P_CAE_SEC_ID_N,
    P_FM_SEC_CODE_C => P_FM_SEC_CODE_C,
    P_PAGE_INDEX => P_PAGE_INDEX,
    P_PAGE_SIZE => P_PAGE_SIZE,
    P_FOF_SEC_REFCUR => v_Return
  );
  -- Modify the code to output the variable
  -- DBMS_OUTPUT.PUT_LINE('P_FOF_SEC_REFCUR = ');
  loop
    fetch v_Return into v_rec;
    exit when v_Return%notfound;
    DBMS_OUTPUT.PUT_LINE('sec_id = ' || v_rec.CAE_SEC_ID || 'sec code = ' ||v_rec.FM_SEC_CODE);
  end loop;

END;

Also making the values categorical and calling summary() would work.

> v = rep(as.factor(c(1,2, 2, 2)), 25)
> summary(v)
 1  2 
25 75 

Your dispatcher servlet does not where to dispatch the request. Issue is your controller bean is not created/working.

Even I faced the same problem. Then added the following under mvc-config.xml

<mvc:annotation-driven/>
<context:component-scan base-package="com.nsv.jsmbaba.teamapp.controller"/>

<bean class="org.springframework.web.servlet.view.InternalResourceViewResolver">
    <property name="prefix"><value>/WEB-INF/view/</value></property>
    <property name="suffix"><value>.jsp</value></property>
</bean>

Hope this helps

I think you might be able to use the ExpandProperty parameter of Select-Object.

For example, to get the list of the current directory and just have the Name property displayed, one would do the following:

ls | select -Property Name

This is still returning DirectoryInfo or FileInfo objects. You can always inspect the type coming through the pipeline by piping to Get-Member (alias gm).

ls | select -Property Name | gm

So, to expand the object to be that of the type of property you're looking at, you can do the following:

ls | select -ExpandProperty Name

In your case, you can just do the following to have a variable be an array of strings, where the strings are the Name property:

$objects = ls | select -ExpandProperty Name

I used the following code to create a temporary file for writing bytes. And its working fine.

File file = new File(Environment.getExternalStorageDirectory() + "/" + File.separator + "test.txt");
file.createNewFile();
byte[] data1={1,1,0,0};
//write the bytes in file
if(file.exists())
{
     OutputStream fo = new FileOutputStream(file);              
     fo.write(data1);
     fo.close();
     System.out.println("file created: "+file);
}               

//deleting the file             
file.delete();
System.out.println("file deleted");

Getting java.nio.file.AccessDeniedException when trying to write to a folder

Unobviously, Comodo antivirus has an "Auto-Containment" setting that can cause this exact error as well. (e.g. the user can write to a location, but the java.exe and javaw.exe processes cannot).

In this edge-case scenario, adding an exception for the process and/or folder should help.

Temporarily disabling the antivirus feature will help understand if Comodo AV is the culprit.

I post this not because I use or prefer Comodo, but because it's a tremendously unobvious symptom to an otherwise functioning Java application and can cost many hours of troubleshooting file permissions that are sane and correct, but being blocked by a 3rd-party application.

I agree that a zip file is a neater solution... But if you have to push multiple file, here's the solution I came up with. It works in IE 9 and up (possibly lower version too - I haven't tested it), Firefox, Safari and Chrome. Chrome will display a message to user to obtain his agreement to download multiple files the first time your site use it.

function deleteIframe (iframe) {
    iframe.remove(); 
}
function createIFrame (fileURL) {
    var iframe = $('<iframe style="display:none"></iframe>');
    iframe[0].src= fileURL;
    $('body').append(iframe);
    timeout(deleteIframe, 60000, iframe);             
 }
 // This function allows to pass parameters to the function in a timeout that are 
 // frozen and that works in IE9
 function timeout(func, time) {
      var args = [];
      if (arguments.length >2) {
           args = Array.prototype.slice.call(arguments, 2);
      }
      return setTimeout(function(){ return func.apply(null, args); }, time);
 }
 // IE will process only the first one if we put no delay
 var wait = (isIE ? 1000 : 0);
 for (var i = 0; i < files.length; i++) {  
      timeout(createIFrame, wait*i, files[i]);
 }

The only side effect of this technique, is that user will see a delay between submit and the download dialog showing. To minimize this effect, I suggest you use the technique describe here and on this question Detect when browser receives file download that consist of setting a cookie with your file to know it has started download. You will have to check for this cookie on client side and to send it on server side. Don't forget to set the proper path for your cookie or you might not see it. You will also have to adapt the solution for multiple file download.

Order Collection Limit :

$orderCollection = Mage::getResourceModel('sales/order_collection'); 
$orderCollection->getSelect()->limit(10);

foreach ($orderCollection->getItems() as $order) :
   $orderModel = Mage::getModel('sales/order');
   $order =   $orderModel->load($order['entity_id']);
   echo $order->getId().'<br>'; 
endforeach; 

You can do something like this in you getJsonResponse function -

jData, err := json.Marshal(Data)
if err != nil {
    // handle error
}
w.Header().Set("Content-Type", "application/json")
w.Write(jData)

You can use a for loop:

for file in * ; do
   echo "$file"
done

Note that if the command in question accepts multiple arguments, then using xargs is almost always more efficient as it only has to spawn the utility in question once instead of multiple times.

The files selected are stored in an array: [input].files

For example, you can access the items

// assuming there is a file input with the ID `my-input`...
var files = document.getElementById("my-input").files;

for (var i = 0; i < files.length; i++)
{
 alert(files[i].name);
}

For jQuery-comfortable people, it's similarly easy

// assuming there is a file input with the ID `my-input`...
var files = $("#my-input")[0].files;

for (var i = 0; i < files.length; i++)
{
 alert(files[i].name);
}

Instead of a functor as Mehrdad proposed, you could use a Lambda function.

sort(numbers.begin(), numbers.end(), [](const int a, const int b) {return a > b; });

You need to set the text to something other than an empty string. In addition, the .html() function should do it while preserving the HTML structure of the div.

$('#one').html($('#one').html().replace('text','replace'));

The secret event that keeps angular ngModel synchronous is the event call input. Hence the best answer to your question should be:

<input type="text" [(ngModel)]="mymodel" (input)="valuechange($event)" />
{{mymodel}}

Real, User and Sys process time statistics

One of these things is not like the other. Real refers to actual elapsed time; User and Sys refer to CPU time used only by the process.

• Real is wall clock time - time from start to finish of the call. This is all elapsed time including time slices used by other processes and time the process spends blocked (for example if it is waiting for I/O to complete).

• User is the amount of CPU time spent in user-mode code (outside the kernel) within the process. This is only actual CPU time used in executing the process. Other processes and time the process spends blocked do not count towards this figure.

• Sys is the amount of CPU time spent in the kernel within the process. This means executing CPU time spent in system calls within the kernel, as opposed to library code, which is still running in user-space. Like 'user', this is only CPU time used by the process. See below for a brief description of kernel mode (also known as 'supervisor' mode) and the system call mechanism.

User+Sys will tell you how much actual CPU time your process used. Note that this is across all CPUs, so if the process has multiple threads (and this process is running on a computer with more than one processor) it could potentially exceed the wall clock time reported by Real (which usually occurs). Note that in the output these figures include the User and Sys time of all child processes (and their descendants) as well when they could have been collected, e.g. by wait(2) or waitpid(2), although the underlying system calls return the statistics for the process and its children separately.

Origins of the statistics reported by time (1)

The statistics reported by time are gathered from various system calls. 'User' and 'Sys' come from wait (2) (POSIX) or times (2) (POSIX), depending on the particular system. 'Real' is calculated from a start and end time gathered from the gettimeofday (2) call. Depending on the version of the system, various other statistics such as the number of context switches may also be gathered by time.

On a multi-processor machine, a multi-threaded process or a process forking children could have an elapsed time smaller than the total CPU time - as different threads or processes may run in parallel. Also, the time statistics reported come from different origins, so times recorded for very short running tasks may be subject to rounding errors, as the example given by the original poster shows.

A brief primer on Kernel vs. User mode

On Unix, or any protected-memory operating system, 'Kernel' or 'Supervisor' mode refers to a privileged mode that the CPU can operate in. Certain privileged actions that could affect security or stability can only be done when the CPU is operating in this mode; these actions are not available to application code. An example of such an action might be manipulation of the MMU to gain access to the address space of another process. Normally, user-mode code cannot do this (with good reason), although it can request shared memory from the kernel, which could be read or written by more than one process. In this case, the shared memory is explicitly requested from the kernel through a secure mechanism and both processes have to explicitly attach to it in order to use it.

The privileged mode is usually referred to as 'kernel' mode because the kernel is executed by the CPU running in this mode. In order to switch to kernel mode you have to issue a specific instruction (often called a trap) that switches the CPU to running in kernel mode and runs code from a specific location held in a jump table. For security reasons, you cannot switch to kernel mode and execute arbitrary code - the traps are managed through a table of addresses that cannot be written to unless the CPU is running in supervisor mode. You trap with an explicit trap number and the address is looked up in the jump table; the kernel has a finite number of controlled entry points.

The 'system' calls in the C library (particularly those described in Section 2 of the man pages) have a user-mode component, which is what you actually call from your C program. Behind the scenes, they may issue one or more system calls to the kernel to do specific services such as I/O, but they still also have code running in user-mode. It is also quite possible to directly issue a trap to kernel mode from any user space code if desired, although you may need to write a snippet of assembly language to set up the registers correctly for the call.

More about 'sys'

There are things that your code cannot do from user mode - things like allocating memory or accessing hardware (HDD, network, etc.). These are under the supervision of the kernel, and it alone can do them. Some operations like malloc orfread/fwrite will invoke these kernel functions and that then will count as 'sys' time. Unfortunately it's not as simple as "every call to malloc will be counted in 'sys' time". The call to malloc will do some processing of its own (still counted in 'user' time) and then somewhere along the way it may call the function in kernel (counted in 'sys' time). After returning from the kernel call, there will be some more time in 'user' and then malloc will return to your code. As for when the switch happens, and how much of it is spent in kernel mode... you cannot say. It depends on the implementation of the library. Also, other seemingly innocent functions might also use malloc and the like in the background, which will again have some time in 'sys' then.

Just for the record, it's possible to configure Maven's Simple SLF4J logging directly in the command line to output only what we need by configuring:

• org.slf4j.simpleLogger.defaultLogLevel=WARN and
• org.slf4j.simpleLogger.log.org.apache.maven.plugins.help=INFO

as documented at http://www.slf4j.org/api/org/slf4j/impl/SimpleLogger.html

MAVEN_OPTS="\
-Dorg.slf4j.simpleLogger.defaultLogLevel=WARN \
-Dorg.slf4j.simpleLogger.log.org.apache.maven.plugins.help=INFO" \
mvn help:evaluate -o -Dexpression=project.version

As a result, one can run simply tail -1 and get:

$ MAVEN_OPTS="\
-Dorg.slf4j.simpleLogger.defaultLogLevel=WARN \
-Dorg.slf4j.simpleLogger.log.org.apache.maven.plugins.help=INFO" \
mvn help:evaluate -o -Dexpression=project.version | tail -1

1.0.0-SNAPSHOT

Note that this is a one-liner. MAVEN_OPTS are being rewritten only for this particular mvn execution.

I had the same issue when upgrading from Tomcat 7 to 8: a continuous large flood of log warnings about cache.

1. Short Answer

Add this within the Context xml element of your $CATALINA_BASE/conf/context.xml:

<!-- The default value is 10240 kbytes, even when not added to context.xml.
So increase it high enough, until the problem disappears, for example set it to 
a value 5 times as high: 51200. -->
<Resources cacheMaxSize="51200" />

So the default is 10240 (10 mbyte), so set a size higher than this. Than tune for optimum settings where the warnings disappear. Note that the warnings may come back under higher traffic situations.

1.1 The cause (short explanation)

The problem is caused by Tomcat being unable to reach its target cache size due to cache entries that are less than the TTL of those entries. So Tomcat didn't have enough cache entries that it could expire, because they were too fresh, so it couldn't free enough cache and thus outputs warnings.

The problem didn't appear in Tomcat 7 because Tomcat 7 simply didn't output warnings in this situation. (Causing you and me to use poor cache settings without being notified.)

The problem appears when receiving a relative large amount of HTTP requests for resources (usually static) in a relative short time period compared to the size and TTL of the cache. If the cache is reaching its maximum (10mb by default) with more than 95% of its size with fresh cache entries (fresh means less than less than 5 seconds in cache), than you will get a warning message for each webResource that Tomcat tries to load in the cache.

1.2 Optional info

Use JMX if you need to tune cacheMaxSize on a running server without rebooting it.

The quickest fix would be to completely disable cache: <Resources cachingAllowed="false" />, but that's suboptimal, so increase cacheMaxSize as I just described.

2. Long Answer

2.1 Background information

A WebSource is a file or directory in a web application. For performance reasons, Tomcat can cache WebSources. The maximum of the static resource cache (all resources in total) is by default 10240 kbyte (10 mbyte). A webResource is loaded into the cache when the webResource is requested (for example when loading a static image), it's then called a cache entry. Every cache entry has a TTL (time to live), which is the time that the cache entry is allowed to stay in the cache. When the TTL expires, the cache entry is eligible to be removed from the cache. The default value of the cacheTTL is 5000 milliseconds (5 seconds).

There is more to tell about caching, but that is irrelevant for the problem.

2.2 The cause

The following code from the Cache class shows the caching policy in detail:

152  // Content will not be cached but we still need metadata size
153  long delta = cacheEntry.getSize();
154  size.addAndGet(delta);
156  if (size.get() > maxSize) {
157      // Process resources unordered for speed. Trades cache
158      // efficiency (younger entries may be evicted before older
159      // ones) for speed since this is on the critical path for
160      // request processing
161      long targetSize =
162              maxSize * (100 - TARGET_FREE_PERCENT_GET) / 100;
163      long newSize = evict(
164              targetSize, resourceCache.values().iterator());
165      if (newSize > maxSize) {
166          // Unable to create sufficient space for this resource
167          // Remove it from the cache
168          removeCacheEntry(path);
169          log.warn(sm.getString("cache.addFail", path));
170      }
171  }

When loading a webResource, the code calculates the new size of the cache. If the calculated size is larger than the default maximum size, than one or more cached entries have to be removed, otherwise the new size will exceed the maximum. So the code will calculate a "targetSize", which is the size the cache wants to stay under (as an optimum), which is by default 95% of the maximum. In order to reach this targetSize, entries have to be removed/evicted from the cache. This is done using the following code:

215  private long evict(long targetSize, Iterator<CachedResource> iter) {
217      long now = System.currentTimeMillis();
219      long newSize = size.get();
221      while (newSize > targetSize && iter.hasNext()) {
222          CachedResource resource = iter.next();
224          // Don't expire anything that has been checked within the TTL
225          if (resource.getNextCheck() > now) {
226              continue;
227          }
229          // Remove the entry from the cache
230          removeCacheEntry(resource.getWebappPath());
232          newSize = size.get();
233      }
235      return newSize;
236  }

So a cache entry is removed when its TTL is expired and the targetSize hasn't been reached yet.

After the attempt to free cache by evicting cache entries, the code will do:

165  if (newSize > maxSize) {
166      // Unable to create sufficient space for this resource
167      // Remove it from the cache
168      removeCacheEntry(path);
169      log.warn(sm.getString("cache.addFail", path));
170  }

So if after the attempt to free cache, the size still exceeds the maximum, it will show the warning message about being unable to free:

cache.addFail=Unable to add the resource at [{0}] to the cache for web application [{1}] because there was insufficient free space available after evicting expired cache entries - consider increasing the maximum size of the cache

2.3 The problem

So as the warning message says, the problem is

insufficient free space available after evicting expired cache entries - consider increasing the maximum size of the cache

If your web application loads a lot of uncached webResources (about maximum of cache, by default 10mb) within a short time (5 seconds), then you'll get the warning.

The confusing part is that Tomcat 7 didn't show the warning. This is simply caused by this Tomcat 7 code:

1606  // Add new entry to cache
1607  synchronized (cache) {
1608      // Check cache size, and remove elements if too big
1609      if ((cache.lookup(name) == null) && cache.allocate(entry.size)) {
1610          cache.load(entry);
1611      }
1612  }

combined with:

231  while (toFree > 0) {
232      if (attempts == maxAllocateIterations) {
233          // Give up, no changes are made to the current cache
234          return false;
235      }

So Tomcat 7 simply doesn't output any warning at all when it's unable to free cache, whereas Tomcat 8 will output a warning.

So if you are using Tomcat 8 with the same default caching configuration as Tomcat 7, and you got warnings in Tomcat 8, than your (and mine) caching settings of Tomcat 7 were performing poorly without warning.

2.4 Solutions

There are multiple solutions:

1. Increase cache (recommended)
2. Lower the TTL (not recommended)
3. Suppress cache log warnings (not recommended)
4. Disable cache

2.4.1. Increase cache (recommended)

As described here: http://tomcat.apache.org/tomcat-8.0-doc/config/resources.html

By adding <Resources cacheMaxSize="XXXXX" /> within the Context element in $CATALINA_BASE/conf/context.xml, where "XXXXX" stands for an increased cache size, specified in kbytes. The default is 10240 (10 mbyte), so set a size higher than this.

You'll have to tune for optimum settings. Note that the problem may come back when you suddenly have an increase in traffic/resource requests.

To avoid having to restart the server every time you want to try a new cache size, you can change it without restarting by using JMX.

To enable JMX, add this to $CATALINA_BASE/conf/server.xml within the Server element: <Listener className="org.apache.catalina.mbeans.JmxRemoteLifecycleListener" rmiRegistryPortPlatform="6767" rmiServerPortPlatform="6768" /> and download catalina-jmx-remote.jar from https://tomcat.apache.org/download-80.cgi and put it in $CATALINA_HOME/lib. Then use jConsole (shipped by default with the Java JDK) to connect over JMX to the server and look through the settings for settings to increase the cache size while the server is running. Changes in these settings should take affect immediately.

2.4.2. Lower the TTL (not recommended)

Lower the cacheTtl value by something lower than 5000 milliseconds and tune for optimal settings.

For example: <Resources cacheTtl="2000" />

This comes effectively down to having and filling a cache in ram without using it.

2.4.3. Suppress cache log warnings (not recommended)

Configure logging to disable the logger for org.apache.catalina.webresources.Cache.

For more info about logging in Tomcat: http://tomcat.apache.org/tomcat-8.0-doc/logging.html

2.4.4. Disable cache

You can disable the cache by setting cachingAllowed to false. <Resources cachingAllowed="false" />

Although I can remember that in a beta version of Tomcat 8, I was using JMX to disable the cache. (Not sure why exactly, but there may be a problem with disabling the cache via server.xml.)

How about DateUtil.daysBetween(). It's Java and it returns a number (difference in days).

You can combine selectors like this

$(".alert-box.warn, .alert-box.dead");

Or if you want a wildcard use the attribute-contains selector

$("[class*='alert-box']");

Note: Preferably you would know the element type or tag when using the selectors above. Knowing the tag can make the selector more efficient.

$("div.alert-box.warn, div.alert-box.dead");
$("div[class*='alert-box']");

You didn't mention what type std::vector<...> myVec is, but if it's a simple type or struct/class that doesn't include pointers, and you want the best efficiency, then you can do a direct memory copy (which I think will be faster than the other answers provided). Here is a general example for std::vector<type> myVec where type in this case is int:

typedef int type; //choose your custom type/struct/class
int iFirst = 100000; //first index to copy
int iLast = 101000; //last index + 1
int iLen = iLast - iFirst;
std::vector<type> newVec;
newVec.resize(iLen); //pre-allocate the space needed to write the data directly
memcpy(&newVec[0], &myVec[iFirst], iLen*sizeof(type)); //write directly to destination buffer from source buffer

If you want to do this purely in CSS then you'd have a class that you'd assign to each alternate list item. E.g.

<ul>
    <li class="alternate"><a href="link">Link 1</a></li>
    <li><a href="link">Link 2</a></li>
    <li class="alternate"><a href="link">Link 3</a></li>
    <li><a href="link">Link 4</a></li>
    <li class="alternate"><a href="link">Link 5</a></li>
</ul>

If your list is dynamically generated, this task would be much easier.

If you don't want to have to manually update this content each time, you could use the jQuery library and apply a style alternately to each <li> item in your list:

<ul id="myList">
    <li><a href="link">Link 1</a></li>
    <li><a href="link">Link 2</a></li>
    <li><a href="link">Link 3</a></li>
    <li><a href="link">Link 4</a></li>
    <li><a href="link">Link 5</a></li>
</ul>

And your jQuery code:

$(document).ready(function(){
  $('#myList li:nth-child(odd)').addClass('alternate');
});

Format(Now(), "yyyy-MM-dd hh:mm:ss")

You could use Trigger.io's UI module to use the native Android date / time picker with a regular HTML5 input. Doing that does require using the overall framework though (so won't work as a regular mobile web page).

You can see before and after screenshots in this blog post: date time picker

In C, a string, as you know, is a character pointer (char *). If you want to swap two strings, you're swapping two char pointers, i.e. just two addresses. In order to do any swap in a function, you need to give it the addresses of the two things you're swapping. So in the case of swapping two pointers, you need a pointer to a pointer. Much like to swap an int, you just need a pointer to an int.

The reason your last code snippet doesn't work is because you're expecting it to swap two char pointers -- it's actually written to swap two characters!

Edit: In your example above, you're trying to swap two int pointers incorrectly, as R. Martinho Fernandes points out. That will swap the two ints, if you had:

int a, b;
intSwap(&a, &b);

I'm not a C programmer so I can't give you information on the uses of static in a C program properly, but when it comes to Object Oriented programming static basically declares a variable, or a function or a class to be the same throughout the life of the program. Take for example.

class A
{
public:
    A();
    ~A();
    void somePublicMethod();
private:
    void somePrivateMethod();
};

When you instantiate this class in your Main you do something like this.

int main()
{
   A a1;
   //do something on a1
   A a2;
   //do something on a2
}

These two class instances are completely different from each other and operate independently from one another. But if you were to recreate the class A like this.

class A
{
public:
    A();
    ~A();
    void somePublicMethod();
    static int x;
private:
    void somePrivateMethod();
};

Lets go back to the main again.

int main()
{
   A a1;
   a1.x = 1;
   //do something on a1
   A a2;
   a2.x++;
   //do something on a2
}

Then a1 and a2 would share the same copy of int x whereby any operations on x in a1 would directly influence the operations of x in a2. So if I was to do this

int main()
{
   A a1;
   a1.x = 1;
   //do something on a1
   cout << a1.x << endl; //this would be 1
   A a2;
   a2.x++;
   cout << a2.x << endl; //this would be 2 
   //do something on a2
}

Both instances of the class A share static variables and functions. Hope this answers your question. My limited knowledge of C allows me to say that defining a function or variable as static means it is only visible to the file that the function or variable is defined as static in. But this would be better answered by a C guy and not me. C++ allows both C and C++ ways of declaring your variables as static because its completely backwards compatible with C.

You may need to install Hotfix KB980368.

This article describes a update that enables certain Internet Information Services (IIS) 7.0 or IIS 7.5 handlers to handle requests whose URLs do not end with a period. Specifically, these handlers are mapped to "." request paths. Currently, a handler that is mapped to a "." request path handles only requests whose URLs end with a period. For example, the handler handles only requests whose URLs resemble the following URL:

http://www.example.com/ExampleSite/ExampleFile.

After you apply this update, handlers that are mapped to a "*." request path can handle requests whose URLs end with a period and requests whose URLs do not end with a period. For example, the handler can now handle requests that resemble the following URLs:

http://www.example.com/ExampleSite/ExampleFile

http://www.example.com/ExampleSite/ExampleFile.

After this patch is applied, ASP.NET 4 applications can handle requests for extensionless URLs. Therefore, managed HttpModules that run prior to handler execution will run. In some cases, the HttpModules can return errors for extensionless URLs. For example, an HttpModule that was written to expect only .aspx requests may now return errors when it tries to access the HttpContext.Session property.

As offer_date is an number, and is of lower accuracy than your real dates, this may work...
- Convert your real date to a string of format YYYYMM
- Conver that value to an INT
- Compare the result you your offer_date

SELECT
  *
FROM
  offers
WHERE
    offer_date = (SELECT CAST(to_char(create_date, 'YYYYMM') AS INT) FROM customers where id = '12345678')
AND offer_rate > 0 

Also, by doing all the manipulation on the create_date you only do the processing on one value.

Additionally, had you manipulated the offer_date you would not be able to utilise any index on that field, and so force SCANs instead of SEEKs.

I had the same issue after a reboot, I followed the guide described here

So try the following:

$ sudo -i
# swapoff -a
# exit
$ strace -eopenat kubectl version

After that it works fine.

for(int a=0;a<myArray.size();a++){
        for(int b=a+1;b<myArray.size();b++){
            if(myArray.get(a).equalsIgnoreCase(myArray.get(b))){
                myArray.remove(b); 
                dups++;
                b--;
            }
        }
}

Heads up, you might need to restart LAMP, Apache or whatever your using to make this take affect. Racked our brains for a while on this one, seemed to make no affect until services were restarted, presumably because the website was caching.

You have your print() statement in the for() loop, It should be after so that it only prints once. the way it currently is, every time the max changes it prints a max.

Realise this is an old thread, but after seeing lots of scripts like this I noticed that you can do this just using conditional formatting.

Assuming the "Status" was Column D:

Highlight cells > right click > conditional formatting. Select "Custom Formula Is" and set the formula as

=RegExMatch($D2,"Complete")

or

=OR(RegExMatch($D2,"Complete"),RegExMatch($D2,"complete"))

Edit (thanks to Frederik Schøning)

=RegExMatch($D2,"(?i)Complete") then set the range to cover all the rows e.g. A2:Z10. This is case insensitive, so will match complete, Complete or CoMpLeTe.

You could then add other rules for "Not Started" etc. The $ is very important. It denotes an absolute reference. Without it cell A2 would look at D2, but B2 would look at E2, so you'd get inconsistent formatting on any given row.

In my case:

Project properties ? Project Facets. Make sure "Dynamic Web Module" is checked. Finally, I enter the version number "2.3" instead of "3.0". After that, the Apache Tomcat 5.5 runtime is listed in the "Runtimes" tab.

Inline elements can't be transformed, and pseudo elements are inline by default, so you must apply display: block or display: inline-block to transform them:

#whatever:after {
  content: "\24B6";
  display: inline-block;
  transform: rotate(30deg);
}

<div id="whatever">Some text </div>

Compiled from above answers, worked for me, did following steps to catch it.

Step #1 - Add following snippet to config file

<configuration>
   <runtime>
      <legacyCorruptedStateExceptionsPolicy enabled="true" />
   </runtime>
</configuration>

Step #2

Add -

[HandleProcessCorruptedStateExceptions]

[SecurityCritical]

on the top of function you are tying catch the exception

source: http://www.gisremotesensing.com/2017/03/catch-exception-attempted-to-read-or.html

Jano's answer is the easiest way to find it.. another way would be if you click on the scheme drop down bar -> edit scheme -> arguments tab and then add NSZombieEnabled in the Environment Variables column and YES in the value column...

The error code indicates the function definition cannot be found. Make sure you're calling the function from the same workspace as the divrat.m file is stored. And make sure divrat function is not a subfunction, it should be first function declaration in the file. You can also try to call the function from the same divrat.m file in order to see if the problem is with workspace selection or the function.

By the way, why didn't you simply say

s = sqrt(diag(C));

Wouldn't it be the same?

You will need to Build and Archive your project. You may need to check what code signing settings you have in the project and executable.

Use the Organiser to select your archive version and then you can Share that version of your project. You will need to select the correct code signing again. It will allow you to save the .ipa file where you want.

Drag and drop the .ipa file into iTunes and then sync with your iPhone.

EDIT: Here are some more detailed instructions including screenshots;

How to distribute ad hoc build with XCode 4

If you would like to modify file names in an editor (such as vim), the click library comes with the command click.edit(), which can be used to receive user input from an editor. Here is an example of how it can be used to refactor files in a directory.

import click
from pathlib import Path

# current directory
direc_to_refactor = Path(".")

# list of old file paths
old_paths = list(direc_to_refactor.iterdir())

# list of old file names
old_names = [str(p.name) for p in old_paths]

# modify old file names in an editor,
# and store them in a list of new file names
new_names = click.edit("\n".join(old_names)).split("\n")

# refactor the old file names
for i in range(len(old_paths)):
    old_paths[i].replace(direc_to_refactor / new_names[i])

I wrote a command line application that uses the same technique, but that reduces the volatility of this script, and comes with more options, such as recursive refactoring. Here is the link to the github page. This is useful if you like command line applications, and are interested in making some quick edits to file names. (My application is similar to the "bulkrename" command found in ranger).

I had an applicationhost.config inside my project folder. It seems IISExpress uses this folder, even though it displays a different file in my c:\users folder

.vs\config\applicationhost.config

You can't have a link to SCSS File in your HTML page.You have to compile it down to CSS First. No there are lots of video tutorials you might want to check out. Lynda provides great video tutorials on SASS. there are also free screencasts you can google...

For official documentation visit this site http://sass-lang.com/documentation/file.SASS_REFERENCE.html And why have you chosen notepad to write Sass?? you can easily download some free text editors for better code handling.

You are plotting all your images on one axis. What you want ist to get a handle for each axis individually and plot your images there. Like so:

fig = plt.figure()
ax1 = fig.add_subplot(2,2,1)
ax1.imshow(...)
ax2 = fig.add_subplot(2,2,2)
ax2.imshow(...)
ax3 = fig.add_subplot(2,2,3)
ax3.imshow(...)
ax4 = fig.add_subplot(2,2,4)
ax4.imshow(...)

For more info have a look here: http://matplotlib.org/examples/pylab_examples/subplots_demo.html

For complex layouts, you should consider using gridspec: http://matplotlib.org/users/gridspec.html

Figured it out by testing all the stuff by myself. Couldn't find any topics about it tho, so I'll just leave the solution here. This might not be the only or even the best solution, but it works for my purposes (within getch's limits) and is better than nothing.

Note: proper keyDown() which would recognize all the keys and actual key presses, is still valued.

Solution: using ord()-function to first turn the getch() into an integer (I guess they're virtual key codes, but not too sure) works fine, and then comparing the result to the actual number representing the wanted key. Also, if I needed to, I could add an extra chr() around the number returned so that it would convert it to a character. However, I'm using mostly down arrow, esc, etc. so converting those to a character would be stupid. Here's the final code:

from msvcrt import getch
while True:
    key = ord(getch())
    if key == 27: #ESC
        break
    elif key == 13: #Enter
        select()
    elif key == 224: #Special keys (arrows, f keys, ins, del, etc.)
        key = ord(getch())
        if key == 80: #Down arrow
            moveDown()
        elif key == 72: #Up arrow
            moveUp()

Also if someone else needs to, you can easily find out the keycodes from google, or by using python and just pressing the key:

from msvcrt import getch
while True:
    print(ord(getch()))

If you are here because of Wordpress giving 404's when loading the environment, this should fix the problem:

define('WP_USE_THEMES', false);
require('../wp-blog-header.php');
status_header( 200 );
//$wp_query->is_404=false; // if necessary

The problem is due to it sending a Status: 404 Not Found header. You have to override that. This will also work:

define('WP_USE_THEMES', false);
require('../wp-blog-header.php');
header("HTTP/1.1 200 OK");
header("Status: 200 All rosy");

EDIT: Only after posting this, did I realize that @350D's answer does the exact same thing.

Surprisingly, this can be done in just 4 lines of code:

const canvas = document.getElementById("canvas"),
  preview = document.getElementById("preview"),
  ctx = canvas.getContext("2d");

canvas.width = 1;
canvas.height = 1;

preview.width = 400;
preview.height = 400;

function getDominantColor(imageObject) {
  //draw the image to one pixel and let the computer find the dominant color
  ctx.drawImage(imageObject, 0, 0, 1, 1);

  //get pixel color
  const i = ctx.getImageData(0, 0, 1, 1).data;

  console.log(`rgba(${i[0]},${i[1]},${i[2]},${i[3]})`);

  console.log("#" + ((1 << 24) + (i[0] << 16) + (i[1] << 8) + i[2]).toString(16).slice(1));
}



// vvv all of this is to just get the uploaded image vvv
const input = document.getElementById("input");
input.type = "file";
input.accept = "image/*";

input.onchange = event => {
  const file = event.target.files[0];
  const reader = new FileReader();

  reader.onload = readerEvent => {
    const image = new Image();
    image.onload = function() {
      //shows preview of uploaded image
      preview.getContext("2d").drawImage(
        image,
        0,
        0,
        preview.width,
        preview.height,
      );
      getDominantColor(image);
    };
    image.src = readerEvent.target.result;
  };
  reader.readAsDataURL(file, "UTF-8");
};

canvas {
  width: 200px;
  height: 200px;
  outline: 1px solid #000000;
}

<canvas id="preview"></canvas>
<canvas id="canvas"></canvas>
<input id="input" type="file" />

How it works:

Create the canvas context

const context = document.createElement("canvas").getContext("2d");

This will draw the image to only one canvas pixel, making the browser find the dominant color for you.

context.drawImage(imageObject, 0, 0, 1, 1);

After that, just get the image data for the pixel:

const i = context.getImageData(0, 0, 1, 1).data;

Finally, convert to rgba or HEX:

const rgba = `rgba(${i[0]},${i[1]},${i[2]},${i[3]})`;

const HEX = "#" + ((1 << 24) + (i[0] << 16) + (i[1] << 8) + i[2]).toString(16).slice(1);

There is one problem with this method though, and that is that getImageData will sometimes throw errors Unable to get image data from canvas because the canvas has been tainted by cross-origin data., which is the reason you need to upload images in the demo instead of inputting a URL for example.

This method can also be used for pixelating images by increasing the width and height to draw the image.

This works on chrome but may not on other browsers.

Use stmt.setDate(1, new java.sql.Date(cal.getTimeInMillis()))

Replace [True | False (default)]

Effect

1.  Replace the directive element. 

Dependency:

1. When replace: true, the template or templateUrl must be required. 

Naive question: Is it possible to somehow download GLIBC 2.15, put it in any folder (e.g. /tmp/myglibc) and then point to this path ONLY when executing something that needs this specific version of glibc?

Yes, it's possible.

If every input asks the same question, you should use a for loop and an array of inputs:

Scanner dd = new Scanner(System.in);
int[] vars = new int[3];

for(int i = 0; i < vars.length; i++) {
  System.out.println("Enter next var: ");
  vars[i] = dd.nextInt();
}

Or as Chip suggested, you can parse the input from one line:

Scanner in = new Scanner(System.in);
int[] vars = new int[3];

System.out.println("Enter "+vars.length+" vars: ");
for(int i = 0; i < vars.length; i++)
  vars[i] = in.nextInt();

You were on the right track, and what you did works. This is just a nicer and more flexible way of doing things.

Based on @Sid's good answer.

Before concatenating, you can load csv files into an intermediate dictionary which gives access to each data set based on the file name (in the form dict_of_df['filename.csv']). Such a dictionary can help you identify issues with heterogeneous data formats, when column names are not aligned for example.

Import modules and locate file paths:

import os
import glob
import pandas
from collections import OrderedDict
path =r'C:\DRO\DCL_rawdata_files'
filenames = glob.glob(path + "/*.csv")

Note: OrderedDict is not necessary, but it'll keep the order of files which might be useful for analysis.

Load csv files into a dictionary. Then concatenate:

dict_of_df = OrderedDict((f, pandas.read_csv(f)) for f in filenames)
pandas.concat(dict_of_df, sort=True)

Keys are file names f and values are the data frame content of csv files. Instead of using f as a dictionary key, you can also use os.path.basename(f) or other os.path methods to reduce the size of the key in the dictionary to only the smaller part that is relevant.

In addition to already suggested p and puts — well, actually in most cases you do can write logger.info "blah" just as you suggested yourself. It works in console too, not only in server mode.

But if all you want is console debugging, puts and p are much shorter to write, anyway.

In the doReflection() is the overhead because of Class.forName("misc.A") (that would require a class lookup, potentially scanning the class path on the filsystem), rather than the newInstance() called on the class. I am wondering what the stats would look like if the Class.forName("misc.A") is done only once outside the for-loop, it doesn't really have to be done for every invocation of the loop.

If one has to call pd.Series.between(l,r) repeatedly (for different bounds l and r), a lot of work is repeated unnecessarily. In this case, it's beneficial to sort the frame/series once and then use pd.Series.searchsorted(). I measured a speedup of up to 25x, see below.

def between_indices(x, lower, upper, inclusive=True):
    """
    Returns smallest and largest index i for which holds 
    lower <= x[i] <= upper, under the assumption that x is sorted.
    """
    i = x.searchsorted(lower, side="left" if inclusive else "right")
    j = x.searchsorted(upper, side="right" if inclusive else "left")
    return i, j

# Sort x once before repeated calls of between()
x = x.sort_values().reset_index(drop=True)
# x = x.sort_values(ignore_index=True) # for pandas>=1.0
ret1 = between_indices(x, lower=0.1, upper=0.9)
ret2 = between_indices(x, lower=0.2, upper=0.8)
ret3 = ...

Benchmark

Measure repeated evaluations (n_reps=100) of pd.Series.between() as well as the method based on pd.Series.searchsorted(), for different arguments lower and upper. On my MacBook Pro 2015 with Python v3.8.0 and Pandas v1.0.3, the below code results in the following outpu

# pd.Series.searchsorted()
# 5.87 ms ± 321 µs per loop (mean ± std. dev. of 7 runs, 100 loops each)
# pd.Series.between(lower, upper)
# 155 ms ± 6.08 ms per loop (mean ± std. dev. of 7 runs, 10 loops each)
# Logical expressions: (x>=lower) & (x<=upper)
# 153 ms ± 3.52 ms per loop (mean ± std. dev. of 7 runs, 10 loops each)

import numpy as np
import pandas as pd

def between_indices(x, lower, upper, inclusive=True):
    # Assumption: x is sorted.
    i = x.searchsorted(lower, side="left" if inclusive else "right")
    j = x.searchsorted(upper, side="right" if inclusive else "left")
    return i, j

def between_fast(x, lower, upper, inclusive=True):
    """
    Equivalent to pd.Series.between() under the assumption that x is sorted.
    """
    i, j = between_indices(x, lower, upper, inclusive)
    if True:
        return x.iloc[i:j]
    else:
        # Mask creation is slow.
        mask = np.zeros_like(x, dtype=bool)
        mask[i:j] = True
        mask = pd.Series(mask, index=x.index)
        return x[mask]

def between(x, lower, upper, inclusive=True):
    mask = x.between(lower, upper, inclusive=inclusive)
    return x[mask]

def between_expr(x, lower, upper, inclusive=True):
    if inclusive:
        mask = (x>=lower) & (x<=upper)
    else:
        mask = (x>lower) & (x<upper)
    return x[mask]

def benchmark(func, x, lowers, uppers):
    for l,u in zip(lowers, uppers):
        func(x,lower=l,upper=u)

n_samples = 1000
n_reps = 100
x = pd.Series(np.random.randn(n_samples))
# Sort the Series.
# For pandas>=1.0:
# x = x.sort_values(ignore_index=True)
x = x.sort_values().reset_index(drop=True)

# Assert equivalence of different methods.
assert(between_fast(x, 0, 1, True ).equals(between(x, 0, 1, True)))
assert(between_expr(x, 0, 1, True ).equals(between(x, 0, 1, True)))
assert(between_fast(x, 0, 1, False).equals(between(x, 0, 1, False)))
assert(between_expr(x, 0, 1, False).equals(between(x, 0, 1, False)))

# Benchmark repeated evaluations of between().
uppers = np.linspace(0, 3, n_reps)
lowers = -uppers
%timeit benchmark(between_fast, x, lowers, uppers)
%timeit benchmark(between, x, lowers, uppers)
%timeit benchmark(between_expr, x, lowers, uppers)

val() is more like a shortcut for attr('value'). For your usage use text() or html() instead

Don't use while True and break statements. It's bad programming.

Imagine you come to debug someone else's code and you see a while True on line 1 and then have to trawl your way through another 200 lines of code with 15 break statements in it, having to read umpteen lines of code for each one to work out what actually causes it to get to the break. You'd want to kill them...a lot.

The condition that causes a while loop to stop iterating should always be clear from the while loop line of code itself without having to look elsewhere.

Phil has the "correct" solution, as it has a clear end condition right there in the while loop statement itself.

Inspired by Expire cache on require.js data-main we updated our deploy script with the following ant task:

<target name="deployWebsite">
    <untar src="${temp.dir}/website.tar.gz" dest="${website.dir}" compression="gzip" />       
    <!-- fetch latest buildNumber from build agent -->
    <replace file="${website.dir}/js/main.js" token="@Revision@" value="${buildNumber}" />
</target>

Where the beginning of main.js looks like:

require.config({
    baseUrl: '/js',
    urlArgs: 'bust=@Revision@',
    ...
});

I have "experienced" this error on Windows because my name (and hence %HOMEPATH%) contains a non ascii character (é). Either git or cmd.exe or anything else could not cope with this.

protected override bool ProcessCmdKey(ref Message msg, Keys dataKey)
    {
        if (dataKey == Keys.Escape)
        {
            this.Close();
            //this.Visible = false;
            //Plus clear values from form, if Visible false.
        }
        return base.ProcessCmdKey(ref msg, dataKey);
    }

It is possible to reference an intellij 'Path Variable' in an intellij 'Run Configuration'.

In 'Path Variables' create a variable for example ANALYTICS_VERSION.

In a 'Run Configuration' under 'Environment Variables' add for example the following:

ANALYTICS_LOAD_LOCATION=$MAVEN_REPOSITORY$\com\my\company\analytics\$ANALYTICS_VERSION$\bin

To answer the original question you would need to add an APP_HOME environment variable to your run configuration which references the path variable:

APP_HOME=$APP_HOME$

Try This

  Intent intent = new Intent(Intent.ACTION_MAIN);
  intent.setComponent(new ComponentName(appDetails.packageName,"com.android.packageinstaller.permission.ui.ManagePermissionsActivity"));
  startActivity(intent);

Instead of tf.placeholder(shape=[None, 2], dtype=tf.float32) use something like tf.compat.v1.placeholder(shape=[None, 2], dtype=tf.float32) if you don't want to disable v2 completely.

this might be repetitive. I was trying to use pycharm to run flask - had anaconda 3, pycharm 2019.1.1 and windows 10. Created a new conda environment - it threw errors. Followed these steps -

1. Used the cmd to install python and flask after creating environment as suggested above.

2. Followed this answer.

3. As suggested above, went to Run -> Edit Configurations and changed the environment there as well as in (2).

Obviously kept the correct python interpreter (the one in the environment) everywhere.

Normally you'd just stick the code in Page_Load in your .aspx page's code-behind.

if (someVar) {
    Item1.Visible = true;
    Item2.Visible = false;
} else {
    Item1.Visible = false;
    Item2.Visible = true;
}

This assumes you've got Item1 and Item2 laid out on the page already.

Skype is usually the culprit because it uses port 80 by default. Just close it or uncheck "Use port 80 and 443 as alternatives for incoming connections" under tools > options... > advanced > connection and then restart Skype.

This answer to a similar question describes how to extend the properties plugin so it can use a remote descriptor for the properties file. The descriptor is basically a jar artifact containing a properties file (the properties file is included under src/main/resources).

The descriptor is added as a dependency to the extended properties plugin so it is on the plugin's classpath. The plugin will search the classpath for the properties file, read the file''s contents into a Properties instance, and apply those properties to the project's configuration so they can be used elsewhere.

The book is a bit dated with respect to subclass-superclass calling. It's also a little dated with respect to subclassing built-in classes.

It looks like this nowadays:

class FileInfo(dict):
    """store file metadata"""
    def __init__(self, filename=None):
        super(FileInfo, self).__init__()
        self["name"] = filename

Note the following:

1. We can directly subclass built-in classes, like dict, list, tuple, etc.

2. The super function handles tracking down this class's superclasses and calling functions in them appropriately.

Unfortunately most of the other answers do not provide information specific to Windows 10.

Windows 10 has versions of its own: 1507, 1511, 1607, 1703, etc. This is what winver shows.

Powershell:
(Get-ItemProperty "HKLM:\SOFTWARE\Microsoft\Windows NT\CurrentVersion").ReleaseId

Command prompt (CMD.EXE):
Reg Query "HKLM\SOFTWARE\Microsoft\Windows NT\CurrentVersion" /v ReleaseId

See also related question on superuser.

As for other Windows versions use systeminfo. Powershell wrapper:

PS C:\> systeminfo /fo csv | ConvertFrom-Csv | select OS*, System*, Hotfix* | Format-List


OS Name             : Microsoft Windows 7 Enterprise
OS Version          : 6.1.7601 Service Pack 1 Build 7601
OS Manufacturer     : Microsoft Corporation
OS Configuration    : Standalone Workstation
OS Build Type       : Multiprocessor Free
System Type         : x64-based PC
System Locale       : ru;Russian
Hotfix(s)           : 274 Hotfix(s) Installed.,[01]: KB2849697,[02]: KB2849697,[03]:...

Windows 10 output for the same command:

OS Name             : Microsoft Windows 10 Enterprise N 2016 LTSB
OS Version          : 10.0.14393 N/A Build 14393
OS Manufacturer     : Microsoft Corporation
OS Configuration    : Standalone Workstation
OS Build Type       : Multiprocessor Free
System Type         : x64-based PC
System Directory    : C:\Windows\system32
System Locale       : en-us;English (United States)
Hotfix(s)           : N/A

Check out this link (Running react-native run-ios occurs an error?). It appears to be a problem with the location of Command line tools.

In Xcode, select Xcode menu, then Preferences, then Locations tab. Select your Xcode version from the dropdown and exit Xcode.

You should implement a Custom List View, such that you define a Layout once and draw it for every row in the list view.

Sure. On my (pretty vanilla) Cygwin setup, bash is in c:\cygwin\bin so I can run a bash script (say testit.sh) from a Windows batch file using a command like:

C:\cygwin\bin\bash testit.sh

... which can be included in a .bat file as easily as it can be typed at the command line, and with the same effect.

Use below script and change the body of the script accordingly.

       <script>
        //Re-Create for on page postbacks
        var prm = Sys.WebForms.PageRequestManager.getInstance();
        prm.add_endRequest(function () {
           //your codes here!
        });
    </script>

Though Dan Lowe's excellent answer very thoroughly answered the OP's question about the difference between PUT and PATCH, its answer to the question of why PATCH is not idempotent is not quite correct.

To show why PATCH isn't idempotent, it helps to start with the definition of idempotence (from Wikipedia):

The term idempotent is used more comprehensively to describe an operation that will produce the same results if executed once or multiple times [...] An idempotent function is one that has the property f(f(x)) = f(x) for any value x.

In more accessible language, an idempotent PATCH could be defined as: After PATCHing a resource with a patch document, all subsequent PATCH calls to the same resource with the same patch document will not change the resource.

Conversely, a non-idempotent operation is one where f(f(x)) != f(x), which for PATCH could be stated as: After PATCHing a resource with a patch document, subsequent PATCH calls to the same resource with the same patch document do change the resource.

To illustrate a non-idempotent PATCH, suppose there is a /users resource, and suppose that calling GET /users returns a list of users, currently:

[{ "id": 1, "username": "firstuser", "email": "[email protected]" }]

Rather than PATCHing /users/{id}, as in the OP's example, suppose the server allows PATCHing /users. Let's issue this PATCH request:

PATCH /users
[{ "op": "add", "username": "newuser", "email": "[email protected]" }]

Our patch document instructs the server to add a new user called newuser to the list of users. After calling this the first time, GET /users would return:

[{ "id": 1, "username": "firstuser", "email": "[email protected]" },
 { "id": 2, "username": "newuser", "email": "[email protected]" }]

Now, if we issue the exact same PATCH request as above, what happens? (For the sake of this example, let's assume that the /users resource allows duplicate usernames.) The "op" is "add", so a new user is added to the list, and a subsequent GET /users returns:

[{ "id": 1, "username": "firstuser", "email": "[email protected]" },
 { "id": 2, "username": "newuser", "email": "[email protected]" },
 { "id": 3, "username": "newuser", "email": "[email protected]" }]

The /users resource has changed again, even though we issued the exact same PATCH against the exact same endpoint. If our PATCH is f(x), f(f(x)) is not the same as f(x), and therefore, this particular PATCH is not idempotent.

Although PATCH isn't guaranteed to be idempotent, there's nothing in the PATCH specification to prevent you from making all PATCH operations on your particular server idempotent. RFC 5789 even anticipates advantages from idempotent PATCH requests:

A PATCH request can be issued in such a way as to be idempotent, which also helps prevent bad outcomes from collisions between two PATCH requests on the same resource in a similar time frame.

In Dan's example, his PATCH operation is, in fact, idempotent. In that example, the /users/1 entity changed between our PATCH requests, but not because of our PATCH requests; it was actually the Post Office's different patch document that caused the zip code to change. The Post Office's different PATCH is a different operation; if our PATCH is f(x), the Post Office's PATCH is g(x). Idempotence states that f(f(f(x))) = f(x), but makes no guarantes about f(g(f(x))).

I have never done this, but from what I understand you can use the PathShape class: http://developer.android.com/reference/android/graphics/drawable/shapes/PathShape.html

<script language="JavaScript">
function importXML()
{
    if (document.implementation && document.implementation.createDocument)
    {
            xmlDoc = document.implementation.createDocument("", "", null);
            xmlDoc.onload = createTable;
    }
    else if (window.ActiveXObject)
    {
            xmlDoc = new ActiveXObject("Microsoft.XMLDOM");
            xmlDoc.onreadystatechange = function () {
                    if (xmlDoc.readyState == 4) createTable()
            };
    }
    else
    {
            alert('Your browser can\'t handle this script');
            return;
    }
    xmlDoc.load("emperors.xml");
}

function createTable()
{
    var theData="";
    var x = xmlDoc.getElementsByTagName('emperor');
    var newEl = document.createElement('TABLE');
    newEl.setAttribute('cellPadding',3);
    newEl.setAttribute('cellSpacing',0);
    newEl.setAttribute('border',1);
    var tmp = document.createElement('TBODY');
    newEl.appendChild(tmp);
    var row = document.createElement('TR');
    for (j=0;j<x[0].childNodes.length;j++)
    {
            if (x[0].childNodes[j].nodeType != 1) continue;
            var container = document.createElement('TH');
            theData = document.createTextNode(x[0].childNodes[j].nodeName);
            container.appendChild(theData);
            row.appendChild(container);
    }
    tmp.appendChild(row);
    for (i=0;i<x.length;i++)
    {
            var row = document.createElement('TR');
            for (j=0;j<x[i].childNodes.length;j++)
            {
                    if (x[i].childNodes[j].nodeType != 1) continue;
                    var container = document.createElement('TD');
                    var theData = document.createTextNode(x[i].childNodes[j].firstChild.nodeValue);
                    container.appendChild(theData);
                    row.appendChild(container);
            }
            tmp.appendChild(row);
    }
    document.getElementById('writeroot').appendChild(newEl);
}
</script>
</HEAD>

<BODY onLoad="javascript:importXML();">
<p id=writeroot> </p>
</BODY>

For more info refer this http://www.easycodingclub.com/xml-parser-in-javascript/javascript-tutorials/

As commented in caf's answer above, ps and in some cases pidstat will give you the lifetime average of the pCPU. To get more accurate results use top. If you need to run top once you can run:

top -b -n 1 -p <PID>

or for process only data and header:

top -b -n 1 -p <PID> | tail -3 | head -2

without headers:

top -b -n 1 -p <PID> | tail -2 | head -1

in my problem I want the text of anchor <a>text</a> inside my view to be based on some value and that text is retrieved form App string Resources

so, this @() is the solution

<a href='#'>
      @(Model.ID == 0 ? Resource_en.Back : Resource_en.Department_View_DescartChanges)
</a>

if the text is not from App string Resources use this

@(Model.ID == 0 ? "Back" :"Descart Changes")

Using .multiply() (ufunc multiply)

a_1 = np.array([1.0, 2.0, 3.0])
a_2 = np.array([[1., 2.], [3., 4.]])
b = 2.0 

np.multiply(a_1,b)
# array([2., 4., 6.])
np.multiply(a_2,b)
# array([[2., 4.],[6., 8.]])

Make sure you don't add or uncomment the AllowNoPassword option after the $i++ line.

/* Uncomment the following to enable logging in to passwordless accounts, * after taking note of the associated security risks. */

 $cfg['Servers'][$i]['AllowNoPassword'] = TRUE;

You can use subprocess, but I always felt that it was not a 'Pythonic' way of doing it. So I created Sultan (shameless plug) that makes it easy to run command line functions.

https://github.com/aeroxis/sultan

setInterval returns an ID which you then use to clear the interval.

var intervalId;
on.onclick = function() {
    if (intervalId) {
        clearInterval(intervalId);
    }
    intervalId = setInterval(fontChange, 500);
};

off.onclick = function() {
    clearInterval(intervalId);
}; 

Try this

TextWatcher changeText = new TextWatcher() {
     @Override
            public void onTextChanged(CharSequence s, int start, int before, int count) {
                tv3.setText(et.getText().toString());
                tv3.post(new Runnable() {           
                    @Override
                    public void run() {
                    while(tv3.getLineCount() >= 3){                     
                            tv3.setTextSize((tv3.getTextSize())-1);                     
                        }
                    }
                });
            }

            @Override public void beforeTextChanged(CharSequence s, int start, int count, int after) {}

            @Override public void afterTextChanged(Editable s) { }
        };

If you need pure Java without any dependencies, you can use build in Nashorn API from Java 8. It is deprecated in Java 11.

This is working for me:

...
import javax.script.ScriptEngine;
import javax.script.ScriptEngineManager;
import javax.script.ScriptException;
...

public class JsonUtils {

    public static Map parseJSON(String json) throws ScriptException {
        ScriptEngineManager sem = new ScriptEngineManager();
        ScriptEngine engine = sem.getEngineByName("javascript");

        String script = "Java.asJSONCompatible(" + json + ")";

        Object result = engine.eval(script);

        return (Map) result;
    }
}

Sample usage

JSON:

{
    "data":[
        {"id":1,"username":"bruce"},
        {"id":2,"username":"clark"},
        {"id":3,"username":"diana"}
    ]
}

Code:

...
import jdk.nashorn.internal.runtime.JSONListAdapter;
...

public static List<String> getUsernamesFromJson(Map json) {
    List<String> result = new LinkedList<>();

    JSONListAdapter data = (JSONListAdapter) json.get("data");

    for(Object obj : data) {
        Map map = (Map) obj;
        result.add((String) map.get("username"));
    }

    return result;
}

Upgrading Python to version 3 fixed my problem. Nothing else did.

#include <stdio.h>

int main(void)
{
    printf("%d\n", __WORDSIZE);
    return 0;
}

Like that we can do....

int main(int argc, char *argv[]) {

    int a, b, c;
    *// Converting string type to integer type
    // using function "atoi( argument)"* 

    a = atoi(argv[1]);     
    b = atoi(argv[2]);
    c = atoi(argv[3]);

 }

Maybe it can be useful for someone:

// Printing the SQL with real values
$vals = $query->getFlattenedParams();
foreach(explode('?', $query->getSqlQuery()) as $i => $part) {
    $sql = (isset($sql) ? $sql : null) . $part;
    if (isset($vals[$i])) $sql .= $vals[$i];
}

echo $sql;

tar.gz file is just a tar file that's been gzipped. Both tar and gzip are available for windows.

If you like GUIs (Graphical user interface), 7zip can pack with both tar and gzip.

I saw this somewhere else. I am using 2016 Excel. What worked for me was to use YYYY Quarters (I was looking for quarterly data). So, I had the source data sorted as YYYY xQ. 2016 1Q, 2016 2Q, 2016 3Q, 2016, 4Q, 2017 1Q, 2017 2Q... You get the idea.

if (str.indexOf(' ') >= 0)

would be (slightly) faster.

This feature is included as part of jquery ui http://docs.jquery.com/UI/Effects/Slide if you want to extend it with your own names you can use this.

jQuery.fn.extend({
  slideRightShow: function() {
    return this.each(function() {
        $(this).show('slide', {direction: 'right'}, 1000);
    });
  },
  slideLeftHide: function() {
    return this.each(function() {
      $(this).hide('slide', {direction: 'left'}, 1000);
    });
  },
  slideRightHide: function() {
    return this.each(function() {
      $(this).hide('slide', {direction: 'right'}, 1000);
    });
  },
  slideLeftShow: function() {
    return this.each(function() {
      $(this).show('slide', {direction: 'left'}, 1000);
    });
  }
});

you will need the following references

<script src="http://code.jquery.com/jquery-latest.js"></script>
<script src="http://jquery-ui.googlecode.com/svn/tags/latest/ui/jquery.effects.core.js"></script>
<script src="http://jquery-ui.googlecode.com/svn/tags/latest/ui/jquery.effects.slide.js"></script>

Please remove /bin and even semi colon ; from JAVA_HOME to resolve.

Say you want to remove the second object by it's field property.

With ES6 it's as easy as this.

myArray.splice(myArray.findIndex(item => item.field === "cStatus"), 1)

I found a workaround.

1. Export with a white background (or other color that will work with transparent graphics). This will be out "whitescreen" layer.
2. Export with "bluescreen" background, or some terrible other color that will make it easy to select out the background from foreground.
3. Open the bluescreen version as a layer on top of the white screen.
4. Use the bluescreen layer to select out only the parts you want to use.
5. Create a mask for the whitescreen layer with the selection made from the bluescreen layer.

This will get good results for edges and aliasing, whilst retaining a good color for the see-

In layman terms it's strongly structured vs unstructured, which implies that you have different degrees of adaptability for your DB. Differences arise in indexation particularly as you need to ensure that a certain reference index can link to a another item -> this a relation. The more strict structure of relational DB comes from this requirement.

To note that NosDB apaprently provides both relational and non relational DBs and a way to query both http://www.alachisoft.com/nosdb/sql-cheat-sheet.html

this problem arises when we use vectors in place of scalars for example in a for loop the range should be a scalar, in case you have given a vector in that place you get error. So to avoid the problem use the length of the vector you have used

One important point that people forget to mention is the life time of the Broadcast Receiver. The difference of programmatically registering it from registering in AndroidManifest.xml is that. In the manifest file, it doesn't depend on application life time. While when programmatically registering it it does depend on the application life time. This means that if you register in AndroidManifest.xml, you can catch the broadcasted intents even when your application is not running.

Edit: The mentioned note is no longer true as of Android 3.1, the Android system excludes all receiver from receiving intents by default if the corresponding application has never been started by the user or if the user explicitly stopped the application via the Android menu (in Manage ? Application). https://developer.android.com/about/versions/android-3.1.html

This is an additional security feature as the user can be sure that only the applications he started will receive broadcast intents.

So it can be understood as receivers programmatically registered in Application's onCreate() would have same effect with ones declared in AndroidManifest.xml from Android 3.1 above.

You could do something like this:

Convert.ToDate(month + " 01, 1900").Month

Just return the following:

return Unauthorized();

tl;dr Set the editor to something nicer, like Sublime or Atom

Here nice is used in the meaning of an editor you like or find more user friendly.

The underlying problem is that Git by default uses an editor that is too unintuitive to use for most people: Vim. Now, don't get me wrong, I love Vim, and while you could set some time aside (like a month) to learn Vim and try to understand why some people think Vim is the greatest editor in existence, there is a quicker way of fixing this problem :-)

The fix is not to memorize cryptic commands, like in the accepted answer, but configuring Git to use an editor that you like and understand! It's really as simple as configuring either of these options

1. the git config setting core.editor (per project, or globally)
2. the VISUAL or EDITOR environment variable (this works for other programs as well)

I'll cover the first option for a couple of popular editors, but GitHub has an excellent guide on this for many editors as well.

To use Atom

Straight from its docs, enter this in a terminal: git config --global core.editor "atom --wait"

Git normally wait for the editor command to finish, but since Atom forks to a background process immediately, this won't work, unless you give it the --wait option.

To use Sublime Text

For the same reasons as in the Atom case, you need a special flag to signal to the process that it shouldn't fork to the background:

git config --global core.editor "subl -n -w"

On Linux

(tested in Ununtu KDE)

There is the option in the menu, under Edit > Lines > Auto Indent or press Cmd + Shift + p, search for Editor: Auto Indent by entering just "ai"

Note: In KDE ctrl-alt-l is already globally set for "lock screen" so better use ctrl-alt-i instead.

You can add a key mapping in Atom:

• Cmd + Shift + p, search for "Settings View: Show Keybindings"
• click on "your keymap file"

• Add a section there like this one:

   'atom-text-editor':
      'ctrl-alt-i': 'editor:auto-indent'
  

If the indention is not working, it can be a reason, that the file-ending is not recognized by Atom. Add the support for your language then, for example for "Lua" install the package "language-lua".

If a File is not recognized for your language:

• open the ~/.atom/config.cson file (by CTRL+SHIFT+p: type ``open config'')

• add/edit a customFileTypes section under core for example like the following:

  core:
    customFileTypes:
      "source.lua": [
        "conf"
      ]
      "text.html.php": [
        "thtml"
      ]
  

(You find the languages scope names ("source.lua", "text.html.php"...) in the language package settings see here)

Screen Size Class

-

1. Hidden on all .d-none

2. Hidden only on xs .d-none .d-sm-block

3. Hidden only on sm .d-sm-none .d-md-block

4. Hidden only on md .d-md-none .d-lg-block

5. Hidden only on lg .d-lg-none .d-xl-block

6. Hidden only on xl .d-xl-none

7. Visible on all .d-block

8. Visible only on xs .d-block .d-sm-none

9. Visible only on sm .d-none .d-sm-block .d-md-none

10. Visible only on md .d-none .d-md-block .d-lg-none

11. Visible only on lg .d-none .d-lg-block .d-xl-none

12. Visible only on xl .d-none .d-xl-block

Refer this link http://getbootstrap.com/docs/4.0/utilities/display/#hiding-elements

4.5 link: https://getbootstrap.com/docs/4.5/utilities/display/#hiding-elements

Data for last month-

select count(distinct switch_id)
  from [email protected]
 where dealer_name =  'XXXX'
   and to_char(CREATION_DATE,'MMYYYY') = to_char(add_months(trunc(sysdate),-1),'MMYYYY');

The error is pretty straightforward - the line starting with check_exists_sql isn't indented properly. From the context of your code, I'd indent it and the following lines to match the line before it:

   #open db connection
   db = MySQLdb.connect("localhost","root","str0ng","TESTDB")

   #prepare a cursor object using cursor() method
   cursor = db.cursor()

   #see if any links in the DB match the crawled link
   check_exists_sql = "SELECT * FROM LINKS WHERE link = '%s' LIMIT 1" % item['link']

   cursor.execute(check_exists_sql)

And keep indenting it until the for loop ends (all the way through to and including items.append(item).

You can't just return an array of objects because there's nothing telling React how to render that. You'll need to return an array of components or elements like:

render: function() {
  return (
    <span>
      // This will go through all the elements in arrayFromJson and
      // render each one as a <SomeComponent /> with data from the object
      {this.state.arrayFromJson.map(function(object) {
        return (
          <SomeComponent key={object.id} data={object} />
        );
      })}
    </span>
  );
}

REPEAT
    ...
UNTIL cond

Is equivalent to

while True:
    ...
    if cond:
        break

You should use the CameraUpdate class to do (probably) all programmatic map movements.

To do this, first calculate the bounds of all the markers like so:

LatLngBounds.Builder builder = new LatLngBounds.Builder();
for (Marker marker : markers) {
    builder.include(marker.getPosition());
}
LatLngBounds bounds = builder.build();

Then obtain a movement description object by using the factory: CameraUpdateFactory:

int padding = 0; // offset from edges of the map in pixels
CameraUpdate cu = CameraUpdateFactory.newLatLngBounds(bounds, padding);

Finally move the map:

googleMap.moveCamera(cu);

Or if you want an animation:

googleMap.animateCamera(cu);

That's all :)

Clarification 1

Almost all movement methods require the Map object to have passed the layout process. You can wait for this to happen using the addOnGlobalLayoutListener construct. Details can be found in comments to this answer and remaining answers. You can also find a complete code for setting map extent using addOnGlobalLayoutListener here.

Clarification 2

One comment notes that using this method for only one marker results in map zoom set to a "bizarre" zoom level (which I believe to be maximum zoom level available for given location). I think this is expected because:

1. The LatLngBounds bounds instance will have northeast property equal to southwest, meaning that the portion of area of the earth covered by this bounds is exactly zero. (This is logical since a single marker has no area.)
2. By passing bounds to CameraUpdateFactory.newLatLngBounds you essentially request a calculation of such a zoom level that bounds (having zero area) will cover the whole map view.
3. You can actually perform this calculation on a piece of paper. The theoretical zoom level that is the answer is +∞ (positive infinity). In practice the Map object doesn't support this value so it is clamped to a more reasonable maximum level allowed for given location.

Another way to put it: how can Map object know what zoom level should it choose for a single location? Maybe the optimal value should be 20 (if it represents a specific address). Or maybe 11 (if it represents a town). Or maybe 6 (if it represents a country). API isn't that smart and the decision is up to you.

So, you should simply check if markers has only one location and if so, use one of:

• CameraUpdate cu = CameraUpdateFactory.newLatLng(marker.getPosition()) - go to marker position, leave current zoom level intact.
• CameraUpdate cu = CameraUpdateFactory.newLatLngZoom(marker.getPosition(), 12F) - go to marker position, set zoom level to arbitrarily chosen value 12.

As you can see, there are several ways to achieve this.

I personally use:

Expand all: CTRL + M + L

Collapse all: CTRL + M + O

Bonus:

Expand/Collapse on cursor location: CTRL + M + M

const styles = StyleSheet.create({
        navigationView: {
        height: 44,
        width: '100%',
        backgroundColor:'darkgray',
        justifyContent: 'center', 
        alignItems: 'center' 
    },
    titleText: {
        fontSize: 20,
        fontWeight: 'bold',
        color: 'white',
        textAlign: 'center',
    },
})


render() {
    return (
        <View style = { styles.navigationView }>
            <Text style = { styles.titleText } > Title name here </Text>
        </View>
    )

}

The Original Question

Why is one loop so much slower than two loops?

Conclusion:

Case 1 is a classic interpolation problem that happens to be an inefficient one. I also think that this was one of the leading reasons why many machine architectures and developers ended up building and designing multi-core systems with the ability to do multi-threaded applications as well as parallel programming.

Looking at it from this kind of an approach without involving how the hardware, OS, and compiler(s) work together to do heap allocations that involve working with RAM, cache, page files, etc.; the mathematics that is at the foundation of these algorithms shows us which of these two is the better solution.

We can use an analogy of a Boss being a Summation that will represent a For Loop that has to travel between workers A & B.

We can easily see that Case 2 is at least half as fast if not a little more than Case 1 due to the difference in the distance that is needed to travel and the time taken between the workers. This math lines up almost virtually and perfectly with both the benchmark times as well as the number of differences in assembly instructions.

I will now begin to explain how all of this works below.

Assessing The Problem

The OP's code:

const int n=100000;

for(int j=0;j<n;j++){
    a1[j] += b1[j];
    c1[j] += d1[j];
}

And

for(int j=0;j<n;j++){
    a1[j] += b1[j];
}
for(int j=0;j<n;j++){
    c1[j] += d1[j];
}

The Consideration

Considering the OP's original question about the two variants of the for loops and his amended question towards the behavior of caches along with many of the other excellent answers and useful comments; I'd like to try and do something different here by taking a different approach about this situation and problem.

The Approach

Considering the two loops and all of the discussion about cache and page filing I'd like to take another approach as to looking at this from a different perspective. One that doesn't involve the cache and page files nor the executions to allocate memory, in fact, this approach doesn't even concern the actual hardware or the software at all.

The Perspective

After looking at the code for a while it became quite apparent what the problem is and what is generating it. Let's break this down into an algorithmic problem and look at it from the perspective of using mathematical notations then apply an analogy to the math problems as well as to the algorithms.

What We Do Know

We know is that this loop will run 100,000 times. We also know that a1, b1, c1 & d1 are pointers on a 64-bit architecture. Within C++ on a 32-bit machine, all pointers are 4 bytes and on a 64-bit machine, they are 8 bytes in size since pointers are of a fixed length.

We know that we have 32 bytes in which to allocate for in both cases. The only difference is we are allocating 32 bytes or two sets of 2-8 bytes on each iteration wherein the second case we are allocating 16 bytes for each iteration for both of the independent loops.

Both loops still equal 32 bytes in total allocations. With this information let's now go ahead and show the general math, algorithms, and analogy of these concepts.

We do know the number of times that the same set or group of operations that will have to be performed in both cases. We do know the amount of memory that needs to be allocated in both cases. We can assess that the overall workload of the allocations between both cases will be approximately the same.

What We Don't Know

We do not know how long it will take for each case unless if we set a counter and run a benchmark test. However, the benchmarks were already included from the original question and from some of the answers and comments as well; and we can see a significant difference between the two and this is the whole reasoning for this proposal to this problem.

Let's Investigate

It is already apparent that many have already done this by looking at the heap allocations, benchmark tests, looking at RAM, cache, and page files. Looking at specific data points and specific iteration indices were also included and the various conversations about this specific problem have many people starting to question other related things about it. How do we begin to look at this problem by using mathematical algorithms and applying an analogy to it? We start off by making a couple of assertions! Then we build out our algorithm from there.

Our Assertions:

• We will let our loop and its iterations be a Summation that starts at 1 and ends at 100000 instead of starting with 0 as in the loops for we don't need to worry about the 0 indexing scheme of memory addressing since we are just interested in the algorithm itself.
• In both cases we have four functions to work with and two function calls with two operations being done on each function call. We will set these up as functions and calls to functions as the following: F1(), F2(), f(a), f(b), f(c) and f(d).

The Algorithms:

1st Case: - Only one summation but two independent function calls.

Sum n=1 : [1,100000] = F1(), F2();
                       F1() = { f(a) = f(a) + f(b); }
                       F2() = { f(c) = f(c) + f(d); }

2nd Case: - Two summations but each has its own function call.

Sum1 n=1 : [1,100000] = F1();
                        F1() = { f(a) = f(a) + f(b); }

Sum2 n=1 : [1,100000] = F1();
                        F1() = { f(c) = f(c) + f(d); }

If you noticed F2() only exists in Sum from Case1 where F1() is contained in Sum from Case1 and in both Sum1 and Sum2 from Case2. This will be evident later on when we begin to conclude that there is an optimization that is happening within the second algorithm.

The iterations through the first case Sum calls f(a) that will add to its self f(b) then it calls f(c) that will do the same but add f(d) to itself for each 100000 iterations. In the second case, we have Sum1 and Sum2 that both act the same as if they were the same function being called twice in a row.

In this case we can treat Sum1 and Sum2 as just plain old Sum where Sum in this case looks like this: Sum n=1 : [1,100000] { f(a) = f(a) + f(b); } and now this looks like an optimization where we can just consider it to be the same function.

Summary with Analogy

With what we have seen in the second case it almost appears as if there is optimization since both for loops have the same exact signature, but this isn't the real issue. The issue isn't the work that is being done by f(a), f(b), f(c), and f(d). In both cases and the comparison between the two, it is the difference in the distance that the Summation has to travel in each case that gives you the difference in execution time.

Think of the for loops as being the summations that does the iterations as being a Boss that is giving orders to two people A & B and that their jobs are to meat C & D respectively and to pick up some package from them and return it. In this analogy, the for loops or summation iterations and condition checks themselves don't actually represent the Boss. What actually represents the Boss is not from the actual mathematical algorithms directly but from the actual concept of Scope and Code Block within a routine or subroutine, method, function, translation unit, etc. The first algorithm has one scope where the second algorithm has two consecutive scopes.

Within the first case on each call slip, the Boss goes to A and gives the order and A goes off to fetch B's package then the Boss goes to C and gives the orders to do the same and receive the package from D on each iteration.

Within the second case, the Boss works directly with A to go and fetch B's package until all packages are received. Then the Boss works with C to do the same for getting all of D's packages.

Since we are working with an 8-byte pointer and dealing with heap allocation let's consider the following problem. Let's say that the Boss is 100 feet from A and that A is 500 feet from C. We don't need to worry about how far the Boss is initially from C because of the order of executions. In both cases, the Boss initially travels from A first then to B. This analogy isn't to say that this distance is exact; it is just a useful test case scenario to show the workings of the algorithms.

In many cases when doing heap allocations and working with the cache and page files, these distances between address locations may not vary that much or they can vary significantly depending on the nature of the data types and the array sizes.

The Test Cases:

First Case: On first iteration the Boss has to initially go 100 feet to give the order slip to A and A goes off and does his thing, but then the Boss has to travel 500 feet to C to give him his order slip. Then on the next iteration and every other iteration after the Boss has to go back and forth 500 feet between the two.

Second Case: The Boss has to travel 100 feet on the first iteration to A, but after that, he is already there and just waits for A to get back until all slips are filled. Then the Boss has to travel 500 feet on the first iteration to C because C is 500 feet from A. Since this Boss( Summation, For Loop ) is being called right after working with A he then just waits there as he did with A until all of C's order slips are done.

The Difference In Distances Traveled

const n = 100000
distTraveledOfFirst = (100 + 500) + ((n-1)*(500 + 500);
// Simplify
distTraveledOfFirst = 600 + (99999*100);
distTraveledOfFirst = 600 + 9999900;
distTraveledOfFirst =  10000500;
// Distance Traveled On First Algorithm = 10,000,500ft

distTraveledOfSecond = 100 + 500 = 600;
// Distance Traveled On Second Algorithm = 600ft;

The Comparison of Arbitrary Values

We can easily see that 600 is far less than 10 million. Now, this isn't exact, because we don't know the actual difference in distance between which address of RAM or from which cache or page file each call on each iteration is going to be due to many other unseen variables. This is just an assessment of the situation to be aware of and looking at it from the worst-case scenario.

From these numbers it would almost appear as if algorithm one should be 99% slower than algorithm two; however, this is only the Boss's part or responsibility of the algorithms and it doesn't account for the actual workers A, B, C, & D and what they have to do on each and every iteration of the Loop. So the boss's job only accounts for about 15 - 40% of the total work being done. The bulk of the work that is done through the workers has a slightly bigger impact towards keeping the ratio of the speed rate differences to about 50-70%

The Observation: - The differences between the two algorithms

In this situation, it is the structure of the process of the work being done. It goes to show that Case 2 is more efficient from both the partial optimization of having a similar function declaration and definition where it is only the variables that differ by name and the distance traveled.

We also see that the total distance traveled in Case 1 is much farther than it is in Case 2 and we can consider this distance traveled our Time Factor between the two algorithms. Case 1 has considerable more work to do than Case 2 does.

This is observable from the evidence of the assembly instructions that were shown in both cases. Along with what was already stated about these cases, this doesn't account for the fact that in Case 1 the boss will have to wait for both A & C to get back before he can go back to A again for each iteration. It also doesn't account for the fact that if A or B is taking an extremely long time then both the Boss and the other worker(s) are idle waiting to be executed.

In Case 2 the only one being idle is the Boss until the worker gets back. So even this has an impact on the algorithm.

The OP's Amended Question(s)

EDIT: The question turned out to be of no relevance, as the behavior severely depends on the sizes of the arrays (n) and the CPU cache. So if there is further interest, I rephrase the question:

Could you provide some solid insight into the details that lead to the different cache behaviors as illustrated by the five regions on the following graph?

It might also be interesting to point out the differences between CPU/cache architectures, by providing a similar graph for these CPUs.

Regarding These Questions

As I have demonstrated without a doubt, there is an underlying issue even before the Hardware and Software becomes involved.

Now as for the management of memory and caching along with page files, etc. which all work together in an integrated set of systems between the following:

• The architecture (hardware, firmware, some embedded drivers, kernels and assembly instruction sets).
• The OS (file and memory management systems, drivers and the registry).
• The compiler (translation units and optimizations of the source code).
• And even the source code itself with its set(s) of distinctive algorithms.

We can already see that there is a bottleneck that is happening within the first algorithm before we even apply it to any machine with any arbitrary architecture, OS, and programmable language compared to the second algorithm. There already existed a problem before involving the intrinsics of a modern computer.

The Ending Results

However; it is not to say that these new questions are not of importance because they themselves are and they do play a role after all. They do impact the procedures and the overall performance and that is evident with the various graphs and assessments from many who have given their answer(s) and or comment(s).

If you paid attention to the analogy of the Boss and the two workers A & B who had to go and retrieve packages from C & D respectively and considering the mathematical notations of the two algorithms in question; you can see without the involvement of the computer hardware and software Case 2 is approximately 60% faster than Case 1.

When you look at the graphs and charts after these algorithms have been applied to some source code, compiled, optimized, and executed through the OS to perform their operations on a given piece of hardware, you can even see a little more degradation between the differences in these algorithms.

If the Data set is fairly small it may not seem all that bad of a difference at first. However, since Case 1 is about 60 - 70% slower than Case 2 we can look at the growth of this function in terms of the differences in time executions:

DeltaTimeDifference approximately = Loop1(time) - Loop2(time)
//where
Loop1(time) = Loop2(time) + (Loop2(time)*[0.6,0.7]) // approximately
// So when we substitute this back into the difference equation we end up with
DeltaTimeDifference approximately = (Loop2(time) + (Loop2(time)*[0.6,0.7])) - Loop2(time)
// And finally we can simplify this to
DeltaTimeDifference approximately = [0.6,0.7]*Loop2(time)

This approximation is the average difference between these two loops both algorithmically and machine operations involving software optimizations and machine instructions.

When the data set grows linearly, so does the difference in time between the two. Algorithm 1 has more fetches than algorithm 2 which is evident when the Boss has to travel back and forth the maximum distance between A & C for every iteration after the first iteration while algorithm 2 the Boss has to travel to A once and then after being done with A he has to travel a maximum distance only one time when going from A to C.

Trying to have the Boss focusing on doing two similar things at once and juggling them back and forth instead of focusing on similar consecutive tasks is going to make him quite angry by the end of the day since he had to travel and work twice as much. Therefore do not lose the scope of the situation by letting your boss getting into an interpolated bottleneck because the boss's spouse and children wouldn't appreciate it.

Amendment: Software Engineering Design Principles

-- The difference between local Stack and heap allocated computations within iterative for loops and the difference between their usages, their efficiencies, and effectiveness --

The mathematical algorithm that I proposed above mainly applies to loops that perform operations on data that is allocated on the heap.

• Consecutive Stack Operations:
  • If the loops are performing operations on data locally within a single code block or scope that is within the stack frame it will still sort of apply, but the memory locations are much closer where they are typically sequential and the difference in distance traveled or execution time is almost negligible. Since there are no allocations being done within the heap, the memory isn't scattered, and the memory isn't being fetched through ram. The memory is typically sequential and relative to the stack frame and stack pointer.
• When consecutive operations are being done on the stack, a modern processor will cache repetitive values and addresses keeping these values within local cache registers. The time of operations or instructions here is on the order of nano-seconds.
• Consecutive Heap Allocated Operations:
  • When you begin to apply heap allocations and the processor has to fetch the memory addresses on consecutive calls, depending on the architecture of the CPU, the bus controller, and the RAM modules the time of operations or execution can be on the order of micro to milliseconds. In comparison to cached stack operations, these are quite slow.
  • The CPU will have to fetch the memory address from RAM and typically anything across the system bus is slow compared to the internal data paths or data buses within the CPU itself.

So when you are working with data that needs to be on the heap and you are traversing through them in loops, it is more efficient to keep each data set and its corresponding algorithms within its own single loop. You will get better optimizations compared to trying to factor out consecutive loops by putting multiple operations of different data sets that are on the heap into a single loop.

It is okay to do this with data that is on the stack since they are frequently cached, but not for data that has to have its memory address queried every iteration.

This is where software engineering and software architecture design comes into play. It is the ability to know how to organize your data, knowing when to cache your data, knowing when to allocate your data on the heap, knowing how to design and implement your algorithms, and knowing when and where to call them.

You might have the same algorithm that pertains to the same data set, but you might want one implementation design for its stack variant and another for its heap-allocated variant just because of the above issue that is seen from its O(n) complexity of the algorithm when working with the heap.

From what I've noticed over the years, many people do not take this fact into consideration. They will tend to design one algorithm that works on a particular data set and they will use it regardless of the data set being locally cached on the stack or if it was allocated on the heap.

If you want true optimization, yes it might seem like code duplication, but to generalize it would be more efficient to have two variants of the same algorithm. One for stack operations, and the other for heap operations that are performed in iterative loops!

Here's a pseudo example: Two simple structs, one algorithm.

struct A {
    int data;
    A() : data{0}{}
    A(int a) : data{a}{}
};
struct B {
    int data;
    B() : data{0}{}
    A(int b) : data{b}{}
}

template<typename T>
void Foo( T& t ) {
    // Do something with t
}

// Some looping operation: first stack then heap.

// Stack data:
A dataSetA[10] = {};
B dataSetB[10] = {};

// For stack operations this is okay and efficient
for (int i = 0; i < 10; i++ ) {
   Foo(dataSetA[i]);
   Foo(dataSetB[i]);
}

// If the above two were on the heap then performing
// the same algorithm to both within the same loop
// will create that bottleneck
A* dataSetA = new [] A();
B* dataSetB = new [] B();
for ( int i = 0; i < 10; i++ ) {
    Foo(dataSetA[i]); // dataSetA is on the heap here
    Foo(dataSetB[i]); // dataSetB is on the heap here
} // this will be inefficient.

// To improve the efficiency above, put them into separate loops...

for (int i = 0; i < 10; i++ ) {
    Foo(dataSetA[i]);
}
for (int i = 0; i < 10; i++ ) {
    Foo(dataSetB[i]);
}
// This will be much more efficient than above.
// The code isn't perfect syntax, it's only psuedo code
// to illustrate a point.

This is what I was referring to by having separate implementations for stack variants versus heap variants. The algorithms themselves don't matter too much, it's the looping structures that you will use them in that do.

In GitHub Wiki markdown I used hash marks (#) followed by two spaces to make the line break larger. It doesn't actually give you multiple line breaks but it made one large line break and served me well for my needs.

Instead of:

text
(space)(space)
more text

I did:

text
(hash mark)(space)(space)
more text

I have an answer:

<a href="#">
    <div class="widget">  
        <div class="title" style="text-decoration: none;">Underlined. Why?</div>  
    </div>
</a>?

It works.

It depends what you are looking for, if you are just looking to see if it is empty just use empty as it checks whether it is set as well, if you want to know whether something is set or not use isset.

Empty checks if the variable is set and if it is it checks it for null, "", 0, etc

Isset just checks if is it set, it could be anything not null

With empty, the following things are considered empty:

• "" (an empty string)
• 0 (0 as an integer)
• 0.0 (0 as a float)
• "0" (0 as a string)
• NULL
• FALSE
• array() (an empty array)
• var $var; (a variable declared, but without a value in a class)

From http://php.net/manual/en/function.empty.php

As mentioned in the comments the lack of warning is also important with empty()

PHP Manual says

empty() is the opposite of (boolean) var, except that no warning is generated when the variable is not set.

Regarding isset

PHP Manual says

isset() will return FALSE if testing a variable that has been set to NULL

Your code would be fine as:

<?php
    $var = '23';
    if (!empty($var)){
        echo 'not empty';
    }else{
        echo 'is not set or empty';
    }
?>

For example:

$var = "";

if(empty($var)) // true because "" is considered empty
 {...}
if(isset($var)) //true because var is set 
 {...}

if(empty($otherVar)) //true because $otherVar is null
 {...}
if(isset($otherVar)) //false because $otherVar is not set 
 {...}

You might need to collect the stats as you go, but @@ROWCOUNT captures this:

declare @Fish table (
Name varchar(32)
)

insert into @Fish values ('Cod')
insert into @Fish values ('Salmon')
insert into @Fish values ('Butterfish')
update @Fish set Name = 'LurpackFish' where Name = 'Butterfish'
select @@ROWCOUNT  --gives 1

update @Fish set Name = 'Dinner'
select @@ROWCOUNT -- gives 3

All the above answers are correct. Just providing with your dataset to find perfect divisor:

#include <stdio.h>

int main() 
{

int arr[7] = {3,5,7,8,9,17,19};
int j = 51;
int i = 0;

for (i=0 ; i < 7; i++) {
    if (j % arr[i] == 0)
        printf("%d is the perfect divisor of %d\n", arr[i], j);
}

return 0;
}

I disagree with that this:

String s = myObj == null ? "" : myObj.ToString();

is a hack in any way. I think it's a good example of clear code. It's absolutely obvious what you want to achieve and that you're expecting null.

UPDATE:

I see now that you were not saying that this was a hack. But it's implied in the question that you think this way is not the way to go. In my mind it's definitely the clearest solution.

CODE

$('#montant-total-prevu').on("change", function() {

var taille = $('#montant-total-prevu').val().length;

    if (taille > 9) {

//TODO

}

});

Execute function FetchData() once after 1000 milliseconds:

setTimeout(FetchData,1000);

Execute function FetchData() repeatedly every 1000 milliseconds:

setInterval(FetchData,1000);

That is because in Python 3, they have replaced the print statement with the print function.

The syntax is now more or less the same as before, but it requires parens:

From the "what's new in python 3" docs:

Old: print "The answer is", 2*2
New: print("The answer is", 2*2)

Old: print x,           # Trailing comma suppresses newline
New: print(x, end=" ")  # Appends a space instead of a newline

Old: print              # Prints a newline
New: print()            # You must call the function!

Old: print >>sys.stderr, "fatal error"
New: print("fatal error", file=sys.stderr)

Old: print (x, y)       # prints repr((x, y))
New: print((x, y))      # Not the same as print(x, y)!

I came up with this function which works for me, hope this will help somebody

$word_list = 'word1, word2, word3, word4';
$str = 'This string contains word1 in it';

function checkStringAgainstList($str, $word_list)
{
  $word_list = explode(', ', $word_list);
  $str = explode(' ', $str);

  foreach ($str as $word):
    if (in_array(strtolower($word), $word_list)) {
        return TRUE;
    }
  endforeach;

  return false;
}

Also, note that answers with strpos() will return true if the matching word is a part of other word. For example if word list contains 'st' and if your string contains 'street', strpos() will return true

A primary key defines a unique identifier for every row in a table: when a table has a primary key, you have a guranteed way to select any row from it.

A unique constraint does not necessarily identify every row; it just specifies that if a row has values in its columns, then they must be unique. This is not sufficient to uniquely identify every row, which is what a primary key must do.

Say you have a download function to download a file from network, and want to be notified when download task has finished.

typealias CompletionHandler = (success:Bool) -> Void

func downloadFileFromURL(url: NSURL,completionHandler: CompletionHandler) {

    // download code.

    let flag = true // true if download succeed,false otherwise

    completionHandler(success: flag)
}

// How to use it.

downloadFileFromURL(NSURL(string: "url_str")!, { (success) -> Void in

    // When download completes,control flow goes here.
    if success {
        // download success
    } else {
        // download fail
    }
})

Hope it helps.

Things are about to change on the "eol conversion" front, with the upcoming Git 1.7.2:

A new config setting core.eol is being added/evolved:

This is a replacement for the 'Add "core.eol" config variable' commit that's currently in pu (the last one in my series).
Instead of implying that "core.autocrlf=true" is a replacement for "* text=auto", it makes explicit the fact that autocrlf is only for users who want to work with CRLFs in their working directory on a repository that doesn't have text file normalization.
When it is enabled, "core.eol" is ignored.

Introduce a new configuration variable, "core.eol", that allows the user to set which line endings to use for end-of-line-normalized files in the working directory.
It defaults to "native", which means CRLF on Windows and LF everywhere else. Note that "core.autocrlf" overrides core.eol.
This means that:

[core]
  autocrlf = true

puts CRLFs in the working directory even if core.eol is set to "lf".

core.eol:

Sets the line ending type to use in the working directory for files that have the text property set.
Alternatives are 'lf', 'crlf' and 'native', which uses the platform's native line ending.
The default value is native.

Other evolutions are being considered:

For 1.8, I would consider making core.autocrlf just turn on normalization and leave the working directory line ending decision to core.eol, but that will break people's setups.

git 2.8 (March 2016) improves the way core.autocrlf influences the eol:

See commit 817a0c7 (23 Feb 2016), commit 6e336a5, commit df747b8, commit df747b8 (10 Feb 2016), commit df747b8, commit df747b8 (10 Feb 2016), and commit 4b4024f, commit bb211b4, commit 92cce13, commit 320d39c, commit 4b4024f, commit bb211b4, commit 92cce13, commit 320d39c (05 Feb 2016) by Torsten Bögershausen (tboegi).
^{(Merged by Junio C Hamano -- gitster -- in commit c6b94eb, 26 Feb 2016)}

convert.c: refactor crlf_action

Refactor the determination and usage of crlf_action.
Today, when no "crlf" attribute are set on a file, crlf_action is set to CRLF_GUESS. Use CRLF_UNDEFINED instead, and search for "text" or "eol" as before.

Replace the old CRLF_GUESS usage:

CRLF_GUESS && core.autocrlf=true -> CRLF_AUTO_CRLF
CRLF_GUESS && core.autocrlf=false -> CRLF_BINARY
CRLF_GUESS && core.autocrlf=input -> CRLF_AUTO_INPUT

Make more clear, what is what, by defining:

- CRLF_UNDEFINED : No attributes set. Temparally used, until core.autocrlf
                   and core.eol is evaluated and one of CRLF_BINARY,
                   CRLF_AUTO_INPUT or CRLF_AUTO_CRLF is selected
- CRLF_BINARY    : No processing of line endings.
- CRLF_TEXT      : attribute "text" is set, line endings are processed.
- CRLF_TEXT_INPUT: attribute "input" or "eol=lf" is set. This implies text.
- CRLF_TEXT_CRLF : attribute "eol=crlf" is set. This implies text.
- CRLF_AUTO      : attribute "auto" is set.
- CRLF_AUTO_INPUT: core.autocrlf=input (no attributes)
- CRLF_AUTO_CRLF : core.autocrlf=true  (no attributes)

As torek adds in the comments:

all these translations (any EOL conversion from eol= or autocrlf settings, and "clean" filters) are run when files move from work-tree to index, i.e., during git add rather than at git commit time.
(Note that git commit -a or --only or --include do add files to the index at that time, though.)

For more on that, see "What is difference between autocrlf and eol".

The full list of the 66 unreserved characters is in RFC3986, here: http://tools.ietf.org/html/rfc3986#section-2.3

This is any character in the following regex set:

[A-Za-z0-9_.\-~]

There is no guarantee that your threads are executing simultaneously regardless of any trivial example anyone else posts. If your OS only gives the java process one processor to work on, your java threads will still be scheduled for each time slice in a round robin fashion. Meaning, no two will ever be executing simultaneously, but the work they do will be interleaved. You can use monitoring tools like Java's Visual VM (standard in the JDK) to observe the threads executing in a Java process.

The column names which are mixed case or uppercase have to be double quoted in PostgresQL. So best convention will be to follow all small case with underscore.

Many different ways have already been described by the other answers. However, many of them either involved ShowDialog() or that form1 stay open but hidden. The best and most intuitive way in my opinion is to simply close form1 and then create form2 from an outside location (i.e. not from within either of those forms). In the case where form1 was created in Main, form2 can simply be created using Application.Run just like form1 before. Here's an example scenario:

I need the user to enter their credentials in order for me to authenticate them somehow. Afterwards, if authentication was successful, I want to show the main application to the user. In order to accomplish this, I'm using two forms: LogingForm and MainForm. The LoginForm has a flag that determines whether authentication was successful or not. This flag is then used to decide whether to create the MainForm instance or not. Neither of these forms need to know about the other and both forms can be opened and closed gracefully. Here's the code for this:

class LoginForm : Form
{
    public bool UserSuccessfullyAuthenticated { get; private set; }

    void LoginButton_Click(object s, EventArgs e)
    {
        if(AuthenticateUser(/* ... */))
        {
            UserSuccessfullyAuthenticated = true;
            Close();
        }
    }
}

static class Program
{
    [STAThread]
    static void Main()
    {
        LoginForm loginForm = new LoginForm();
        Application.Run(loginForm);

        if(loginForm.UserSuccessfullyAuthenticated)
        {
            // MainForm is defined elsewhere
            Application.Run(new MainForm());
        }
    }
}

apt repo have this extension, just run the below command from your terminal ::

sudo apt-get install php7.2-bcmath*

I suggest you use null to represent a null value.

What is the exception you get?

BTW:

There is no year called 0 or 0000. (Though some dates allow this year)

And there is no 0 month of the year or 0 day of the month. (Which may be the cause of your problem)

If this is for something that will be called a lot in an application instance, it's a lot faster to compile and cache dynamic code instead of using the activator or ConstructorInfo.Invoke(). Two easy options for dynamic compilation are compiled Linq Expressions or some simple IL opcodes and DynamicMethod. Either way, the difference is huge when you start getting into tight loops or multiple calls.

I think you can also use BufferedWriter :

BufferedWriter writer = new BufferedWriter(new FileWriter(new File("note.txt")));

String stuffToWrite = info;

writer.write(stuffToWrite);

writer.close();

and before that remember too add

import java.io.BufferedWriter;