Search Results

Search found 12530 results on 502 pages for 'icc cricket world cup 2011'.

Page 445/502 | < Previous Page | 441 442 443 444 445 446 447 448 449 450 451 452  | Next Page >

  • BlackBerry Deployment Strategies

    - by cagreen
    I'm new to large scale BB app deployment and I'm looking for some clarification on the various methods of deployment. Please bear with me as I'm sure there is more to it than my naive view would lead me to believe. My app is very targeted to corporate users and requires a subscription to some additional services before it can be used. In other words, it's not targeted towards the consumer market, so I'm not worried about people not being able to easily find it online. What do I need to be aware of when looking at deployment strategies? Any gotchas? From my understanding my choices are: - App World small upfront vendor fee users can easily search for and find my app billing handled by RIM 4 licensing models (static, single, pool, dynamic). Though I'm not sure I've seen enough info on the pool and dynamic to fully appreciate how it might help me. - Download from my website billing is handled by me can I enforce the number of licenses that are in use within an organization? is this easier/harder for a user? - What else am I missing?

    Read the article

  • Glitch when moving camera in OpenGL

    - by CG
    I am writing a tile-based game engine for the iPhone and it works in general apart from the following glitch. Basically, the camera will always keep the player in the centre of the screen, and it moves to follow the player correctly and draws everything correctly when stationary. However whilst the player is moving, the tiles of the surface the player is walking on glitch as shown: Compared to the stationary (correct): Does anyone have any idea why this could be? Thanks for the responses so far. Floating point error was my first thought also and I tried slightly increasing the size of the tiles but this did not help. Changing glClearColor to red still leaves black gaps so maybe it isn't floating point error. Since the tiles in general will use different textures, I don't know if vertex arrays can be used (I always thought that the same texture had to be applied to everything in the array, correct me if I'm wrong), and I don't think VBO is available in OpenGL ES. Setting the filtering to nearest neighbour improved things but the glitch still happens every ten frames or so, and the pixelly result means that this solution is not viable anyway. The main difference between what I'm doing now and what I've done in the past is that this time I am moving the camera rather than the stationary objects in the world (i.e. the tiles, the player is still being moved). The code I'm using to move the camera is: void Camera::CentreAtPoint( GLfloat x, GLfloat y ) { glMatrixMode(GL_PROJECTION); glLoadIdentity(); glOrthof(x - size.x / 2.0f, x + size.x / 2.0f, y + size.y / 2.0f, y - size.y / 2.0f, 0.01f, 5.0f); glMatrixMode(GL_MODELVIEW); } Is there a problem with doing things this way and if so is there a solution?

    Read the article

  • Why is Chrome miscalculating jQuery submenu dimensions?

    - by chunkymonkey
    I'm trying to implement this dropdown menu with flyouts: http://jsfiddle.net/chunkymonkey/fr6x4/ In Chrome certain categories can be expanded to show their subcategories while others show nothing when opened up. For example: Alternative Rock can be expanded to show its multiple subcategories . . . BUT . . . World Music, which has as many subcategories, shows no subcategories when expanded. (SCREENSHOT: http://i.imgur.com/0WorR.jpg) I thought I had tracked this problem down to a problem with they way the dimensions of the dropdown elements are calculated in the original code: First change: - var newLeftVal = - ($('.fg-menu-current').parents('ul').size() - 1) * 180; + var newLeftVal = - ($('.fg-menu-current').parents('ul').size() - 1) * container.width(); Second change: Remove: var checkMenuHeight = function(el) { if (el.height() > options.maxHeight) { el.addClass('fg-menu-scroll') }; el.css({ height: options.maxHeight }); }; Add: var checkMenuHeight = function(el) { var max_height = options.maxHeight - breadcrumb.getTotalHeight(); if (el.height() > max_height) { el.addClass('fg-menu-scroll'); el.height(max_height); topList.height(max_height); } else { if (topList.height() < el.height()) { topList.height(el.height()); } } }; But it's still not working only on Chrome (version 8, Windows & Mac) (not sure why Chrome is different).

    Read the article

  • Haskell Cons Operator (:)

    - by Carson Myers
    I am really new to Haskell (Actually I saw "Real World Haskell" from O'Reilly and thought "hmm, I think I'll learn functional programming" yesterday) and I am wondering: I can use the construct operator to add an item to the beginning of a list: 1 : [2,3] [1,2,3] I tried making an example data type I found in the book and then playing with it: --in a file data BillingInfo = CreditCard Int String String | CashOnDelivery | Invoice Int deriving (Show) --in ghci $ let order_list = [Invoice 2345] $ order_list [Invoice 2345] $ let order_list = CashOnDelivery : order_list $ order_list [CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, CashOnDelivery, ...- etc... it just repeats forever, is this because it uses lazy evaluation? -- EDIT -- okay, so it is being pounded into my head that let order_list = CashOnDelivery:order_list doesn't add CashOnDelivery to the original order_list and then set the result to order_list, but instead is recursive and creates an infinite list, forever adding CashOnDelivery to the beginning of itself. Of course now I remember that Haskell is a functional language and I can't change the value of the original order_list, so what should I do for a simple "tack this on to the end (or beginning, whatever) of this list?" Make a function which takes a list and BillingInfo as arguments, and then return a list? -- EDIT 2 -- well, based on all the answers I'm getting and the lack of being able to pass an object by reference and mutate variables (such as I'm used to)... I think that I have just asked this question prematurely and that I really need to delve further into the functional paradigm before I can expect to really understand the answers to my questions... I guess what i was looking for was how to write a function or something, taking a list and an item, and returning a list under the same name so the function could be called more than once, without changing the name every time (as if it was actually a program which would add actual orders to an order list, and the user wouldn't have to think of a new name for the list each time, but rather append an item to the same list).

    Read the article

  • How to get the most out of a 3 month intern?

    - by firoso
    We've got a software engineering intern coming in who's fairly competent and shows promise. There's one catch: we have him for 3 months full time and can't count on anything past that. He still has a year of school left, which is why we can't say for sure that we have him past 3 months. We have a specific project we're putting him on. How can we maximize his productivity while still giving him a positive learning experience? He wants to learn about development cycles and real-world software engineering. Anything that you think would be critical that you wish you had learned earlier? Nearly six months later: He's preformed admirably and even I have learned a lot from him. Thank you all for the input. Now I want to provide feedback to YOU! He has benefited most from sitting down and writing code. However, he has had a nasty history of bad software engineering practices which I'm trying to replace with good habits (properly finishing a method before moving on, not hacking code together, proper error channeling, etc). He has also really gained a lot by feeling involved in design decisions, even if most of the time they're related to my own design plans.

    Read the article

  • Drawing an image in Java, slow as hell on a netbook.

    - by Norswap
    In follow-up to my previous questions (especially this one : http://stackoverflow.com/questions/2684123/java-volatileimage-slower-than-bufferedimage), i have noticed that simply drawing an Image (it doesn't matter if it's buffered or volatile, since the computer has no accelerated memory*, and tests shows it's doesn't change anything), tends to be very long. (*) System.out.println(GraphicsEnvironment.getLocalGraphicsEnvironment() .getDefaultScreenDevice().getAvailableAcceleratedMemory()); --> 0 How long ? For a 500x400 image, about 0.04 seconds. This is only drawing the image on the backbuffer (obtained via buffer strategy). Now considering that world of warcraft runs on that netbook (tough it is quite laggy) and that online java games seems to have no problem whatsoever, this is quite thought provoking. I'm quite certain I didn't miss something obvious, I've searched extensively the web, but nothing will do. So do any of you java whiz have an idea of what obscure problem might be causing this (or maybe it is normal, tough I doubt it) ? PS : As I'm writing this I realized this might be cause by my Linux installation (archlinux) tough I have the correct Intel driver. But my computer normally has "Integrated Intel Graphics Media Accelerator 950", which would mean it should have accelerated video memory somehow. Any ideas about this side of things ?

    Read the article

  • No GPS Update retrieved? Problem in Code?

    - by poeschlorn
    Hello mates, I've got a serious problem with my GPS on my Nexus One: I wrote a kind of hello world with GPS, but the Toast that should be displayed isn't :( I don't know what I'm doing wrong...maybe you could help me getting this work. Here's my code: package gps.test; import android.app.Activity; import android.content.Context; import android.location.Location; import android.location.LocationListener; import android.location.LocationManager; import android.os.Bundle; import android.widget.Toast; public class GPS extends Activity { private LocationManager lm; private LocationListener locationListener; /** Called when the activity is first created. */ @Override public void onCreate(Bundle savedInstanceState) { super.onCreate(savedInstanceState); setContentView(R.layout.main); // ---use the LocationManager class to obtain GPS locations--- lm = (LocationManager) getSystemService(Context.LOCATION_SERVICE); locationListener = new MyLocationListener(); lm.requestLocationUpdates(LocationManager.GPS_PROVIDER, 100, 1, locationListener); } private class MyLocationListener implements LocationListener { @Override public void onLocationChanged(Location loc) { if (loc != null) { Toast.makeText( getBaseContext(), "Location changed : Lat: " + loc.getLatitude() + " Lng: " + loc.getLongitude(), Toast.LENGTH_SHORT).show(); } } @Override public void onProviderDisabled(String provider) { // TODO Auto-generated method stub } @Override public void onProviderEnabled(String provider) { // TODO Auto-generated method stub } @Override public void onStatusChanged(String provider, int status, Bundle extras) { // TODO Auto-generated method stub } } } Theoretically there should be a new toast every 100 milliseconds, shouldn't it? Or at least, when I change my position by one meter!? I've no idea why it doesn't. I must admit I'm new to the topic, maybe I've missed something? It would be great if you could give me a hint :) nice greetings, poeschlorn

    Read the article

  • Session variables not getting set but only in Internet Explorer and not on all machines

    - by gaoshan88
    Logging into a site I'm working on functions as expected on my local machine but fails on the remote server but ONLY in Internet Explorer. The kicker is that it works in IE locally, just not on the remote machine. What in the world could cause this? I have stepped through the code on the remote machine and can see the entered login values being checked in the database, they are found and then a login function is called. This sets two $_SESSION variables and redirects to the main admin page. However, in IE only (and not when run on local machine... this is key) the $_SESSION variables are not present by the time you get to the main admin page. var_dump($_SESSION) gives me what I expect on every browser when I am running this in my local environment and in every browser except IE 6, 7 and 8 when run on the remote server (where I get a null value as if nothing has been set for $_SESSION). This really has me stumped so any advice is appreciated. For an example... in IE, run locally, var_dump gives me: array 'Username' => string 'theusername' length=11 'UserID' => string 'somevalue' length=9 Run on the remote server (IE only... works fine in other browsers) var_dump gives me: array(0){} Code: $User = GetUser($Username, $Password); if ($User->UserID <> "") { // this works so we call Login()... Login($User); // this also works and gives expected results. on to redirect... header("Location: index.php"); // a var_dump at index.php shows that there is no session data at all in IE, remotely. } else { header("Location: login.php"); } function Login($data) { $_SESSION['Username'] = $data->Username; $_SESSION['UserID'] = $data->UserID; // a var dump here gives the expected data in every browser }

    Read the article

  • Where does "new" fit in the flex creation cycle?

    - by deux11
    In the following code, the call to myChild.bar() results in an exception because myChild is null. myParent is a valid object. What I don't understand is why myChild has not been created yet. I have read the following document related to object creation sequence, but I am unsure how "new" is related: http://livedocs.adobe.com/flex/3/html/help.html?content=layoutperformance_03.html Any help is appreciated! // Main.mxml <mx:Application xmlns:mx="http://www.adobe.com/2006/mxml" creationComplete="created()"> <mx:Script> <![CDATA[ public var myParent:Parent = new Parent(); public function created():void { myParent.foo(); } ]]> </mx:Script> </mx:Application> // Parent.mxml <mx:Canvas xmlns:mx="http://www.adobe.com/2006/mxml" xmlns="*"> <mx:Script> <![CDATA[ public function foo():void { myChild.bar(); } ]]> </mx:Script> <Child id="myChild"/> </mx:Canvas> // Child.mxml <mx:Canvas xmlns:mx="http://www.adobe.com/2006/mxml"> <mx:Script> <![CDATA[ public function bar():void { trace("Hello World"); } ]]> </mx:Script> </mx:Canvas>

    Read the article

  • Quality of TFS 2008 merged code

    - by paologios
    Does the quality of code merged by TFS 2008 depend on the used programming language? I know merging in Java / Subversion, and merging a branch to its trunk usually does not create much conflicts. Now in my company, we use VB.NET. When I merge two files TFS does not always get code blocks right, e.g. does not find the right If..then / end if lines. To give you an example, I mean: File 2 is created as a branch of File 1. Both files were changed later, now I'm going to merge those files and am recieving conficts: The marked end-if lines (1) are detected as corresponding, meaning the added event handler Button1_Click is being deleted. Now I wonder if this behavior is by language (C# vs. VB.NET) or are other source control solutions just better than TFS? (And I really liked TFS up to now :) ) File 1: Protected Sub Page_Load(ByVal sender As Object, ByVal e As System.EventArgs) Handles Me.Load If Not Page.IsPostBack Then Label1.Text = "Hello" Label2.Text = "World" End If End Sub Protected Sub Button2_Click(ByVal sender, ByVal e as System.EventArgs) Handles Button2.Click // .... If Page.IsValid Then Label3.Text = "Hello Button 2" End If // .... End Sub File 2 (Branch of File 1): Protected Sub Page_Load(ByVal sender As Object, ByVal e As System.EventArgs) Handles Me.Load If Not Page.IsPostBack Then fillTableFromDatabase() End If // (1) End Sub Protected Sub Button1_Click(ByVal sender, ByVal e as System.EventArgs) Handles Button1.Click // do something here End Sub Protected Sub Button2_Click(ByVal sender, ByVal e as System.EventArgs) Handles Button2.Click // .... If Page.IsValid Then End If // (1) // .... End Sub

    Read the article

  • DOM manipulation in PHP

    - by bluedaniel
    Hello everyone, Im trying to use the DOM in PHP to do a pretty specific job and Ive got no luck so far, the objective is to take a string of HTML from a Wordpress blog post (from the DB, this is a wordpress plugin). And then out of that HTML replace <div id="do_not_edit">old content</div>" with <div id="do_not_edit">new content</div>" in its place. Saving anything above and below that div in its structure. Then save the HTML back into the DB, should be simple really, I have read that a regex wouldnt be the right way to go here so Ive turned to the DOM instead. The problem is I just cant get it to work, cant extract the div or anything. Help me!! UPDATE The HTML coming out of the wordpress table looks like: Congratulations on finding us here on the world wide web, we are on a mission to create a website that will show off your culinary skills better than any other website does. <div id="do_not_edit">blah blah</div> We want this website to be fun and easy to use, we strive for simple elegance and incredible functionality.We aim to provide a 'complete package'. By this we want to create a website where people can meet, share ideas and help each other out. After several different (incorrect) workings all Ive got below is: $content = ($wpdb->get_var( "SELECT `post_content` FROM $wpdb->posts WHERE ID = {$article[post_id]}" )); $doc = new DOMDocument(); $doc->validateOnParse = true; $doc->loadHTMLFile($content); $element = $doc->getElementById('do_not_edit'); echo $element;

    Read the article

  • How to track auto-generated id's in select-insert statement

    - by k rey
    I have two tables detail and head. The detail table will be written first. Later, the head table will be written. The head is a summary of the detail table. I would like to keep a reference from the detail to the head table. I have a solution but it is not elegant and requires duplicating the joins and filters that were used during summation. I am looking for a better solution. The below is an example of what I currently have. In this example, I have simplified the table structure. In the real world, the summation is very complex. -- Preparation create table #detail ( detail_id int identity(1,1) , code char(4) , amount money , head_id int null ); create table #head ( head_id int identity(1,1) , code char(4) , subtotal money ); insert into #detail ( code, amount ) values ( 'A', 5 ); insert into #detail ( code, amount ) values ( 'A', 5 ); insert into #detail ( code, amount ) values ( 'B', 2 ); insert into #detail ( code, amount ) values ( 'B', 2 ); -- I would like to somehow simplify the following two queries insert into #head ( code, subtotal ) select code, sum(amount) from #detail group by code update #detail set head_id = h.head_id from #detail d inner join #head h on d.code = h.code -- This is the desired end result select * from #detail Desired end result of detail table: detail_id code amount head_id 1 A 5.00 1 2 A 5.00 1 3 B 2.00 2 4 B 2.00 2

    Read the article

  • Typical Hadoop setup for remote job submission

    - by Artii
    So I am still a bit new to hadoop and am currently in the process of setting up a small test cluster on Amazonaws. So my question relates to some tips on the structuring of the cluster so it is possible to work submit jobs from remote machines. Currently I have 5 machines. 4 are basically the Hadoop cluster with the NameNodes, Yarn etc. One machine is used as a manager machine( Cloudera Manager). I am gonna describe my thinking process on the setup and if anyone can chime in the points I am not clear with, that would be great. I was thinking what was the best setup for a small cluster. So I decided to expose only one manager machine and probably use that to submit all the jobs through it. The other machines will see each other etc, but not be accessible from the outside world. I am have conceptual idea on how to do this,but I am not sure how to properly go about doing this though, if anyone could point me in the right direction that would great. Also another big point is, I want to be able to submit jobs to the cluster through exposed machine from a client machine (might be Windows). I am not so clear on this setup as well. Do I need to have Hadoop installed on the machine in order to use the normal hadoop commands, and to write/submit jobs say from Eclipse or something similar. So to sum it up my questions are, Is this an ok setup for a small test cluster How can I go about using one exposed machine to submit/route jobs to the cluster, without having any of the Hadoop nodes on it. How do I setup a client machine to submit jobs to a remote cluster, and an example on how to do it on Windows. Also if there are any reason not to use Windows as a client machine in this setup. Thanks I would greatly appreciate any advice or help on this.

    Read the article

  • rabbitmq-erlang-client, using rebar friendly pkg, works on dev env fails on rebar release

    - by lfurrea
    I am successfully using the rebar-friendly package of rabbitmq-erlang-client for a simple Hello World rebarized and OTP "compliant" app and things work fine on the dev environment. I am able to fire up an erl console and do my application:start(helloworld). and connect to the broker, open up a channel and communicate to queues. However, then I proceed to do rebar generate and it builds up the release just fine, but when I try to fire up from the self contained release package then things suddenly explode. I know rebar releases are known to be an obscure art, but I would like to know what are my options as far as deployment for an app using the rabbitmq-erlang-client. Below you will find the output of the console on the crash: =INFO REPORT==== 18-Dec-2012::16:41:35 === application: session_record exited: {{{badmatch, {error, {'EXIT', {undef, [{amqp_connection_sup,start_link, [{amqp_params_network,<<"guest">>,<<"guest">>,<<"/">>, "127.0.0.1",5672,0,0,0,infinity,none, [#Fun<amqp_auth_mechanisms.plain.3>, #Fun<amqp_auth_mechanisms.amqplain.3>], [],[]}], []}, {supervisor2,do_start_child_i,3, [{file,"src/supervisor2.erl"},{line,391}]}, {supervisor2,handle_call,3, [{file,"src/supervisor2.erl"},{line,413}]}, {gen_server,handle_msg,5, [{file,"gen_server.erl"},{line,588}]}, {proc_lib,init_p_do_apply,3, [{file,"proc_lib.erl"},{line,227}]}]}}}}, [{amqp_connection,start,1, [{file,"src/amqp_connection.erl"},{line,164}]}, {hello_qp,start_link,0,[{file,"src/hello_qp.erl"},{line,10}]}, {session_record_sup,init,1, [{file,"src/session_record_sup.erl"},{line,55}]}, {supervisor_bridge,init,1, [{file,"supervisor_bridge.erl"},{line,79}]}, {gen_server,init_it,6,[{file,"gen_server.erl"},{line,304}]}, {proc_lib,init_p_do_apply,3, [{file,"proc_lib.erl"},{line,227}]}]}, {session_record_app,start,[normal,[]]}} type: permanent

    Read the article

  • Does it ever make sense to make a fundamental (non-pointer) parameter const?

    - by Scott Smith
    I recently had an exchange with another C++ developer about the following use of const: void Foo(const int bar); He felt that using const in this way was good practice. I argued that it does nothing for the caller of the function (since a copy of the argument was going to be passed, there is no additional guarantee of safety with regard to overwrite). In addition, doing this prevents the implementer of Foo from modifying their private copy of the argument. So, it both mandates and advertises an implementation detail. Not the end of the world, but certainly not something to be recommended as good practice. I'm curious as to what others think on this issue. Edit: OK, I didn't realize that const-ness of the arguments didn't factor into the signature of the function. So, it is possible to mark the arguments as const in the implementation (.cpp), and not in the header (.h) - and the compiler is fine with that. That being the case, I guess the policy should be the same for making local variables const. One could make the argument that having different looking signatures in the header and source file would confuse others (as it would have confused me). While I try to follow the Principle of Least Astonishment with whatever I write, I guess it's reasonable to expect developers to recognize this as legal and useful.

    Read the article

  • DOMNode reference doesn't work anymore after appending it to another element twice

    - by Robbie Groenewoudt
    Hi, I'm using the a wrapper around the PHP5-class DOMDocument to generate my HTML. This makes it easy to modify the HTML by using the DOM. An example is creating element #1 and adding it to the element #2 and still be able to modify element #1 directly. A problem arises however with the following: Element #1 is added to element #2 Element #2 is added to element #3 Element #1 is modified but no changes are visible in the DOM of element #3 (which contains #1 and #2) A simplified sample code: <?php $doc1 = new DOMDocument(); $el1 = $doc1->createElement('h1', 'Hello'); $doc1->appendChild($el1); $doc2 = new DOMDocument(); $el2 = $doc2->createElement('h2', 'World'); $doc2->appendChild($el2); $doc3 = new DOMDocument(); $el3 = $doc3->createElement('h3', 'Today'); $doc3->appendChild($el3); // Import el1 into el2 $el1 = $doc2->importNode($el1, true); $el2->appendChild( $el1 ); $doc1 = $doc2; // Import el2 into el3 $el2 = $doc3->importNode($el2, true); //$el1 = $doc3->importNode($el1, true); Necessary? $el3->appendChild($el2); $doc2 = $doc3; // Modify el1 $el1->nodeValue = "Boo"; // This doesn't work? //$el2->nodeValue = "Boo"; // Changing element2 or 3 works... // Display result echo $doc3->saveHTML(); ?>` Any idea's on why modifying $el1 won't work? (While $el2 works fine) Or an easy way to set $el1 to the right element?

    Read the article

  • HOWTO: implement a jQuery version of ASP.Net MVC "Strongly Typed Partial Views"

    - by Sam Carleton
    I am working on a multi-page assessment form where the questions/responses are database driven. Currently I the basic system working with Html.BeginForm via standard ASP.Net MVC. At this point in time, the key to the whole system is the 'Strongly Typed Partial Views'. When the question/response is read from the database, the response type determines which derived model is created and added to the collection. The main view it iterates through the collection and uses the 'Strongly Typed Partial Views' system of ASP.Net MVC to determine which view to render the correct type of response (radio button, drop down, or text box). I would like to change this process from a Html.BeginForm to Ajax.BeginForm. The problem is I don't have a clue as to how to implement the dynamic creation of the question/response in the JavaScript/jQuery world. Any thoughts and/or suggestions? Here is the current code to generate the dynamic form: @using (Html.BeginForm(new { mdsId = @Model.MdsId, sectionId = @Model.SectionId })) { <div class="SectionTitle"> <span>Section @Model.SectionName - @Model.SectionDescription</span> <span style="float: right">@Html.CheckBoxFor(x => x.ShowUnansweredQuestions) Show only unaswered questions</span> </div> @Html.HiddenFor(x => x.PrevSectionId) @Html.HiddenFor(x => x.NextSectionId) for (var i = 0; i < Model.answers.Count(); i++) { @Html.EditorFor(m => m.answers[i]); } }

    Read the article

  • How do you slow down the output from a DOS command

    - by JW
    I have lots of experience of writing php scripts that are run in the context of a webserver and almost no epxerience of writing php scripts for CLI or GUI output. I have used the command line for linux but do not have much expereince with DOS. Lets say I have php script that is: <?php echo('Hello world'); for ($idx = 0 ; $idx < 100 ; $idx++ ) { echo 'I am line '. $idx . PHP_EOL; } Then, I run it in my DOS Command prompt: # php helloworld.php Now this will spurt out the output quckly and i have to scroll the DOS command window up to see the output. I want to see the output one 'screen full' at a time. How do you do that from the perspective of a DOS user? Furthermore, although this is not my main main question, I would be also interested in knowing how to make the php script 'wait for input' from the command prompt.

    Read the article

  • Why are (almost) all the on-line games written in ActionScript (Flash) not Java?

    - by MasterPeter
    I absolutely love good defender games (e.g. Gemcraft, Protector: reclaiming the throne) as they can be intellectually quite challenging; it's like playing chess but a little less thinking a bit more action. Sadly, there are not that many good ones out there and I thought I would create one myself and share it with the rest of the world by making it available on-line. I have never worked with ActionScript but when it comes to on-line games, this is the main choice. I have tried to find a decent 2D game in the form of a Java applet but to no avail. Why is this so? I could write the game, most comfortably, in Delphi for Win32 but then people would need to download the executable, which could deter some form downloading it, and also it would only work on Windows. I am also familiar with Java, having worked with Java for the last four years or so. Although I don't have much experience with games programming. Should I note be deterred by the fact that all online games are written for in Flash and create my defender game as a Java applet, or should I consider learning ActionScript and games development for the ActionScript Virtual Machine (AS3 looks very much like Java... but still, it's an entirely new technology to me and I might never use it professionally.) Could you, please, just answer the the question in the title? Why Flash, not Java applets? Is it only 'politics'?

    Read the article

  • PHP/MySQL time zone migration

    - by El Yobo
    I have an application that currently stores timestamps in MySQL DATETIME and TIMESTAMP values. However, the application needs to be able to accept data from users in multiple time zones and show the timestamps in the time zone of other users. As such, this is how I plan to amend the application; I would appreciate any suggestions to improve the approach. Database modifications All TIMESTAMPs will be converted to DATETIME values; this is to ensure consistency in approach and to avoid having MySQL try to do clever things and convert time zones (I want to keep the conversion in PHP, as it involves less modification to the application, and will be more portable when I eventually manage to escape from MySQL). All DATETIME values will be adjusted to convert them to UTC time (currently all in Australian EST) Query modifications All usage of NOW() to be replaced with UTC_TIMESTAMP() in queries, triggers, functions, etc. Application modifications The application must store the time zone and preferred date format (e.g. US vs the rest of the world) All timestamps will be converted according to the user settings before being displayed All input timestamps will be converted to UTC according to the user settings before being input Additional notes Converting formats will be done at the application level for several main reasons The approach to converting time zones varies from DB to DB, so handing it there will be non-portable (and I really hope to be migrating away from MySQL some time in the not-to-distant future). MySQL TIMESTAMPs have limited ranges to the permitted dates (~1970 to ~2038) MySQL TIMESTAMPs have other undesirable attributes, including bizarre auto-update behaviour (if not carefully disabled) and sensitivity to the server zone settings (and I suspect I might screw these up when I migrate to Amazon later in the year). Is there anything that I'm missing here, or does anyone have better suggestions for the approach?

    Read the article

  • How can I lookup data about a book from its barcode number?

    - by Joel Spolsky
    I'm building the world's simplest library application. All I want to be able to do is scan in a book's UPC (barcode) using a typical scanner (which just types the numbers of the barcode into a field) and then use it to look up data about the book... at a minimum, title, author, year published, and either the Dewey Decimal or Library of Congress catalog number. The goal is to print out a tiny sticker ("spine label") with the card catalog number that I can stick on the spine of the book, and then I can sort the books by card catalog number on the shelves in our company library. That way books on similar subjects will tend to be near each other, for example, if you know you're looking for a book about accounting, all you have to do is find SOME book about accounting and you'll see the other half dozen that we have right next to it which makes it convenient to browse the library. There seem to be lots of web APIs to do this, including Amazon and the Library of Congress. But those are all extremely confusing to me. What I really just want is a single higher level function that takes a UPC barcode number and returns some basic data about the book.

    Read the article

  • Creating a RESTful API - HELP!

    - by Martin Cox
    Hi Chaps Over the last few weeks I've been learning about iOS development, which has naturally led me into the world of APIs. Now, searching around on the Internet, I've come to the conclusion that using the REST architecture is very much recommended - due to it's supposed simplicity and ease of implementation. However, I'm really struggling with the implementation side of REST. I understand the concept; using HTTP methods as verbs to describe the action of a request and responding with suitable response codes, and so on. It's just, I don't understand how to code it. I don't get how I map a URI to an object. I understand that a GET request for domain.com/api/user/address?user_id=999 would return the address of user 999 - but I don't understand where or how that mapping from /user/address to some method that queries a database has taken place. Is this all coded in one php script? Would I just have a method that grabs the URI like so: $array = explode("/", ltrim(rtrim($_SERVER['REQUEST_URI'], "/"), "/")) And then cycle through that array, first I would have a request for a "user", so the PHP script would direct my request to the user object and then invoke the address method. Is that what actually happens? I've probably not explained my thought process very well there. The main thing I'm not getting is how that URI /user/address?id=999 somehow is broken down and executed - does it actually resolve to code? class user(id) { address() { //get user address } } I doubt I'm making sense now, so I'll call it a day trying to explain further. I hope someone out there can understand what I'm trying to say! Thanks Chaps, look forward to your responses. Martin p.s - I'm not a developer yet, I'm learning :)

    Read the article

  • Deploying a Rails app on an Ubuntu server using Git

    - by NudeCanalTroll
    I'm completely new to Linux, but today I find myself setting up a server (Ubuntu 10.04 LTS lucid) from scratch to host a Rails application. Anyway, I managed to get a Rails app up and running on the server itself, but I had to scrap that because I want to use Git. So I setup a git repository on the server, then pushed all the code from my local machine to the repository. Buuuut, of course Git doesn't actually store the files themselves in the repository -- all the code for my Rails app is now only on my local machine. How am I supposed to tell the server to host that? Right now my solution is to have the server use git to pull the code from its own repository. That's the code I'll host for all the world to see. In order to update the code, I guess I'll have to do something like this: Update the code on my local machine. Do some git adds, git commits, and a git push. On the server, do a git pull to update the code. So my question is, am I doing this the right way? enter code here

    Read the article

  • How can I invoke a .Net DLL from a LabView 6.1 VI?

    - by tw1k
    I work in a manufacturing company that uses LabView for testing the devices we make. Most of the test engineers are using 7.1 which can natively reference a .Net assembly. However, there is a group that is stuck on LabView 6.1. I would like for them to be able to use my .Net assembly which is basically a proxy to some web services. I have created a test assembly that is nothing more than Hello World, and I'm trying to consume it in a VI. I made it COM visible, and registered it with regasm.exe and created a type library, which I'm not sure I need. I can see it in Visual Studio in the list of COM objects when I open the Add Reference window, so I know it's registered properly. I'm very unfamiliar with VI's. I'm only looking at it because no one I have spoken to in manufacturing knows anything about invoking a COM object in a VI. I'm basically looking for some names of controls or menu options to get the test engineers pointed in the right direction. I did a bunch of web searching on Google and the NI forums, but didn't find much. Alternatively, would it be easier to write a C or C++ DLL that acts as a proxy to my .Net DLL? Or is there a simple way to invoke a web service from a VI? That might obviate the need for a DLL altogether. I'm currently reading through this document from NI for help, but it obviously knows nothing about .Net and might not be able to help me choose the best path forward.

    Read the article

  • starting a windows executable via batch script, exe not in Program Files

    - by Anthony
    This is probably batch scripting 101, but I can't find any clear explanation/documentation on why this is happening or if my workaround is actually the solution. So basically any terminology or links to good sources is really appreciated. So I have a program I want to execute via batch script (along with several other programs). It's the only one where the exe is not in a Program Files folder. I can get it to start like this: C:\WeirdProgram\WeirdProgramModule\weirdmodule.exe But I get an error along the lines of: Run-time Error '3024': Could not find file C:\Users\MyUserName\Desktop\ModuleSettings.mdb So it seems that the program is looking for its settings files from the same location that the batch script starts up. Given that I finally got everything to work by doing the following: cd C:\WeirdProgram\WeirdProgramModule\ weirdmodule.exe That works fine, and it's not the end of the world to have to go this route (just one extra line), but I've convinced myself that I'm doing something wrong based on lack of basic understanding. Anybody know or can point me to why it works this way? Oh, and doing the following: start "C:\WeirdProgram\WeirdProgramModule\weirdmodule.exe" doesn't do anything at all. Thanks,

    Read the article

< Previous Page | 441 442 443 444 445 446 447 448 449 450 451 452  | Next Page >