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  • BAM Data Control in multiple ADF Faces Components

    - by [email protected]
    As we know Oracle BAM data control instance sharing is not supported.When two or more ADF Faces components must display the same data, and are bound to the same Oracle BAM data control definition, we have to make sure that we wrap each ADF Faces component in an ADF task flow, and set the Data Control Scope to isolated. This blog will show a small sample to demonstrate this. In this sample we will create a Pie and Bar using same BAM DC, such that both components use same Data control but have isolated scope.This sample can be downloaded  fromSample1.zip Set-up: Create a BAM data control using employees DO (sample) Steps: Right click on View Controller project and select "New->ADF Task Flow" Check "Create Bounded Task Flow" and give some meaningful name (ex:EmpPieTF.xml ) to the TaskFlow(TF) and click on "OK"CreateTF.bmpFrom the "Components Palette", drag and drop "View" into the task flow diagram. Give a meaningful name to the view. Double Click and Click "Ok" for  "Create New JSF Page Fragment" From "Data Controls" drag and drop "Employees->Query"  into this jsff page as "Graph->Pie" (Pie: Sales_Number and Slices: Salesperson) Repeat step 1 through 4 for another Task Flow (ex: EmpBarTF). From "Data Controls" drag and drop "Employees->Query"  into this jsff page as "Graph->Bar" (Bars :Sales_Number and X-axis : Salesperson). Open the Taskflow created in step 2. In the Structure Pane, right click on "Task Flow Definition -EmpPieTF" Click "Insert inside Task Flow Definition - EmpPieTF -> ADF Task Flow -> Data Control Scope". Click "OK"TFDCScope.bmpFor the "Data Control Scope", In the Property Inspector ->General section, change data control scope from Shared to Isolated. Repeat step 8 through 11 for the 2nd Task flow created. Now create a new jspx page example: Main.jspxDrag and drop both the Task flows (ex: "EmpPieTF" and "EmpBarTF") as regions. Surround with panel components as needed.Run the page Main.jspxMainPage.bmpNow when the page runs although both components are created using same Data control the bindings are not shared and each component will have a separate instance of the data control.

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  • Problems with Level Architect, Citrus Engine, Flash

    - by Idan
    I am using the Citrus Engine to make a Flash game, and the Level Architect doesn't work well for me. Firstly, when I first launch it and open my project and my level, nothing is shown, no assets and not anything I have previously done with my level. To fix it, I open another project. The other project works fine, meaning I can see the assets and the level. Then I go back to the actual project I am working on, and the problem is fixed, only it does not fix the second problem: I can't add my own assests. I follow the manual and add tags like this: [Property(value="0")] But it doesn't change a thing in the level architect window (even after I close and reopen it). Any ideas? Thanks! Here's the code of the class I want to be shown in the Level Architect: package { import com.citrusengine.objects.PhysicsObject; import com.citrusengine.objects.platformer.Sensor; import flash.utils.clearTimeout; import flash.utils.setTimeout; /** * @author Aymeric */ public class Teleporter extends Sensor { [Property(value="0")] public var endX:Number=0; [Property(value="0")] public var endY:Number=0; public var object:PhysicsObject; [Property(value="0")] public var time:Number = 0; public var needToTeleport:Boolean = false; protected var _teleporting:Boolean = false; private var _teleportTimeoutID:uint; public function Teleporter(name:String, params:Object = null) { super(name, params); } override public function destroy():void { clearTimeout(_teleportTimeoutID); super.destroy(); } override public function update(timeDelta:Number):void { super.update(timeDelta); if (needToTeleport) { _teleporting = true; _teleportTimeoutID = setTimeout(_teleport, time); needToTeleport = false; } _updateAnimation(); } protected function _teleport():void { _teleporting = false; object.x = endX; object.y = endY; clearTimeout(_teleportTimeoutID); } protected function _updateAnimation():void { if (_teleporting) { _animation = "teleport"; } else { _animation = "normal"; } } } }

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  • Slow Firefox Javascript Canvas Performance?

    - by jujumbura
    As a followup from a previous post, I have been trying to track down some slowdown I am having when drawing a scene using Javascript and the canvas element. I decided to narrow down my focus to a REALLY barebones animation that only clears the canvas and draws a single image, once per-frame. This of course runs silky smooth in Chrome, but it still stutters in Firefox. I added a simple FPS calculator, and indeed it appears that my page is typically getting an FPS in the 50's when running Firefox. This doesn't seem right to me, I must be doing something wrong here. Can anybody see anything I might be doing that is causing this drop in FPS? <!DOCTYPE HTML> <html> <head> </head> <body bgcolor=silver> <canvas id="myCanvas" width="600" height="400"></canvas> <img id="myHexagon" src="Images/Hexagon.png" style="display: none;"> <script> window.requestAnimFrame = (function(callback) { return window.requestAnimationFrame || window.webkitRequestAnimationFrame || window.mozRequestAnimationFrame || window.oRequestAnimationFrame || window.msRequestAnimationFrame || function(callback) { window.setTimeout(callback, 1000 / 60); }; })(); var animX = 0; var frameCounter = 0; var fps = 0; var time = new Date(); function animate() { var canvas = document.getElementById("myCanvas"); var context = canvas.getContext("2d"); context.clearRect(0, 0, canvas.width, canvas.height); animX += 1; if (animX == canvas.width) { animX = 0; } var image = document.getElementById("myHexagon"); context.drawImage(image, animX, 128); context.lineWidth=1; context.fillStyle="#000000"; context.lineStyle="#ffffff"; context.font="18px sans-serif"; context.fillText("fps: " + fps, 20, 20); ++frameCounter; var currentTime = new Date(); var elapsedTimeMS = currentTime - time; if (elapsedTimeMS >= 1000) { fps = frameCounter; frameCounter = 0; time = currentTime; } // request new frame requestAnimFrame(function() { animate(); }); } window.onload = function() { animate(); }; </script> </body> </html>

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  • Creating metadata value relationships

    - by kyle.hatlestad
    I was recently asked an question about an interesting use case. They wanted content to be submitted into UCM with a particular ID in a custom metadata field. But they wanted that ID to be translated during submission into an employee name in another metadata field upon submission. My initial thought was that this could be done with a dependent choice list (DCL). One option list field driving the choices in another. But this didn't work in this case for a couple of reasons. First, the number of IDs could potentially be very large. So making that into a drop-down list would not be practical. The preference would be for that field to simply be a text field to type in the ID. Secondly, data could be submitted through different methods other then the web-based check-in form. And without an interface to select the DCL choices, the system needed a way to determine and populate the name field. So instead I went the approach of having the value of the ID field drive the value of the Name field using the derived field approach in my rule. In looking at it though, it was easy to simply copy the value of the ID field into the Name field...but to have it look up and translate the value proved to be the tricky part. So here is the approach I took... First I created my two metadata fields as standard text fields in the Configuration Manager applet. Next I create a table that stores the relationship between the IDs and Names. I then create a View into that table and set the column to the EmployeeID. I now create a new Application Field and set it as an option list using the View I created in the previous step. The reason I create it as an Application field is because I don't need to display the field or store a value in it. I simply need to make use of the option list in the next step... Finally, I create a Rule in which I select the Employee Name field and turn on the 'Is derived field' checkbox. I edit the derived value and add a new condition. Because the option list is a Application field and not an Information field, I can't use the Compute button. Instead, I insert this line directly in the Value field: @getFieldViewValue("EmployeeMapping",#active.xEmployeeID, "EmployeeName") The "EmployeeMapping" parameter designates that the value should be pulled from the EmployeeMapping Application field that I had created in the previous step. The #active.xEmployeeID field is the ID value that should be pulled from what the user entered. "EmployeeName" is the column name in the table which has the value which corresponds to the ID. The extracted name then becomes the value within our Employee Name field. That's it. You can then add additional Rules to make the Name field read-only/hidden on the check-in page and such.

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  • C#/.NET Little Wonders: The Generic Func Delegates

    - by James Michael Hare
    Once again, in this series of posts I look at the parts of the .NET Framework that may seem trivial, but can help improve your code by making it easier to write and maintain. The index of all my past little wonders posts can be found here. Back in one of my three original “Little Wonders” Trilogy of posts, I had listed generic delegates as one of the Little Wonders of .NET.  Later, someone posted a comment saying said that they would love more detail on the generic delegates and their uses, since my original entry just scratched the surface of them. Last week, I began our look at some of the handy generic delegates built into .NET with a description of delegates in general, and the Action family of delegates.  For this week, I’ll launch into a look at the Func family of generic delegates and how they can be used to support generic, reusable algorithms and classes. Quick Delegate Recap Delegates are similar to function pointers in C++ in that they allow you to store a reference to a method.  They can store references to either static or instance methods, and can actually be used to chain several methods together in one delegate. Delegates are very type-safe and can be satisfied with any standard method, anonymous method, or a lambda expression.  They can also be null as well (refers to no method), so care should be taken to make sure that the delegate is not null before you invoke it. Delegates are defined using the keyword delegate, where the delegate’s type name is placed where you would typically place the method name: 1: // This delegate matches any method that takes string, returns nothing 2: public delegate void Log(string message); This delegate defines a delegate type named Log that can be used to store references to any method(s) that satisfies its signature (whether instance, static, lambda expression, etc.). Delegate instances then can be assigned zero (null) or more methods using the operator = which replaces the existing delegate chain, or by using the operator += which adds a method to the end of a delegate chain: 1: // creates a delegate instance named currentLogger defaulted to Console.WriteLine (static method) 2: Log currentLogger = Console.Out.WriteLine; 3:  4: // invokes the delegate, which writes to the console out 5: currentLogger("Hi Standard Out!"); 6:  7: // append a delegate to Console.Error.WriteLine to go to std error 8: currentLogger += Console.Error.WriteLine; 9:  10: // invokes the delegate chain and writes message to std out and std err 11: currentLogger("Hi Standard Out and Error!"); While delegates give us a lot of power, it can be cumbersome to re-create fairly standard delegate definitions repeatedly, for this purpose the generic delegates were introduced in various stages in .NET.  These support various method types with particular signatures. Note: a caveat with generic delegates is that while they can support multiple parameters, they do not match methods that contains ref or out parameters. If you want to a delegate to represent methods that takes ref or out parameters, you will need to create a custom delegate. We’ve got the Func… delegates Just like it’s cousin, the Action delegate family, the Func delegate family gives us a lot of power to use generic delegates to make classes and algorithms more generic.  Using them keeps us from having to define a new delegate type when need to make a class or algorithm generic. Remember that the point of the Action delegate family was to be able to perform an “action” on an item, with no return results.  Thus Action delegates can be used to represent most methods that take 0 to 16 arguments but return void.  You can assign a method The Func delegate family was introduced in .NET 3.5 with the advent of LINQ, and gives us the power to define a function that can be called on 0 to 16 arguments and returns a result.  Thus, the main difference between Action and Func, from a delegate perspective, is that Actions return nothing, but Funcs return a result. The Func family of delegates have signatures as follows: Func<TResult> – matches a method that takes no arguments, and returns value of type TResult. Func<T, TResult> – matches a method that takes an argument of type T, and returns value of type TResult. Func<T1, T2, TResult> – matches a method that takes arguments of type T1 and T2, and returns value of type TResult. Func<T1, T2, …, TResult> – and so on up to 16 arguments, and returns value of type TResult. These are handy because they quickly allow you to be able to specify that a method or class you design will perform a function to produce a result as long as the method you specify meets the signature. For example, let’s say you were designing a generic aggregator, and you wanted to allow the user to define how the values will be aggregated into the result (i.e. Sum, Min, Max, etc…).  To do this, we would ask the user of our class to pass in a method that would take the current total, the next value, and produce a new total.  A class like this could look like: 1: public sealed class Aggregator<TValue, TResult> 2: { 3: // holds method that takes previous result, combines with next value, creates new result 4: private Func<TResult, TValue, TResult> _aggregationMethod; 5:  6: // gets or sets the current result of aggregation 7: public TResult Result { get; private set; } 8:  9: // construct the aggregator given the method to use to aggregate values 10: public Aggregator(Func<TResult, TValue, TResult> aggregationMethod = null) 11: { 12: if (aggregationMethod == null) throw new ArgumentNullException("aggregationMethod"); 13:  14: _aggregationMethod = aggregationMethod; 15: } 16:  17: // method to add next value 18: public void Aggregate(TValue nextValue) 19: { 20: // performs the aggregation method function on the current result and next and sets to current result 21: Result = _aggregationMethod(Result, nextValue); 22: } 23: } Of course, LINQ already has an Aggregate extension method, but that works on a sequence of IEnumerable<T>, whereas this is designed to work more with aggregating single results over time (such as keeping track of a max response time for a service). We could then use this generic aggregator to find the sum of a series of values over time, or the max of a series of values over time (among other things): 1: // creates an aggregator that adds the next to the total to sum the values 2: var sumAggregator = new Aggregator<int, int>((total, next) => total + next); 3:  4: // creates an aggregator (using static method) that returns the max of previous result and next 5: var maxAggregator = new Aggregator<int, int>(Math.Max); So, if we were timing the response time of a web method every time it was called, we could pass that response time to both of these aggregators to get an idea of the total time spent in that web method, and the max time spent in any one call to the web method: 1: // total will be 13 and max 13 2: int responseTime = 13; 3: sumAggregator.Aggregate(responseTime); 4: maxAggregator.Aggregate(responseTime); 5:  6: // total will be 20 and max still 13 7: responseTime = 7; 8: sumAggregator.Aggregate(responseTime); 9: maxAggregator.Aggregate(responseTime); 10:  11: // total will be 40 and max now 20 12: responseTime = 20; 13: sumAggregator.Aggregate(responseTime); 14: maxAggregator.Aggregate(responseTime); The Func delegate family is useful for making generic algorithms and classes, and in particular allows the caller of the method or user of the class to specify a function to be performed in order to generate a result. What is the result of a Func delegate chain? If you remember, we said earlier that you can assign multiple methods to a delegate by using the += operator to chain them.  So how does this affect delegates such as Func that return a value, when applied to something like the code below? 1: Func<int, int, int> combo = null; 2:  3: // What if we wanted to aggregate the sum and max together? 4: combo += (total, next) => total + next; 5: combo += Math.Max; 6:  7: // what is the result? 8: var comboAggregator = new Aggregator<int, int>(combo); Well, in .NET if you chain multiple methods in a delegate, they will all get invoked, but the result of the delegate is the result of the last method invoked in the chain.  Thus, this aggregator would always result in the Math.Max() result.  The other chained method (the sum) gets executed first, but it’s result is thrown away: 1: // result is 13 2: int responseTime = 13; 3: comboAggregator.Aggregate(responseTime); 4:  5: // result is still 13 6: responseTime = 7; 7: comboAggregator.Aggregate(responseTime); 8:  9: // result is now 20 10: responseTime = 20; 11: comboAggregator.Aggregate(responseTime); So remember, you can chain multiple Func (or other delegates that return values) together, but if you do so you will only get the last executed result. Func delegates and co-variance/contra-variance in .NET 4.0 Just like the Action delegate, as of .NET 4.0, the Func delegate family is contra-variant on its arguments.  In addition, it is co-variant on its return type.  To support this, in .NET 4.0 the signatures of the Func delegates changed to: Func<out TResult> – matches a method that takes no arguments, and returns value of type TResult (or a more derived type). Func<in T, out TResult> – matches a method that takes an argument of type T (or a less derived type), and returns value of type TResult(or a more derived type). Func<in T1, in T2, out TResult> – matches a method that takes arguments of type T1 and T2 (or less derived types), and returns value of type TResult (or a more derived type). Func<in T1, in T2, …, out TResult> – and so on up to 16 arguments, and returns value of type TResult (or a more derived type). Notice the addition of the in and out keywords before each of the generic type placeholders.  As we saw last week, the in keyword is used to specify that a generic type can be contra-variant -- it can match the given type or a type that is less derived.  However, the out keyword, is used to specify that a generic type can be co-variant -- it can match the given type or a type that is more derived. On contra-variance, if you are saying you need an function that will accept a string, you can just as easily give it an function that accepts an object.  In other words, if you say “give me an function that will process dogs”, I could pass you a method that will process any animal, because all dogs are animals.  On the co-variance side, if you are saying you need a function that returns an object, you can just as easily pass it a function that returns a string because any string returned from the given method can be accepted by a delegate expecting an object result, since string is more derived.  Once again, in other words, if you say “give me a method that creates an animal”, I can pass you a method that will create a dog, because all dogs are animals. It really all makes sense, you can pass a more specific thing to a less specific parameter, and you can return a more specific thing as a less specific result.  In other words, pay attention to the direction the item travels (parameters go in, results come out).  Keeping that in mind, you can always pass more specific things in and return more specific things out. For example, in the code below, we have a method that takes a Func<object> to generate an object, but we can pass it a Func<string> because the return type of object can obviously accept a return value of string as well: 1: // since Func<object> is co-variant, this will access Func<string>, etc... 2: public static string Sequence(int count, Func<object> generator) 3: { 4: var builder = new StringBuilder(); 5:  6: for (int i=0; i<count; i++) 7: { 8: object value = generator(); 9: builder.Append(value); 10: } 11:  12: return builder.ToString(); 13: } Even though the method above takes a Func<object>, we can pass a Func<string> because the TResult type placeholder is co-variant and accepts types that are more derived as well: 1: // delegate that's typed to return string. 2: Func<string> stringGenerator = () => DateTime.Now.ToString(); 3:  4: // This will work in .NET 4.0, but not in previous versions 5: Sequence(100, stringGenerator); Previous versions of .NET implemented some forms of co-variance and contra-variance before, but .NET 4.0 goes one step further and allows you to pass or assign an Func<A, BResult> to a Func<Y, ZResult> as long as A is less derived (or same) as Y, and BResult is more derived (or same) as ZResult. Sidebar: The Func and the Predicate A method that takes one argument and returns a bool is generally thought of as a predicate.  Predicates are used to examine an item and determine whether that item satisfies a particular condition.  Predicates are typically unary, but you may also have binary and other predicates as well. Predicates are often used to filter results, such as in the LINQ Where() extension method: 1: var numbers = new[] { 1, 2, 4, 13, 8, 10, 27 }; 2:  3: // call Where() using a predicate which determines if the number is even 4: var evens = numbers.Where(num => num % 2 == 0); As of .NET 3.5, predicates are typically represented as Func<T, bool> where T is the type of the item to examine.  Previous to .NET 3.5, there was a Predicate<T> type that tended to be used (which we’ll discuss next week) and is still supported, but most developers recommend using Func<T, bool> now, as it prevents confusion with overloads that accept unary predicates and binary predicates, etc.: 1: // this seems more confusing as an overload set, because of Predicate vs Func 2: public static SomeMethod(Predicate<int> unaryPredicate) { } 3: public static SomeMethod(Func<int, int, bool> binaryPredicate) { } 4:  5: // this seems more consistent as an overload set, since just uses Func 6: public static SomeMethod(Func<int, bool> unaryPredicate) { } 7: public static SomeMethod(Func<int, int, bool> binaryPredicate) { } Also, even though Predicate<T> and Func<T, bool> match the same signatures, they are separate types!  Thus you cannot assign a Predicate<T> instance to a Func<T, bool> instance and vice versa: 1: // the same method, lambda expression, etc can be assigned to both 2: Predicate<int> isEven = i => (i % 2) == 0; 3: Func<int, bool> alsoIsEven = i => (i % 2) == 0; 4:  5: // but the delegate instances cannot be directly assigned, strongly typed! 6: // ERROR: cannot convert type... 7: isEven = alsoIsEven; 8:  9: // however, you can assign by wrapping in a new instance: 10: isEven = new Predicate<int>(alsoIsEven); 11: alsoIsEven = new Func<int, bool>(isEven); So, the general advice that seems to come from most developers is that Predicate<T> is still supported, but we should use Func<T, bool> for consistency in .NET 3.5 and above. Sidebar: Func as a Generator for Unit Testing One area of difficulty in unit testing can be unit testing code that is based on time of day.  We’d still want to unit test our code to make sure the logic is accurate, but we don’t want the results of our unit tests to be dependent on the time they are run. One way (of many) around this is to create an internal generator that will produce the “current” time of day.  This would default to returning result from DateTime.Now (or some other method), but we could inject specific times for our unit testing.  Generators are typically methods that return (generate) a value for use in a class/method. For example, say we are creating a CacheItem<T> class that represents an item in the cache, and we want to make sure the item shows as expired if the age is more than 30 seconds.  Such a class could look like: 1: // responsible for maintaining an item of type T in the cache 2: public sealed class CacheItem<T> 3: { 4: // helper method that returns the current time 5: private static Func<DateTime> _timeGenerator = () => DateTime.Now; 6:  7: // allows internal access to the time generator 8: internal static Func<DateTime> TimeGenerator 9: { 10: get { return _timeGenerator; } 11: set { _timeGenerator = value; } 12: } 13:  14: // time the item was cached 15: public DateTime CachedTime { get; private set; } 16:  17: // the item cached 18: public T Value { get; private set; } 19:  20: // item is expired if older than 30 seconds 21: public bool IsExpired 22: { 23: get { return _timeGenerator() - CachedTime > TimeSpan.FromSeconds(30.0); } 24: } 25:  26: // creates the new cached item, setting cached time to "current" time 27: public CacheItem(T value) 28: { 29: Value = value; 30: CachedTime = _timeGenerator(); 31: } 32: } Then, we can use this construct to unit test our CacheItem<T> without any time dependencies: 1: var baseTime = DateTime.Now; 2:  3: // start with current time stored above (so doesn't drift) 4: CacheItem<int>.TimeGenerator = () => baseTime; 5:  6: var target = new CacheItem<int>(13); 7:  8: // now add 15 seconds, should still be non-expired 9: CacheItem<int>.TimeGenerator = () => baseTime.AddSeconds(15); 10:  11: Assert.IsFalse(target.IsExpired); 12:  13: // now add 31 seconds, should now be expired 14: CacheItem<int>.TimeGenerator = () => baseTime.AddSeconds(31); 15:  16: Assert.IsTrue(target.IsExpired); Now we can unit test for 1 second before, 1 second after, 1 millisecond before, 1 day after, etc.  Func delegates can be a handy tool for this type of value generation to support more testable code.  Summary Generic delegates give us a lot of power to make truly generic algorithms and classes.  The Func family of delegates is a great way to be able to specify functions to calculate a result based on 0-16 arguments.  Stay tuned in the weeks that follow for other generic delegates in the .NET Framework!   Tweet Technorati Tags: .NET, C#, CSharp, Little Wonders, Generics, Func, Delegates

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  • Wordpress Installation (on IIS and SQL Server)

    - by Davide Mauri
    To proceed with the installation of Wordpress on SQL Server and IIS, first of all, you need to do the following steps Create a database on SQL Server that will be used by Wordpress Create login that can access to the just created database and put the user into ddladmin, db_datareader, db_datawriter roles Download and unpack Wordpress 3.3.2 (latest version as of 27 May 2012) zip file into a directory of your choice Download the wp-db-abstraction 1.1.4 (latest version as of 27 May 2012) plugin from wordpress.org website Now that the basic action has been done, you can start to setup and configure your Wordpress installation. Unpack and follow the instructions in the README.TXT file to install the Database Abstraction Layer. Mainly you have to: Upload wp-db-abstraction.php and the wp-db-abstraction directory to wp-content/mu-plugins.  This should be parallel to your regular plugins directory.  If the mu-plugins directory does not exist, you must create it. Put the db.php file from inside the wp-db-abstraction.php directory to wp-content/db.php Now you can create an application pool in IIS like the following one Create a website, using the above Application Pool, that points to the folder where you unpacked Wordpress files. Be sure to give the “Write” permission to the IIS account, as pointed out in this (old, but still quite valid) installation manual: http://wordpress.visitmix.com/development/installing-wordpress-on-sql-server#iis Now you’re ready to go. Point your browser to the configured website and the Wordpress installation screen will be there for you. When you’re requested to enter information to connect to MySQL database, simply skip that page, leaving the default values. If you have installed the Database Abstraction Layer, another database installation screen will appear after the one used by MySQL, and here you can enter the configuration information needed to connect to SQL Server. After having finished the installation steps, you should be able to access and navigate your wordpress site.  A final touch, and it’s done: just add the needed rewrite rules http://wordpress.visitmix.com/development/installing-wordpress-on-sql-server#urlrewrite and that’s it! Well. Not really. Unfortunately the current (as of 27 May 2012) version of the Database Abstraction Layer (1.1.4) has some bugs. Luckily they can be quickly fixed: Backslash Fix http://wordpress.org/support/topic/plugin-wp-db-abstraction-fix-problems-with-backslash-usage Select Top 0 Fix Make the change to the file “.\wp-content\mu-plugins\wp-db-abstraction\translations\sqlsrv\translations.php” suggested by “debettap”   http://sourceforge.net/tracker/?func=detail&aid=3485384&group_id=315685&atid=1328061 And now you have a 100% working Wordpress installation on SQL Server! Since I also wanted to take advantage of SQL Server Full Text Search, I’ve created a very simple wordpress plugin to setup full-text search and to use it as website search engine: http://wpfts.codeplex.com/ Enjoy!

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  • Returning null vs Throwing exceptions

    - by Svish
    Is in a bit of disagreement with a more experienced developer on this issue, and was wondering what you guys here think about this. Environment is Java, EJB 3, services, etc. The code I wrote calls a service to get things and to create things. Problem was that I got null pointer exceptions in places that didn't make sense. For example when I asked the service to create an object, I got null back. And when I tried to look up an object with an id I knew existed, I still got null back. Was like it was ignoring me. Spent some time trying to figure out what was wrong in my code (since I'm less experienced I usually assume I have messed up). Turns out the reason was security. If the user principal using my service didn't have the right permissions to use the service I called from my service, then that service simply returned null. The services that are here already are usually not documented either, so this is just something you have to know... somehow... So here is the thing: I mean that this is rather confusing as a developer interacting with this service. To me it would make much more sense if that service thew an exception which would tell me that hey, you don't have the proper permissions to get info about this thing or to create this new thing. I would then immediately know why my service wasn't working as expected. However, he argued that asking is not wrong. Exceptions should only be thrown when there is an error and asking for a thing is not an error. Even if you don't have permission to "see" that the thing you asked for. The things are often looked up in a GUI by users and for those users not having the right permissions, these things simply "do not exist". So, in short: Asking is not wrong, hence no exception. Get methods return null because to those users those things "doesn't exist". Create methods return null because nothing was created, since the user wasn't allowed to create anything. So, what do you guys think? Is this normal and/or good practice? I prefer exceptions as I prefer throwing and catching exceptions because I find it much easier to know what's going on. So I would for example also prefer to throw a NotFoundException if you asked for an id which didn't exist, rather than returning null. Anyways, just curious to what others think about this as I'm not the most experienced developer yet.

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  • Wordpress Installation (on IIS and SQL Server)

    - by Davide Mauri
    To proceed with the installation of Wordpress on SQL Server and IIS, first of all, you need to do the following steps Create a database on SQL Server that will be used by Wordpress Create login that can access to the just created database and put the user into ddladmin, db_datareader, db_datawriter roles Download and unpack Wordpress 3.3.2 (latest version as of 27 May 2012) zip file into a directory of your choice Download the wp-db-abstraction 1.1.4 (latest version as of 27 May 2012) plugin from wordpress.org website Now that the basic action has been done, you can start to setup and configure your Wordpress installation. Unpack and follow the instructions in the README.TXT file to install the Database Abstraction Layer. Mainly you have to: Upload wp-db-abstraction.php and the wp-db-abstraction directory to wp-content/mu-plugins.  This should be parallel to your regular plugins directory.  If the mu-plugins directory does not exist, you must create it. Put the db.php file from inside the wp-db-abstraction.php directory to wp-content/db.php Now you can create an application pool in IIS like the following one Create a website, using the above Application Pool, that points to the folder where you unpacked Wordpress files. Be sure to give the “Write” permission to the IIS account, as pointed out in this (old, but still quite valid) installation manual: http://wordpress.visitmix.com/development/installing-wordpress-on-sql-server#iis Now you’re ready to go. Point your browser to the configured website and the Wordpress installation screen will be there for you. When you’re requested to enter information to connect to MySQL database, simply skip that page, leaving the default values. If you have installed the Database Abstraction Layer, another database installation screen will appear after the one used by MySQL, and here you can enter the configuration information needed to connect to SQL Server. After having finished the installation steps, you should be able to access and navigate your wordpress site.  A final touch, and it’s done: just add the needed rewrite rules http://wordpress.visitmix.com/development/installing-wordpress-on-sql-server#urlrewrite and that’s it! Well. Not really. Unfortunately the current (as of 27 May 2012) version of the Database Abstraction Layer (1.1.4) has some bugs. Luckily they can be quickly fixed: Backslash Fix http://wordpress.org/support/topic/plugin-wp-db-abstraction-fix-problems-with-backslash-usage Select Top 0 Fix Make the change to the file “.\wp-content\mu-plugins\wp-db-abstraction\translations\sqlsrv\translations.php” suggested by “debettap”   http://sourceforge.net/tracker/?func=detail&aid=3485384&group_id=315685&atid=1328061 And now you have a 100% working Wordpress installation on SQL Server! Since I also wanted to take advantage of SQL Server Full Text Search, I’ve created a very simple wordpress plugin to setup full-text search and to use it as website search engine: http://wpfts.codeplex.com/ Enjoy!

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  • ODI 11g - Cleaning control characters and User Functions

    - by David Allan
    In ODI user functions have a poor name really, they should be user expressions - a way of wrapping common expressions that you may wish to reuse many times - across many different technologies is an added bonus. To illustrate look at the problem of how to remove control characters from text. Users ask these types of questions over all technologies - Microsoft SQL Server, Oracle, DB2 and for many years - how do I clean a string, how do I tokenize a string and so on. After some searching around you will find a few ways of doing this, in Oracle there is a convenient way of using the TRANSLATE and REPLACE functions. So you can convert some text using the following SQL; replace( translate('This is my string'||chr(9)||' which has a control character', chr(3)||chr(4)||chr(5)||chr(9), chr(3) ), chr(3), '' ) If you had many columns to perform this kind of transformation on, in the Oracle database the natural solution you'd go to would be to code this as a PLSQL function since you don't want the code splattered everywhere. Someone tells you that there is another control character that needs added equals a maintenance headache. Coding it as a PLSQL function will incur a context switch between SQL and PLSQL which could prove costly. In ODI user functions let you capture this expression text and reference it many times across your mappings. This will protect the expression from being copy-pasted by developers and make maintenance much simpler - change the expression definition in one place. Firstly define a name and a syntax for the user function, I am calling it UF_STRIP_BAD_CHARACTERS and it has one parameter an input string;  We then can define an implementation for each technology we will use it, I will define Oracle's using the inputString parameter and the TRANSLATE and REPLACE functions with whatever control characters I want to replace; I can then use this inside mapping expressions in ODI, below I am cleaning the ENAME column - a fabricated example but you get the gist.  Note when I use the user function the function name remains in the text of the mapping, the actual expression is not substituted until I generate the scenario. If you generate the scenario and export the scenario you can have a peak at the code that is processed in the runtime - below you can see a snippet of my export scenario;  That's all for now, hopefully a useful snippet of info.

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  • TomEE Integration in NetBeans Next

    - by Geertjan
    At JavaOne 2013, there was a lot of buzz around the TomEE server, e.g., many Tweets, nice party, and a new TomEE consulting company. For those tracking TomEE developments, it is interesting to note that recently the NetBeans IDE development builds have had added to them... TomEE support. Note: The TomEE support described here is not in NetBeans IDE 7.4, but in development builds for the next release of NetBeans IDE.For example, with NetBeans IDE development builds you're able to: register TomEE as a server in the Services window (TomEE has several distributions, e.g., one can use the "with JAX-RS" one, for example) create a Java EE 6 web project (e.g., Maven based) against this server create JPA entities from database create JAX-RS classes from JPA entities create JSF pages from JPA entities the IDE lets you create a new data source for TomEE and deploy it to the server the IDE figures out the components that are already packaged in TomEE, and the fact that (unlike with regular Tomcat), it does not need to package any components such as JSF implementation, persistence provider, or JAX-RS runtime, so that the resulting WAR file is very small the IDE can also do "deploy on save" with TomEE, so that your development cycle is very fast Adam Bien blogged about how he set up TomEE sometime ago, here. The official support in NetBeans IDE will be much more tightly integrated, simplifying the steps Adam describes. For example, the IDE does step 2 from Adam's blog for you, i.e., it sets up TomEE deployment roles. Moreover, it knows about all the technologies included in TomEE so that it can optimize the packaging; it knows about TomEE's persistence setup; it can work with TomEE data sources, etc. Below you see a Maven-based Java EE 6 PrimeFaces application (all entities and JSF pages generated from a database) deployed to TomEE in NetBeans IDE: And here's the management console for configuring and finetuning TomEE in NetBeans IDE: When I tried out the NetBeans IDE development build and TomEE, to see how everything fits together, I was surprised at how fast TomEE started up. Not sure what they did to it, but seems like a server on steroids. And setting it up in NetBeans IDE was trivial. Add the simple set up of TomEE in NetBeans IDE to the many benefits that the widely praised out of the box NetBeans Maven tools make possible, together with the fact that not one single plugin had to be installed to get everything you see described here up and running... and you have a really powerful combination of dev tools, all for free.

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  • New Slides - and a discussion about Dictionary Statistics

    - by Mike Dietrich
    First of all we have just upoaded a new version of the Upgrade and Migration Workshop slides with some added information. So please feel free to download them from here.The slides have one new interesting information which lead to a discussion I've had in the past days with a very large customer regarding their upgrades - and internally on the mailing list targeting an EBS database upgrade from Oracle 10.2 to Oracle 11.2. Why are we creating dictionary statistics during upgrade? I'd believe this forced dictionary statistics creation got introduced with the desupport of the Rule Based Optimizer in Oracle 10g. The goal: as RBO is not supported anymore we have to make sure that the data dictionary has fresh and non-stale statistics. Actually that would have led in Oracle 9i to strange behaviour in some databases - so in Oracle 9i this was strongly disrecommended. The upgrade scripts got hardcoded to create these stats. But during tests we had the following findings: It's important to create dictionary statistics the night before the upgrade. Not two weeks before, not 60 minutes before your downtime begins. But very close to the upgrade. From Oracle 10g onwards you'd just say: $ execute DBMS_STATS.GATHER_DICTIONARY_STATS; This is important to make sure you have fresh dictionary statistics during upgrade for performance reasons. Tests have shown that running an upgrade without valid dictionary statistics might slow down the whole upgrade by factors of 2x-3x. And it would be also a great idea post upgrade to create again fresh dictionary statistics when you've did suppress the stats creation during the upgrade process. Suppress? Yes, you could set this underscore parameter in the init.ora: _optim_dict_stats_at_db_cr_upg=FALSE to suppress the forced dictionary statistics collection during an upgrade. We believe strongly that (a) people using the default statistics creation process which will create dictionary statistics by default and (b) create fresh stats before upgrade on the dictionary. Therefore we find it save once you have followed our advice to use the underscore during upgrade. And we've taken out that forced statistics collection during upgrade in the next release of the database. Please note: If you are using the DBUA for the upgrade it will remove underscore parameters for the upgrade run to improve performance - which is generally a good idea. So you'll have to start the DBUA with that call: $ dbua -initParam "_optim_dict_stats_at_cb_cr_upg"=FALSE -Mike

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  • New Oracles VM RAC template with support for oracle vm 3 built-in

    - by wcoekaer
    The RAC team did it again (thanks Saar!) - another awesome set of Oracle VM templates published and uploaded to My Oracle Support. You can find the main page here. What's special about the latest version of DeployCluster is that it integrates tightly with Oracle VM 3 manager. It basically is an Oracle VM frontend that helps start VMs, pass arguments down automatically and there is absolutely no need to log into the Oracle VM servers or the guests. Once it completes, you have an entire Oracle RAC database setup ready to go. Here's a short summary of the steps : Set up an Oracle VM 3 server pool Download the Oracle VM RAC template from oracle.com Import the template into Oracle VM using Oracle VM Manager repository - import Create a public and private network in Oracle VM Manager in the network tab Configure the template with the right public and private virtual networks Create a set of shared disks (physical or virtual) to assign to the VMs you want to create (for ASM/at least 5) Clone a set of VMs from the template (as many RAC nodes as you plan to configure) With Oracle VM 3.1 you can clone with a number so one clone command for, say 8 VMs is easy. Assign the shared devices/disks to the cloned VMs Create a netconfig.ini file on your manager node or a client where you plan to run DeployCluster This little text file just contains the IP addresses, hostnames etc for your cluster. It is a very simple small textfile. Run deploycluster.py with the VM names as argument Done. At this point, the tool will connect to Oracle VM Manager, start the VMs and configure each one, Configure the OS (Oracle Linux) Configure the disks with ASM Configure the clusterware (CRS) Configure ASM Create database instances on each node. Now you are ready to log in, and use your x node database cluster. x No need to download various products from various websites, click on trial licenses for the OS, go to a Virtual Machine store with sample and test versions only - this is production ready and supported. Software. Complete. example netconfig.ini : # Node specific information NODE1=racnode1 NODE1VIP=racnode1-vip NODE1PRIV=racnode1-priv NODE1IP=192.168.1.2 NODE1VIPIP=192.168.1.22 NODE1PRIVIP=10.0.0.22 NODE2=racnode2 NODE2VIP=racnode2-vip NODE2PRIV=racnode2-priv NODE2IP=192.168.1.3 NODE2VIPIP=192.168.1.23 NODE2PRIVIP=10.0.0.23 # Common data PUBADAP=eth0 PUBMASK=255.255.255.0 PUBGW=192.168.1.1 PRIVADAP=eth1 PRIVMASK=255.255.255.0 RACCLUSTERNAME=raccluster DOMAINNAME=mydomain.com DNSIP= # Device used to transfer network information to second node # in interview mode NETCONFIG_DEV=/dev/xvdc # 11gR2 specific data SCANNAME=racnode12-scan SCANIP=192.168.1.50

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  • Domain Models (PHP)

    - by Calum Bulmer
    I have been programming in PHP for several years and have, in the past, adopted methods of my own to handle data within my applications. I have built my own MVC, in the past, and have a reasonable understanding of OOP within php but I know my implementation needs some serious work. In the past I have used an is-a relationship between a model and a database table. I now know after doing some research that this is not really the best way forward. As far as I understand it I should create models that don't really care about the underlying database (or whatever storage mechanism is to be used) but only care about their actions and their data. From this I have established that I can create models of lets say for example a Person an this person object could have some Children (human children) that are also Person objects held in an array (with addPerson and removePerson methods, accepting a Person object). I could then create a PersonMapper that I could use to get a Person with a specific 'id', or to save a Person. This could then lookup the relationship data in a lookup table and create the associated child objects for the Person that has been requested (if there are any) and likewise save the data in the lookup table on the save command. This is now pushing the limits to my knowledge..... What if I wanted to model a building with different levels and different rooms within those levels? What if I wanted to place some items in those rooms? Would I create a class for building, level, room and item with the following structure. building can have 1 or many level objects held in an array level can have 1 or many room objects held in an array room can have 1 or many item objects held in an array and mappers for each class with higher level mappers using the child mappers to populate the arrays (either on request of the top level object or lazy load on request) This seems to tightly couple the different objects albeit in one direction (ie. a floor does not need to be in a building but a building can have levels) Is this the correct way to go about things? Within the view I am wanting to show a building with an option to select a level and then show the level with an option to select a room etc.. but I may also want to show a tree like structure of items in the building and what level and room they are in. I hope this makes sense. I am just struggling with the concept of nesting objects within each other when the general concept of oop seems to be to separate things. If someone can help it would be really useful. Many thanks

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  • JavaScript Class Patterns &ndash; In CoffeeScript

    - by Liam McLennan
    Recently I wrote about JavaScript class patterns, and in particular, my favourite class pattern that uses closure to provide encapsulation. A class to represent a person, with a name and an age, looks like this: var Person = (function() { // private variables go here var name,age; function constructor(n, a) { name = n; age = a; } constructor.prototype = { toString: function() { return name + " is " + age + " years old."; } }; return constructor; })(); var john = new Person("John Galt", 50); console.log(john.toString()); Today I have been experimenting with coding for node.js in CoffeeScript. One of the first things I wanted to do was to try and implement my class pattern in CoffeeScript and then see how it compared to CoffeeScript’s built-in class keyword. The above Person class, implemented in CoffeeScript, looks like this: # JavaScript style class using closure to provide private methods Person = (() -> [name,age] = [{},{}] constructor = (n, a) -> [name,age] = [n,a] null constructor.prototype = toString: () -> "name is #{name} age is #{age} years old" constructor )() I am satisfied with how this came out, but there are a few nasty bits. To declare the two private variables in javascript is as simple as var name,age; but in CoffeeScript I have to assign a value, hence [name,age] = [{},{}]. The other major issue occurred because of CoffeeScript’s implicit function returns. The last statement in any function is returned, so I had to add null to the end of the constructor to get it to work. The great thing about the technique just presented is that it provides encapsulation ie the name and age variables are not visible outside of the Person class. CoffeeScript classes do not provide encapsulation, but they do provide nicer syntax. The Person class using native CoffeeScript classes is: # CoffeeScript style class using the class keyword class CoffeePerson constructor: (@name, @age) -> toString: () -> "name is #{@name} age is #{@age} years old" felix = new CoffeePerson "Felix Hoenikker", 63 console.log felix.toString() So now I have a trade-off: nice syntax against encapsulation. I think I will experiment with both strategies in my project and see which works out better.

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  • How to Enable JavaScript file API in IE8 [closed]

    - by saeed
    i have developed a web application in asp.net , there is a page in this project which user should choose a file in picture format (jpeg,jpg,bmp,...) and i want to preview image in the page but i don't want to post file to server i want to handle it in client i have done it with java scripts functions via file API but it only works in IE9 but most of costumers use IE8 the reason is that IE8 doesn't support file API is there any way to make IE8 upgrade or some patches in code behind i mean that check if the browser is IE and not support file API call a function which upgrades IE8 to IE9 automatically. i don't want to ask user to do it in message i want to do it programmatic !! even if it is possible install a special patch that is required for file API because customers thought it is a bug in my application and their computer knowledge is low what am i supposed to do with this? i also use Async File Upload Ajax Control But it post the file to server any way with ajax solution and http handler but java scripts do it all in client browser!!! following script checks the browser supports API or not <script> if (window.File && window.FileReader && window.FileList && window.Blob) document.write("<b>File API supported.</b>"); else document.write('<i>File API not supported by this browser.</i>'); </script> following scripts do the read and Load Image function readfile(e1) { var filename = e1.target.files[0]; var fr = new FileReader(); fr.onload = readerHandler; fr.readAsText(filename); } HTML code: <input type="file" id="getimage"> <fieldset><legend>Your image here</legend> <div id="imgstore"></div> </fieldset> JavaScript code: <script> function imageHandler(e2) { var store = document.getElementById('imgstore'); store.innerHTML='<img src="' + e2.target.result +'">'; } function loadimage(e1) { var filename = e1.target.files[0]; var fr = new FileReader(); fr.onload = imageHandler; fr.readAsDataURL(filename); } window.onload=function() { var x = document.getElementById("filebrowsed"); x.addEventListener('change', readfile, false); var y = document.getElementById("getimage"); y.addEventListener('change', loadimage, false); } </script>

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  • 2D water with dynamic waves

    - by user1103457
    New Super Mario Bros has really cool 2D water that I'd like to learn how to create. Here's a video showing it. When something hits the water, it creates a wave. There are also constant "background" waves. You can get a good look at the constant waves just after 00:50 when the camera isn't moving. I assume the splashes in NSMB work as in the first part of this tutorial. But in NSMB the water also has constant waves on the surface, and the splashes look very different. Another difference is that in the tutorial, if you create a splash, it first creates a deep "hole" in the water at the origin of the splash. In new super mario bros this hole is absent or much smaller. I am referring to the splashes that the player creates when jumping in and out of the water. How do they create the constant waves and the splashes? I am especially interested in the splashes, and how they work together with the constant waves. I am programming in XNA. I've tried this myself, but couldn't really get it all to work well together. Bonus questions: How do they create the light spots just under the surface of the waves and how do they texture the deeper parts of the water? This is the first time I try to create water like this. EDIT: I assume the constant waves are created using a sine function. The splashes are probably created in a way like in the tutorial. (But they are not the same, so I am still interested in how to make this kind of splashes) But I have a lot of trouble combining those things. I know I can use the sine function to set the height of a specific watercolumn but the splashes are using the speed, to determine the new height. I can't figure out how to combine those. Not that I am not asking how the developers of new super mario bros did this exactly. I am just interested in ways to recreate an effect like it. This week I have an examweek so I don't have time to work on the code. After this week I will spend a lot of time on it. But I am constantly thinking about it, so that's why I will be checking comments etc. I just won't be looking at the code since it might be too time-consuming.

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  • Issues with shooting in a HTML5 platformer game

    - by fnx
    I'm coding a 2D sidescroller using only JavaScript and HTML5 canvas, and in my game I have two problems with shooting: 1) Player shoots continous stream of bullets. I want that player can shoot only a single bullet even though the shoot-button is being held down. 2) Also, I get an error "Uncaught TypeError: Cannot call method 'draw' of undefined" when all the bullets are removed. My shooting code goes like this: When player shoots, I do game.bullets.push(new Bullet(this, this.scale)); and after that: function Bullet(source, dir) { this.id = "bullet"; this.width = 10; this.height = 3; this.dir = dir; if (this.dir == 1) { this.x = source.x + source.width - 5; this.y = source.y + 16; } if (this.dir == -1) { this.x = source.x; this.y = source.y + 16; } } Bullet.prototype.update = function() { if (this.dir == 1) this.x += 8; if (this.dir == -1) this.x -= 8; for (var i in game.enemies) { checkCollisions(this, game.enemies[i]); } // Check if bullet leaves the viewport if (this.x < game.viewX * 32 || this.x > (game.viewX + game.tilesX) * 32) { removeFromList(game.bullets, this); } } Bullet.prototype.draw = function() { // bullet flipping uses orientation of the player var posX = game.player.scale == 1 ? this.x : (this.x + this.width) * -1; game.ctx.scale(game.player.scale, 1); game.ctx.drawImage(gameData.getGfx("bullet"), posX, this.y); } I handle removing with this function: function removeFromList(list, object) { for (i in list) { if (object == list[i]) { list.splice(i, 1); break; } } } And finally, in the main game loop I have this: for (var i in game.bullets) { game.bullets[i].update(); game.bullets[i].draw(); } I have tried adding if (game.bullets.length > 0) to the main game loop before the above draw&update calls, but I still get the same error.

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  • How to remove the boundary effects arising due to zero padding in scipy/numpy fft?

    - by Omkar
    I have made a python code to smoothen a given signal using the Weierstrass transform, which is basically the convolution of a normalised gaussian with a signal. The code is as follows: #Importing relevant libraries from __future__ import division from scipy.signal import fftconvolve import numpy as np def smooth_func(sig, x, t= 0.002): N = len(x) x1 = x[-1] x0 = x[0] # defining a new array y which is symmetric around zero, to make the gaussian symmetric. y = np.linspace(-(x1-x0)/2, (x1-x0)/2, N) #gaussian centered around zero. gaus = np.exp(-y**(2)/t) #using fftconvolve to speed up the convolution; gaus.sum() is the normalization constant. return fftconvolve(sig, gaus/gaus.sum(), mode='same') If I run this code for say a step function, it smoothens the corner, but at the boundary it interprets another corner and smoothens that too, as a result giving unnecessary behaviour at the boundary. I explain this with a figure shown in the link below. Boundary effects This problem does not arise if we directly integrate to find convolution. Hence the problem is not in Weierstrass transform, and hence the problem is in the fftconvolve function of scipy. To understand why this problem arises we first need to understand the working of fftconvolve in scipy. The fftconvolve function basically uses the convolution theorem to speed up the computation. In short it says: convolution(int1,int2)=ifft(fft(int1)*fft(int2)) If we directly apply this theorem we dont get the desired result. To get the desired result we need to take the fft on a array double the size of max(int1,int2). But this leads to the undesired boundary effects. This is because in the fft code, if size(int) is greater than the size(over which to take fft) it zero pads the input and then takes the fft. This zero padding is exactly what is responsible for the undesired boundary effects. Can you suggest a way to remove this boundary effects? I have tried to remove it by a simple trick. After smoothening the function I am compairing the value of the smoothened signal with the original signal near the boundaries and if they dont match I replace the value of the smoothened func with the input signal at that point. It is as follows: i = 0 eps=1e-3 while abs(smooth[i]-sig[i])> eps: #compairing the signals on the left boundary smooth[i] = sig[i] i = i + 1 j = -1 while abs(smooth[j]-sig[j])> eps: # compairing on the right boundary. smooth[j] = sig[j] j = j - 1 There is a problem with this method, because of using an epsilon there are small jumps in the smoothened function, as shown below: jumps in the smooth func Can there be any changes made in the above method to solve this boundary problem?

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  • With AMD style modules in JavaScript is there any benefit to namespaces?

    - by gman
    Coming from C++ originally and seeing lots of Java programmers doing the same we brought namespaces to JavaScript. See Google's closure library as an example where they have a main namespace, goog and under that many more namespaces like goog.async, goog.graphics But now, having learned the AMD style of requiring modules it seems like namespaces are kind of pointless in JavaScript. Not only pointless but even arguably an anti-pattern. What is AMD? It's a way of defining and including modules that removes all direct dependencies. Effectively you do this // some/module.js define([ 'name/of/needed/module', 'name/of/someother/needed/module', ], function( RefToNeededModule, RefToSomeOtherNeededModule) { ...code... return object or function }); This format lets the AMD support code know that this module needs name/of/needed/module.js and name/of/someother/needed/module.js loaded. The AMD code can load all the modules and then, assuming no circular dependencies, call the define function on each module in the correct order, record the object/function returned by the module as it calls them, and then call any other modules' define function with references to those modules. This seems to remove any need for namespaces. In your own code you can call the reference to any other module anything you want. For example if you had 2 string libraries, even if they define similar functions, as long as they follow the AMD pattern you can easily use both in the same module. No need for namespaces to solve that. It also means there's no hard coded dependencies. For example in Google's closure any module could directly reference another module with something like var value = goog.math.someMathFunc(otherValue) and if you're unlucky it will magically work where as with AMD style you'd have to explicitly include the math library otherwise the module wouldn't have a reference to it since there are no globals with AMD. On top of that dependency injection for testing becomes easy. None of the code in the AMD module references things by namespace so there is no hardcoded namespace paths, you can easily mock classes at testing time. Is there any other point to namespaces or is that something that C++ / Java programmers are bringing to JavaScript that arguably doesn't really belong?

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  • 2D SAT Collision Detection not working when using certain polygons (With example)

    - by sFuller
    My SAT algorithm falsely reports that collision is occurring when using certain polygons. I believe this happens when using a polygon that does not contain a right angle. Here is a simple diagram of what is going wrong: Here is the problematic code: std::vector<vec2> axesB = polygonB->GetAxes(); //loop over axes B for(int i = 0; i < axesB.size(); i++) { float minA,minB,maxA,maxB; polygonA->Project(axesB[i],&minA,&maxA); polygonB->Project(axesB[i],&minB,&maxB); float intervalDistance = polygonA->GetIntervalDistance(minA, maxA, minB, maxB); if(intervalDistance >= 0) return false; //Collision not occurring } This function retrieves axes from the polygon: std::vector<vec2> Polygon::GetAxes() { std::vector<vec2> axes; for(int i = 0; i < verts.size(); i++) { vec2 a = verts[i]; vec2 b = verts[(i+1)%verts.size()]; vec2 edge = b-a; axes.push_back(vec2(-edge.y,edge.x).GetNormailzed()); } return axes; } This function returns the normalized vector: vec2 vec2::GetNormailzed() { float mag = sqrt( x*x + y*y ); return *this/mag; } This function projects a polygon onto an axis: void Polygon::Project(vec2* axis, float* min, float* max) { float d = axis->DotProduct(&verts[0]); float _min = d; float _max = d; for(int i = 1; i < verts.size(); i++) { d = axis->DotProduct(&verts[i]); _min = std::min(_min,d); _max = std::max(_max,d); } *min = _min; *max = _max; } This function returns the dot product of the vector with another vector. float vec2::DotProduct(vec2* other) { return (x*other->x + y*other->y); } Could anyone give me a pointer in the right direction to what could be causing this bug? Edit: I forgot this function, which gives me the interval distance: float Polygon::GetIntervalDistance(float minA, float maxA, float minB, float maxB) { float intervalDistance; if (minA < minB) { intervalDistance = minB - maxA; } else { intervalDistance = minA - maxB; } return intervalDistance; //A positive value indicates this axis can be separated. } Edit 2: I have recreated the problem in HTML5/Javascript: Demo

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  • Oops, I left my kernel zone configuration behind!

    - by mgerdts
    Most people use boot environments to move in one direction.  A system starts with an initial installation and from time to time new boot environments are created - typically as a result of pkg update - and then the new BE is booted.  This post is of little interest to those people as no hackery is needed.  This post is about some mild hackery. During development, I commonly test different scenarios across multiple boot environments.  Many times, those tests aren't related to the act of configuring or installing zone and I so it's kinda handy to avoid the effort involved of zone configuration and installation.  A somewhat common order of operations is like the following: # beadm create -e golden -a test1 # reboot Once the system is running in the test1 BE, I install a kernel zone. # zonecfg -z a178 create -t SYSsolaris-kz # zoneadm -z a178 install Time passes, and I do all kinds of stuff to the test1 boot environment and want to test other scenarios in a clean boot environment.  So then I create a new one from my golden BE and reboot into it. # beadm create -e golden -a test2 # reboot Since the test2 BE was created from the golden BE, it doesn't have the configuration for the kernel zone that I configured and installed.  Getting that zone over to the test2 BE is pretty easy.  My test1 BE is really known as s11fixes-2. root@vzl-212:~# beadm mount s11fixes-2 /mnt root@vzl-212:~# zonecfg -R /mnt -z a178 export | zonecfg -z a178 -f - root@vzl-212:~# beadm unmount s11fixes-2 root@vzl-212:~# zoneadm -z a178 attach root@vzl-212:~# zoneadm -z a178 boot On the face of it, it would seem as though it would have been easier to just use zonecfg -z a178 create -t SYSolaris-kz within the test2 BE to get the new configuration over.  That would almost work, but it would have left behind the encryption key required for access to host data and any suspend image.  See solaris-kz(5) for more info on host data.  I very commonly have more complex configurations that contain many storage URIs and non-default resource controls.  Retyping them would be rather tedious.

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  • Obtaining the correct Client IP address when a Physical Load Balancer and a Web Server Configured With Proxy Plug-in Are Between The Client And Weblogic

    - by adejuanc
    Some Load Balancers like Big-IP have build in interoperability with Weblogic Cluster, this means they know how Weblogic understand a header named 'WL-Proxy-Client-IP' to identify the original client ip.The problem comes when you have a Web Server configured with weblogic plug-in between the Load Balancer and the back-end weblogic servers - WL-Proxy-Client-IP this is not designed to go to Web server proxy plug-in. The plug-in will not use a WL-Proxy-Client-IP header that came in from the previous hop (which is this case is the Physical Load Balancer but could be anything), in order to prevent IP spoofing, therefore the plug-in won't pass on what Load Balancer has set for it.So unfortunately under this Architecture the header will be useless. To get the client IP from Weblogic you need to configure extended log format and create a custom field that gets the appropriate header containing the IP of the client.On WLS versions prior to 10.3.3 use these instructions:You can also create user-defined fields for inclusion in an HTTP access log file that uses the extended log format. To create a custom field you identify the field in the ELF log file using the Fields directive and then you create a matching Java class that generates the desired output. You can create a separate Java class for each field, or the Java class can output multiple fields. For a sample of the Java source for such a class, seeJava Class for Creating a Custom ELF Field to import weblogic.servlet.logging.CustomELFLogger;import weblogic.servlet.logging.FormatStringBuffer;import weblogic.servlet.logging.HttpAccountingInfo;/* This example outputs the X-Forwarded-For field into a custom field called MyOriginalClientIPField */public class MyOriginalClientIPField implements CustomELFLogger{ public void logField(HttpAccountingInfo metrics,  FormatStringBuffer buff) {   buff.appendValueOrDash(metrics.getHeader("X-Forwarded-For");  }}In this case we are using 'X-Forwarded-For' but it could be changed for the header that contains the data you need to use.Compile the class, jar it, and prepend it to the classpath.In order to compile and package the class: 1. Navigate to <WLS_HOME>/user_projects/domains/<SOME_DOMAIN>/bin2. Set up an environment by executing: $ . ./setDomainEnv.sh This will include weblogic.jar into classpath, in order to use any of the libraries included under package weblogic.*3. Compile the class by copying the content of the code above and naming the file as:MyOriginalClientIPField.java4. Run javac to compile the class.$javac MyOriginalClientIPField.java5. Package the compiled class into a jar file by executing:$jar cvf0 MyOriginalClientIPField.jar MyOriginalClientIPField.classExpected output is:added manifestadding: MyOriginalClientIPField.class(in = 711) (out= 711)(stored 0%)6. This will produce a file called:MyOriginalClientIPField.jar This way you will be able to get the real client IP when the request is passing through a Load Balancer and a Web server before reaching WLS. Since 10.3.3 it is possible to configure a specific header that WLS will check when getRemoteAddr is called. That can be set on the WebServer Mbean. In this case, set that to be X-Forwarded-For header coming from Load Balancer as well.

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  • Is my JavaScript/jQuery methodology good? [migrated]

    - by absentx
    I am seeking critique on what has become my normal methodology of writing JavaScript code. I have become heavily reliant on the jQuery library, but I think this has helped me learn the native language better also. Anyway, please critique the following style of JavaScript coding... Buried are a lot of questions of scope; if you could point out the strengths and weaknesses of this style I would appreciate it. var critique ={ start: function(){ globalness = 'GLOBAL-GLOBAL'; //Available to all critique's methods var notglobalness = 'LOCAL-LOCAL'; // Only available to critiques start method //Am I using the "method" teminology properly here?? $('#stuff').on('click','a.closer-target',function(){ $target = $(this); if($target.hasClass('active')){ $target.removeClass('active'); } else{ $target.addClass('active'); critique.madness($target); } }) console.log(notglobalness+': at least I am useful at home'); console.log('note here that: '+notglobalness+' is no longer available after this point, lets continue on:'); critique.madness(notglobalness); }, madness: function($e){ //Do a bunch of awesomeness with $e, //but continue to keep it seperate because you think its best to keep things isolated. //Send to the next function when complete here console.log('Here is globalness, which is still available from the start method of critique!! ' + globalness); console.log('Let us see if the globalness carries on to a new var object!!'); console.log('The locally isolated variable of NOTGLOBALNESS is available here, because it was passed to this method. Let us show it:'+$e); carryOn.start(); } } //end critique var carryOn={ start: function(){ console.log('any chance critique.globalness will work here??? lets see: ' +globalness); console.log('it absolutely does'); } } $(document).ready(critique.start); (I always struggle with which of the Stack Exchange sites is best to post "questions of theory" like this, but I think Programmers is the best, if not, as usual a mod will move it, etc...)

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  • Advantages to Multiple Methods over Switch

    - by tandu
    I received a code review from a senior developer today asking "By the way, what is your objection to dispatching functions by way of a switch statement?" I have read in many places about how pumping an argument through switch to call methods is bad OOP, not as extensible, etc. However, I can't really come up with a definitive answer for him. I would like to settle this for myself once and for all. Here are our competing code suggestions (php used as an example, but can apply more universally): class Switch { public function go($arg) { switch ($arg) { case "one": echo "one\n"; break; case "two": echo "two\n"; break; case "three": echo "three\n"; break; default: throw new Exception("Unknown call: $arg"); break; } } } class Oop { public function go_one() { echo "one\n"; } public function go_two() { echo "two\n"; } public function go_three() { echo "three\n"; } public function __call($_, $__) { throw new Exception("Unknown call $_ with arguments: " . print_r($__, true)); } } Part of his argument was "It (switch method) has a much cleaner way of handling default cases than what you have in the generic __call() magic method." I disagree about the cleanliness and in fact prefer call, but I would like to hear what others have to say. Arguments I can come up with in support of Oop scheme: A bit cleaner in terms of the code you have to write (less, easier to read, less keywords to consider) Not all actions delegated to a single method. Not much difference in execution here, but at least the text is more compartmentalized. In the same vein, another method can be added anywhere in the class instead of a specific spot. Methods are namespaced, which is nice. Does not apply here, but consider a case where Switch::go() operated on a member rather than a parameter. You would have to change the member first, then call the method. For Oop you can call the methods independently at any time. Arguments I can come up with in support of Switch scheme: For the sake of argument, cleaner method of dealing with a default (unknown) request Seems less magical, which might make unfamiliar developers feel more comfortable Anyone have anything to add for either side? I'd like to have a good answer for him.

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  • Setting the SlideShowExtender's Index

    - by Bunch
    The AJAX SlideShowExtender is pretty useful. It does what it says and works without much fuss. There was one trick I needed it to perform that I could not find natively within the control. That was to set the slide’s current index. With a little JavaScript however I could make the control do what I wanted. The example below assumes a few things. First you already have a SlideShowExtender setup and working (or see this post). Second this SlideShowExtender is on a page all by itself so the index to set the slide to is passed in the URL. The scenario I had was this SSE was showing full images, the index was passed from another page that had a SSE showing thumbnails. JavaScript in <head> <script type="text/javascript">      function pageLoad() {          var slider = $find("sse");          var photoIndex = GetQuerystring('Index', 0);          slider._currentIndex = photoIndex - 1;          slider._slides = '';          slider.setCurrentImage();      }      function GetQuerystring(key, default_) {          if (default_ == null) default_ = '0';          key = key.replace(/[\[]/, "\\\[").replace(/[\]]/, "\\\]");          var regex = new RegExp("[\\?&]" + key + "=([^&#]*)");          var qs = regex.exec(window.location.href);          if (qs == null)              return default_;          else              return qs[1];      } </script> The GetQuerystring function is what grabs the Index value I pass from the page with the thumbnails. It does not have anything else to do with setting the slide index. The code in the pageLoad function sets the index on the slide_currentIndex line. The slider.setCurrentImage() line does pretty much what it says. I added the slider._slider = ‘’ to avoid an error (not a show stopper just a bit annoying). Control in <body> <cc1:SlideShowExtender ID="ssePhotos" runat="server" TargetControlID="imgFull" AutoPlay="false"          PreviousButtonID="btnPrev" NextButtonID="btnNext" SlideShowServicePath="PlacePhotos.asmx"           SlideShowServiceMethod="GetPlaceFullPhotos" BehaviorID="sse" ImageDescriptionLabelID="lblPictureDescription"> </cc1:SlideShowExtender> The main property to set with the SSE is the BehaviorID. This is what a JavaScript function can use to find the control rather than the control’s ID value. Technorati Tags: AJAX,ASP.Net,JavaScript

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