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  • Which is the most practical way to add functionality to this piece of code?

    - by Adam Arold
    I'm writing an open source library which handles hexagonal grids. It mainly revolves around the HexagonalGrid and the Hexagon class. There is a HexagonalGridBuilder class which builds the grid which contains Hexagon objects. What I'm trying to achieve is to enable the user to add arbitrary data to each Hexagon. The interface looks like this: public interface Hexagon extends Serializable { // ... other methods not important in this context <T> void setSatelliteData(T data); <T> T getSatelliteData(); } So far so good. I'm writing another class however named HexagonalGridCalculator which adds some fancy pieces of computation to the library like calculating the shortest path between two Hexagons or calculating the line of sight around a Hexagon. My problem is that for those I need the user to supply some data for the Hexagon objects like the cost of passing through a Hexagon, or a boolean flag indicating whether the object is transparent/passable or not. My question is how should I implement this? My first idea was to write an interface like this: public interface HexagonData { void setTransparent(boolean isTransparent); void setPassable(boolean isPassable); void setPassageCost(int cost); } and make the user implement it but then it came to my mind that if I add any other functionality later all code will break for those who are using the old interface. So my next idea is to add annotations like @PassageCost, @IsTransparent and @IsPassable which can be added to fields and when I'm doing the computation I can look for the annotations in the satelliteData supplied by the user. This looks flexible enough if I take into account the possibility of later changes but it uses reflection. I have no benchmark of the costs of using annotations so I'm a bit in the dark here. I think that in 90-95% of the cases the efficiency is not important since most users wont't use a grid where this is significant but I can imagine someone trying to create a grid with a size of 5.000.000.000 X 5.000.000.000. So which path should I start walking on? Or are there some better alternatives? Note: These ideas are not implemented yet so I did not pay too much attention to good names.

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  • Using multiple diagrams per model in Entity Framework 5.0

    - by nikolaosk
    I have downloaded .Net framework 4.5 and Visual Studio 2012 since it was released to MSDN subscribers on the 15th of August.For people that do not know about that yet please have a look at Jason Zander's excellent blog post .Since then I have been investigating the many new features that have been introduced in this release.In this post I will be looking into theIn order to follow along this post you must have Visual Studio 2012 and .Net Framework 4.5 installed in your machine.Download and install VS 20120 using this link.My machine runs on Windows 8 and Visual Studio 2012 works just fine. I have also installed in my machine SQL Server 2012 developer edition. I have also downloaded and installed AdventureWorksLT2012 database.You can download this database from the codeplex website.   Before I start showcasing the demo I want to say that I strongly believe that Entity Framework is maturing really fast and now at version 5.0 can be used as your data access layer in all your .Net projects.I have posted extensively about Entity Framework in my blog.Please find all the EF related posts here. In this demo I will show you how to split an entity model into multiple diagrams using the new enhanced EF designer. We will not build an application in this demo.Sometimes our model can become too large to edit or view.In earlier versions we could only have one diagram per EDMX file.In EF 5.0 we can split the model into more diagrams.1) Launch VS 2012. Express edition will work fine.2) Create a New Project. From the available templates choose a Web Forms application  3) Add a new item in your project, an ADO.Net Entity Data Model. I have named it AdventureWorksLT.edmx.Then we will create the model from the database and click Next.Create a new connection by specifying the SQL Server instance and the database name and click OK.Then click Next in the wizard.In the next screen of the wizard select all the tables from the database and hit Finish.4) It will take a while for our .edmx diagram to be created. When I select an Entity (e.g Customer) from my diagram and right click on it,a new option appears "Move to new Diagram".Make sure you have the Model Browser window open.Have a look at the picture below 5) When we do that a new diagram is created and our new Entity is moved there.Have a look at the picture below  6) We can also right-click and include the related entities. Have a look at the picture below. 7) When we do that the related entities are copied to the new diagram.Have a look at the picture below  8) Now we can cut (CTRL+X) the entities from Diagram2 and paste them back to Diagram1.9) Finally another great enhancement of the EF 5.0 designer is that you can change colors in the various entities that make up the model.Select the entities you want to change color, then in the Properties window choose the color of your choice. Have a look at the picture below. To recap we have demonstrated how to split your entity model in multiple diagrams which comes handy in EF models that have a large number of entities in them Hope it helps!!!!

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  • How can a solo programmer become a good team player?

    - by Nick
    I've been programming (obsessively) since I was 12. I am fairly knowledgeable across the spectrum of languages out there, from assembly, to C++, to Javascript, to Haskell, Lisp, and Qi. But all of my projects have been by myself. I got my degree in chemical engineering, not CS or computer engineering, but for the first time this fall I'll be working on a large programming project with other people, and I have no clue how to prepare. I've been using Windows all of my life, but this project is going to be very unix-y, so I purchased a Mac recently in the hopes of familiarizing myself with the environment. I was fortunate to participate in a hackathon with some friends this past year -- both CS majors -- and excitingly enough, we won. But I realized as I worked with them that their workflow was very different from mine. They used Git for version control. I had never used it at the time, but I've since learned all that I can about it. They also used a lot of frameworks and libraries. I had to learn what Rails was pretty much overnight for the hackathon (on the other hand, they didn't know what lexical scoping or closures were). All of our code worked well, but they didn't understand mine, and I didn't understand theirs. I hear references to things that real programmers do on a daily basis -- unit testing, code reviews, but I only have the vaguest sense of what these are. I normally don't have many bugs in my little projects, so I have never needed a bug tracking system or tests for them. And the last thing is that it takes me a long time to understand other people's code. Variable naming conventions (that vary with each new language) are difficult (__mzkwpSomRidicAbbrev), and I find the loose coupling difficult. That's not to say I don't loosely couple things -- I think I'm quite good at it for my own work, but when I download something like the Linux kernel or the Chromium source code to look at it, I spend hours trying to figure out how all of these oddly named directories and files connect. It's a programming sin to reinvent the wheel, but I often find it's just quicker to write up the functionality myself than to spend hours dissecting some library. Obviously, people who do this for a living don't have these problems, and I'll need to get to that point myself. Question: What are some steps that I can take to begin "integrating" with everyone else? Thanks!

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  • Using a SQL Prompt snippet with template parameters

    - by SQLDev
    As part of my product management role I regularly attend trade shows and man the Red Gate booth in the vendor exhibition hall. Amongst other things this involves giving product demos to customers. Our latest demo involves SQL Source Control and SQL Test in a continuous integration environment. In order to demonstrate quite how easy it is to set up our tools from scratch we start the demo by creating an entirely new database to link to source control, using an individual database name for each conference attendee. In SQL Server Management Studio this can be done either by selecting New Database from the Object Explorer or by executing “CREATE DATABASE DemoDB_John” in a query window. We recently extended the demo to include SQL Test. This uses an open source SQL Server unit testing framework called tSQLt (www.tsqlt.org), which has a CLR object that requires EXTERNAL_ACCESS to be set as follows: ALTER DATABASE DemoDB_John SET TRUSTWORTHY ON This isn’t hard to do, but if you’re giving demo after demo, this two-step process soon becomes tedious. This is where SQL Prompt snippets come into their own. I can create a snippet named create_demo_db for this following: CREATE DATABASE DemoDB_John GO USE DemoDB_John GO ALTER DATABASE DemoDB_John SET TRUSTWORTHY ON Now I just have to type the first few characters of the snippet name, select the snippet from SQL Prompt’s candidate list, and execute the code. Simple! The problem is that this can only work once due to the hard-coded database name. Luckily I can leverage a nice feature in SQL Server Management Studio called Template Parameters. If I modify my snippet to be: CREATE DATABASE <DBName,, DemoDB_> GO USE <DBName,, DemoDB_> GO ALTER DATABASE <DBName,, DemoDB_> SET TRUSTWORTHY ON Once I’ve invoked the snippet, I can press Ctrl-Shift-M, which calls up the Specify Values for Template Parameters dialog, where I can type in my database name just once. Now you can click OK and run the query. Easy. Ideally I’d like for SQL Prompt to auto-invoke the Template Parameter dialog for all snippets where it detects the angled bracket syntax, but typing in the keyboard shortcut is a small price to pay for the time savings.

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  • Precise Touch Screen Dragging Issue: Trouble Aligning with the Finger due to Different Screen Resolution

    - by David Dimalanta
    Please, I need your help. I'm trying to make a game that will drag-n-drop a sprite/image while my finger follows precisely with the image without being offset. When I'm trying on a 900x1280 (in X [900] and Y [1280]) screen resolution of the Google Nexus 7 tablet, it follows precisely. However, if I try testing on a phone smaller than 900x1280, my finger and the image won't aligned properly and correctly except it still dragging. This is the code I used for making a sprite dragging with my finger under touchDragged(): x = ((screenX + Gdx.input.getX())/2) - (fruit.width/2); y = ((camera_2.viewportHeight * multiplier) - ((screenY + Gdx.input.getY())/2) - (fruit.width/2)); This code above will make the finger and the image/sprite stays together in place while dragging but only works on 900x1280. You'll be wondering there's camera_2.viewportHeight in my code. Here are for two reasons: to prevent inverted drag (e.g. when you swipe with your finger downwards, the sprite moves upward instead) and baseline for reading coordinate...I think. Now when I'm adding another orthographic camera named camera_1 and changing its setting, I recently used it for adjusting the falling object by meter per pixel. Also, it seems effective independently for smartphones that has smaller resolution and this is what I used here: show() camera_1 = new OrthographicCamera(); camera_1.viewportHeight = 280; // --> I set it to a smaller view port height so that the object would fall faster, decreasing the chance of drag force. camera_1.viewportWidth = 196; // --> Make it proportion to the original screen view size as possible. camera_1.position.set(camera_1.viewportWidth * 0.5f, camera_1.viewportHeight * 0.5f, 0f); camera_1.update(); touchDragged() x = ((screenX + (camera_1.viewportWidth/Gdx.input.getX()))/2) - (fruit.width/2); y = ((camera_1.viewportHeight * multiplier) - ((screenY + (camera_1.viewportHeight/Gdx.input.getY()))/2) - (fruit.width/2)); But the result instead of just following the image/sprite closely to my finger, it still has a space/gap between the sprite/image and the finger. It is possibly dependent on coordinates based on the screen resolution. I'm trying to drag the blueberry sprite with my finger. My expectation did not met since I want my finger and the sprite/image (blueberry) to stay close together while dragging until I release it. Here's what it looks like: I got to figure it out how to make independent on all screen sizes by just following the image/sprite closely to my finger while dragging even on most different screen sizes instead.

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  • Unable to Mount an external hard drive (NTFS)

    - by Mediterran81
    Ubuntu 11.10. When I plug my external Drive Western Digital MyPassport (500Go NTFS) I named WD. I get the following error: Unable to mount WD Error mounting: mount exited with exit code 1: helper failed with: mount: according to mtab, /dev/sdb1 is already mounted on /media/WD mount failed I have no problem with the internal NTFS partitions that auto-mounts on startup (ntfs-config does that). If I plug the WD before I boot Ubuntu, upon login, it's recognized and I can access without no problem. But if I remove it using (Safely remove) and then replug it, I get the error above. Here is my fstab: # /etc/fstab: static file system information. # # <file system> <mount point> <type> <options> <dump> <pass> proc /proc proc nodev,noexec,nosuid 0 0 #Entry for /dev/sda5 : UUID=24540d0f-5803-493c-ace9-e3b3c0cedb26 / ext4 errors=remount-ro 0 1 #Entry for /dev/sda3 : UUID=E4C43F7EC43F51D2 /media/OS ntfs-3g defaults,locale=en_US.UTF-8 0 0 #Entry for /dev/sda2 : UUID=6A0070F10070C61B /media/RECOVERY ntfs-3g defaults,locale=en_US.UTF-8 0 0 #Entry for /dev/sdb1 : UUID=EA6854D268549F5F /media/WD ntfs-3g defaults,nosuid,nodev,locale=en_US.UTF-8 0 0 #Entry for /dev/sda6 : UUID=ed077c52-c50e-406c-9120-9cb6f86ec204 none swap sw 0 0 Here is my mtab /dev/sda5 / ext4 rw,errors=remount-ro,commit=0 0 0 proc /proc proc rw,noexec,nosuid,nodev 0 0 sysfs /sys sysfs rw,noexec,nosuid,nodev 0 0 fusectl /sys/fs/fuse/connections fusectl rw 0 0 none /sys/kernel/debug debugfs rw 0 0 none /sys/kernel/security securityfs rw 0 0 udev /dev devtmpfs rw,mode=0755 0 0 devpts /dev/pts devpts rw,noexec,nosuid,gid=5,mode=0620 0 0 tmpfs /run tmpfs rw,noexec,nosuid,size=10%,mode=0755 0 0 none /run/lock tmpfs rw,noexec,nosuid,nodev,size=5242880 0 0 none /run/shm tmpfs rw,nosuid,nodev 0 0 /dev/sda3 /media/OS fuseblk rw,nosuid,nodev,allow_other,blksize=4096 0 0 /dev/sda2 /media/RECOVERY fuseblk rw,nosuid,nodev,allow_other,blksize=4096 0 0 /dev/sdb1 /media/WD fuseblk rw,nosuid,nodev,allow_other,blksize=4096 0 0 binfmt_misc /proc/sys/fs/binfmt_misc binfmt_misc rw,noexec,nosuid,nodev 0 0 gvfs-fuse-daemon /home/hanine/.gvfs fuse.gvfs-fuse-daemon rw,nosuid,nodev,user=hanine 0 0 Appearently it cannot be mounted because upon login, it finds that it is already mounted. Some sort of conflict. Does anyone have a clue on how to solve this. Thanks.

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  • How to become a good team player?

    - by Nick
    I've been programming (obsessively) since I was 12. I am fairly knowledgeable across the spectrum of languages out there, from assembly, to C++, to Javascript, to Haskell, Lisp, and Qi. But all of my projects have been by myself. I got my degree in chemical engineering, not CS or computer engineering, but for the first time this fall I'll be working on a large programming project with other people, and I have no clue how to prepare. I've been using Windows all of my life, but this project is going to be very unix-y, so I purchased a Mac recently in the hopes of familiarizing myself with the environment. I was fortunate to participate in a hackathon with some friends this past year -- both CS majors -- and excitingly enough, we won. But I realized as I worked with them that their workflow was very different from mine. They used Git for version control. I had never used it at the time, but I've since learned all that I can about it. They also used a lot of frameworks and libraries. I had to learn what Rails was pretty much overnight for the hackathon (on the other hand, they didn't know what lexical scoping or closures were). All of our code worked well, but they didn't understand mine, and I didn't understand theirs. I hear references to things that real programmers do on a daily basis -- unit testing, code reviews, but I only have the vaguest sense of what these are. I normally don't have many bugs in my little projects, so I have never needed a bug tracking system or tests for them. And the last thing is that it takes me a long time to understand other people's code. Variable naming conventions (that vary with each new language) are difficult (__mzkwpSomRidicAbbrev), and I find the loose coupling difficult. That's not to say I don't loosely couple things -- I think I'm quite good at it for my own work, but when I download something like the Linux kernel or the Chromium source code to look at it, I spend hours trying to figure out how all of these oddly named directories and files connect. It's a programming sin to reinvent the wheel, but I often find it's just quicker to write up the functionality myself than to spend hours dissecting some library. Obviously, people who do this for a living don't have these problems, and I'll need to get to that point myself. Question: What are some steps that I can take to begin "integrating" with everyone else? Thanks!

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  • Saddling your mountain lion with JDeveloper

    - by Blueberry Coder
    Last October, Apple released Java Update 2012-006. This patch brought the Apple-provided JDK for OS X Lion v10.7 and OS X Mountain Lion v10.8 to version 1.6.0_37. At the same time, it disabled the Apple Java plugins and removed the Java Preferences panel that enabled users to manage the various Java releases on their computer. On the Windows and Linux platforms, JDeveloper 11g R1 has been certified  to run on Java 7 since patch set 5. This is not the case on OS X.   ( The above is not a typo. Apple's OS for personal computer is now known as OS X; the « Mac » prefix has been dropped with the 10.8 release. And it's pronounced « Oh-Ess-Ten », by the way. Yes, I am a nitpicker. I know... ) Please note JDeveloper 11g R2 is not certified either. On any platform. It will generally work, but there are known issues with ADF Mobile. Personally, I would recommend to wait for 12c before going to JDK 7.  Now, suppose you have installed Oracle's JDK 7 on your Mac. JDeveloper will not run on it. It will even not install. Susan and I discovered this the hard way while setting up the ADF Mobile hands-on lab we ran at the UKOUG 2012 conference. The lab was a great success nevertheless, attracting nearly a hundred delegates. It was great to see the interest ADF Mobile already generates, especially among PL/SQL Developers and DBAs. But what did we do to make it work?  While Java Update 2012-006 removed the Java Preferences panel, it leaved in place OS X's command-line Java infrastructure. Thus, it is possible to invoke the Apple JDK 6 to start the JDeveloper installer. Suppose your user is named « Fred », and that the JDeveloper installer is on your desktop. You can execute the following command in a terminal window (on a single line) to start the installer:  /usr/libexec/java_home --version 1.6.0  --exec java -jar /Users/Fred/Desktop/jdevstudio11116install.jar  The JDeveloper installer, being provided a valid JDK reference, will set up the IDE and embedded WebLogic Server instance accordingly. Clever engineering at its finest!

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  • OData &ndash; The easiest service I can create

    - by Jon Dalberg
    I wanted to create an OData service with the least amount of code so I fired up Visual Studio and got cracking. I decided to serve up a list of naughty words and make them read-only. Create a new web project. I created an empty MVC 2 application but MVC is not required for OData. Add a new WCF Data Service to the project. I named mine NastyWords.svc since I’m serving up a list of nasty words. Add a class to expose via the service: NastyWord 1: [DataServiceKey("Word")] 2: public class NastyWord 3: { 4: public string Word { get; set; } 5: }   I need to be able to uniquely identify instances of NastyWords for the DataService so I used the DataServiceKey attribute with the “Word” property as the key. I could have added an “ID” property which would have uniquely identified them and would then not need the “DataServiceKey” attribute because the DataService would apply some reflection and heuristics to guess at which property would be the unique identifier. However, the words themselves are unique so adding an “ID” property would be redundantly repetitive. Then I created a data source to expose my NastyWord objects to the service. This is just a simple class with IQueryable<T> properties exposing the entities for my service: 1: public class NastyWordsDataSource 2: { 3: private static IList<NastyWord> words = new List<NastyWord> 4: { 5: new NastyWord{ Word="crap"}, 6: new NastyWord{ Word="darn"}, 7: new NastyWord{ Word="hell"}, 8: new NastyWord{ Word="shucks"} 9: }; 10:   11: public NastyWordsDataSource() 12: { 13: NastyWords = words.AsQueryable(); 14: } 15:   16: public IQueryable<NastyWord> NastyWords { get; private set; } 17: }   Now I can go to the NastyWords.svc class and tell it which data source to use and which entities to expose: 1: public class NastyWords : DataService<NastyWordsDataSource> 2: { 3: // This method is called only once to initialize service-wide policies. 4: public static void InitializeService(DataServiceConfiguration config) 5: { 6: config.SetEntitySetAccessRule("*", EntitySetRights.AllRead); 7: config.DataServiceBehavior.MaxProtocolVersion = DataServiceProtocolVersion.V2; 8: } 9: }   Compile and browse to my NastWords.svc and weep with joy Now I can query my service just like any other OData service. Next time, I’ll modify this service to allow updates to sent so I can build up my list of nasty words. Enjoy!

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  • Inline template efficiency

    - by Darryl Gove
    I like inline templates, and use them quite extensively. Whenever I write code with them I'm always careful to check the disassembly to see that the resulting output is efficient. Here's a potential cause of inefficiency. Suppose we want to use the mis-named Leading Zero Detect (LZD) instruction on T4 (this instruction does a count of the number of leading zero bits in an integer register - so it should really be called leading zero count). So we put together an inline template called lzd.il looking like: .inline lzd lzd %o0,%o0 .end And we throw together some code that uses it: int lzd(int); int a; int c=0; int main() { for(a=0; a<1000; a++) { c=lzd(c); } return 0; } We compile the code with some amount of optimisation, and look at the resulting code: $ cc -O -xtarget=T4 -S lzd.c lzd.il $ more lzd.s .L77000018: /* 0x001c 11 */ lzd %o0,%o0 /* 0x0020 9 */ ld [%i1],%i3 /* 0x0024 11 */ st %o0,[%i2] /* 0x0028 9 */ add %i3,1,%i0 /* 0x002c */ cmp %i0,999 /* 0x0030 */ ble,pt %icc,.L77000018 /* 0x0034 */ st %i0,[%i1] What is surprising is that we're seeing a number of loads and stores in the code. Everything could be held in registers, so why is this happening? The problem is that the code is only inlined at the code generation stage - when the actual instructions are generated. Earlier compiler phases see a function call. The called functions can do all kinds of nastiness to global variables (like 'a' in this code) so we need to load them from memory after the function call, and store them to memory before the function call. Fortunately we can use a #pragma directive to tell the compiler that the routine lzd() has no side effects - meaning that it does not read or write to memory. The directive to do that is #pragma no_side_effect(<routine name), and it needs to be placed after the declaration of the function. The new code looks like: int lzd(int); #pragma no_side_effect(lzd) int a; int c=0; int main() { for(a=0; a<1000; a++) { c=lzd(c); } return 0; } Now the loop looks much neater: /* 0x0014 10 */ add %i1,1,%i1 ! 11 ! { ! 12 ! c=lzd(c); /* 0x0018 12 */ lzd %o0,%o0 /* 0x001c 10 */ cmp %i1,999 /* 0x0020 */ ble,pt %icc,.L77000018 /* 0x0024 */ nop

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  • NHibernate Session Load vs Get when using Table per Hierarchy. Always use ISession.Get&lt;T&gt; for TPH to work.

    - by Rohit Gupta
    Originally posted on: http://geekswithblogs.net/rgupta/archive/2014/06/01/nhibernate-session-load-vs-get-when-using-table-per-hierarchy.aspxNHibernate ISession has two methods on it : Load and Get. Load allows the entity to be loaded lazily, meaning the actual call to the database is made only when properties on the entity being loaded is first accessed. Additionally, if the entity has already been loaded into NHibernate Cache, then the entity is loaded directly from the cache instead of querying the underlying database. ISession.Get<T> instead makes the call to the database, every time it is invoked. With this background, it is obvious that we would prefer ISession.Load<T> over ISession.Get<T> most of the times for performance reasons to avoid making the expensive call to the database. let us consider the impact of using ISession.Load<T> when we are using the Table per Hierarchy implementation of NHibernate. Thus we have base class/ table Animal, there is a derived class named Snake with the Discriminator column being Type which in this case is “Snake”. If we load This Snake entity using the Repository for Animal, we would have a entity loaded, as shown below: public T GetByKey(object key, bool lazy = false) { if (lazy) return CurrentSession.Load<T>(key); return CurrentSession.Get<T>(key); } var tRepo = new NHibernateReadWriteRepository<TPHAnimal>(); var animal = tRepo.GetByKey(new Guid("602DAB56-D1BD-4ECC-B4BB-1C14BF87F47B"), true); var snake = animal as Snake; snake is null As you can see that the animal entity retrieved from the database cannot be cast to Snake even though the entity is actually a snake. The reason being ISession.Load prevents the entity to be cast to Snake and will throw the following exception: System.InvalidCastException :  Message=Unable to cast object of type 'TPHAnimalProxy' to type 'NHibernateChecker.Model.Snake'. Thus we can see that if we lazy load the entity using ISession.Load<TPHAnimal> then we get a TPHAnimalProxy and not a snake. =============================================================== However if do not lazy load the same cast works perfectly fine, this is since we are loading the entity from database and the entity being loaded is not a proxy. Thus the following code does not throw any exceptions, infact the snake variable is not null: var tRepo = new NHibernateReadWriteRepository<TPHAnimal>(); var animal = tRepo.GetByKey(new Guid("602DAB56-D1BD-4ECC-B4BB-1C14BF87F47B"), false); var snake = animal as Snake; if (snake == null) { var snake22 = (Snake) animal; }

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  • Partial Shader Signatures HLSL D3D11 C++

    - by ThePhD
    I had been debugging a problem I was having in a single shader file with 2 functions in it. I'm using DirectX 11, vs_5_0 and ps_5_0. I have stripped it down to its basic components to understand what was going wrong with the shaders, because the different named components of the Pixel and Vertex shaders were swapping the data being input: void QuadVertex ( inout float4 position : SV_Position, inout float4 color : COLOR0, inout float2 tex : TEXCOORD0 ) { // ViewProject is a 4x4 matrix, // just included here to show the simple passthrough of the data position = mul(position, ViewProjection); } And a Pixel Shader: float4 QuadPixel ( float4 color : COLOR0, float2 tex : TEXCOORD0 ) : SV_Target0 { // Color is filled with position data and tex is // filled with color values from the Vertex Shader return color; } The ID3D11InputLayout and associated C++ code correctly compiles the shaders and sets them up with some simple primitive data: data[0].Position.x = 0.0f * 210; data[0].Position.y = 1.0f * 160; data[0].Position.z = 0.0f; data[1].Position.x = 0.0f * 210; data[1].Position.y = 0.0f * 160; data[1].Position.z = 0.0f; data[2].Position.x = 1.0f * 210; data[2].Position.y = 1.0f * 160; data[2].Position.z = 0.0f; data[0].Colour = Colors::Red; data[1].Colour = Colors::Red; data[2].Colour = Colors::Red; data[0].Texture = Vector2::Zero; data[1].Texture = Vector2::Zero; data[2].Texture = Vector2::Zero; When used with the shader, the float4 color always ended up with the position data, and the float2 tex always ended up with the color data. After a moment, I figured out that the shader's input and output signatures needed to be in the correct order and the correct format and be laid out in the exact order of the output from the Vertex Shader, regardless of the semantics: float4 QuadPixel ( float4 pos : SV_Position, float4 color : COLOR0, float2 tex : TEXCOORD0 ) : SV_Target0 { return color; } After finding this out, My question is: Why don't the semantics map the appropriate components when going from Vertex Shader to Pixel Shader? Is there any way that I can make it so certain semantics are always mapped to other semantics, or do I always have to follow the rigid Shader Signature (in this case, Position, Color, and Texture) ? As a side note for why I'm asking: I know that when using XNA, my shader signatures for functions could differ in position and even drop items from Vertex Shader to Pixel Shader function parameters, having only the COLOR0 and TEXCOORD0 components being used (and it would still match up correctly). However, I also know that XNA relied on DX9 (and maybe a little DX10) implementation, and that maybe this kind of flexibility no longer exists in DX11?

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  • Getting a virus is *very* annoying

    - by bconlon
    I spent most of yesterday removing an annoying virus from my PC. I feel slightly foolish for getting one in the first place, but after so many years I guess I was always going to eventually succumb. I was also a little surprised at the failure of various tools at removing it. The virus would redirect the browser to websites including ‘licosearch’, ‘hugosearch’ and ‘facebook’, and the disk would be thrashing away infecting dlls in some way. I had the full up to date version of McAfee installed. This identified that there was an issue in some dlls on the system and was able to ‘fix’ them. But they kept getting re-infected. So I installed Microsoft Security Essentials and this too was able to identify and ‘fix’ the infected dlls. The system scans take forever and I really expected better results. I also tried Malwarebytes, Hitman Pro, AVG and Sophos to no avail. Eventually I thought I’d investigate myself. It turned out that on reboot, the virus would start 3 instances of Firefox.exe which I’m guessing would do bad things including infecting as many dlls on the system as possible. I removed Firefox and the virus cleverly then launched 3 instances of Chrome! So I uninstalled Chrome and yes, it then started to launch 3 instances of iexplore.exe. If I’m honest, by this stage I was just seeing if it would be able to use any of the browsers! As it was starting these on reboot, I looked in my User Startup folder and there was a <randomly named>.exe and several log files. I deleted these and rebooted. When I looked they had been recreated. So I then looked in the registry Run and RunOnce entries: HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Windows\CurrentVersion\Run. Sure enough there were entries to run a file in C:\Program Files\<random name folder>\<random name file>.exe. I deleted this and rebooted and it was fixed. I also looked in the event log and found a warning that Winlogon had failed to start the file C:\Program Files\<random name folder>\<random name file>.exe So I also checked HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Windows NT\CurrentVersion\Winlogon and this entry had also been changed. Finally I ran a full system scan to clean up any infected dlls. I hope it’s gone for good!  #

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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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  • Brain Teaser: How Did I Do This (Part 1: The Solution)

    - by Geertjan
    In Part 1: The Challenge, published this time last week, I introduced a "brain teaser". The brain teaser asks you to figure out how to allow images and other files to be meaningfully dropped onto a NetBeans Platform application, i.e., on the drop something useful should happen with the dropped file: if the file is an image, the image should open in the IDE; if the file is a PDF document, the PDF viewer should open externally; if the file is a text file, it should open as a text in the IDE, etc. Solution. And here is the solution: http://bits.netbeans.org/dev/javadoc/org-openide-windows/org/openide/windows/ExternalDropHandler.html When an implementation of the "ExternalDropHandler" class is available in the global Lookup, and an object is being dragged over some part of the main window, the window system may call the methods of this class to decide whether it can accept or reject the drag operation. And when the object is actually dropped, this class will be asked to handle the drop. OK, so go ahead and implement the above class and put it into the Lookup. Or... guess what? The NetBeans Platform has a default implementation of the above class, appropriately named "DefaultExternalDropHandler". Not only is this useful to learn about how to implement the ExternalDropHandler class (i.e., by reading the source here): you can simply include the module that contains this class in your own NetBeans Platform application and then your application will be able to receive external drag/drop events and do something meaningful with them thanks to the DefaultExternalDropHandler. Do this: Open your NetBeans Platform application in NetBeans IDE. Right-click the application in the Projects window and choose Properties. In the Libraries tab, expand the "ide" cluster, and select "User Utilities". (That's where "DefaultExternalDropHandler.java" is found and registered in the Lookup.) Now click the "Resolve" button, if it appears, because some additional related modules need to now be included, if they haven't been included yet. Again in the "ide" cluster in the Libraries tab, select "Image". That's the Image Editor. Click OK. Run the application. Drag an image or some other type of file into your application, from outside the application, and you'll see the application tries to handle the drop. If the file being dragged is an image, it will open in the Image Editor, which you included in the previous step of these instructions. Hurray, you're done. Without any programming at all, you've added a cool new feature to your application.

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  • Applying DDD principles in a RESTish web service

    - by Andy
    I am developing an RESTish web service. I think I got the idea of the difference between aggregation and composition. Aggregation does not enforce lifecycle/scope on the objects it references. Composition does enforce lifecycle/scope on the objects it contain/own. If I delete a composite object then all the objects it contain/own are deleted as well, while the deleting an aggregate root does not delete referenced objects. 1) If it is true that deleting aggregate roots does not necessary delete referenced objects, what sense does it make to not have a repository for the references objects? Or are aggregate roots as a term referring to what is known as composite object? 2) When you create an web service you will have multiple endpoints, in my case I have one entity Book and another named Comment. It does not make sense to leave the comments in my application if the book is deleted. Therefore, book is a composite object. I guess I should not have a repository for comments since that would break the enforcement of lifecycle and rules that the book class may have. However I have URL such as (examples only): GET /books/1/comments POST /books/1/comments Now, if I do not have a repository for comments, does that mean I have to load the book object and then return the referenced comments? Am I allowed to return a list of Comment entities from the BookRepository, does that make sense? The repository for Book may eventually become rather big with all sorts of methods. Am I allowed to write JPQL (JPA queries) that targets comments and not books inside the repository? What about pagination and filtering of comments. When adding a new comment triggered by the POST endpoint, do you need to load the book, add the comment to the book, and then update the whole book object? What I am currently doing is having a own CommentRepository, even though the comments are deleted with the book. I could need some direction on how to do it correct. Since you are exposing not only root objects in RESTish services I wonder how to handle this at the backend. I am using Hibernate and Spring.

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  • How to Send the Contents of the Clipboard to a Text File via the Send to Menu

    - by Jason Faulkner
    We have previously covered how to send the contents of a text file to the Windows Clipboard with a simple Send To shortcut, but what if you want to do the opposite? That is: send the contents of the clipboard to a text file with a simple shortcut. No problem. Here’s how. Copy the ClipOut Utility While Windows offers the command line tool ‘clip’ as a way to direct console output to the clipboard, it does not have a tool to direct the clipboard contents to the console. To do this, we are going to use a small utility named ClipOut (download link at the bottom). Simply download and extract this file to a location in your Windows PATH variable (if you don’t know what this means, just extract the EXE to your C:\Windows folder) and you are ready to go. Add the Send To Shortcut Open your Send To folder location by going to Run > shell:sendto Create a new shortcut with the command: CMD /C ClipOut > Note the above command will overwrite the contents of the selected file. If you would like to append to the contents of the selected file, use this command instead: CMD /C ClipOut >> Of course, you could make shortcuts for both. Give a descriptive name to the shortcut. You’re finished. Using this shortcut will now send the text contents copied to your Windows Clipboard to the selected file. It is important to note that the ClipOut tool only supports outputting text. If you had binary data copied to your clipboard, then the output would be empty. Changing the Icon By default, the icon for the shortcut will appear as a command prompt, but you can easily change this by editing the properties of the shortcut and clicking the Change Icon button. We used an icon located in “%SystemRoot%\System32\shell32.dll”, but any icon of your liking will do. As an additional tweak, you can set the properties of the shortcut to run minimized. This will prevent the command window from “blinking” when the send to command is run (instead it will blink in your taskbar, which is hardly noticeable). Links Download ClipOut Utility     

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  • How to use multiple search keys?

    - by user32565
    I have a database wherein the files are named abcd100.00b, abcd101.00b, etc. I need a code where when the user enters abcd separate then 100 to 110, all the files with the name abcd and in the range 100 to 110 should get displayed now the following code can display only the first four characters. How do I implement this? <?php //capture search term and remove spaces at its both ends if the is any $searchTerm = trim($_GET['keyname']) ; //check whether the name parsed is empty if($searchTerm == "rinex_file") { echo "Enter name you are searching for."; exit(); } if($searchTerm == "rinex_file") { echo "Enter name you are searching for."; exit(); } //database connection info $host = "localhost"; //server $db = "rinex"; //database name $user = "m"; //dabases user name $pwd = "c"; //password //connecting to server and creating link to database $link = mysqli_connect($host, $user, $pwd, $db); //MYSQL search statement $query = "SELECT * FROM rinexo WHERE rinex_file LIKE '%$searchTerm%'"; $results = mysqli_query($link, $query) ; /* check whethere there were matching records in the table by counting the number of results returned */ if(mysqli_num_rows($results) >= 1){ echo '<table border="1"> <tr> <th>rinex version</th> <th>program</th> <th>date</th> <th>maker name</th> <th>maker number</th> <th>observer</th> <th>agency</th> <th>position_X_Y_Z</th> </tr>'; while($row = mysqli_fetch_array($results)){ echo '<tr> <td>'.$row['rinex_version'].'</td> <td>'.$row['pgm'].'</td> <td>'.$row['date'].'</td> <td>'.$row['marker_name'].'</td> <td>'.$row['marker_no'].'</td> <td>'.$row['observer'].'</td> <td>'.$row['agency'].'</td> <td>'.$row['position_X_Y_Z'].'</td> </tr>'; } echo '</table>'; }else{ echo "There was no matching record for the name " . $searchTerm; }

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  • Configuring trace file size and number in WebCenter Content 11g

    - by Kyle Hatlestad
    Lately I've been doing a lot of debugging using the System Output tracing in WebCenter Content 11g.  This is built-in tracing in the content server which provides a great level of detail on what's happening under the hood.  You can access the settings as well as a view of the tracing by going to Administration -> System Audit Information.  From here, you can select the tracing sections to include.  Some of my personal favorites are searchquery,  systemdatabase, userstorage, and indexer.  Usually I'm trying to find out some information regarding a search, database query, or user information.  Besides debugging, it's also very helpful for performance tuning. One of the nice tricks with the tracing is it honors the wildcard (*) character.  So you can put in 'schema*' and gather all of the schema related tracing.  And you can notice if you select 'all' and update, it changes to just a *.   To view the tracing in real-time, you simply go to the 'View Server Output' page and the latest tracing information will be at the bottom. This works well if you're looking at something pretty discrete and the system isn't getting much activity.  But if you've got a lot of tracing going on, it would be better to go after the trace log file itself.  By default, the log files can be found in the <content server instance directory>/data/trace directory. You'll see it named 'idccs_<managed server name>_current.log.  You may also find previous trace logs that have rolled over.  In this case they will identified by a date/time stamp in the name.  By default, the server will rotate the logs after they reach 1MB in size.  And it will keep the most recent 10 logs before they roll off and get deleted.  If your server is in a cluster, then the trace file should be configured to be local to the node per the recommended configuration settings. If you're doing some extensive tracing and need to capture all of the information, there are a couple of configuration flags you can set to control the logs. #Change log size to 10MB and number of logs to 20FileSizeLimit=10485760FileCountLimit=20 This is set by going to Admin Server -> General Configuration and entering them in the Additional Configuration Variables: section.  Restart the server and it should take on the new logging settings. 

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  • How can you tell whether to use Composite Pattern or a Tree Structure, or a third implementation?

    - by Aske B.
    I have two client types, an "Observer"-type and a "Subject"-type. They're both associated with a hierarchy of groups. The Observer will receive (calendar) data from the groups it is associated with throughout the different hierarchies. This data is calculated by combining data from 'parent' groups of the group trying to collect data (each group can have only one parent). The Subject will be able to create the data (that the Observers will receive) in the groups they're associated with. When data is created in a group, all 'children' of the group will have the data as well, and they will be able to make their own version of a specific area of the data, but still linked to the original data created (in my specific implementation, the original data will contain time-period(s) and headline, while the subgroups specify the rest of the data for the receivers directly linked to their respective groups). However, when the Subject creates data, it has to check if all affected Observers have any data that conflicts with this, which means a huge recursive function, as far as I can understand. So I think this can be summed up to the fact that I need to be able to have a hierarchy that you can go up and down in, and some places be able to treat them as a whole (recursion, basically). Also, I'm not just aiming at a solution that works. I'm hoping to find a solution that is relatively easy to understand (architecture-wise at least) and also flexible enough to be able to easily receive additional functionality in the future. Is there a design pattern, or a good practice to go by, to solve this problem or similar hierarchy problems? EDIT: Here's the design I have: The "Phoenix"-class is named that way because I didn't think of an appropriate name yet. But besides this I need to be able to hide specific activities for specific observers, even though they are attached to them through the groups. A little Off-topic: Personally, I feel that I should be able to chop this problem down to smaller problems, but it escapes me how. I think it's because it involves multiple recursive functionalities that aren't associated with each other and different client types that needs to get information in different ways. I can't really wrap my head around it. If anyone can guide me in a direction of how to become better at encapsulating hierarchy problems, I'd be very glad to receive that as well.

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  • LibGDX - Textures rendering at wrong position

    - by ACluelessGuy
    Update 2: Let me further explain my problem since I think that i didn't make it clear enough: The Y-coordinates on the bottom of my screen should be 0. Instead it is the height of my screen. That means the "higher" i touch/click the screen the less my y-coordinate gets. Above that the origin is not inside my screen, atleast not the 0 y-coordinate. Original post: I'm currently developing a tower defence game for fun by using LibGDX. There are places on my map where the player is or is not allowed to put towers on. So I created different ArrayLists holding rectangles representing a tile on my map. (towerPositions) for(int i = 0; i < map.getLayers().getCount(); i++) { curLay = (TiledMapTileLayer) map.getLayers().get(i); //For all Cells of current Layer for(int k = 0; k < curLay.getWidth(); k++) { for(int j = 0; j < curLay.getHeight(); j++) { curCell = curLay.getCell(k, j); //If there is a actual cell if(curCell != null) { tileWidth = curLay.getTileWidth(); tileHeight = curLay.getTileHeight(); xTileKoord = tileWidth*k; yTileKoord = tileHeight*j; switch(curLay.getName()) { //If layer named "TowersAllowed" picked case "TowersAllowed": towerPositions.add(new Rectangle(xTileKoord, yTileKoord, tileWidth, tileHeight)); // ... AND SO ON If the player clicks on a "allowed" field later on he has the opportunity to build a tower of his coice via a menu. Now here is the problem: The towers render, but they render at wrong position. (They appear really random on the map, no certain pattern for me) for(Rectangle curRect : towerPositions) { if(curRect.contains(xCoord, yCoord)) { //Using a certain tower in this example (left the menu out if(gameControl.createTower("towerXY")) { //RenderObject is just a class holding the Texture and x/y coordinates renderList.add(new RenderObject(new Texture(Gdx.files.internal("TowerXY.png")), curRect.x, curRect.y)); } } } Later on i render it: game.batch.begin(); for(int i = 0; i < renderList.size() ; i++) { game.batch.draw(renderList.get(i).myTexture, renderList.get(i).x, renderList.get(i).y); } game.batch.end(); regards

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  • How to configure a longer version Number in artifactory

    - by claudine
    The version-numbers for our jars have to be longer them x.x.x. We would rather need x.x.x.x to integrate some old-fashioned self-made mechanism. This is, because we tag our software with x.x.x and as soon as we have a delivery to a customer one specific jar has to be build exactly at this point of time to fit to another backend, which communicates with our program. For that reason this one jar has the version 2.3.4.1, when generated and in next delivery of the same Version it is build and named 2.3.4.2. Now artifactory cannot handle this an doesn't save more than x.x.x.2 in some cases. So we thought of maybe edit the regular expression in the maven repository layout (see attached Screenshot) Because testing the path in the field below shows, that it cannot handle the version number. Of course for the rest of our jars still x.x.x has to work.. For Example here is the maven-metadata.xml <?xml version="1.0" encoding="UTF-8"?> <metadata> <groupId>com.firm</groupId> <artifactId>someid</artifactId> <version>1.5.1</version> <versioning> <latest>1.5.1</latest> <release>1.5.1</release> <versions> <version>1.4.62</version> </versions> <lastUpdated>20120926073942</lastUpdated> </versioning> </metadata> The folder structure looks like: someid 1.4.62 1.4.62.1 1.4.62.2 1.4.62.3 If we deploy an new artifact version (1.4.62.1), the maven-metadata.xml contains the 1.4.62.1 version. But the artifactory overrides the version number (1.4.62.x) to (1.4.62) after an unspecified time. It seems that the artifactory only support major, minor and revision numbers, and deletes the buildnumber. Now we looking for a solution do disable this behavior. We use the JFrog Artifactory version 2.5.0 (rev. 13086). Any ideas, maybe? Thanks in andvance

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  • How to have mulptiple search keys using one filename

    - by user107020
    I have a database wherein the files are named abcd100.00b, abcd101.00b,..... i need a code where wen the user enters abcd separate then 100 to 110 . all the files with the name abcd and in the range 100 to 110 should get displayed now the following code can display pnly the first four characters how do i implement this?????? <?php //capture search term and remove spaces at its both ends if the is any $searchTerm = trim($_GET['keyname']) ; //check whether the name parsed is empty if($searchTerm == "rinex_file") { echo "Enter name you are searching for."; exit(); } if($searchTerm == "rinex_file") { echo "Enter name you are searching for."; exit(); } //database connection info $host = "localhost"; //server $db = "rinex"; //database name $user = "m"; //dabases user name $pwd = "c"; //password //connecting to server and creating link to database $link = mysqli_connect($host, $user, $pwd, $db); //MYSQL search statement $query = "SELECT * FROM rinexo WHERE rinex_file LIKE '%$searchTerm%'"; $results = mysqli_query($link, $query) ; /* check whethere there were matching records in the table by counting the number of results returned */ if(mysqli_num_rows($results) >= 1){ echo '<table border="1"> <tr> <th>rinex version</th> <th>program</th> <th>date</th> <th>maker name</th> <th>maker number</th> <th>observer</th> <th>agency</th> <th>position_X_Y_Z</th> </tr>'; while($row = mysqli_fetch_array($results)){ echo '<tr> <td>'.$row['rinex_version'].'</td> <td>'.$row['pgm'].'</td> <td>'.$row['date'].'</td> <td>'.$row['marker_name'].'</td> <td>'.$row['marker_no'].'</td> <td>'.$row['observer'].'</td> <td>'.$row['agency'].'</td> <td>'.$row['position_X_Y_Z'].'</td> </tr>'; } echo '</table>'; }else{ echo "There was no matching record for the name " . $searchTerm; }

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  • Is this a valid implementation of the repository pattern?

    - by user1578653
    I've been reading up about the repository pattern, with a view to implementing it in my own application. Almost all examples I've found on the internet use some kind of existing framework rather than showing how to implement it 'from scratch'. Here's my first thoughts of how I might implement it - I was wondering if anyone could advise me on whether this is correct? I have two tables, named CONTAINERS and BITS. Each CONTAINER can contain any number of BITs. I represent them as two classes: class Container{ private $bits; private $id; //...and a property for each column in the table... public function __construct(){ $this->bits = array(); } public function addBit($bit){ $this->bits[] = $bit; } //...getters and setters... } class Bit{ //some properties, methods etc... } Each class will have a property for each column in its respective table. I then have a couple of 'repositories' which handle things to do with saving/retrieving these objects from the database: //repository to control saving/retrieving Containers from the database class ContainerRepository{ //inject the bit repository for use later public function __construct($bitRepo){ $this->bitRepo = $bitRepo; } public function getById($id){ //talk directly to Oracle here to all column data into the object //get all the bits in the container $bits = $this->bitRepo->getByContainerId($id); foreach($bits as $bit){ $container->addBit($bit); } //return an instance of Container } public function persist($container){ //talk directly to Oracle here to save it to the database //if its ID is NULL, create a new container in database, otherwise update the existing one //use BitRepository to save each of the Bits inside the Container $bitRepo = $this->bitRepo; foreach($container->bits as $bit){ $bitRepo->persist($bit); } } } //repository to control saving/retrieving Bits from the database class BitRepository{ public function getById($id){} public function getByContainerId($containerId){} public function persist($bit){} } Therefore, the code I would use to get an instance of Container from the database would be: $bitRepo = new BitRepository(); $containerRepo = new ContainerRepository($bitRepo); $container = $containerRepo->getById($id); Or to create a new one and save to the database: $bitRepo = new BitRepository(); $containerRepo = new ContainerRepository($bitRepo); $container = new Container(); $container->setSomeProperty(1); $bit = new Bit(); $container->addBit($bit); $containerRepo->persist($container); Can someone advise me as to whether I have implemented this pattern correctly? Thanks!

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  • SSTP client disconnects shortly after successfully connected to VPN

    - by Eran Betzalel
    I'm successfully authenticating and connecting to a SSTP VPN (on windows 2008) from my windows 7 machine, but for some reason, the connection is disconnected about a 1-2 seconds after it's established. I've done the following: Defined a SSTP VPN on my windows server 2008. Defined the same machine as CA. Issued the needed certificates and published them on the client. I'm currently testing this VPN inside my LAN so all the needed ports are opened. Here are the event log entries when trying to connect: Error Log (Client): The user HOME\User dialed a connection named Home VPN which has terminated. The reason code returned on termination is 829. Error Log (Server-VPN): The user HOME\User connected on port VPN0-0 on 7/27/2012 at 1:57 AM and disconnected on 7/27/2012 at 1:57 AM. The user was active for 0 minutes 0 seconds. 312 bytes were sent and 4528 bytes were received. The reason for disconnecting was user request. What would be the issue? How can I resolve or debug it? UPDATE: I've found an event log (Log=System, Source=RasSstp) message on the windows 7 machine that tries to connect to the VPN: The SSTP-based VPN connection to the remote access server was terminated because of a security check failure. Security settings on the remote access server do not match settings on this computer. Contact the system administrator of the remote access server and relay the following information: SHA1 Certificate Hash: 065D681...520375552F SHA256 Certificate Hash: 18DED363...EEEE28CFD00

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