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• IIS 7.0 informational HTTP status codes

  - by Samir R. Bhogayta

  1xx - Informational These HTTP status codes indicate a provisional response. The client computer receives one or more 1xx responses before the client computer receives a regular response. IIS 7.0 uses the following informational HTTP status codes: 100 - Continue. 101 - Switching protocols. 2xx - Success These HTTP status codes indicate that the server successfully accepted the request. IIS 7.0 uses the following success HTTP status codes: 200 - OK. The client request has succeeded. 201 - Created. 202 - Accepted. 203 - Nonauthoritative information. 204 - No content. 205 - Reset content. 206 - Partial content. 3xx - Redirection These HTTP status codes indicate that the client browser must take more action to fulfill the request. For example, the client browser may have to request a different page on the server. Or, the client browser may have to repeat the request by using a proxy server. IIS 7.0 uses the following redirection HTTP status codes: 301 - Moved permanently. 302 - Object moved. 304 - Not modified. 307 - Temporary redirect. 4xx - Client error These HTTP status codes indicate that an error occurred and that the client browser appears to be at fault. For example, the client browser may have requested a page that does not exist. Or, the client browser may not have provided valid authentication information. IIS 7.0 uses the following client error HTTP status codes: 400 - Bad request. The request could not be understood by the server due to malformed syntax. The client should not repeat the request without modifications. IIS 7.0 defines the following HTTP status codes that indicate a more specific cause of a 400 error: 400.1 - Invalid Destination Header. 400.2 - Invalid Depth Header. 400.3 - Invalid If Header. 400.4 - Invalid Overwrite Header. 400.5 - Invalid Translate Header. 400.6 - Invalid Request Body. 400.7 - Invalid Content Length. 400.8 - Invalid Timeout. 400.9 - Invalid Lock Token. 401 - Access denied. IIS 7.0 defines several HTTP status codes that indicate a more specific cause of a 401 error. The following specific HTTP status codes are displayed in the client browser but are not displayed in the IIS log: 401.1 - Logon failed. 401.2 - Logon failed due to server configuration. 401.3 - Unauthorized due to ACL on resource. 401.4 - Authorization failed by filter. 401.5 - Authorization failed by ISAPI/CGI application. 403 - Forbidden. IIS 7.0 defines the following HTTP status codes that indicate a more specific cause of a 403 error: 403.1 - Execute access forbidden. 403.2 - Read access forbidden. 403.3 - Write access forbidden. 403.4 - SSL required. 403.5 - SSL 128 required. 403.6 - IP address rejected. 403.7 - Client certificate required. 403.8 - Site access denied. 403.9 - Forbidden: Too many clients are trying to connect to the Web server. 403.10 - Forbidden: Web server is configured to deny Execute access. 403.11 - Forbidden: Password has been changed. 403.12 - Mapper denied access. 403.13 - Client certificate revoked. 403.14 - Directory listing denied. 403.15 - Forbidden: Client access licenses have exceeded limits on the Web server. 403.16 - Client certificate is untrusted or invalid. 403.17 - Client certificate has expired or is not yet valid. 403.18 - Cannot execute requested URL in the current application pool. 403.19 - Cannot execute CGI applications for the client in this application pool. 403.20 - Forbidden: Passport logon failed. 403.21 - Forbidden: Source access denied. 403.22 - Forbidden: Infinite depth is denied. 404 - Not found. IIS 7.0 defines the following HTTP status codes that indicate a more specific cause of a 404 error: 404.0 - Not found. 404.1 - Site Not Found. 404.2 - ISAPI or CGI restriction. 404.3 - MIME type restriction. 404.4 - No handler configured. 404.5 - Denied by request filtering configuration. 404.6 - Verb denied. 404.7 - File extension denied. 404.8 - Hidden namespace. 404.9 - File attribute hidden. 404.10 - Request header too long. 404.11 - Request contains double escape sequence. 404.12 - Request contains high-bit characters. 404.13 - Content length too large. 404.14 - Request URL too long. 404.15 - Query string too long. 404.16 - DAV request sent to the static file handler. 404.17 - Dynamic content mapped to the static file handler via a wildcard MIME mapping. 404.18 - Querystring sequence denied. 404.19 - Denied by filtering rule. 405 - Method Not Allowed. 406 - Client browser does not accept the MIME type of the requested page. 408 - Request timed out. 412 - Precondition failed. 5xx - Server error These HTTP status codes indicate that the server cannot complete the request because the server encounters an error. IIS 7.0 uses the following server error HTTP status codes: 500 - Internal server error. IIS 7.0 defines the following HTTP status codes that indicate a more specific cause of a 500 error: 500.0 - Module or ISAPI error occurred. 500.11 - Application is shutting down on the Web server. 500.12 - Application is busy restarting on the Web server. 500.13 - Web server is too busy. 500.15 - Direct requests for Global.asax are not allowed. 500.19 - Configuration data is invalid. 500.21 - Module not recognized. 500.22 - An ASP.NET httpModules configuration does not apply in Managed Pipeline mode. 500.23 - An ASP.NET httpHandlers configuration does not apply in Managed Pipeline mode. 500.24 - An ASP.NET impersonation configuration does not apply in Managed Pipeline mode. 500.50 - A rewrite error occurred during RQ_BEGIN_REQUEST notification handling. A configuration or inbound rule execution error occurred. Note Here is where the distributed rules configuration is read for both inbound and outbound rules. 500.51 - A rewrite error occurred during GL_PRE_BEGIN_REQUEST notification handling. A global configuration or global rule execution error occurred. Note Here is where the global rules configuration is read. 500.52 - A rewrite error occurred during RQ_SEND_RESPONSE notification handling. An outbound rule execution occurred. 500.53 - A rewrite error occurred during RQ_RELEASE_REQUEST_STATE notification handling. An outbound rule execution error occurred. The rule is configured to be executed before the output user cache gets updated. 500.100 - Internal ASP error. 501 - Header values specify a configuration that is not implemented. 502 - Web server received an invalid response while acting as a gateway or proxy. IIS 7.0 defines the following HTTP status codes that indicate a more specific cause of a 502 error: 502.1 - CGI application timeout. 502.2 - Bad gateway. 503 - Service unavailable. IIS 7.0 defines the following HTTP status codes that indicate a more specific cause of a 503 error: 503.0 - Application pool unavailable. 503.2 - Concurrent request limit exceeded.

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• C#/.NET Little Wonders: Comparer&lt;T&gt;.Default

  - by James Michael Hare

  I’ve been working with a wonderful team on a major release where I work, which has had the side-effect of occupying most of my spare time preparing, testing, and monitoring.  However, I do have this Little Wonder tidbit to offer today. Introduction The IComparable<T> interface is great for implementing a natural order for a data type.  It’s a very simple interface with a single method: 1: public interface IComparer<in T> 2: { 3: // Compare two instances of same type. 4: int Compare(T x, T y); 5: }  So what do we expect for the integer return value?  It’s a pseudo-relative measure of the ordering of x and y, which returns an integer value in much the same way C++ returns an integer result from the strcmp() c-style string comparison function: If x == y, returns 0. If x > y, returns > 0 (often +1, but not guaranteed) If x < y, returns < 0 (often –1, but not guaranteed) Notice that the comparison operator used to evaluate against zero should be the same comparison operator you’d use as the comparison operator between x and y.  That is, if you want to see if x > y you’d see if the result > 0. The Problem: Comparing With null Can Be Messy This gets tricky though when you have null arguments.  According to the MSDN, a null value should be considered equal to a null value, and a null value should be less than a non-null value.  So taking this into account we’d expect this instead: If x == y (or both null), return 0. If x > y (or y only is null), return > 0. If x < y (or x only is null), return < 0. But here’s the problem – if x is null, what happens when we attempt to call CompareTo() off of x? 1: // what happens if x is null? 2: x.CompareTo(y); It’s pretty obvious we’ll get a NullReferenceException here.  Now, we could guard against this before calling CompareTo(): 1: int result; 2:  3: // first check to see if lhs is null. 4: if (x == null) 5: { 6: // if lhs null, check rhs to decide on return value. 7: if (y == null) 8: { 9: result = 0; 10: } 11: else 12: { 13: result = -1; 14: } 15: } 16: else 17: { 18: // CompareTo() should handle a null y correctly and return > 0 if so. 19: result = x.CompareTo(y); 20: } Of course, we could shorten this with the ternary operator (?:), but even then it’s ugly repetitive code: 1: int result = (x == null) 2: ? ((y == null) ? 0 : -1) 3: : x.CompareTo(y); Fortunately, the null issues can be cleaned up by drafting in an external Comparer.  The Soltuion: Comparer<T>.Default You can always develop your own instance of IComparer<T> for the job of comparing two items of the same type.  The nice thing about a IComparer is its is independent of the things you are comparing, so this makes it great for comparing in an alternative order to the natural order of items, or when one or both of the items may be null. 1: public class NullableIntComparer : IComparer<int?> 2: { 3: public int Compare(int? x, int? y) 4: { 5: return (x == null) 6: ? ((y == null) ? 0 : -1) 7: : x.Value.CompareTo(y); 8: } 9: }  Now, if you want a custom sort -- especially on large-grained objects with different possible sort fields -- this is the best option you have.  But if you just want to take advantage of the natural ordering of the type, there is an easier way.  If the type you want to compare already implements IComparable<T> or if the type is System.Nullable<T> where T implements IComparable, there is a class in the System.Collections.Generic namespace called Comparer<T> which exposes a property called Default that will create a singleton that represents the default comparer for items of that type.  For example: 1: // compares integers 2: var intComparer = Comparer<int>.Default; 3:  4: // compares DateTime values 5: var dateTimeComparer = Comparer<DateTime>.Default; 6:  7: // compares nullable doubles using the null rules! 8: var nullableDoubleComparer = Comparer<double?>.Default;  This helps you avoid having to remember the messy null logic and makes it to compare objects where you don’t know if one or more of the values is null. This works especially well when creating say an IComparer<T> implementation for a large-grained class that may or may not contain a field.  For example, let’s say you want to create a sorting comparer for a stock open price, but if the market the stock is trading in hasn’t opened yet, the open price will be null.  We could handle this (assuming a reasonable Quote definition) like: 1: public class Quote 2: { 3: // the opening price of the symbol quoted 4: public double? Open { get; set; } 5:  6: // ticker symbol 7: public string Symbol { get; set; } 8:  9: // etc. 10: } 11:  12: public class OpenPriceQuoteComparer : IComparer<Quote> 13: { 14: // Compares two quotes by opening price 15: public int Compare(Quote x, Quote y) 16: { 17: return Comparer<double?>.Default.Compare(x.Open, y.Open); 18: } 19: } Summary Defining a custom comparer is often needed for non-natural ordering or defining alternative orderings, but when you just want to compare two items that are IComparable<T> and account for null behavior, you can use the Comparer<T>.Default comparer generator and you’ll never have to worry about correct null value sorting again.     Technorati Tags: C#,.NET,Little Wonders,BlackRabbitCoder,IComparable,Comparer

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• JSP Precompilation for ADF Applications

  - by Duncan Mills

  A question that comes up from time to time, particularly in relation to build automation, is how to best pre-compile the .jspx and .jsff files in an ADF application. Thus ensuring that the app is ready to run as soon as it's installed into WebLogic. In the normal run of things, the first poor soul to hit a page pays the price and has to wait a little whilst the JSP is compiled into a servlet. Everyone else subsequently gets a free lunch. So it's a reasonable thing to want to do... Let Me List the Ways So forth to Google (other search engines are available)... which lead me to a fairly old article on WLDJ - Removing Performance Bottlenecks Through JSP Precompilation. Technololgy wise, it's somewhat out of date, but the one good point that it made is that it's really not very useful to try and use the precompile option in the weblogic.xml file. That's a really good observation - particularly if you're trying to integrate a pre-compile step into a Hudson Continuous Integration process. That same article mentioned an alternative approach for programmatic pre-compilation using weblogic.jspc. This seemed like a much more useful approach for a CI environment. However, weblogic.jspc is now obsoleted by weblogic.appc so we'll use that instead.  Thanks to Steve for the pointer there. And So To APPC APPC has documentation - always a great place to start, and supports usage both from Ant via the wlappc task and from the command line using the weblogic.appc command. In my testing I took the latter approach. Usage, as the documentation will show you, is superficially pretty simple.  The nice thing here, is that you can pass an existing EAR file (generated of course using OJDeploy) and that EAR will be updated in place with the freshly compiled servlet classes created from the JSPs. Appc takes care of all the unpacking, compiling and re-packing of the EAR for you. Neat.  So we're done right...? Not quite. The Devil is in the Detail  OK so I'm being overly dramatic but it's not all plain sailing, so here's a short guide to using weblogic.appc to compile a simple ADF application without pain.  Information You'll Need The following is based on the assumption that you have a stand-alone WLS install with the Application Development  Runtime installed and a suitable ADF enabled domain created. This could of course all be run off of a JDeveloper install as well 1. Your Weblogic home directory. Everything you need is relative to this so make a note.  In my case it's c:\builds\wls_ps4. 2. Next deploy your EAR as normal and have a peek inside it using your favourite zip management tool. First of all look at the weblogic-application.xml inside the EAR /META-INF directory. Have a look for any library references. Something like this: <library-ref>    <library-name>adf.oracle.domain</library-name> </library-ref>   Make a note of the library ref (adf.oracle.domain in this case) , you'll need that in a second. 3. Next open the nested WAR file within the EAR and then have a peek inside the weblogic.xml file in the /WEB-INF directory. Again  make a note of the library references. 4. Now start the WebLogic as per normal and run the WebLogic console app (e.g. http://localhost:7001/console). In the Domain Structure navigator, select Deployments. 5. For each of the libraries you noted down drill into the library definition and make a note of the .war, .ear or .jar that defines the library. For example, in my case adf.oracle.domain maps to "C:\ builds\ WLS_PS4\ oracle_common\ modules\ oracle. adf. model_11. 1. 1\ adf. oracle. domain. ear". Note the extra spaces that are salted throughout this string as it is displayed in the console - just to make it annoying, you'll have to strip these out. 6. Finally you'll need the location of the adfsharebean.jar. We need to pass this on the classpath for APPC so that the ADFConfigLifeCycleCallBack listener can be found. In a more complex app of your own you may need additional classpath entries as well.  Now we're ready to go, and it's a simple matter of applying the information we have gathered into the relevant command line arguments for the utility A Simple CMD File to Run APPC  Here's the stub .cmd file I'm using on Windows to run this. @echo offREM Stub weblogic.appc Runner setlocal set WLS_HOME=C:\builds\WLS_PS4 set ADF_LIB_ROOT=%WLS_HOME%\oracle_common\modulesset COMMON_LIB_ROOT=%WLS_HOME%\wlserver_10.3\common\deployable-libraries set ADF_WEBAPP=%ADF_LIB_ROOT%\oracle.adf.view_11.1.1\adf.oracle.domain.webapp.war set ADF_DOMAIN=%ADF_LIB_ROOT%\oracle.adf.model_11.1.1\adf.oracle.domain.ear set JSTL=%COMMON_LIB_ROOT%\jstl-1.2.war set JSF=%COMMON_LIB_ROOT%\jsf-1.2.war set ADF_SHARE=%ADF_LIB_ROOT%\oracle.adf.share_11.1.1\adfsharembean.jar REM Set up the WebLogic Environment so appc can be found call %WLS_HOME%\wlserver_10.3\server\bin\setWLSEnv.cmd CLS REM Now compile away!java weblogic.appc -verbose -library %ADF_WEBAPP%,%ADF_DOMAIN%,%JSTL%,%JSF% -classpath %ADF_SHARE% %1 endlocal Running the above on a target ADF .ear  file will zip through and create all of the relevant compiled classes inside your nested .war file in the \WEB-INF\classes\jsp_servlet\ directory (but don't take my word for it, run it and take a look!) And So... In the immortal words of  the Pet Shop Boys, Was It Worth It? Well, here's where you'll have to do your own testing. In  my case here, with a simple ADF application, pre-compilation shaved an non-scientific "3 Elephants" off of the initial page load time for the first access of each page. That's a pretty significant payback for such a simple step to add into your CI process, so why not give it a go.

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• How to Identify Which Hardware Component is Failing in Your Computer

  - by Chris Hoffman

  Concluding that your computer has a hardware problem is just the first step. If you’re dealing with a hardware issue and not a software issue, the next step is determining what hardware problem you’re actually dealing with. If you purchased a laptop or pre-built desktop PC and it’s still under warranty, you don’t need to care about this. Have the manufacturer fix the PC for you — figuring it out is their problem. If you’ve built your own PC or you want to fix a computer that’s out of warranty, this is something you’ll need to do on your own. Blue Screen 101: Search for the Error Message This may seem like obvious advice, but searching for information about a blue screen’s error message can help immensely. Most blue screens of death you’ll encounter on modern versions of Windows will likely be caused by hardware failures. The blue screen of death often displays information about the driver that crashed or the type of error it encountered. For example, let’s say you encounter a blue screen that identified “NV4_disp.dll” as the driver that caused the blue screen. A quick Google search will reveal that this is the driver for NVIDIA graphics cards, so you now have somewhere to start. It’s possible that your graphics card is failing if you encounter such an error message. Check Hard Drive SMART Status Hard drives have a built in S.M.A.R.T. (Self-Monitoring, Analysis, and Reporting Technology) feature. The idea is that the hard drive monitors itself and will notice if it starts to fail, providing you with some advance notice before the drive fails completely. This isn’t perfect, so your hard drive may fail even if SMART says everything is okay. If you see any sort of “SMART error” message, your hard drive is failing. You can use SMART analysis tools to view the SMART health status information your hard drives are reporting. Test Your RAM RAM failure can result in a variety of problems. If the computer writes data to RAM and the RAM returns different data because it’s malfunctioning, you may see application crashes, blue screens, and file system corruption. To test your memory and see if it’s working properly, use Windows’ built-in Memory Diagnostic tool. The Memory Diagnostic tool will write data to every sector of your RAM and read it back afterwards, ensuring that all your RAM is working properly. Check Heat Levels How hot is is inside your computer? Overheating can rsult in blue screens, crashes, and abrupt shut downs. Your computer may be overheating because you’re in a very hot location, it’s ventilated poorly, a fan has stopped inside your computer, or it’s full of dust. Your computer monitors its own internal temperatures and you can access this information. It’s generally available in your computer’s BIOS, but you can also view it with system information utilities such as SpeedFan or Speccy. Check your computer’s recommended temperature level and ensure it’s within the appropriate range. If your computer is overheating, you may see problems only when you’re doing something demanding, such as playing a game that stresses your CPU and graphics card. Be sure to keep an eye on how hot your computer gets when it performs these demanding tasks, not only when it’s idle. Stress Test Your CPU You can use a utility like Prime95 to stress test your CPU. Such a utility will fore your computer’s CPU to perform calculations without allowing it to rest, working it hard and generating heat. If your CPU is becoming too hot, you’ll start to see errors or system crashes. Overclockers use Prime95 to stress test their overclock settings — if Prime95 experiences errors, they throttle back on their overclocks to ensure the CPU runs cooler and more stable. It’s a good way to check if your CPU is stable under load. Stress Test Your Graphics Card Your graphics card can also be stress tested. For example, if your graphics driver crashes while playing games, the games themselves crash, or you see odd graphical corruption, you can run a graphics benchmark utility like 3DMark. The benchmark will stress your graphics card and, if it’s overheating or failing under load, you’ll see graphical problems, crashes, or blue screens while running the benchmark. If the benchmark seems to work fine but you have issues playing a certain game, it may just be a problem with that game. Swap it Out Not every hardware problem is easy to diagnose. If you have a bad motherboard or power supply, their problems may only manifest through occasional odd issues with other components. It’s hard to tell if these components are causing problems unless you replace them completely. Ultimately, the best way to determine whether a component is faulty is to swap it out. For example, if you think your graphics card may be causing your computer to blue screen, pull the graphics card out of your computer and swap in a new graphics card. If everything is working well, it’s likely that your previous graphics card was bad. This isn’t easy for people who don’t have boxes of components sitting around, but it’s the ideal way to troubleshoot. Troubleshooting is all about trial and error, and swapping components out allows you to pin down which component is actually causing the problem through a process of elimination. This isn’t a complete guide to everything that could likely go wrong and how to identify it — someone could write a full textbook on identifying failing components and still not cover everything. But the tips above should give you some places to start dealing with the more common problems. Image Credit: Justin Marty on Flickr     

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• How To: Using spatial data with Entity Framework and Connector/Net

  - by GABMARTINEZ

  One of the new features introduced in Entity Framework 5.0 is the incorporation of some new types of data within an Entity Data Model: the spatial data types. These types allow us to perform operations on coordinates values in an easier way. There's no need to add stored routines or functions for every operation among these geometry types, now the user can have the alternative to put this logic on his application or keep it in the database. In the new 6.7.4 version there's also this new feature incorporated to Connector/Net library so our users can start exploring it and could provide us some feedback or comments about this new functionality. Through this tutorial on how to create a Code First Entity Model with a geometry column, we'll show an example on using Geometry types and some common operations when using geometry types inside an application. Requirements: - Connector/Net 6.7.4 - Entity Framework 5.0 version - .NET Framework 4.5 version - Basic understanding on Entity Framework and C# language. - An installed and running instance of MySQL Server 5.5.x or 5.6.10 version- Visual Studio 2012. Step One: Create a new Console Application  Inside Visual Studio select File->New Project menu option and select the Console Application template. Also make sure the .Net 4.5 version is selected so the new features for EF 5.0 will work with the application. Step Two: Add the Entity Framework Package For adding the Entity Framework Package there is more than one option: the package manager console or the Manage Nuget Packages option dialog. If you want to open the Package Manager Console, go to the Tools Menu -> Library Package Manager -> Package Manager Console. On the Package Manager Console Type:Install-Package EntityFrameworkThis will add the reference to the project of the latest released No alpha version of Entity Framework. Step Three: Adding Entity class and DBContext We'll add a simple class that represents a table entity to save some places and its location using a DBGeometry column that will be mapped to a Geometry type in MySQL. After that some operations can be performed using this data. public class MyPlace { [Key] public int Id { get; set; } public string name { get; set; } public DbGeometry location { get; set; } } public class JourneyDb : DbContext { public DbSet<MyPlace> MyPlaces { get; set; } }  Also make sure to add the connection string to the App.Config file as in the example: <?xml version="1.0" encoding="utf-8"?> <configuration>   <configSections>     <!-- For more information on Entity Framework configuration, visit http://go.microsoft.com/fwlink/?LinkID=237468 -->     <section name="entityFramework" type="System.Data.Entity.Internal.ConfigFile.EntityFrameworkSection, EntityFramework, Version=5.0.0.0, Culture=neutral, PublicKeyToken=b77a5c561934e089" requirePermission="false" />   </configSections>   <startup>     <supportedRuntime version="v4.0" sku=".NETFramework,Version=v4.5" />   </startup>   <connectionStrings>     <add name="JourneyDb" connectionString="server=localhost;userid=root;pwd=;database=journeydb" providerName="MySql.Data.MySqlClient"/>   </connectionStrings>   <entityFramework>     </entityFramework> </configuration> Note also that the <entityFramework> section is empty.Step Four: Adding some new records.On the Program.cs file add the following code for the Main method so the Database gets created and also some new data can be added to the new table. This code adds some records containing some determinate locations. After being added a distance function will be used to know how much distance has each location in reference to the Queens Village Station in New York. static void Main(string[] args)    {     using (JourneyDb cxt = new JourneyDb())      {        cxt.Database.Delete();        cxt.Database.Create();         cxt.MyPlaces.Add(new MyPlace()        {          name = "JFK INTERNATIONAL AIRPORT OF NEW YORK",          location = DbGeometry.FromText("POINT(40.644047 -73.782291)"),        });         cxt.MyPlaces.Add(new MyPlace()        {          name = "ALLEY POND PARK",          location = DbGeometry.FromText("POINT(40.745696 -73.742638)"),        });       cxt.MyPlaces.Add(new MyPlace()        {          name = "CUNNINGHAM PARK",          location = DbGeometry.FromText("POINT(40.735031 -73.768387)"),        });         cxt.MyPlaces.Add(new MyPlace()        {          name = "QUEENS VILLAGE STATION",          location = DbGeometry.FromText("POINT(40.717957 -73.736501)"),        });         cxt.SaveChanges();         var points = (from p in cxt.MyPlaces                      select new { p.name, p.location });        foreach (var item in points)       {         Console.WriteLine("Location " + item.name + " has a distance in Km from Queens Village Station " + DbGeometry.FromText("POINT(40.717957 -73.736501)").Distance(item.location) * 100);       }       Console.ReadKey();      }  }}Output : Location JFK INTERNATIONAL AIRPORT OF NEW YORK has a distance from Queens Village Station 8.69448802402959 Km. Location ALLEY POND PARK has a distance from Queens Village Station 2.84097675104912 Km. Location CUNNINGHAM PARK has a distance from Queens Village Station 3.61695793727275 Km. Location QUEENS VILLAGE STATION has a distance from Queens Village Station 0 Km. Conclusion:Adding spatial data to a table is easier than before when having Entity Framework 5.0. This new Entity Framework feature that handles spatial data columns within the Data layer has a lot of integrated functions and methods toease this type of tasks.Notes:This version of Connector/Net is just released as GA so is preatty much stable to be used on a ProductionEnvironment. Please send us your comments or questions using this blog or at the Forums where we keep answering any questions you have about Connector/Net and MySQL Server.A copy of this sample project can be downloaded here. This application does not include any library so you will haveto add them before running it. Happly MySQL/.Net Coding.

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• Light following me around the room. Something is wrong with my shader!

  - by Robinson

  I'm trying to do a spot (Blinn) light, with falloff and attenuation. It seems to be working OK except I have a bit of a space problem. That is, whenever I move the camera the light moves to maintain the same relative position, rather than changing with the camera. This results in the light moving around, i.e. not always falling on the same surfaces. It's as if there's a flashlight attached to the camera. I'm transforming the lights beforehand into view space, so Light_Position and Light_Direction are already in eye space (I hope!). I made a little movie of what it looks like here: My camera rotating around a point inside a box. The light is fixed in the centre up and its "look at" point in a fixed position in front of it. As you can see, as the camera rotates around the origin (always looking at the centre), so don't think the box is rotating (!). The lighting follows it around. To start, some code. This is how I'm transforming the light into view space (it gets passed into the shader already in view space): // Compute eye-space light position. Math::Vector3d eyeSpacePosition = MyCamera->ViewMatrix() * MyLightPosition; MyShaderVariables->Set(MyLightPositionIndex, eyeSpacePosition); // Compute eye-space light direction vector. Math::Vector3d eyeSpaceDirection = Math::Unit(MyLightLookAt - MyLightPosition); MyCamera->ViewMatrixInverseTranspose().TransformNormal(eyeSpaceDirection); MyShaderVariables->Set(MyLightDirectionIndex, eyeSpaceDirection); Can anyone give me a clue as to what I'm doing wrong here? I think the light should remain looking at a fixed point on the box, regardless of the camera orientation. Here are the vertex and pixel shaders: /////////////////////////////////////////////////// // Vertex Shader /////////////////////////////////////////////////// #version 420 /////////////////////////////////////////////////// // Uniform Buffer Structures /////////////////////////////////////////////////// // Camera. layout (std140) uniform Camera { mat4 Camera_View; mat4 Camera_ViewInverseTranspose; mat4 Camera_Projection; }; // Matrices per model. layout (std140) uniform Model { mat4 Model_World; mat4 Model_WorldView; mat4 Model_WorldViewInverseTranspose; mat4 Model_WorldViewProjection; }; // Spotlight. layout (std140) uniform OmniLight { float Light_Intensity; vec3 Light_Position; vec3 Light_Direction; vec4 Light_Ambient_Colour; vec4 Light_Diffuse_Colour; vec4 Light_Specular_Colour; float Light_Attenuation_Min; float Light_Attenuation_Max; float Light_Cone_Min; float Light_Cone_Max; }; /////////////////////////////////////////////////// // Streams (per vertex) /////////////////////////////////////////////////// layout(location = 0) in vec3 attrib_Position; layout(location = 1) in vec3 attrib_Normal; layout(location = 2) in vec3 attrib_Tangent; layout(location = 3) in vec3 attrib_BiNormal; layout(location = 4) in vec2 attrib_Texture; /////////////////////////////////////////////////// // Output streams (per vertex) /////////////////////////////////////////////////// out vec3 attrib_Fragment_Normal; out vec4 attrib_Fragment_Position; out vec2 attrib_Fragment_Texture; out vec3 attrib_Fragment_Light; out vec3 attrib_Fragment_Eye; /////////////////////////////////////////////////// // Main /////////////////////////////////////////////////// void main() { // Transform normal into eye space attrib_Fragment_Normal = (Model_WorldViewInverseTranspose * vec4(attrib_Normal, 0.0)).xyz; // Transform vertex into eye space (world * view * vertex = eye) vec4 position = Model_WorldView * vec4(attrib_Position, 1.0); // Compute vector from eye space vertex to light (light is in eye space already) attrib_Fragment_Light = Light_Position - position.xyz; // Compute vector from the vertex to the eye (which is now at the origin). attrib_Fragment_Eye = -position.xyz; // Output texture coord. attrib_Fragment_Texture = attrib_Texture; // Compute vertex position by applying camera projection. gl_Position = Camera_Projection * position; } and the pixel shader: /////////////////////////////////////////////////// // Pixel Shader /////////////////////////////////////////////////// #version 420 /////////////////////////////////////////////////// // Samplers /////////////////////////////////////////////////// uniform sampler2D Map_Diffuse; /////////////////////////////////////////////////// // Global Uniforms /////////////////////////////////////////////////// // Material. layout (std140) uniform Material { vec4 Material_Ambient_Colour; vec4 Material_Diffuse_Colour; vec4 Material_Specular_Colour; vec4 Material_Emissive_Colour; float Material_Shininess; float Material_Strength; }; // Spotlight. layout (std140) uniform OmniLight { float Light_Intensity; vec3 Light_Position; vec3 Light_Direction; vec4 Light_Ambient_Colour; vec4 Light_Diffuse_Colour; vec4 Light_Specular_Colour; float Light_Attenuation_Min; float Light_Attenuation_Max; float Light_Cone_Min; float Light_Cone_Max; }; /////////////////////////////////////////////////// // Input streams (per vertex) /////////////////////////////////////////////////// in vec3 attrib_Fragment_Normal; in vec3 attrib_Fragment_Position; in vec2 attrib_Fragment_Texture; in vec3 attrib_Fragment_Light; in vec3 attrib_Fragment_Eye; /////////////////////////////////////////////////// // Result /////////////////////////////////////////////////// out vec4 Out_Colour; /////////////////////////////////////////////////// // Main /////////////////////////////////////////////////// void main(void) { // Compute N dot L. vec3 N = normalize(attrib_Fragment_Normal); vec3 L = normalize(attrib_Fragment_Light); vec3 E = normalize(attrib_Fragment_Eye); vec3 H = normalize(L + E); float NdotL = clamp(dot(L,N), 0.0, 1.0); float NdotH = clamp(dot(N,H), 0.0, 1.0); // Compute ambient term. vec4 ambient = Material_Ambient_Colour * Light_Ambient_Colour; // Diffuse. vec4 diffuse = texture2D(Map_Diffuse, attrib_Fragment_Texture) * Light_Diffuse_Colour * Material_Diffuse_Colour * NdotL; // Specular. float specularIntensity = pow(NdotH, Material_Shininess) * Material_Strength; vec4 specular = Light_Specular_Colour * Material_Specular_Colour * specularIntensity; // Light attenuation (so we don't have to use 1 - x, we step between Max and Min). float d = length(-attrib_Fragment_Light); float attenuation = smoothstep(Light_Attenuation_Max, Light_Attenuation_Min, d); // Adjust attenuation based on light cone. float LdotS = dot(-L, Light_Direction), CosI = Light_Cone_Min - Light_Cone_Max; attenuation *= clamp((LdotS - Light_Cone_Max) / CosI, 0.0, 1.0); // Final colour. Out_Colour = (ambient + diffuse + specular) * Light_Intensity * attenuation; }

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• Checking who is connected to your server, with PowerShell.

  - by Fatherjack

  There are many occasions when, as a DBA, you want to see who is connected to your SQL Server, along with how they are connecting and what sort of activities they are carrying out. I’m going to look at a couple of ways of getting this information and compare the effort required and the results achieved of each. SQL Server comes with a couple of stored procedures to help with this sort of task – sp_who and its undocumented counterpart sp_who2. There is also the pumped up version of these called sp_whoisactive, written by Adam Machanic which does way more than these procedures. I wholly recommend you try it out if you don’t already know how it works. When it comes to serious interrogation of your SQL Server activity then it is absolutely indispensable. Anyway, back to the point of this blog, we are going to look at getting the information from sp_who2 for a remote server. I wrote this Powershell script a week or so ago and was quietly happy with it for a while. I’m relatively new to Powershell so forgive both my rather low threshold for entertainment and the fact that something so simple is a moderate achievement for me. $Server = 'SERVERNAME' $SMOServer = New-Object Microsoft.SQLServer.Management.SMO.Server $Server # connection and query stuff         $ConnectionStr = "Server=$Server;Database=Master;Integrated Security=True" $Query = "EXEC sp_who2" $Connection = new-object system.Data.SQLClient.SQLConnection $Table = new-object "System.Data.DataTable" $Connection.connectionstring = $ConnectionStr try{ $Connection.open() $Command = $Connection.CreateCommand() $Command.commandtext = $Query $result = $Command.ExecuteReader() $Table.Load($result) } catch{ # Show error $error[0] | format-list -Force } $Title = "Data access processes (" + $Table.Rows.Count + ")" $Table | Out-GridView -Title $Title $Connection.close() So this is pretty straightforward, create an SMO object that represents our chosen server, define a connection to the database and a table object for the results when we get them, execute our query over the connection, load the results into our table object and then, if everything is error free display these results to the PowerShell grid viewer. The query simply gets the results of ‘EXEC sp_who2′ for us. Depending on how many connections there are will influence how long the query runs. The grid viewer lets me sort and search the results so it can be a pretty handy way to locate troublesome connections. Like I say, I was quite pleased with this, it seems a pretty simple script and was working well for me, I have added a few parameters to control the output and give me more specific details but then I see a script that uses the $SMOServer object itself to provide the process information and saves having to define the connection object and query specifications. $Server = 'SERVERNAME' $SMOServer = New-Object Microsoft.SQLServer.Management.SMO.Server $Server $Processes = $SMOServer.EnumProcesses() $Title = "SMO processes (" + $Processes.Rows.Count + ")" $Processes | Out-GridView -Title $Title Create the SMO object of our server and then call the EnumProcesses method to get all the process information from the server. Staggeringly simple! The results are a little different though. Some columns are the same and we can see the same basic information so my first thought was to which runs faster – so that I can get my results more quickly and also so that I place less stress on my server(s). PowerShell comes with a great way of testing this – the Measure-Command function. All you have to do is wrap your piece of code in Measure-Command {[your code here]} and it will spit out the time taken to execute the code. So, I placed both of the above methods of getting SQL Server process connections in two Measure-Command wrappers and pressed F5! The Powershell console goes blank for a while as the code is executed internally when Measure-Command is used but the grid viewer windows appear and the console shows this. You can take the output from Measure-Command and format it for easier reading but in a simple comparison like this we can simply cross refer the TotalMilliseconds values from the two result sets to see how the two methods performed. The query execution method (running EXEC sp_who2 ) is the first set of timings and the SMO EnumProcesses is the second. I have run these on a variety of servers and while the results vary from execution to execution I have never seen the SMO version slower than the other. The difference has varied and the time for both has ranged from sub-second as we see above to almost 5 seconds on other systems. This difference, I would suggest is partly due to the cost overhead of having to construct the data connection and so on where as the SMO EnumProcesses method has the connection to the server already in place and just needs to call back the process information. There is also the difference in the data sets to consider. Let’s take a look at what we get and where the two methods differ Query execution method (sp_who2) SMO EnumProcesses Description - Urn What looks like an XML or JSON representation of the server name and the process ID SPID Spid The process ID Status Status The status of the process Login Login The login name of the user executing the command HostName Host The name of the computer where the  process originated BlkBy BlockingSpid The SPID of a process that is blocking this one DBName Database The database that this process is connected to Command Command The type of command that is executing CPUTime Cpu The CPU activity related to this process DiskIO - The Disk IO activity related to this process LastBatch - The time the last batch was executed from this process. ProgramName Program The application that is facilitating the process connection to the SQL Server. SPID1 - In my experience this is always the same value as SPID. REQUESTID - In my experience this is always 0 - Name In my experience this is always the same value as SPID and so could be seen as analogous to SPID1 from sp_who2 - MemUsage An indication of the memory used by this process but I don’t know what it is measured in (bytes, Kb, Mb…) - IsSystem True or False depending on whether the process is internal to the SQL Server instance or has been created by an external connection requesting data. - ExecutionContextID In my experience this is always 0 so could be analogous to REQUESTID from sp_who2. Please note, these are my own very brief descriptions of these columns, detail can be found from MSDN for columns in the sp_who results here http://msdn.microsoft.com/en-GB/library/ms174313.aspx. Where the columns are common then I would use that description, in other cases then the information returned is purely for interpretation by the reader. Rather annoyingly both result sets have useful information that the other doesn’t. sp_who2 returns Disk IO and LastBatch information which is really useful but the SMO processes method give you IsSystem and MemUsage which have their place in fault diagnosis methods too. So which is better? On reflection I think I prefer to use the sp_who2 method primarily but knowing that the SMO Enumprocesses method is there when I need it is really useful and I’m sure I’ll use it regularly. I’m OK with the fact that it is the slower method because Measure-Command has shown me how close it is to the other option and that it really isn’t a large enough margin to matter.

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• Can Google Employees See My Saved Google Chrome Passwords?

  - by Jason Fitzpatrick

  Storing your passwords in your web browser seems like a great time saver, but are the passwords secure and inaccessible to others (even employees of the browser company) when squirreled away? Today’s Question & Answer session comes to us courtesy of SuperUser—a subdivision of Stack Exchange, a community-driven grouping of Q&A web sites. The Question SuperUser reader MMA is curious if Google employees have (or could have) access to the passwords he stores in Google Chrome: I understand that we are really tempted to save our passwords in Google Chrome. The likely benefit is two fold, You don’t need to (memorize and) input those long and cryptic passwords. These are available wherever you are once you log in to your Google account. The last point sparked my doubt. Since the password is available anywhere, the storage must in some central location, and this should be at Google. Now, my simple question is, can a Google employee see my passwords? Searching over the Internet revealed several articles/messages. Do you save passwords in Chrome? Maybe you should reconsider: Talks about your passwords being stolen by someone who has access to your computer account. Nothing mentioned about the central storage security and vulnerability. There is even a response from Chrome browser security tech lead about the first issue. Chrome’s insane password security strategy: Mostly along the same line. You can steal password from somebody if you have access to the computer account. How to Steal Passwords Saved in Google Chrome in 5 Simple Steps: Teaches you how to actually perform the act mentioned in the previous two when you have access to somebody else’s account. There are many more (including this one at this site), mostly along the same line, points, counter-points, huge debates. I refrain from mentioning them here, simply carry a search if you want to find them. Coming back to my original query, can a Google employee see my password? Since I can view the password using a simple button, definitely they can be unhashed (decrypted) even if encrypted. This is very different from the passwords saved in Unix-like OS’s where the saved password can never be seen in plain text. They use a one-way encryption algorithm to encrypt your passwords. This encrypted password is then stored in the passwd or shadow file. When you attempt to login, the password you type in is encrypted again and compared with the entry in the file that stores your passwords. If they match, it must be the same password, and you are allowed access. Thus, a superuser can change my password, can block my account, but he can never see my password. So are his concerns well founded or will a little insight dispel his worry? The Answer SuperUser contributor Zeel helps put his mind at ease: Short answer: No* Passwords stored on your local machine can be decrypted by Chrome, as long as your OS user account is logged in. And then you can view those in plain text. At first this seems horrible, but how did you think auto-fill worked? When that password field gets filled in, Chrome must insert the real password into the HTML form element – or else the page wouldn’t work right, and you could not submit the form. And if the connection to the website is not over HTTPS, the plain text is then sent over the internet. In other words, if chrome can’t get the plain text passwords, then they are totally useless. A one way hash is no good, because we need to use them. Now the passwords are in fact encrypted, the only way to get them back to plain text is to have the decryption key. That key is your Google password, or a secondary key you can set up. When you sign into Chrome and sync the Google servers will transmit the encrypted passwords, settings, bookmarks, auto-fill, etc, to your local machine. Here Chrome will decrypt the information and be able to use it. On Google’s end all that info is stored in its encrpyted state, and they do not have the key to decrypt it. Your account password is checked against a hash to log in to Google, and even if you let chrome remember it, that encrypted version is hidden in the same bundle as the other passwords, impossible to access. So an employee could probably grab a dump of the encrypted data, but it wouldn’t do them any good, since they would have no way to use it.* So no, Google employees can not** access your passwords, since they are encrypted on their servers. * However, do not forget that any system that can be accessed by an authorized user can be accessed by an unauthorized user. Some systems are easier to break than other, but none are fail-proof. . . That being said, I think I will trust Google and the millions they spend on security systems, over any other password storage solution. And heck, I’m a wimpy nerd, it would be easier to beat the passwords out of me than break Google’s encryption. ** I am also assuming that there isn’t a person who just happens to work for Google gaining access to your local machine. In that case you are screwed, but employment at Google isn’t actually a factor any more. Moral: Hit Win + L before leaving machine. While we agree with zeel that it’s a pretty safe bet (as long as your computer is not compromised) that your passwords are in fact safe while stored in Chrome, we prefer to encrypt all our logins and passwords in a LastPass vault. Have something to add to the explanation? Sound off in the the comments. Want to read more answers from other tech-savvy Stack Exchange users? Check out the full discussion thread here.     

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• Why does Google mark one e-mail as spam while does not the other?

  - by nKn

  I've a Postfix installation which works fine, I don't get any trouble with mails sent through a mail client (in my case, Thunderbird or RoundCube) when the To: address is a GMail account. However, I recently needed to use the PHPMailer tool to send some e-mails to some GMail accounts, so I configured an account to be used via SASL authentication + TLS. I don't mean mass mailing, just 2-3 mails. If I send the e-mail from the Thunderbird or RoundCube clients, the mail is not marked as spam. However, if I use PHPMailer, it always gets catalogued as spam. So I compared both headers and I just can't find the reason why the second is marked as spam while the first one is just ok. The first header sent from a mail client which is not marked as spam: Delivered-To: [email protected] Received: by 10.76.153.102 with SMTP id vf6csp230573oab; Tue, 19 Aug 2014 11:08:19 -0700 (PDT) X-Received: by 10.60.23.39 with SMTP id j7mr45544050oef.20.1408471699715; Tue, 19 Aug 2014 11:08:19 -0700 (PDT) Return-Path: <[email protected]> Received: from mail.mydomain.com (X.ip-92-222-X.eu. [92.222.X.X]) by mx.google.com with ESMTPS id t5si27115082oej.10.2014.08.19.11.08.18 for <[email protected]> (version=TLSv1.2 cipher=ECDHE-RSA-AES128-GCM-SHA256 bits=128/128); Tue, 19 Aug 2014 11:08:19 -0700 (PDT) Received-SPF: pass (google.com: domain of [email protected] designates 92.222.X.X as permitted sender) client-ip=92.222.X.X; Authentication-Results: mx.google.com; spf=pass (google.com: domain of [email protected] designates 92.222.X.X as permitted sender) [email protected]; dkim=pass (test mode) [email protected] Received: by mail.mydomain.com (Postfix, from userid 111) id D8F69120293D; Tue, 19 Aug 2014 19:08:17 +0100 (BST) DKIM-Signature: v=1; a=rsa-sha256; c=relaxed/relaxed; d=mydomain.com; s=mail; t=1408471697; bh=wKMX9gkQ7tCLv8ezrG5t4bICm/SSLQsNfTdZMToksWw=; h=Date:From:To:Subject:From; b=qRNcYVdmk+n3D1uuv0FInTx7/LzH2ojck9DgCmabFPvfke233lkojUOjezCUGx7iV DL8EayZ28mzzzHpB7ETeMzop/5OS3BmvFtGKVD9gzc78cDIFXTDoRFAnkRWDR2IOxI SOn5tiyODTFpkbDgJOndzQ6qL5K0S9ASNGCZrNL4= X-Spam-Checker-Version: SpamAssassin 3.4.0 (2014-02-07) on vpsX.ovh.net X-Spam-Level: X-Spam-Status: No, score=-1.0 required=3.0 tests=ALL_TRUSTED,T_DKIM_INVALID autolearn=ham autolearn_force=no version=3.4.0 Received: from [192.168.1.111] (unknown [77.231.X.X]) (using TLSv1 with cipher ECDHE-RSA-AES128-SHA (128/128 bits)) (No client certificate requested) (Authenticated sender: [email protected]) by mail.mydomain.com (Postfix) with ESMTPSA id 910341202624 for <[email protected]>; Tue, 19 Aug 2014 19:08:17 +0100 (BST) DKIM-Signature: v=1; a=rsa-sha256; c=relaxed/relaxed; d=mydomain.com; s=mail; t=1408471697; bh=wKMX9gkQ7tCLv8ezrG5t4bICm/SSLQsNfTdZMToksWw=; h=Date:From:To:Subject:From; b=qRNcYVdmk+n3D1uuv0FInTx7/LzH2ojck9DgCmabFPvfke233lkojUOjezCUGx7iV DL8EayZ28mzzzHpB7ETeMzop/5OS3BmvFtGKVD9gzc78cDIFXTDoRFAnkRWDR2IOxI SOn5tiyODTFpkbDgJOndzQ6qL5K0S9ASNGCZrNL4= Message-ID: <[email protected]> Date: Tue, 19 Aug 2014 19:08:24 +0100 From: My Name <[email protected]> User-Agent: Mozilla/5.0 (Windows NT 6.1; WOW64; rv:24.0) Gecko/20100101 Thunderbird/24.6.0 MIME-Version: 1.0 To: My other account <[email protected]> Subject: . Content-Type: text/plain; charset=ISO-8859-1; format=flowed Content-Transfer-Encoding: 7bit . The second header sent from PHPMailer which is always marked as spam: Delivered-To: [email protected] Received: by 10.76.153.102 with SMTP id vf6csp230832oab; Tue, 19 Aug 2014 11:12:10 -0700 (PDT) X-Received: by 10.60.121.67 with SMTP id li3mr44086252oeb.17.1408471930520; Tue, 19 Aug 2014 11:12:10 -0700 (PDT) Return-Path: <[email protected]> Received: from mail.mydomain.com (X.ip-92-222-X.eu. [92.222.X.X]) by mx.google.com with ESMTPS id w8si27103806obn.30.2014.08.19.11.12.10 for <[email protected]> (version=TLSv1.2 cipher=ECDHE-RSA-AES128-GCM-SHA256 bits=128/128); Tue, 19 Aug 2014 11:12:10 -0700 (PDT) Received-SPF: pass (google.com: domain of [email protected] designates 92.222.X.X as permitted sender) client-ip=92.222.X.X; Authentication-Results: mx.google.com; spf=pass (google.com: domain of [email protected] designates 92.222.X.X as permitted sender) [email protected]; dkim=pass (test mode) [email protected] Received: by mail.mydomain.com (Postfix, from userid 111) id 1999D120293D; Tue, 19 Aug 2014 19:12:09 +0100 (BST) DKIM-Signature: v=1; a=rsa-sha256; c=relaxed/relaxed; d=mydomain.com; s=mail; t=1408471929; bh=N1JuHq1S+8GrjHcEK3xn8P1JS+ygEBv5LKe0BiXuVJo=; h=Date:To:From:Reply-to:Subject:From; b=K7tcPyArzSTY91VEw6mAAFtDurSGwgTLGkfUZdC5mqsg0g/1LzmZkgwdjj4NdJa6M E2kDz3dwYN8FcZmbampJYFXxj4NQVtSnzjiWV40rpfOFqD2rXDGNIyB2QOjBZZ4WK3 7s4lyoJ/BrdQH4en8ctLVsDHed/KpHD4iGFEl67E= X-Spam-Checker-Version: SpamAssassin 3.4.0 (2014-02-07) on vpsX.ovh.net X-Spam-Level: X-Spam-Status: No, score=-1.0 required=3.0 tests=ALL_TRUSTED,T_DKIM_INVALID autolearn=ham autolearn_force=no version=3.4.0 Received: from rpi.mydomain.com (unknown [77.231.X.X]) (using TLSv1 with cipher ECDHE-RSA-AES256-SHA (256/256 bits)) (No client certificate requested) (Authenticated sender: [email protected]) by mail.mydomain.com (Postfix) with ESMTPSA id B42AF1202624 for <[email protected]>; Tue, 19 Aug 2014 19:12:08 +0100 (BST) DKIM-Signature: v=1; a=rsa-sha256; c=relaxed/relaxed; d=mydomain.com; s=mail; t=1408471928; bh=N1JuHq1S+8GrjHcEK3xn8P1JS+ygEBv5LKe0BiXuVJo=; h=Date:To:From:Reply-to:Subject:From; b=iXPM0tS36swudPTT4FOHHtPi5Ll6LbR60kNqCinZ8utcWoFE31SFTpoMEq5aCM5ux wQMdFiN8c6vkjRGabmvqFTTIbwJsrToHo/4+Lt5HEBoQQE2Y3T+xGmnmGAHCS6stKB yb7SVmtrIAsVtSMKA8VYIbmu2oYqV3afYt7g0OMQ= Date: Tue, 19 Aug 2014 20:12:07 +0200 To: [email protected] From: Trying another account <[email protected]> Reply-to: Trying another account <[email protected]> Subject: . Message-ID: <[email protected]> X-Priority: 3 X-Mailer: PHPMailer 5.1 (phpmailer.sourceforge.net) MIME-Version: 1.0 Content-Transfer-Encoding: 8bit Content-Type: text/plain; charset="UTF-8" . I also tried: Adding a User-Agent header to match the first one. Removing the X-Mailer header. No one of them made a difference. Is there some significant difference which is making the second e-mail to be marked as spam by Google?

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• MapReduce in DryadLINQ and PLINQ

  - by JoshReuben

  MapReduce See http://en.wikipedia.org/wiki/Mapreduce The MapReduce pattern aims to handle large-scale computations across a cluster of servers, often involving massive amounts of data. "The computation takes a set of input key/value pairs, and produces a set of output key/value pairs. The developer expresses the computation as two Func delegates: Map and Reduce. Map - takes a single input pair and produces a set of intermediate key/value pairs. The MapReduce function groups results by key and passes them to the Reduce function. Reduce - accepts an intermediate key I and a set of values for that key. It merges together these values to form a possibly smaller set of values. Typically just zero or one output value is produced per Reduce invocation. The intermediate values are supplied to the user's Reduce function via an iterator." the canonical MapReduce example: counting word frequency in a text file.     MapReduce using DryadLINQ see http://research.microsoft.com/en-us/projects/dryadlinq/ and http://connect.microsoft.com/Dryad DryadLINQ provides a simple and straightforward way to implement MapReduce operations. This The implementation has two primary components: A Pair structure, which serves as a data container. A MapReduce method, which counts word frequency and returns the top five words. The Pair Structure - Pair has two properties: Word is a string that holds a word or key. Count is an int that holds the word count. The structure also overrides ToString to simplify printing the results. The following example shows the Pair implementation. public struct Pair { private string word; private int count; public Pair(string w, int c) { word = w; count = c; } public int Count { get { return count; } } public string Word { get { return word; } } public override string ToString() { return word + ":" + count.ToString(); } } The MapReduce function  that gets the results. the input data could be partitioned and distributed across the cluster. 1. Creates a DryadTable<LineRecord> object, inputTable, to represent the lines of input text. For partitioned data, use GetPartitionedTable<T> instead of GetTable<T> and pass the method a metadata file. 2. Applies the SelectMany operator to inputTable to transform the collection of lines into collection of words. The String.Split method converts the line into a collection of words. SelectMany concatenates the collections created by Split into a single IQueryable<string> collection named words, which represents all the words in the file. 3. Performs the Map part of the operation by applying GroupBy to the words object. The GroupBy operation groups elements with the same key, which is defined by the selector delegate. This creates a higher order collection, whose elements are groups. In this case, the delegate is an identity function, so the key is the word itself and the operation creates a groups collection that consists of groups of identical words. 4. Performs the Reduce part of the operation by applying Select to groups. This operation reduces the groups of words from Step 3 to an IQueryable<Pair> collection named counts that represents the unique words in the file and how many instances there are of each word. Each key value in groups represents a unique word, so Select creates one Pair object for each unique word. IGrouping.Count returns the number of items in the group, so each Pair object's Count member is set to the number of instances of the word. 5. Applies OrderByDescending to counts. This operation sorts the input collection in descending order of frequency and creates an ordered collection named ordered. 6. Applies Take to ordered to create an IQueryable<Pair> collection named top, which contains the 100 most common words in the input file, and their frequency. Test then uses the Pair object's ToString implementation to print the top one hundred words, and their frequency.   public static IQueryable<Pair> MapReduce( string directory, string fileName, int k) { DryadDataContext ddc = new DryadDataContext("file://" + directory); DryadTable<LineRecord> inputTable = ddc.GetTable<LineRecord>(fileName); IQueryable<string> words = inputTable.SelectMany(x => x.line.Split(' ')); IQueryable<IGrouping<string, string>> groups = words.GroupBy(x => x); IQueryable<Pair> counts = groups.Select(x => new Pair(x.Key, x.Count())); IQueryable<Pair> ordered = counts.OrderByDescending(x => x.Count); IQueryable<Pair> top = ordered.Take(k);   return top; }   To Test: IQueryable<Pair> results = MapReduce(@"c:\DryadData\input", "TestFile.txt", 100); foreach (Pair words in results) Debug.Print(words.ToString());   Note: DryadLINQ applications can use a more compact way to represent the query: return inputTable         .SelectMany(x => x.line.Split(' '))         .GroupBy(x => x)         .Select(x => new Pair(x.Key, x.Count()))         .OrderByDescending(x => x.Count)         .Take(k);     MapReduce using PLINQ The pattern is relevant even for a single multi-core machine, however. We can write our own PLINQ MapReduce in a few lines. the Map function takes a single input value and returns a set of mapped values àLINQ's SelectMany operator. These are then grouped according to an intermediate key à LINQ GroupBy operator. The Reduce function takes each intermediate key and a set of values for that key, and produces any number of outputs per key à LINQ SelectMany again. We can put all of this together to implement MapReduce in PLINQ that returns a ParallelQuery<T> public static ParallelQuery<TResult> MapReduce<TSource, TMapped, TKey, TResult>( this ParallelQuery<TSource> source, Func<TSource, IEnumerable<TMapped>> map, Func<TMapped, TKey> keySelector, Func<IGrouping<TKey, TMapped>, IEnumerable<TResult>> reduce) { return source .SelectMany(map) .GroupBy(keySelector) .SelectMany(reduce); } the map function takes in an input document and outputs all of the words in that document. The grouping phase groups all of the identical words together, such that the reduce phase can then count the words in each group and output a word/count pair for each grouping: var files = Directory.EnumerateFiles(dirPath, "*.txt").AsParallel(); var counts = files.MapReduce( path => File.ReadLines(path).SelectMany(line => line.Split(delimiters)), word => word, group => new[] { new KeyValuePair<string, int>(group.Key, group.Count()) });

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• General Purpose ASP.NET Data Source Control

  - by Ricardo Peres

  OK, you already know about the ObjectDataSource control, so what’s wrong with it? Well, for once, it doesn’t pass any context to the SelectMethod, you only get the parameters supplied on the SelectParameters plus the desired ordering, starting page and maximum number of rows to display. Also, you must have two separate methods, one for actually retrieving the data, and the other for getting the total number of records (SelectCountMethod). Finally, you don’t get a chance to alter the supplied data before you bind it to the target control. I wanted something simple to use, and more similar to ASP.NET 4.5, where you can have the select method on the page itself, so I came up with CustomDataSource. Here’s how to use it (I chose a GridView, but it works equally well with any regular data-bound control): 1: <web:CustomDataSourceControl runat="server" ID="datasource" PageSize="10" OnData="OnData" /> 2: <asp:GridView runat="server" ID="grid" DataSourceID="datasource" DataKeyNames="Id" PageSize="10" AllowPaging="true" AllowSorting="true" /> The OnData event handler receives a DataEventArgs instance, which contains some properties that describe the desired paging location and size, and it’s where you return the data plus the total record count. Here’s a quick example: 1: protected void OnData(object sender, DataEventArgs e) 2: { 3: //just return some data 4: var data = Enumerable.Range(e.StartRowIndex, e.PageSize).Select(x => new { Id = x, Value = x.ToString(), IsPair = ((x % 2) == 0) }); 5: e.Data = data; 6: //the total number of records 7: e.TotalRowCount = 100; 8: } Here’s the code for the DataEventArgs: 1: [Serializable] 2: public class DataEventArgs : EventArgs 3: { 4: public DataEventArgs(Int32 pageSize, Int32 startRowIndex, String sortExpression, IOrderedDictionary parameters) 5: { 6: this.PageSize = pageSize; 7: this.StartRowIndex = startRowIndex; 8: this.SortExpression = sortExpression; 9: this.Parameters = parameters; 10: } 11:  12: public IEnumerable Data 13: { 14: get; 15: set; 16: } 17:  18: public IOrderedDictionary Parameters 19: { 20: get; 21: private set; 22: } 23:  24: public String SortExpression 25: { 26: get; 27: private set; 28: } 29:  30: public Int32 StartRowIndex 31: { 32: get; 33: private set; 34: } 35:  36: public Int32 PageSize 37: { 38: get; 39: private set; 40: } 41:  42: public Int32 TotalRowCount 43: { 44: get; 45: set; 46: } 47: } As you can guess, the StartRowIndex and PageSize receive the starting row and the desired page size, where the page size comes from the PageSize property on the markup. There’s also a SortExpression, which gets passed the sorted-by column and direction (if descending) and a dictionary containing all the values coming from the SelectParameters collection, if any. All of these are read only, and it is your responsibility to fill in the Data and TotalRowCount. The code for the CustomDataSource is very simple: 1: [NonVisualControl] 2: public class CustomDataSourceControl : DataSourceControl 3: { 4: public CustomDataSourceControl() 5: { 6: this.SelectParameters = new ParameterCollection(); 7: } 8:  9: protected override DataSourceView GetView(String viewName) 10: { 11: return (new CustomDataSourceView(this, viewName)); 12: } 13:  14: internal void GetData(DataEventArgs args) 15: { 16: this.OnData(args); 17: } 18:  19: protected virtual void OnData(DataEventArgs args) 20: { 21: EventHandler<DataEventArgs> data = this.Data; 22:  23: if (data != null) 24: { 25: data(this, args); 26: } 27: } 28:  29: [Browsable(false)] 30: [DesignerSerializationVisibility(DesignerSerializationVisibility.Visible)] 31: [PersistenceMode(PersistenceMode.InnerProperty)] 32: public ParameterCollection SelectParameters 33: { 34: get; 35: private set; 36: } 37:  38: public event EventHandler<DataEventArgs> Data; 39:  40: public Int32 PageSize 41: { 42: get; 43: set; 44: } 45: } Also, the code for the accompanying internal – as there is no need to use it from outside of its declaring assembly - data source view: 1: sealed class CustomDataSourceView : DataSourceView 2: { 3: private readonly CustomDataSourceControl dataSourceControl = null; 4:  5: public CustomDataSourceView(CustomDataSourceControl dataSourceControl, String viewName) : base(dataSourceControl, viewName) 6: { 7: this.dataSourceControl = dataSourceControl; 8: } 9:  10: public override Boolean CanPage 11: { 12: get 13: { 14: return (true); 15: } 16: } 17:  18: public override Boolean CanRetrieveTotalRowCount 19: { 20: get 21: { 22: return (true); 23: } 24: } 25:  26: public override Boolean CanSort 27: { 28: get 29: { 30: return (true); 31: } 32: } 33:  34: protected override IEnumerable ExecuteSelect(DataSourceSelectArguments arguments) 35: { 36: IOrderedDictionary parameters = this.dataSourceControl.SelectParameters.GetValues(HttpContext.Current, this.dataSourceControl); 37: DataEventArgs args = new DataEventArgs(this.dataSourceControl.PageSize, arguments.StartRowIndex, arguments.SortExpression, parameters); 38:  39: this.dataSourceControl.GetData(args); 40:  41: arguments.TotalRowCount = args.TotalRowCount; 42: arguments.MaximumRows = this.dataSourceControl.PageSize; 43: arguments.AddSupportedCapabilities(DataSourceCapabilities.Page | DataSourceCapabilities.Sort | DataSourceCapabilities.RetrieveTotalRowCount); 44: arguments.RetrieveTotalRowCount = true; 45:  46: if (!(args.Data is ICollection)) 47: { 48: return (args.Data.OfType<Object>().ToList()); 49: } 50: else 51: { 52: return (args.Data); 53: } 54: } 55: } As always, looking forward to hearing from you!

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• PASS: The Budget Process

  - by Bill Graziano

  Every fiscal year PASS creates a detailed budget.  This helps us set priorities and communicate to our members what we’re going to do in the upcoming year.  You can review the current budget on the PASS Governance page.  That page currently requires you to login but I’m talking with HQ to see if there are any legal issues with opening that up. The Accounting Team The PASS accounting team is two people.  The Executive Vice-President of Finance (“EVP”) and the PASS Accounting Manager.  Sandy Cherry is the accounting manager and works at PASS HQ.  Sandy has been with PASS since we switched management companies in 2007.  Throughout this document when I talk about any actual work related to the budget that’s all Sandy :)  She’s the glue that gets us through this process.  Last year we went through 32 iterations of the budget before the Board approved so it’s a pretty busy time for her us – well, mostly her. Fiscal Year The PASS fiscal year runs from July 1st through June 30th the following year.  Right now we’re in fiscal year 2011.  Our 2010 Summit actually occurred in FY2011.  We switched to this schedule from a calendar year in 2006.  Our goal was to have the Summit occur early in our fiscal year.  That gives us the rest of the year to handle any significant financial impact from the Summit.  If registrations are down we can reduce spending.  If registrations are up we can decide how much to increase our reserves and how much to spend.  Keep in mind that the Summit is budgeted to generate 82% of our revenue this year.  How it performs has a significant impact on our financials.  The other benefit of this fiscal year is that it matches the Microsoft fiscal year.  We sign an annual sponsorship agreement with Microsoft and it’s very helpful that our fiscal years match. This year our budget process will probably start in earnest in March or April.  I’d like to be done in early June so we can publish before July 1st.  I was late publishing it this year and I’m trying not to repeat that. Our Budget Our actual budget is an Excel spreadsheet with 36 sheets.  We remove some of those when we publish it since they include salary information.  The budget is broken up into various portfolios or departments.  We have 20 portfolios.  They include chapters, marketing, virtual chapters, marketing, etc.  Ideally each portfolio is assigned to a Board member.  Each portfolio also typically has a staff person assigned to it.  Portfolios that aren’t assigned to a Board member are monitored by HQ and the ExecVP-Finance (me).  These are typically smaller portfolios such as deferred membership or Summit futures.  (More on those in a later post.)  All portfolios are reviewed by all Board members during the budget approval process, when interim financials are released internally and at year-end. The Process Our first step is to budget revenues.  The Board determines a target attendee number.  We have formulas based on historical performance that convert that to an overall attendee revenue number.  Other revenue projections (such as vendor sponsorships) come from different parts of the organization.  I hope to have another post with more details on how we project revenues. The next step is to budget expenses.  Board members fill out a sample spreadsheet with their budget for the year.  They can add line items and notes describing what the amounts are for.  Each Board portfolio typically has from 10 to 30 line items.  Any new initiatives they want to pursue needs to be budgeted.  The Summit operations budget is managed by HQ.  It includes the cost for food, electrical, internet, etc.  Most of these come from our estimate of attendees and our contract with the convention center.  During this process the Board can ask for more or less to be spent on various line items.  For example, if we weren’t happy with the Internet at the last Summit we can ask them to look into different options and/or increasing the budget.  HQ will also make adjustments to these numbers based on what they see at the events and the feedback we receive on the surveys. After we have all the initial estimates we start reviewing the entire budget.  It is sent out to the Board and we can see what each portfolio requested and what the overall profit and loss number is.  We usually start with too much in expenses and need to cut.  In years past the Board started haggling over these numbers as a group.  This past year they decided I should take a first cut and present them with a reasonable budget and a list of what I changed.  That worked well and I think we’ll continue to do that in the future. We go through a number of iterations on the budget.  If I remember correctly, we went through 32 iterations before we passed the budget.  At each iteration various revenue and expense numbers can change.  Keep in mind that the PASS budget has 200+ line items spread over 20 portfolios.  Many of these depend on other numbers.  For example, if we decide increase the projected attendees that cascades through our budget.  At each iteration we list what changed and the impact.  Ideally these discussions will take place at a face-to-face Board meeting.  Many of them also take place over the phone.  Board members explain any increase they are asking for while performing due diligence on other budget requests.  Eventually a budget emerges and is passed. Publishing After the budget is passed we create a version without the formulas and salaries for posting on the web site.  Sandy also creates some charts to help our members understand the budget.  The EVP writes a nice little letter describing some of the changes from last year’s budget.  You can see my letter and our budget on the PASS Governance page. And then, eight months later, we start all over again.

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• Disaster, or Migration?

  - by Rob Farley

  This post is in two parts – technical and personal. And I should point out that it’s prompted in part by this month’s T-SQL Tuesday, hosted by Allen Kinsel. First, the technical: I’ve had a few conversations with people recently about migration – moving a SQL Server database from one box to another (sometimes, but not primarily, involving an upgrade). One question that tends to come up is that of downtime. Obviously there will be some period of time between the old server being available and the new one. The way that most people seem to think of migration is this: Build a new server. Stop people from using the old server. Take a backup of the old server Restore it on the new server. Reconfigure the client applications (or alternatively, configure the new server to use the same address as the old) Make the new server online. There are other things involved, such as testing, of course. But this is essentially the process that people tell me they’re planning to follow. The bit that I want to look at today (as you’ve probably guessed from my title) is the “backup and restore” section. If a SQL database is using the Simple Recovery Model, then the only restore option is the last database backup. This backup could be full or differential. The transaction log never gets backed up in the Simple Recovery Model. Instead, it truncates regularly to stay small. One that’s using the Full Recovery Model (or Bulk-Logged) won’t truncate its log – the log must be backed up regularly. This provides the benefit of having a lot more option available for restores. It’s a requirement for most systems of High Availability, because if you’re making sure that a spare box is up-and-running, ready to take over, then you have to be interested in the logs that are happening on the current box, rather than truncating them all the time. A High Availability system such as Mirroring, Replication or Log Shipping will initialise the spare machine by restoring a full database backup (and maybe a differential backup if available), and then any subsequent log backups. Once the secondary copy is close, transactions can be applied to keep the two in sync. The main aspect of any High Availability system is to have a redundant system that is ready to take over. So the similarity for migration should be obvious. If you need to move a database from one box to another, then introducing a High Availability mechanism can help. By turning on the Full Recovery Model and then taking a backup (so that the now-interesting logs have some context), logs start being kept, and are therefore available for getting the new box ready (even if it’s an upgraded version). When the migration is ready to occur, a failover can be done, letting the new server take over the responsibility of the old, just as if a disaster had happened. Except that this is a planned failover, not a disaster at all. There’s a fine line between a disaster and a migration. Failovers can be useful in patching, upgrading, maintenance, and more. Hopefully, even an unexpected disaster can be seen as just another failover, and there can be an opportunity there – perhaps to get some work done on the principal server to increase robustness. And if I’ve just set up a High Availability system for even the simplest of databases, it’s not necessarily a bad thing. :) So now the personal: It’s been an interesting time recently... June has been somewhat odd. A court case with which I was involved got resolved (through mediation). I can’t go into details, but my lawyers tell me that I’m allowed to say how I feel about it. The answer is ‘lousy’. I don’t regret pursuing it as long as I did – but in the end I had to make a decision regarding the commerciality of letting it continue, and I’m going to look forward to the days when the kind of money I spent on my lawyers is small change. Mind you, if I had a similar situation with an employer, I’d do the same again, but that doesn’t really stop me feeling frustrated about it. The following day I had to fly to country Victoria to see my grandmother, who wasn’t expected to last the weekend. She’s still around a week later as I write this, but her 92-year-old body has basically given up on her. She’s been a Christian all her life, and is looking forward to eternity. We’ll all miss her though, and it’s hard to see my family grieving. Then on Tuesday, I was driving back to the airport with my family to come home, when something really bizarre happened. We were travelling down the freeway, just pulled out to go past a truck (farm-truck sized, not a semi-trailer), when a car-sized mass of metal fell off it. It was something like an industrial air-conditioner, but from where I was sitting, it was just a mass of spinning metal, like something out of a movie (one friend described it as “holidays by Michael Bay”). Somehow, and I’m really don’t know how, the part of it nearest us bounced high enough to clear the car, and there wasn’t even a scratch. We pulled over the check, and I was just thanking God that we’d changed lanes when we had, and that we remained unharmed. I had all kinds of thoughts about what could’ve happened if we’d had something that size land on the windscreen... All this has drilled home that while I feel that I haven’t provided as well for the family as I could’ve done (like by pursuing an expensive legal case), I shouldn’t even consider that I have proper control over things. I get to live life, and make decisions based on what I feel is right at the time. But I’m not going to get everything right, and there will be things that feel like disasters, some which could’ve been in my control and some which are very much beyond my control. The case feels like something I could’ve pursued differently, a disaster that could’ve been avoided in some way. Gran dying is lousy of course. An accident on the freeway would have been awful. I need to recognise that the worst disasters are ones that I can’t affect, and that I need to look at things in context – perhaps seeing everything that happens as a migration instead. Life is never the same from one day to the next. Every event has a before and an after – sometimes it’s clearly positive, sometimes it’s not. I remember good events in my life (such as my wedding), and bad (such as the loss of my father when I was ten, or the back injury I had eight years ago). I’m not suggesting that I know how to view everything from the “God works all things for good” perspective, but I am trying to look at last week as a migration of sorts. Those things are behind me now, and the future is in God’s hands. Hopefully I’ve learned things, and will be able to live accordingly. I’ve come through this time now, and even though I’ll miss Gran, I’ll see her again one day, and the future is bright.

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• Why do we (really) program to interfaces?

  - by Kyle Burns

  One of the earliest lessons I was taught in Enterprise development was "always program against an interface".  This was back in the VB6 days and I quickly learned that no code would be allowed to move to the QA server unless my business objects and data access objects each are defined as an interface and have a matching implementation class.  Why?  "It's more reusable" was one answer.  "It doesn't tie you to a specific implementation" a slightly more knowing answer.  And let's not forget the discussion ending "it's a standard".  The problem with these responses was that senior people didn't really understand the reason we were doing the things we were doing and because of that, we were entirely unable to realize the intent behind the practice - we simply used interfaces and had a bunch of extra code to maintain to show for it. It wasn't until a few years later that I finally heard the term "Inversion of Control".  Simply put, "Inversion of Control" takes the creation of objects that used to be within the control (and therefore a responsibility of) of your component and moves it to some outside force.  For example, consider the following code which follows the old "always program against an interface" rule in the manner of many corporate development shops: 1: ICatalog catalog = new Catalog(); 2: Category[] categories = catalog.GetCategories(); In this example, I met the requirement of the rule by declaring the variable as ICatalog, but I didn't hit "it doesn't tie you to a specific implementation" because I explicitly created an instance of the concrete Catalog object.  If I want to test the functionality of the code I just wrote I have to have an environment in which Catalog can be created along with any of the resources upon which it depends (e.g. configuration files, database connections, etc) in order to test my functionality.  That's a lot of setup work and one of the things that I think ultimately discourages real buy-in of unit testing in many development shops. So how do I test my code without needing Catalog to work?  A very primitive approach I've seen is to change the line the instantiates catalog to read: 1: ICatalog catalog = new FakeCatalog();   once the test is run and passes, the code is switched back to the real thing.  This obviously poses a huge risk for introducing test code into production and in my opinion is worse than just keeping the dependency and its associated setup work.  Another popular approach is to make use of Factory methods which use an object whose "job" is to know how to obtain a valid instance of the object.  Using this approach, the code may look something like this: 1: ICatalog catalog = CatalogFactory.GetCatalog();   The code inside the factory is responsible for deciding "what kind" of catalog is needed.  This is a far better approach than the previous one, but it does make projects grow considerably because now in addition to the interface, the real implementation, and the fake implementation(s) for testing you have added a minimum of one factory (or at least a factory method) for each of your interfaces.  Once again, developers say "that's too complicated and has me writing a bunch of useless code" and quietly slip back into just creating a new Catalog and chalking any test failures up to "it will probably work on the server". This is where software intended specifically to facilitate Inversion of Control comes into play.  There are many libraries that take on the Inversion of Control responsibilities in .Net and most of them have many pros and cons.  From this point forward I'll discuss concepts from the standpoint of the Unity framework produced by Microsoft's Patterns and Practices team.  I'm primarily focusing on this library because it questions about it inspired this posting. At Unity's core and that of most any IoC framework is a catalog or registry of components.  This registry can be configured either through code or using the application's configuration file and in the most simple terms says "interface X maps to concrete implementation Y".  It can get much more complicated, but I want to keep things at the "what does it do" level instead of "how does it do it".  The object that exposes most of the Unity functionality is the UnityContainer.  This object exposes methods to configure the catalog as well as the Resolve<T> method which is used to obtain an instance of the type represented by T.  When using the Resolve<T> method, Unity does not necessarily have to just "new up" the requested object, but also can track dependencies of that object and ensure that the entire dependency chain is satisfied. There are three basic ways that I have seen Unity used within projects.  Those are through classes directly using the Unity container, classes requiring injection of dependencies, and classes making use of the Service Locator pattern. The first usage of Unity is when classes are aware of the Unity container and directly call its Resolve method whenever they need the services advertised by an interface.  The up side of this approach is that IoC is utilized, but the down side is that every class has to be aware that Unity is being used and tied directly to that implementation. Many developers don't like the idea of as close a tie to specific IoC implementation as is represented by using Unity within all of your classes and for the most part I agree that this isn't a good idea.  As an alternative, classes can be designed for Dependency Injection.  Dependency Injection is where a force outside the class itself manipulates the object to provide implementations of the interfaces that the class needs to interact with the outside world.  This is typically done either through constructor injection where the object has a constructor that accepts an instance of each interface it requires or through property setters accepting the service providers.  When using dependency, I lean toward the use of constructor injection because I view the constructor as being a much better way to "discover" what is required for the instance to be ready for use.  During resolution, Unity looks for an injection constructor and will attempt to resolve instances of each interface required by the constructor, throwing an exception of unable to meet the advertised needs of the class.  The up side of this approach is that the needs of the class are very clearly advertised and the class is unaware of which IoC container (if any) is being used.  The down side of this approach is that you're required to maintain the objects passed to the constructor as instance variables throughout the life of your object and that objects which coordinate with many external services require a lot of additional constructor arguments (this gets ugly and may indicate a need for refactoring). The final way that I've seen and used Unity is to make use of the ServiceLocator pattern, of which the Patterns and Practices team has also provided a Unity-compatible implementation.  When using the ServiceLocator, your class calls ServiceLocator.Retrieve in places where it would have called Resolve on the Unity container.  Like using Unity directly, it does tie you directly to the ServiceLocator implementation and makes your code aware that dependency injection is taking place, but it does have the up side of giving you the freedom to swap out the underlying IoC container if necessary.  I'm not hugely concerned with hiding IoC entirely from the class (I view this as a "nice to have"), so the single biggest problem that I see with the ServiceLocator approach is that it provides no way to proactively advertise needs in the way that constructor injection does, allowing more opportunity for difficult to track runtime errors. This blog entry has not been intended in any way to be a definitive work on IoC, but rather as something to spur thought about why we program to interfaces and some ways to reach the intended value of the practice instead of having it just complicate your code.  I hope that it helps somebody begin or continue a journey away from being a "Cargo Cult Programmer".

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• Per-pixel displacement mapping GLSL

  - by Chris

  Im trying to implement a per-pixel displacement shader in GLSL. I read through several papers and "tutorials" I found and ended up with trying to implement the approach NVIDIA used in their Cascade Demo (http://www.slideshare.net/icastano/cascades-demo-secrets) starting at Slide 82. At the moment I am completly stuck with following problem: When I am far away the displacement seems to work. But as more I move closer to my surface, the texture gets bent in x-axis and somehow it looks like there is a little bent in general in one direction. EDIT: I added a video: click I added some screen to illustrate the problem: Well I tried lots of things already and I am starting to get a bit frustrated as my ideas run out. I added my full VS and FS code: VS: #version 400 layout(location = 0) in vec3 IN_VS_Position; layout(location = 1) in vec3 IN_VS_Normal; layout(location = 2) in vec2 IN_VS_Texcoord; layout(location = 3) in vec3 IN_VS_Tangent; layout(location = 4) in vec3 IN_VS_BiTangent; uniform vec3 uLightPos; uniform vec3 uCameraDirection; uniform mat4 uViewProjection; uniform mat4 uModel; uniform mat4 uView; uniform mat3 uNormalMatrix; out vec2 IN_FS_Texcoord; out vec3 IN_FS_CameraDir_Tangent; out vec3 IN_FS_LightDir_Tangent; void main( void ) { IN_FS_Texcoord = IN_VS_Texcoord; vec4 posObject = uModel * vec4(IN_VS_Position, 1.0); vec3 normalObject = (uModel * vec4(IN_VS_Normal, 0.0)).xyz; vec3 tangentObject = (uModel * vec4(IN_VS_Tangent, 0.0)).xyz; //vec3 binormalObject = (uModel * vec4(IN_VS_BiTangent, 0.0)).xyz; vec3 binormalObject = normalize(cross(tangentObject, normalObject)); // uCameraDirection is the camera position, just bad named vec3 fvViewDirection = normalize( uCameraDirection - posObject.xyz); vec3 fvLightDirection = normalize( uLightPos.xyz - posObject.xyz ); IN_FS_CameraDir_Tangent.x = dot( tangentObject, fvViewDirection ); IN_FS_CameraDir_Tangent.y = dot( binormalObject, fvViewDirection ); IN_FS_CameraDir_Tangent.z = dot( normalObject, fvViewDirection ); IN_FS_LightDir_Tangent.x = dot( tangentObject, fvLightDirection ); IN_FS_LightDir_Tangent.y = dot( binormalObject, fvLightDirection ); IN_FS_LightDir_Tangent.z = dot( normalObject, fvLightDirection ); gl_Position = (uViewProjection*uModel) * vec4(IN_VS_Position, 1.0); } The VS just builds the TBN matrix, from incoming normal, tangent and binormal in world space. Calculates the light and eye direction in worldspace. And finally transforms the light and eye direction into tangent space. FS: #version 400 // uniforms uniform Light { vec4 fvDiffuse; vec4 fvAmbient; vec4 fvSpecular; }; uniform Material { vec4 diffuse; vec4 ambient; vec4 specular; vec4 emissive; float fSpecularPower; float shininessStrength; }; uniform sampler2D colorSampler; uniform sampler2D normalMapSampler; uniform sampler2D heightMapSampler; in vec2 IN_FS_Texcoord; in vec3 IN_FS_CameraDir_Tangent; in vec3 IN_FS_LightDir_Tangent; out vec4 color; vec2 TraceRay(in float height, in vec2 coords, in vec3 dir, in float mipmap){ vec2 NewCoords = coords; vec2 dUV = - dir.xy * height * 0.08; float SearchHeight = 1.0; float prev_hits = 0.0; float hit_h = 0.0; for(int i=0;i<10;i++){ SearchHeight -= 0.1; NewCoords += dUV; float CurrentHeight = textureLod(heightMapSampler,NewCoords.xy, mipmap).r; float first_hit = clamp((CurrentHeight - SearchHeight - prev_hits) * 499999.0,0.0,1.0); hit_h += first_hit * SearchHeight; prev_hits += first_hit; } NewCoords = coords + dUV * (1.0-hit_h) * 10.0f - dUV; vec2 Temp = NewCoords; SearchHeight = hit_h+0.1; float Start = SearchHeight; dUV *= 0.2; prev_hits = 0.0; hit_h = 0.0; for(int i=0;i<5;i++){ SearchHeight -= 0.02; NewCoords += dUV; float CurrentHeight = textureLod(heightMapSampler,NewCoords.xy, mipmap).r; float first_hit = clamp((CurrentHeight - SearchHeight - prev_hits) * 499999.0,0.0,1.0); hit_h += first_hit * SearchHeight; prev_hits += first_hit; } NewCoords = Temp + dUV * (Start - hit_h) * 50.0f; return NewCoords; } void main( void ) { vec3 fvLightDirection = normalize( IN_FS_LightDir_Tangent ); vec3 fvViewDirection = normalize( IN_FS_CameraDir_Tangent ); float mipmap = 0; vec2 NewCoord = TraceRay(0.1,IN_FS_Texcoord,fvViewDirection,mipmap); //vec2 ddx = dFdx(NewCoord); //vec2 ddy = dFdy(NewCoord); vec3 BumpMapNormal = textureLod(normalMapSampler, NewCoord.xy, mipmap).xyz; BumpMapNormal = normalize(2.0 * BumpMapNormal - vec3(1.0, 1.0, 1.0)); vec3 fvNormal = BumpMapNormal; float fNDotL = dot( fvNormal, fvLightDirection ); vec3 fvReflection = normalize( ( ( 2.0 * fvNormal ) * fNDotL ) - fvLightDirection ); float fRDotV = max( 0.0, dot( fvReflection, fvViewDirection ) ); vec4 fvBaseColor = textureLod( colorSampler, NewCoord.xy,mipmap); vec4 fvTotalAmbient = fvAmbient * fvBaseColor; vec4 fvTotalDiffuse = fvDiffuse * fNDotL * fvBaseColor; vec4 fvTotalSpecular = fvSpecular * ( pow( fRDotV, fSpecularPower ) ); color = ( fvTotalAmbient + (fvTotalDiffuse + fvTotalSpecular) ); } The FS implements the displacement technique in TraceRay method, while always using mipmap level 0. Most of the code is from NVIDIA sample and another paper I found on the web, so I guess there cannot be much wrong in here. At the end it uses the modified UV coords for getting the displaced normal from the normal map and the color from the color map. I looking forward for some ideas. Thanks in advance! Edit: Here is the code loading the heightmap: glTexImage2D(GL_TEXTURE_2D, 0, GL_RGBA, mWidth, mHeight, 0, GL_RGBA, GL_UNSIGNED_BYTE, mImageData); glGenerateMipmap(GL_TEXTURE_2D); //glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_MAG_FILTER, GL_LINEAR_MIPMAP_LINEAR); //glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_MIN_FILTER, GL_LINEAR_MIPMAP_LINEAR); //glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_WRAP_S, GL_REPEAT); //glTexParameteri(GL_TEXTURE_2D, GL_TEXTURE_WRAP_T, GL_REPEAT); Maybe something wrong in here?

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• We've completed the first iteration

  - by CliveT

  There are a lot of features in C# that are implemented by the compiler and not by the underlying platform. One such feature is a lambda expression. Since local variables cannot be accessed once the current method activation finishes, the compiler has to go out of its way to generate a new class which acts as a home for any variable whose lifetime needs to be extended past the activation of the procedure. Take the following example:     Random generator = new Random();     Func func = () = generator.Next(10); In this case, the compiler generates a new class called c_DisplayClass1 which is marked with the CompilerGenerated attribute. [CompilerGenerated] private sealed class c__DisplayClass1 {     // Fields     public Random generator;     // Methods     public int b__0()     {         return this.generator.Next(10);     } } Two quick comments on this: (i)    A display was the means that compilers for languages like Algol recorded the various lexical contours of the nested procedure activations on the stack. I imagine that this is what has led to the name. (ii)    It is a shame that the same attribute is used to mark all compiler generated classes as it makes it hard to figure out what they are being used for. Indeed, you could imagine optimisations that the runtime could perform if it knew that classes corresponded to certain high level concepts. We can see that the local variable generator has been turned into a field in the class, and the body of the lambda expression has been turned into a method of the new class. The code that builds the Func object simply constructs an instance of this class and initialises the fields to their initial values.     c__DisplayClass1 class2 = new c__DisplayClass1();     class2.generator = new Random();     Func func = new Func(class2.b__0); Reflector already contains code to spot this pattern of code and reproduce the form containing the lambda expression, so this is example is correctly decompiled. The use of compiler generated code is even more spectacular in the case of iterators. C# introduced the idea of a method that could automatically store its state between calls, so that it can pick up where it left off. The code can express the logical flow with yield return and yield break denoting places where the method should return a particular value and be prepared to resume.         {             yield return 1;             yield return 2;             yield return 3;         } Of course, there was already a .NET pattern for expressing the idea of returning a sequence of values with the computation proceeding lazily (in the sense that the work for the next value is executed on demand). This is expressed by the IEnumerable interface with its Current property for fetching the current value and the MoveNext method for forcing the computation of the next value. The sequence is terminated when this method returns false. The C# compiler links these two ideas together so that an IEnumerator returning method using the yield keyword causes the compiler to produce the implementation of an Iterator. Take the following piece of code.         IEnumerable GetItems()         {             yield return 1;             yield return 2;             yield return 3;         } The compiler implements this by defining a new class that implements a state machine. This has an integer state that records which yield point we should go to if we are resumed. It also has a field that records the Current value of the enumerator and a field for recording the thread. This latter value is used for optimising the creation of iterator instances. [CompilerGenerated] private sealed class d__0 : IEnumerable, IEnumerable, IEnumerator, IEnumerator, IDisposable {     // Fields     private int 1__state;     private int 2__current;     public Program 4__this;     private int l__initialThreadId; The body gets converted into the code to construct and initialize this new class. private IEnumerable GetItems() {     d__0 d__ = new d__0(-2);     d__.4__this = this;     return d__; } When the class is constructed we set the state, which was passed through as -2 and the current thread. public d__0(int 1__state) {     this.1__state = 1__state;     this.l__initialThreadId = Thread.CurrentThread.ManagedThreadId; } The state needs to be set to 0 to represent a valid enumerator and this is done in the GetEnumerator method which optimises for the usual case where the returned enumerator is only used once. IEnumerator IEnumerable.GetEnumerator() {     if ((Thread.CurrentThread.ManagedThreadId == this.l__initialThreadId)               && (this.1__state == -2))     {         this.1__state = 0;         return this;     } The state machine itself is implemented inside the MoveNext method. private bool MoveNext() {     switch (this.1__state)     {         case 0:             this.1__state = -1;             this.2__current = 1;             this.1__state = 1;             return true;         case 1:             this.1__state = -1;             this.2__current = 2;             this.1__state = 2;             return true;         case 2:             this.1__state = -1;             this.2__current = 3;             this.1__state = 3;             return true;         case 3:             this.1__state = -1;             break;     }     return false; } At each stage, the current value of the state is used to determine how far we got, and then we generate the next value which we return after recording the next state. Finally we return false from the MoveNext to signify the end of the sequence. Of course, that example was really simple. The original method body didn't have any local variables. Any local variables need to live between the calls to MoveNext and so they need to be transformed into fields in much the same way that we did in the case of the lambda expression. More complicated MoveNext methods are required to deal with resources that need to be disposed when the iterator finishes, and sometimes the compiler uses a temporary variable to hold the return value. Why all of this explanation? We've implemented the de-compilation of iterators in the current EAP version of Reflector (7). This contrasts with previous version where all you could do was look at the MoveNext method and try to figure out the control flow. There's a fair amount of things we have to do. We have to spot the use of a CompilerGenerated class which implements the Enumerator pattern. We need to go to the class and figure out the fields corresponding to the local variables. We then need to go to the MoveNext method and try to break it into the various possible states and spot the state transitions. We can then take these pieces and put them back together into an object model that uses yield return to show the transition points. After that Reflector can carry on optimising using its usual optimisations. The pattern matching is currently a little too sensitive to changes in the code generation, and we only do a limited analysis of the MoveNext method to determine use of the compiler generated fields. In some ways, it is a pity that iterators are compiled away and there is no metadata that reflects the original intent. Without it, we are always going to dependent on our knowledge of the compiler's implementation. For example, we have noticed that the Async CTP changes the way that iterators are code generated, so we'll have to do some more work to support that. However, with that warning in place, we seem to do a reasonable job of decompiling the iterators that are built into the framework. Hopefully, the EAP will give us a chance to find examples where we don't spot the pattern correctly or regenerate the wrong code, and we can improve things. Please give it a go, and report any problems.

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• Cancelling Route Navigation in AngularJS Controllers

  - by dwahlin

  If you’re new to AngularJS check out my AngularJS in 60-ish Minutes video tutorial or download the free eBook. Also check out The AngularJS Magazine for up-to-date information on using AngularJS to build Single Page Applications (SPAs). Routing provides a nice way to associate views with controllers in AngularJS using a minimal amount of code. While a user is normally able to navigate directly to a specific route, there may be times when a user triggers a route change before they’ve finalized an important action such as saving data. In these types of situations you may want to cancel the route navigation and ask the user if they’d like to finish what they were doing so that their data isn’t lost. In this post I’ll talk about a technique that can be used to accomplish this type of routing task.   The $locationChangeStart Event When route navigation occurs in an AngularJS application a few events are raised. One is named $locationChangeStart and the other is named $routeChangeStart (there are other events as well). At the current time (version 1.2) the $routeChangeStart doesn’t provide a way to cancel route navigation, however, the $locationChangeStart event can be used to cancel navigation. If you dig into the AngularJS core script you’ll find the following code that shows how the $locationChangeStart event is raised as the $browser object’s onUrlChange() function is invoked:   $browser.onUrlChange(function (newUrl) { if ($location.absUrl() != newUrl) { if ($rootScope.$broadcast('$locationChangeStart', newUrl, $location.absUrl()).defaultPrevented) { $browser.url($location.absUrl()); return; } $rootScope.$evalAsync(function () { var oldUrl = $location.absUrl(); $location.$$parse(newUrl); afterLocationChange(oldUrl); }); if (!$rootScope.$$phase) $rootScope.$digest(); } }); .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } The key part of the code is the call to $broadcast. This call broadcasts the $locationChangeStart event to all child scopes so that they can be notified before a location change is made. To handle the $locationChangeStart event you can use the $rootScope.on() function. For this example I’ve added a call to $on() into a function that is called immediately after the controller is invoked:   function init() { //initialize data here.. //Make sure they're warned if they made a change but didn't save it //Call to $on returns a "deregistration" function that can be called to //remove the listener (see routeChange() for an example of using it) onRouteChangeOff = $rootScope.$on('$locationChangeStart', routeChange); } This code listens for the $locationChangeStart event and calls routeChange() when it occurs. The value returned from calling $on is a “deregistration” function that can be called to detach from the event. In this case the deregistration function is named onRouteChangeOff (it’s accessible throughout the controller). You’ll see how the onRouteChangeOff function is used in just a moment.   Cancelling Route Navigation The routeChange() callback triggered by the $locationChangeStart event displays a modal dialog similar to the following to prompt the user:     Here’s the code for routeChange(): function routeChange(event, newUrl) { //Navigate to newUrl if the form isn't dirty if (!$scope.editForm.$dirty) return; var modalOptions = { closeButtonText: 'Cancel', actionButtonText: 'Ignore Changes', headerText: 'Unsaved Changes', bodyText: 'You have unsaved changes. Leave the page?' }; modalService.showModal({}, modalOptions).then(function (result) { if (result === 'ok') { onRouteChangeOff(); //Stop listening for location changes $location.path(newUrl); //Go to page they're interested in } }); //prevent navigation by default since we'll handle it //once the user selects a dialog option event.preventDefault(); return; } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Looking at the parameters of routeChange() you can see that it accepts an event object and the new route that the user is trying to navigate to. The event object is used to prevent navigation since we need to prompt the user before leaving the current view. Notice the call to event.preventDefault() at the end of the function. The modal dialog is shown by calling modalService.showModal() (see my previous post for more information about the custom modalService that acts as a wrapper around Angular UI Bootstrap’s $modal service). If the user selects “Ignore Changes” then their changes will be discarded and the application will navigate to the route they intended to go to originally. This is done by first detaching from the $locationChangeStart event by calling onRouteChangeOff() (recall that this is the function returned from the call to $on()) so that we don’t get stuck in a never ending cycle where the dialog continues to display when they click the “Ignore Changes” button. A call is then made to $location.path(newUrl) to handle navigating to the target view. If the user cancels the operation they’ll stay on the current view. Conclusion The key to canceling routes is understanding how to work with the $locationChangeStart event and cancelling it so that route navigation doesn’t occur. I’m hoping that in the future the same type of task can be done using the $routeChangeStart event but for now this code gets the job done. You can see this code in action in the Customer Manager application available on Github (specifically the customerEdit view). Learn more about the application here.

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• Agilist, Heal Thyself!

  - by Dylan Smith

  I’ve been meaning to blog about a great experience I had earlier in the year at Prairie Dev Con Calgary.  Myself and Steve Rogalsky did a session that we called “Agilist, Heal Thyself!”.  We used a format that was new to me, but that Steve had seen used at another conference.  What we did was start by asking the audience to give us a list of challenges they had had when adopting agile.  We wrote them all down, then had everybody vote on the most interesting ones.  Then we split into two groups, and each group was assigned one of the agile challenges.  We had 20 minutes to discuss the challenge, and suggest solutions or approaches to improve things.  At the end of the 20 minutes, each of the groups gave a brief summary of their discussion and learning's, then we mixed up the groups and repeated with another 2 challenges. The 2 groups I was part of had some really interesting discussions, and suggestions: Unfinished Stories at the end of Sprints The first agile challenge we tackled, was something that every single Scrum team I have worked with has struggled with.  What happens when you get to the end of a Sprint, and there are some stories that are only partially completed.  The team in question was getting very de-moralized as they felt that every Sprint was a failure as they never had a set of fully completed stories. How do you avoid this? and/or what do you do when it happens? There were 2 pieces of advice that were well received: 1. Try to bring stories to completion before starting new ones.  This is advice I give all my Scrum teams.  If you have a 3-week sprint, what happens all too often is you get to the end of week 2, and a lot of stories are almost done; but almost none are completely done.  This is a Bad Thing.  I encourage the teams I work with to only start a new story as a very last resort.  If you finish your task look at the stories in progress and see if there’s anything you can do to help before moving onto a new story.  In the daily standup, put a focus on seeing what stories got completed yesterday, if a few days go by with none getting completed, be sure this fact is visible to the team and do something about it.  Something I’ve been doing recently is introducing WIP (Work In Progress) limits while using Scrum.  My current team has 2-week sprints, and we usually have about a dozen or stories in a sprint.  We instituted a WIP limit of 4 stories.  If 4 stories have been started but not finished then nobody is allowed to start new stories.  This made it obvious very quickly that our QA tasks were our bottleneck (we have 4 devs, but only 1.5 testers).  The WIP limit forced the developers to start to pickup QA tasks before moving onto the next dev tasks, and we ended our sprints with many more stories completely finished than we did before introducing WIP limits. 2. Rather than using time-boxed sprints, why not just do away with them altogether and go to a continuous flow type approach like KanBan.  Limit WIP to keep things under control, but don’t have a fixed time box at the end of which all tasks are supposed to be done.  This eliminates the problem almost entirely.  At some points in the project (releases) you need to be able to burn down all the half finished stories to get a stable release build, but this probably occurs less often than every sprint, and there are alternative approaches to achieve it using branching strategies rather than forcing your team to try to get to Zero WIP every 2-weeks (e.g. when you are ready for a release, create a new branch for any new stories, but finish all existing stories in the current branch and release it). Trying to Introduce Agile into a team with previous Bad Agile Experiences One of the agile adoption challenges somebody described, was he was in a leadership role on a team he had recently joined – lets call him Dave.  This team was currently very waterfall in their ALM process, but they were about to start on a new green-field project.  Dave wanted to use this new project as an opportunity to do things the “right way”, using an Agile methodology like Scrum, adopting TDD, automated builds, proper branching strategies, etc.  The problem he was facing is everybody else on the team had previously gone through an “Agile Adoption” that was a horrible failure.  Dave blamed this failure on the consultant brought in previously to lead this agile transition, but regardless of the reason, the team had very negative feelings towards agile, and was very resistant to trying it out again.  Dave possibly had the authority to try to force the team to adopt Agile practices, but we all know that doesn’t work very well.  What was Dave to do? Ultimately, the best advice was to question *why* did Dave want to adopt all these various practices. Rather than trying to convince his team that these were the “right way” to run a dev project, and trying to do a Big Bang approach to introducing change.  He would be better served by identifying problems the team currently faces, have a discussion with the team to get everybody to agree that specific problems existed, then have an open discussion about ways to address those problems.  This way Dave could incrementally introduce agile practices, and he doesn’t even need to identify them as “agile” practices if he doesn’t want to.  For example, when we discussed with Dave, he said probably the teams biggest problem was long periods without feedback from users, then finding out too late that the software is not going to meet their needs.  Rather than Dave jumping right to introducing Scrum and all it entails, it would be easier to get buy-in from team if he framed it as a discussion of existing problems, and brainstorming possible solutions.  And possibly most importantly, don’t try to do massive changes all at once with a team that has not bought-into those changes.  Taking an incremental approach has a greater chance of success. I see something similar in my day job all the time too.  Clients who for one reason or another claim to not be fans of agile (or not ready for agile yet).  But then they go on to ask me to help them get shorter feedback cycles, quicker delivery cycles, iterative development processes, etc.  It’s kind of funny at times, sometimes you just need to phrase the suggestions in terms they are using and avoid the word “agile”. PS – I haven’t blogged all that much over the past couple of years, but in an attempt to motivate myself, a few of us have accepted a blogger challenge.  There’s 6 of us who have all put some money into a pool, and the agreement is that we each need to blog at least once every 2-weeks.  The first 2-week period that we miss we’re eliminated.  Last person standing gets the money.  So expect at least one blog post every couple of weeks for the near future (I hope!).  And check out the blogs of the other 5 people in this blogger challenge: Steve Rogalsky: http://winnipegagilist.blogspot.ca Aaron Kowall: http://www.geekswithblogs.net/caffeinatedgeek Tyler Doerkson: http://blog.tylerdoerksen.com David Alpert: http://www.spinthemoose.com Dave White: http://www.agileramblings.com (note: site not available yet.  should be shortly or he owes me some money!)

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• Multiple setInterval in a HTML5 Canvas game

  - by kushsolitary

  I'm trying to achieve multiple animations in a game that I am creating using Canvas (it is a simple ping-pong game). This is my first game and I am new to canvas but have created a few experiments before so I have a good knowledge about how canvas work. First, take a look at the game here. The problem is, when the ball hits the paddle, I want a burst of n particles at the point of contact but that doesn't came right. Even if I set the particles number to 1, they just keep coming from the point of contact and then hides automatically after some time. Also, I want to have the burst on every collision but it occurs on first collision only. I am pasting the code here: //Initialize canvas var canvas = document.getElementById("canvas"), ctx = canvas.getContext("2d"), W = window.innerWidth, H = window.innerHeight, particles = [], ball = {}, paddles = [2], mouse = {}, points = 0, fps = 60, particlesCount = 50, flag = 0, particlePos = {}; canvas.addEventListener("mousemove", trackPosition, true); //Set it's height and width to full screen canvas.width = W; canvas.height = H; //Function to paint canvas function paintCanvas() { ctx.globalCompositeOperation = "source-over"; ctx.fillStyle = "black"; ctx.fillRect(0, 0, W, H); } //Create two paddles function createPaddle(pos) { //Height and width this.h = 10; this.w = 100; this.x = W/2 - this.w/2; this.y = (pos == "top") ? 0 : H - this.h; } //Push two paddles into the paddles array paddles.push(new createPaddle("bottom")); paddles.push(new createPaddle("top")); //Setting up the parameters of ball ball = { x: 2, y: 2, r: 5, c: "white", vx: 4, vy: 8, draw: function() { ctx.beginPath(); ctx.fillStyle = this.c; ctx.arc(this.x, this.y, this.r, 0, Math.PI*2, false); ctx.fill(); } }; //Function for creating particles function createParticles(x, y) { this.x = x || 0; this.y = y || 0; this.radius = 0.8; this.vx = -1.5 + Math.random()*3; this.vy = -1.5 + Math.random()*3; } //Draw everything on canvas function draw() { paintCanvas(); for(var i = 0; i < paddles.length; i++) { p = paddles[i]; ctx.fillStyle = "white"; ctx.fillRect(p.x, p.y, p.w, p.h); } ball.draw(); update(); } //Mouse Position track function trackPosition(e) { mouse.x = e.pageX; mouse.y = e.pageY; } //function to increase speed after every 5 points function increaseSpd() { if(points % 4 == 0) { ball.vx += (ball.vx < 0) ? -1 : 1; ball.vy += (ball.vy < 0) ? -2 : 2; } } //function to update positions function update() { //Move the paddles on mouse move if(mouse.x && mouse.y) { for(var i = 1; i < paddles.length; i++) { p = paddles[i]; p.x = mouse.x - p.w/2; } } //Move the ball ball.x += ball.vx; ball.y += ball.vy; //Collision with paddles p1 = paddles[1]; p2 = paddles[2]; if(ball.y >= p1.y - p1.h) { if(ball.x >= p1.x && ball.x <= (p1.x - 2) + (p1.w + 2)){ ball.vy = -ball.vy; points++; increaseSpd(); particlePos.x = ball.x, particlePos.y = ball.y; flag = 1; } } else if(ball.y <= p2.y + 2*p2.h) { if(ball.x >= p2.x && ball.x <= (p2.x - 2) + (p2.w + 2)){ ball.vy = -ball.vy; points++; increaseSpd(); particlePos.x = ball.x, particlePos.y = ball.y; flag = 1; } } //Collide with walls if(ball.x >= W || ball.x <= 0) ball.vx = -ball.vx; if(ball.y > H || ball.y < 0) { clearInterval(int); } if(flag == 1) { setInterval(emitParticles(particlePos.x, particlePos.y), 1000/fps); } } function emitParticles(x, y) { for(var k = 0; k < particlesCount; k++) { particles.push(new createParticles(x, y)); } counter = particles.length; for(var j = 0; j < particles.length; j++) { par = particles[j]; ctx.beginPath(); ctx.fillStyle = "white"; ctx.arc(par.x, par.y, par.radius, 0, Math.PI*2, false); ctx.fill(); par.x += par.vx; par.y += par.vy; par.radius -= 0.02; if(par.radius < 0) { counter--; if(counter < 0) particles = []; } } } var int = setInterval(draw, 1000/fps); Now, my function for emitting particles is on line 156, and I have called this function on line 151. The problem here can be because of I am not resetting the flag variable but I tried doing that and got more weird results. You can check that out here. By resetting the flag variable, the problem of infinite particles gets resolved but now they only animate and appear when the ball collides with the paddles. So, I am now out of any solution.

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• parallel_for_each from amp.h – part 1

  - by Daniel Moth

  This posts assumes that you've read my other C++ AMP posts on index<N> and extent<N>, as well as about the restrict modifier. It also assumes you are familiar with C++ lambdas (if not, follow my links to C++ documentation). Basic structure and parameters Now we are ready for part 1 of the description of the new overload for the concurrency::parallel_for_each function. The basic new parallel_for_each method signature returns void and accepts two parameters: a grid<N> (think of it as an alias to extent) a restrict(direct3d) lambda, whose signature is such that it returns void and accepts an index of the same rank as the grid So it looks something like this (with generous returns for more palatable formatting) assuming we are dealing with a 2-dimensional space: // some_code_A parallel_for_each( g, // g is of type grid<2> [ ](index<2> idx) restrict(direct3d) { // kernel code } ); // some_code_B The parallel_for_each will execute the body of the lambda (which must have the restrict modifier), on the GPU. We also call the lambda body the "kernel". The kernel will be executed multiple times, once per scheduled GPU thread. The only difference in each execution is the value of the index object (aka as the GPU thread ID in this context) that gets passed to your kernel code. The number of GPU threads (and the values of each index) is determined by the grid object you pass, as described next. You know that grid is simply a wrapper on extent. In this context, one way to think about it is that the extent generates a number of index objects. So for the example above, if your grid was setup by some_code_A as follows: extent<2> e(2,3); grid<2> g(e); ...then given that: e.size()==6, e[0]==2, and e[1]=3 ...the six index<2> objects it generates (and hence the values that your lambda would receive) are:    (0,0) (1,0) (0,1) (1,1) (0,2) (1,2) So what the above means is that the lambda body with the algorithm that you wrote will get executed 6 times and the index<2> object you receive each time will have one of the values just listed above (of course, each one will only appear once, the order is indeterminate, and they are likely to call your code at the same exact time). Obviously, in real GPU programming, you'd typically be scheduling thousands if not millions of threads, not just 6. If you've been following along you should be thinking: "that is all fine and makes sense, but what can I do in the kernel since I passed nothing else meaningful to it, and it is not returning any values out to me?" Passing data in and out It is a good question, and in data parallel algorithms indeed you typically want to pass some data in, perform some operation, and then typically return some results out. The way you pass data into the kernel, is by capturing variables in the lambda (again, if you are not familiar with them, follow the links about C++ lambdas), and the way you use data after the kernel is done executing is simply by using those same variables. In the example above, the lambda was written in a fairly useless way with an empty capture list: [ ](index<2> idx) restrict(direct3d), where the empty square brackets means that no variables were captured. If instead I write it like this [&](index<2> idx) restrict(direct3d), then all variables in the some_code_A region are made available to the lambda by reference, but as soon as I try to use any of those variables in the lambda, I will receive a compiler error. This has to do with one of the direct3d restrictions, where only one type can be capture by reference: objects of the new concurrency::array class that I'll introduce in the next post (suffice for now to think of it as a container of data). If I write the lambda line like this [=](index<2> idx) restrict(direct3d), all variables in the some_code_A region are made available to the lambda by value. This works for some types (e.g. an integer), but not for all, as per the restrictions for direct3d. In particular, no useful data classes work except for one new type we introduce with C++ AMP: objects of the new concurrency::array_view class, that I'll introduce in the post after next. Also note that if you capture some variable by value, you could use it as input to your algorithm, but you wouldn’t be able to observe changes to it after the parallel_for_each call (e.g. in some_code_B region since it was passed by value) – the exception to this rule is the array_view since (as we'll see in a future post) it is a wrapper for data, not a container. Finally, for completeness, you can write your lambda, e.g. like this [av, &ar](index<2> idx) restrict(direct3d) where av is a variable of type array_view and ar is a variable of type array - the point being you can be very specific about what variables you capture and how. So it looks like from a large data perspective you can only capture array and array_view objects in the lambda (that is how you pass data to your kernel) and then use the many threads that call your code (each with a unique index) to perform some operation. You can also capture some limited types by value, as input only. When the last thread completes execution of your lambda, the data in the array_view or array are ready to be used in the some_code_B region. We'll talk more about all this in future posts… (a)synchronous Please note that the parallel_for_each executes as if synchronous to the calling code, but in reality, it is asynchronous. I.e. once the parallel_for_each call is made and the kernel has been passed to the runtime, the some_code_B region continues to execute immediately by the CPU thread, while in parallel the kernel is executed by the GPU threads. However, if you try to access the (array or array_view) data that you captured in the lambda in the some_code_B region, your code will block until the results become available. Hence the correct statement: the parallel_for_each is as-if synchronous in terms of visible side-effects, but asynchronous in reality.   That's all for now, we'll revisit the parallel_for_each description, once we introduce properly array and array_view – coming next. Comments about this post by Daniel Moth welcome at the original blog.

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• Passing a parameter so that it cannot be changed – C#

  - by nmarun

  I read this requirement of not allowing a user to change the value of a property passed as a parameter to a method. In C++, as far as I could recall (it’s been over 10 yrs, so I had to refresh memory), you can pass ‘const’ to a function parameter and this ensures that the parameter cannot be changed inside the scope of the function. There’s no such direct way of doing this in C#, but that does not mean it cannot be done!! Ok, so this ‘not-so-direct’ technique depends on the type of the parameter – a simple property or a collection. Parameter as a simple property: This is quite easy (and you might have guessed it already). Bulent Ozkir clearly explains how this can be done here. Parameter as a collection property: Obviously the above does not work if the parameter is a collection of some type. Let’s dig-in. Suppose I need to create a collection of type KeyTitle as defined below. 1: public class KeyTitle 2: { 3: public int Key { get; set; } 4: public string Title { get; set; } 5: } My class is declared as below: 1: public class Class1 2: { 3: public Class1() 4: { 5: MyKeyTitleList = new List<KeyTitle>(); 6: } 7: 8: public List<KeyTitle> MyKeyTitleList { get; set; } 9: public ReadOnlyCollection<KeyTitle> ReadonlyKeyTitleCollection 10: { 11: // .AsReadOnly creates a ReadOnlyCollection<> type 12: get { return MyKeyTitleList.AsReadOnly(); } 13: } 14: } See the .AsReadOnly() method used in the second property? As MSDN says it: “Returns a read-only IList<T> wrapper for the current collection.” Knowing this, I can implement my code as: 1: public static void Main() 2: { 3: Class1 class1 = new Class1(); 4: class1.MyKeyTitleList.Add(new KeyTitle { Key = 1, Title = "abc" }); 5: class1.MyKeyTitleList.Add(new KeyTitle { Key = 2, Title = "def" }); 6: class1.MyKeyTitleList.Add(new KeyTitle { Key = 3, Title = "ghi" }); 7: class1.MyKeyTitleList.Add(new KeyTitle { Key = 4, Title = "jkl" }); 8:  9: TryToModifyCollection(class1.MyKeyTitleList.AsReadOnly()); 10:  11: Console.ReadLine(); 12: } 13:  14: private static void TryToModifyCollection(ReadOnlyCollection<KeyTitle> readOnlyCollection) 15: { 16: // can only read 17: for (int i = 0; i < readOnlyCollection.Count; i++) 18: { 19: Console.WriteLine("{0} - {1}", readOnlyCollection[i].Key, readOnlyCollection[i].Title); 20: } 21: // Add() - not allowed 22: // even the indexer does not have a setter 23: } The output is as expected: The below image shows two things. In the first line, I’ve tried to access an element in my read-only collection through an indexer. It shows that the ReadOnlyCollection<> does not have a setter on the indexer. The second line tells that there’s no ‘Add()’ method for this type of collection. The capture below shows there’s no ‘Remove()’ method either, there-by eliminating all ways of modifying a collection. Mission accomplished… right? Now, even if you have a collection of different type, all you need to do is to somehow cast (used loosely) it to a List<> and then do a .AsReadOnly() to get a ReadOnlyCollection of your custom collection type. As an example, if you have an IDictionary<int, string>, you can create a List<T> of this type with a wrapper class (KeyTitle in our case). 1: public IDictionary<int, string> MyDictionary { get; set; } 2:  3: public ReadOnlyCollection<KeyTitle> ReadonlyDictionary 4: { 5: get 6: { 7: return (from item in MyDictionary 8: select new KeyTitle 9: { 10: Key = item.Key, 11: Title = item.Value, 12: }).ToList().AsReadOnly(); 13: } 14: } Cool huh? Just one thing you need to know about the .AsReadOnly() method is that the only way to modify your ReadOnlyCollection<> is to modify the original collection. So doing: 1: public static void Main() 2: { 3: Class1 class1 = new Class1(); 4: class1.MyKeyTitleList.Add(new KeyTitle { Key = 1, Title = "abc" }); 5: class1.MyKeyTitleList.Add(new KeyTitle { Key = 2, Title = "def" }); 6: class1.MyKeyTitleList.Add(new KeyTitle { Key = 3, Title = "ghi" }); 7: class1.MyKeyTitleList.Add(new KeyTitle { Key = 4, Title = "jkl" }); 8: TryToModifyCollection(class1.MyKeyTitleList.AsReadOnly()); 9:  10: Console.WriteLine(); 11:  12: class1.MyKeyTitleList.Add(new KeyTitle { Key = 5, Title = "mno" }); 13: class1.MyKeyTitleList[2] = new KeyTitle{Key = 3, Title = "GHI"}; 14: TryToModifyCollection(class1.MyKeyTitleList.AsReadOnly()); 15:  16: Console.ReadLine(); 17: } Gives me the output of: See that the second element’s Title is changed to upper-case and the fifth element also gets displayed even though we’re still looping through the same ReadOnlyCollection<KeyTitle>. Verdict: Now you know of a way to implement ‘Method(const param1)’ in your code!

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• “It’s only test code…”

  - by Chris George

  “Let me hack this in, it’s only test code”, “Don’t worry about getting it reviewed, it’s only test code”, “It doesn’t have to be elegant or efficient, it’s only test code”… do these phrases sound familiar? Chances are if you’ve working with test automation, at one point or other you will have heard these phrases, you have probably even used them yourself! What is certain is that code written under this “it’s only test code” mantra will come back and bite you in the arse! I’ve recently encountered a case where a test was giving a false positive, therefore hiding a real product bug because that test code was very badly written. Firstly it was very difficult to understand what the test was actually trying to achieve let alone how it was doing it, and this complexity masked a simple logic error. These issues are real and they do happen. Let’s take a step back from this and look at what we are trying to do. We are writing test code that tests product code, and we do this to create a suite of tests that will help protect our software against regressions. This test code is making sure that the product behaves as it should by employing some sort of expected result verification. The simple cases of these are generally not a problem. However, automation allows us to explore more complex scenarios in many more permutations. As this complexity increases then so does the complexity of the test code. It is at this point that code which has not been architected properly will cause problems.   Keep your friends close… So, how do we make sure we are doing it right? The development teams I have worked on have always had Test Engineers working very closely with their Software Engineers. This is something that I have always tried to take full advantage of. They are coding experts! So run your ideas past them, ask for advice on how to structure your code, help you design your data structures. This may require a shift in your teams viewpoint, as contrary to this section title and folklore, Software Engineers are not actually the mortal enemy of Test Engineers. As time progresses, and test automation becomes more and more ingrained in what we do, the two roles are converging more than ever. Over the 16 years I have spent as a Test Engineer, I have seen the grey area between the two roles grow significantly larger. This serves to strengthen the relationship and common bond between the two roles which helps to make test code activities so much easier!   Pair for the win Possibly the best thing you could do to write good test code is to pair program on the task. This will serve a few purposes. you will get the benefit of the Software Engineers knowledge and experience the Software Engineer will gain knowledge on the testing process. Sharing the love is a wonderful thing! two pairs of eyes are always better than one… And so are two brains. Between the two of you, I will guarantee you will derive more useful test cases than if it was just one of you.   Code reviews Another policy which certainly pays dividends is the practice of code reviews. By having one of your peers review your code before you commit it serves two purposes. Firstly, it forces you to explain your code. Just the act of doing this will often pick up errors in your code. Secondly, it gets yet another pair of eyes on your code! I cannot stress enough how important code reviews are. The benefits they offer apply as much to product code as test code. In short, Software and Test Engineers should all be doing them! It can be extended even further by getting test code reviewed by a Software Engineer and a Test Engineer, and likewise product code. This serves to keep both functions in the loop with changes going on within your code base.   Learn from your devs I briefly touched on this earlier but I’d like to go into more detail here. Pairing with your Software Engineers when writing your test code is such an amazing opportunity to improve your coding skills. As I sit here writing this article waiting to be called into court for jury service, it reminds me that it takes a lot of patience to be a Test Engineer, almost as much as it takes to be a juror! However tempting it is to go rushing in and start writing your automated tests, resist that urge. Discuss what you want to achieve then talk through the approach you’re going to take. Then code it up together. I find it really enlightening to ask questions like ‘is there a better way to do this?’ Or ‘is this how you would code it?’ The latter question, especially, is where I learn the most. I’ve found that most Software Engineers will be reluctant to show you the ‘right way’ to code something when writing tests because they perceive the ‘right way’ to be too complicated for the Test Engineer (e.g. not mentioning LINQ and instead doing something verbose). So by asking how THEY would code it, it unleashes their true dev-ness and advanced code usually ensues! I would like to point out, however, that you don’t have to accept their method as the final answer. On numerous occasions I have opted for the more simple/verbose solution because I found the code written by the Software Engineer too advanced and therefore I would find it unreadable when I return to the code in a months’ time! Always keep the target audience in mind when writing clever code, and in my case that is mostly Test Engineers.  

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• Benefits of Behavior Driven Development

  - by Aligned

  Originally posted on: http://geekswithblogs.net/Aligned/archive/2013/07/26/benefits-of-behavior-driven-development.aspxContinuing my previous article on BDD, I wanted to point out some benefits of BDD and since BDD is an extension of Test Driven Development (TDD), you get those as well. I’ll add another article on some possible downsides of this approach. There are many articles about the benefits of TDD and they apply to BDD. I’ve pointed out some here and copied some of the main points for each article, but there are many more including the book The Art of Unit Testing by Roy Osherove. http://geekswithblogs.net/leesblog/archive/2008/04/30/the-benefits-of-test-driven-development.aspx (Lee Brandt) Stability Accountability Design Ability Separated Concerns Progress Indicator http://tddftw.com/benefits-of-tdd/ Help maintainers understand the intention behind the code Bring validation and proper data handling concerns to the forefront. Writing the tests first is fun. Better APIs come from writing testable code. TDD will make you a better developer. http://www.slideshare.net/dhelper/benefit-from-unit-testing-in-the-real-world (from Typemock). Take a look at the slides, especially the extra time required for TDD (slide 10) and the next one of the bugs avoided using TDD (slide 11). Less bugs (slide 11) about testing and development (13) Increase confidence in code (14) Fearlessly change your code (14) Document Requirements (14) also see http://visualstudiomagazine.com/articles/2013/06/01/roc-rocks.aspx Discover usability issues early (14) All these points and articles are great and there are many more. The following are my additions to the benefits of BDD from using it in real projects for my company. July 2013 on MSDN - Behavior-Driven Design with SpecFlow Scott Allen did a very informative TDD and MVC module, but to me he is doing BDDCompile and Execute Requirements in Microsoft .NET ~ Video from TechEd 2012 Communication I was working through a complicated task that the decision tree kept growing. After writing out the Given, When, Then of the scenario, I was able tell QA what I had worked through for their initial test cases. They were able to add from there. It is also useful to use this language with other developers, managers, or clients to help make informed decisions on if it meets the requirements or if it can simplified to save time (money). Thinking through solutions, before starting to code This was the biggest benefit to me. I like to jump into coding to figure out the problem. Many times I don't understand my path well enough and have to do some parts over. A past supervisor told me several times during reviews that I need to get better at seeing "the forest for the trees". When I sit down and write out the behavior that I need to implement, I force myself to think things out further and catch scenarios before they get to QA. A co-worker that is new to BDD and we’ve been using it in our new project for the last 6 months, said “It really clarifies things”. It took him awhile to understand it all, but now he’s seeing the value of this approach (yes there are some downsides, but that is a different issue). Developers’ Confidence This is huge for me. With tests in place, my confidence grows that I won’t break code that I’m not directly changing. In the past, I’ve worked on projects with out tests and we would frequently find regression bugs (or worse the users would find them). That isn’t fun. We don’t catch all problems with the tests, but when QA catches one, I can write a test to make sure it doesn’t happen again. It’s also good for Releasing code, telling your manager that it’s good to go. As time goes on and the code gets older, how confident are you that checking in code won’t break something somewhere else? Merging code - pre release confidence If you’re merging code a lot, it’s nice to have the tests to help ensure you didn’t merge incorrectly. Interrupted work I had a task that I started and planned out, then was interrupted for a month because of different priorities. When I started it up again, and un-shelved my changes, I had the BDD specs and it helped me remember what I had figured out and what was left to do. It would have much more difficult without the specs and tests. Testing and verifying complicated scenarios Sometimes in the UI there are scenarios that get tricky, because there are a lot of steps involved (click here to open the dialog, enter the information, make sure it’s valid, when I click cancel it should do {x}, when I click ok it should close and do {y}, then do this, etc….). With BDD I can avoid some of the mouse clicking define the scenarios and have them re-run quickly, without using a mouse. UI testing is still needed, but this helps a bunch. The same can be true for tricky server logic. Documentation of Assumptions and Specifications The BDD spec tests (Jasmine or SpecFlow or other tool) also work as documentation and show what the original developer was trying to accomplish. It’s not a different Word document, so developers will keep this up to date, instead of letting it become obsolete. What happens if you leave the project (consulting, new job, etc) with no specs or at the least good comments in the code? Sometimes I think of a new scenario, so I add a failing spec and continue in the same stream of thought (don’t forget it because it was on a piece of paper or in a notepad). Then later I can come back and handle it and have it documented. Jasmine tests and JavaScript –> help deal with the non-typed system I like JavaScript, but I also dislike working with JavaScript. I miss C# telling me if a property doesn’t actually exist at build time. I like the idea of TypeScript and hope to use it more in the future. I also use KnockoutJs, which has observables that need to be called with ending (), since the observable is a function. It’s hard to remember when to use () or not and the Jasmine specs/tests help ensure the correct usage.   This should give you an idea of the benefits that I see in using the BDD approach. I’m sure there are more. It talks a lot of practice, investment and experimentation to figure out how to approach this and to get comfortable with it. I agree with Scott Allen in the video I linked above “Remember that TDD can take some practice. So if you're not doing test-driven design right now? You can start and practice and get better. And you'll reach a point where you'll never want to get back.”

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• Let your Signature Experience drive IT-decision making

  - by Tania Le Voi

  Today’s CIO job description:  ‘’Align IT infrastructure and solutions with business goals and objectives ; AND while doing so reduce costs; BUT ALSO, be innovative, ensure the architectures are adaptable and agile as we need to act today on the changes that we may request tomorrow.”   Sound like an unachievable request? The fact is, reality dictates that CIO’s are put under this type of pressure to deliver more with less. In a past career phase I spent a few years as an IT Relationship Manager for a large Insurance company. This is a role that we see all too infrequently in many of our customers, and it’s a shame.  The purpose of this role was to build a bridge, a relationship between IT and the business. Key to achieving that goal was to ensure the same language was being spoken and more importantly that objectives were commonly understood - hence service and projects were delivered to time, to budget and actually solved the business problems. In reality IT and the business are already married, but the relationship is most often defined as ‘supplier’ of IT rather than a ‘trusted partner’. To deliver business value they need to understand how to work together effectively to attain this next level of partnership. The Business cannot compete if they do not get a new product to market ahead of the competition, or for example act in a timely manner to address a new industry problem such as a legislative change. An even better example is when the Application or Service fails and the Business takes a hit by bad publicity, being trending topics on social media and losing direct revenue from online channels. For this reason alone Business and IT need the alignment of their priorities and deliverables now more than ever! Take a look at Forrester’s recent study that found ‘many IT respondents considering themselves to be trusted partners of the business but their efforts are impaired by the inadequacy of tools and organizations’.  IT Meet the Business; Business Meet IT So what is going on? We talk about aligning the business with IT but the reality is it’s difficult to do. Like any relationship each side has different goals and needs and language can be a barrier; business vs. technology jargon! What if we could translate the needs of both sides into actionable information, backed by data both sides understand, presented in a meaningful way?  Well now we can with the Business-Driven Application Management capabilities in Oracle Enterprise Manager 12cR2! Enterprise Manager’s Business-Driven Application Management capabilities provide the information that IT needs to understand the impact of its decisions on business criteria.  No longer does IT need to be focused solely on speeds and feeds, performance and throughput – now IT can understand IT’s impact on business KPIs like inventory turns, order-to-cash cycle, pipeline-to-forecast, and similar.  Similarly, now the line of business can understand which IT services are most critical for the KPIs they care about. There are a good deal of resources on Oracle Technology Network that describe the functionality of these products, so I won’t’ rehash them here.  What I want to talk about is what you do with these products. What’s next after we meet? Where do you start? Step 1:  Identify the Signature Experience. This is THE business process (or set of processes) that is core to the business, the one that drives the economic engine, the process that a customer recognises the company brand for, reputation, the customer experience, the process that a CEO would state as his number one priority. The crème de la crème of your business! Once you have nailed this it gets easy as Enterprise Manager 12c makes it easy. Step 2:  Map the Signature Experience to underlying IT.  Taking the signature experience, map out the touch points of the components that play a part in ensuring this business transaction is successful end to end, think of it like mapping out a critical path; the applications, middleware, databases and hardware. Use the wealth of Enterprise Manager features such as Systems, Services, Business Application Targets and Business Transaction Management (BTM) to assist you. Adding Real User Experience Insight (RUEI) into the mix will make the end to end customer satisfaction story transparent. Work with the business and define meaningful key performance indicators (KPI’s) and thresholds to enable you to report and action upon. Step 3:  Observe the data over time.  You now have meaningful insight into every step enabling your signature experience and you understand the implication of that experience on your underlying IT.  Watch if for a few months, see what happens and reconvene with your business stakeholders and set clear and measurable targets which can re-define service levels.  Step 4:  Change the information about which you and the business communicate.  It’s amazing what happens when you and the business speak the same language.  You’ll be able to make more informed business and IT decisions. From here IT can identify where/how budget is spent whether on the level of support, performance, capacity, HA, DR, certification etc. IT SLA’s no longer need be focused on metrics such as %availability but structured around business process requirements. The power of this way of thinking doesn’t end here. IT staff get to see and understand how their own role contributes to the business making them accountable for the business service. Take a step further and appraise your staff on the business competencies that are linked to the service availability. For the business, the language barrier is removed by producing targeted reports on the signature experience core to the business and therefore key to the CEO. Chargeback or show back becomes easier to justify as the ‘cost of day per outage’ can be more easily calculated; the business will be able to translate the cost to the business to the cost/value of the underlying IT that supports it. Used this way, Oracle Enterprise Manager 12c is a key enabler to a harmonious relationship between the end customer the business and IT to deliver ultimate service and satisfaction. Just engage with the business upfront, make the signature experience visible and let Enterprise Manager 12c do the rest. In the next blog entry we will cover some of the Enterprise Manager features mentioned to enable you to implement this new way of working.  

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• OS8- AK8- The bad news...

  - by Steve Tunstall

  Ok I told you I would give you the bad news of AK8 to go along with all the cool new stuff, so here it is. It's not that bad, really, just things you need to be aware of. First, the 2013.1 code is being called OS8, AK8 and 2013.1 by different people. I mean different people INSIDE Oracle!! It was supposed to be easy, but it never is. So for the rest of this blog entry, I'm calling it AK8. AK8 is not compatible with the 7x10 series. Ever. The 7x10 series is not supported with AK8, and if you try to upgrade one, it will fail at the healthcheck. All 7x20 series, all of them regardless of age, are supported with AK8. Drive trays. Let's talk about drive trays and SAS cards. The older drive trays for the 7x20 series were called the "Riverwalk 2" or "DS2" trays. They were technically the "J4410" series JBODs that Sun used to sell a la carte before we stopped selling JBODs. Don't get me started on that, it still makes me mad. We used these for many years, and you can still buy them right now until December 15th, 2013, when they will no longer be sold. The DS2 tray only came as a 4u, 24 drive shelf. It held 3.5" drives, and you had a choice of 2TB, 3TB, 300GB or 600GB drives. The SAS HBA in the 7x20 series was called a "Thebe" card, with a part # of 7105394. The 7420, for example, came standard with two of these "Thebe" cards for connecting to the disk trays. Two Thebe cards could handle up to 12 trays, so one would add two more cards to go to 24 trays, or have up to six Thebe cards to handle 36 trays. This card was for external SAS only. It did not connect to the internal OS drives or the Readzillas, both of which used the internal SCSI controller of the server. These Riverwalk 2 trays ARE supported with AK8. You can upgrade your older 7420 or 7320, no problem, as-is. The much older Riverwalk 1 trays or J4400 trays are NOT supported by AK8. However, they were only used by the 7x10 series, and we already said that the 7x10 series was not supported. Here's where it gets tricky. Since last January, we have been selling the new style disk trays. We call them the "DE2-24P" and the "DE2-24C" trays. The "C" tray is for capacity drives, which are 3.5" 3TB or 4TB drives. The "P" trays are for performance drives, which are 2.5" 300GB and 900GB drives. These trays are NOT Riverwalk 2 trays, even though the "C" series may kind of look like it. Different manufacturer and different firmware. They are not new. Like I said, we've been selling them with the 7x20 series since last January. They are the only disk trays we will be selling going forward. Of course, AK8 supports them. So what's the problem? The problem is going to be for people who have to mix drive trays. Remember, your older 7x20 series has Thebe SAS2 HBAs. These have 2 SAS ports per card.  The new ZS3-2 and ZS3-4 systems, however, have the new "Thebe2" SAS2 HBAs. These Thebe2 cards have 4 ports per card. This is very cool, as we can now do more SAS channels with less cards. Instead of needing 4 SAS cards to grow to 24 trays like we did with the old Thebe cards, I can now do 24 trays with only 2 Thebe2 cards. This means more IO slots for fun things like Infiniband and 10G. So far, so good, right? These Thebe2 cards work with any disk tray. You can even mix older DS2 trays with the newer DE2 trays in the same system, as long as you have Thebe2 cards. Ah, there's your problem. You don't have Thebe2 cards in your old 7420, do you? Well, I told you the bad news wasn't that bad, right? We can take out your Thebe cards and replace them with Thebe2. You can then plug your older DS2 trays right back in, and also now get newer DE2 trays going forward. However, it's important that the trays are on different SAS channels. You can mix them in the same system, but not on the same channel. Ask your local SC if you need help with the new cable layout. By the way, the new ZS3-2 and ZS3-4 systems also include a new IO card called "Erie" cards. These are for INTERNAL SAS to the OS drives and the Readzillas. So those are now SAS2 instead of SATA like the older models. Yes, the Erie card uses an IO slot, but that's OK, because the Thebe2 cards allow us to use less SAS HBAs to grow the system, right? That's it. Not too much bad news and really not that bad. AK8 does not support the 7x10 series, and you may need new Thebe2 cards in your older systems if you want to add on newer DE2 trays. I think we can all agree that there are worse things out there. Like our Congress.   Next up.... More good news and cool AK8 tricks. Such as virtual NICS. 

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