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  • A way of doing real-world test-driven development (and some thoughts about it)

    - by Thomas Weller
    Lately, I exchanged some arguments with Derick Bailey about some details of the red-green-refactor cycle of the Test-driven development process. In short, the issue revolved around the fact that it’s not enough to have a test red or green, but it’s also important to have it red or green for the right reasons. While for me, it’s sufficient to initially have a NotImplementedException in place, Derick argues that this is not totally correct (see these two posts: Red/Green/Refactor, For The Right Reasons and Red For The Right Reason: Fail By Assertion, Not By Anything Else). And he’s right. But on the other hand, I had no idea how his insights could have any practical consequence for my own individual interpretation of the red-green-refactor cycle (which is not really red-green-refactor, at least not in its pure sense, see the rest of this article). This made me think deeply for some days now. In the end I found out that the ‘right reason’ changes in my understanding depending on what development phase I’m in. To make this clear (at least I hope it becomes clear…) I started to describe my way of working in some detail, and then something strange happened: The scope of the article slightly shifted from focusing ‘only’ on the ‘right reason’ issue to something more general, which you might describe as something like  'Doing real-world TDD in .NET , with massive use of third-party add-ins’. This is because I feel that there is a more general statement about Test-driven development to make:  It’s high time to speak about the ‘How’ of TDD, not always only the ‘Why’. Much has been said about this, and me myself also contributed to that (see here: TDD is not about testing, it's about how we develop software). But always justifying what you do is very unsatisfying in the long run, it is inherently defensive, and it costs time and effort that could be used for better and more important things. And frankly: I’m somewhat sick and tired of repeating time and again that the test-driven way of software development is highly preferable for many reasons - I don’t want to spent my time exclusively on stating the obvious… So, again, let’s say it clearly: TDD is programming, and programming is TDD. Other ways of programming (code-first, sometimes called cowboy-coding) are exceptional and need justification. – I know that there are many people out there who will disagree with this radical statement, and I also know that it’s not a description of the real world but more of a mission statement or something. But nevertheless I’m absolutely sure that in some years this statement will be nothing but a platitude. Side note: Some parts of this post read as if I were paid by Jetbrains (the manufacturer of the ReSharper add-in – R#), but I swear I’m not. Rather I think that Visual Studio is just not production-complete without it, and I wouldn’t even consider to do professional work without having this add-in installed... The three parts of a software component Before I go into some details, I first should describe my understanding of what belongs to a software component (assembly, type, or method) during the production process (i.e. the coding phase). Roughly, I come up with the three parts shown below:   First, we need to have some initial sort of requirement. This can be a multi-page formal document, a vague idea in some programmer’s brain of what might be needed, or anything in between. In either way, there has to be some sort of requirement, be it explicit or not. – At the C# micro-level, the best way that I found to formulate that is to define interfaces for just about everything, even for internal classes, and to provide them with exhaustive xml comments. The next step then is to re-formulate these requirements in an executable form. This is specific to the respective programming language. - For C#/.NET, the Gallio framework (which includes MbUnit) in conjunction with the ReSharper add-in for Visual Studio is my toolset of choice. The third part then finally is the production code itself. It’s development is entirely driven by the requirements and their executable formulation. This is the delivery, the two other parts are ‘only’ there to make its production possible, to give it a decent quality and reliability, and to significantly reduce related costs down the maintenance timeline. So while the first two parts are not really relevant for the customer, they are very important for the developer. The customer (or in Scrum terms: the Product Owner) is not interested at all in how  the product is developed, he is only interested in the fact that it is developed as cost-effective as possible, and that it meets his functional and non-functional requirements. The rest is solely a matter of the developer’s craftsmanship, and this is what I want to talk about during the remainder of this article… An example To demonstrate my way of doing real-world TDD, I decided to show the development of a (very) simple Calculator component. The example is deliberately trivial and silly, as examples always are. I am totally aware of the fact that real life is never that simple, but I only want to show some development principles here… The requirement As already said above, I start with writing down some words on the initial requirement, and I normally use interfaces for that, even for internal classes - the typical question “intf or not” doesn’t even come to mind. I need them for my usual workflow and using them automatically produces high componentized and testable code anyway. To think about their usage in every single situation would slow down the production process unnecessarily. So this is what I begin with: namespace Calculator {     /// <summary>     /// Defines a very simple calculator component for demo purposes.     /// </summary>     public interface ICalculator     {         /// <summary>         /// Gets the result of the last successful operation.         /// </summary>         /// <value>The last result.</value>         /// <remarks>         /// Will be <see langword="null" /> before the first successful operation.         /// </remarks>         double? LastResult { get; }       } // interface ICalculator   } // namespace Calculator So, I’m not beginning with a test, but with a sort of code declaration - and still I insist on being 100% test-driven. There are three important things here: Starting this way gives me a method signature, which allows to use IntelliSense and AutoCompletion and thus eliminates the danger of typos - one of the most regular, annoying, time-consuming, and therefore expensive sources of error in the development process. In my understanding, the interface definition as a whole is more of a readable requirement document and technical documentation than anything else. So this is at least as much about documentation than about coding. The documentation must completely describe the behavior of the documented element. I normally use an IoC container or some sort of self-written provider-like model in my architecture. In either case, I need my components defined via service interfaces anyway. - I will use the LinFu IoC framework here, for no other reason as that is is very simple to use. The ‘Red’ (pt. 1)   First I create a folder for the project’s third-party libraries and put the LinFu.Core dll there. Then I set up a test project (via a Gallio project template), and add references to the Calculator project and the LinFu dll. Finally I’m ready to write the first test, which will look like the following: namespace Calculator.Test {     [TestFixture]     public class CalculatorTest     {         private readonly ServiceContainer container = new ServiceContainer();           [Test]         public void CalculatorLastResultIsInitiallyNull()         {             ICalculator calculator = container.GetService<ICalculator>();               Assert.IsNull(calculator.LastResult);         }       } // class CalculatorTest   } // namespace Calculator.Test       This is basically the executable formulation of what the interface definition states (part of). Side note: There’s one principle of TDD that is just plain wrong in my eyes: I’m talking about the Red is 'does not compile' thing. How could a compiler error ever be interpreted as a valid test outcome? I never understood that, it just makes no sense to me. (Or, in Derick’s terms: this reason is as wrong as a reason ever could be…) A compiler error tells me: Your code is incorrect, but nothing more.  Instead, the ‘Red’ part of the red-green-refactor cycle has a clearly defined meaning to me: It means that the test works as intended and fails only if its assumptions are not met for some reason. Back to our Calculator. When I execute the above test with R#, the Gallio plugin will give me this output: So this tells me that the test is red for the wrong reason: There’s no implementation that the IoC-container could load, of course. So let’s fix that. With R#, this is very easy: First, create an ICalculator - derived type:        Next, implement the interface members: And finally, move the new class to its own file: So far my ‘work’ was six mouse clicks long, the only thing that’s left to do manually here, is to add the Ioc-specific wiring-declaration and also to make the respective class non-public, which I regularly do to force my components to communicate exclusively via interfaces: This is what my Calculator class looks like as of now: using System; using LinFu.IoC.Configuration;   namespace Calculator {     [Implements(typeof(ICalculator))]     internal class Calculator : ICalculator     {         public double? LastResult         {             get             {                 throw new NotImplementedException();             }         }     } } Back to the test fixture, we have to put our IoC container to work: [TestFixture] public class CalculatorTest {     #region Fields       private readonly ServiceContainer container = new ServiceContainer();       #endregion // Fields       #region Setup/TearDown       [FixtureSetUp]     public void FixtureSetUp()     {        container.LoadFrom(AppDomain.CurrentDomain.BaseDirectory, "Calculator.dll");     }       ... Because I have a R# live template defined for the setup/teardown method skeleton as well, the only manual coding here again is the IoC-specific stuff: two lines, not more… The ‘Red’ (pt. 2) Now, the execution of the above test gives the following result: This time, the test outcome tells me that the method under test is called. And this is the point, where Derick and I seem to have somewhat different views on the subject: Of course, the test still is worthless regarding the red/green outcome (or: it’s still red for the wrong reasons, in that it gives a false negative). But as far as I am concerned, I’m not really interested in the test outcome at this point of the red-green-refactor cycle. Rather, I only want to assert that my test actually calls the right method. If that’s the case, I will happily go on to the ‘Green’ part… The ‘Green’ Making the test green is quite trivial. Just make LastResult an automatic property:     [Implements(typeof(ICalculator))]     internal class Calculator : ICalculator     {         public double? LastResult { get; private set; }     }         One more round… Now on to something slightly more demanding (cough…). Let’s state that our Calculator exposes an Add() method:         ...   /// <summary>         /// Adds the specified operands.         /// </summary>         /// <param name="operand1">The operand1.</param>         /// <param name="operand2">The operand2.</param>         /// <returns>The result of the additon.</returns>         /// <exception cref="ArgumentException">         /// Argument <paramref name="operand1"/> is &lt; 0.<br/>         /// -- or --<br/>         /// Argument <paramref name="operand2"/> is &lt; 0.         /// </exception>         double Add(double operand1, double operand2);       } // interface ICalculator A remark: I sometimes hear the complaint that xml comment stuff like the above is hard to read. That’s certainly true, but irrelevant to me, because I read xml code comments with the CR_Documentor tool window. And using that, it looks like this:   Apart from that, I’m heavily using xml code comments (see e.g. here for a detailed guide) because there is the possibility of automating help generation with nightly CI builds (using MS Sandcastle and the Sandcastle Help File Builder), and then publishing the results to some intranet location.  This way, a team always has first class, up-to-date technical documentation at hand about the current codebase. (And, also very important for speeding up things and avoiding typos: You have IntelliSense/AutoCompletion and R# support, and the comments are subject to compiler checking…).     Back to our Calculator again: Two more R# – clicks implement the Add() skeleton:         ...           public double Add(double operand1, double operand2)         {             throw new NotImplementedException();         }       } // class Calculator As we have stated in the interface definition (which actually serves as our requirement document!), the operands are not allowed to be negative. So let’s start implementing that. Here’s the test: [Test] [Row(-0.5, 2)] public void AddThrowsOnNegativeOperands(double operand1, double operand2) {     ICalculator calculator = container.GetService<ICalculator>();       Assert.Throws<ArgumentException>(() => calculator.Add(operand1, operand2)); } As you can see, I’m using a data-driven unit test method here, mainly for these two reasons: Because I know that I will have to do the same test for the second operand in a few seconds, I save myself from implementing another test method for this purpose. Rather, I only will have to add another Row attribute to the existing one. From the test report below, you can see that the argument values are explicitly printed out. This can be a valuable documentation feature even when everything is green: One can quickly review what values were tested exactly - the complete Gallio HTML-report (as it will be produced by the Continuous Integration runs) shows these values in a quite clear format (see below for an example). Back to our Calculator development again, this is what the test result tells us at the moment: So we’re red again, because there is not yet an implementation… Next we go on and implement the necessary parameter verification to become green again, and then we do the same thing for the second operand. To make a long story short, here’s the test and the method implementation at the end of the second cycle: // in CalculatorTest:   [Test] [Row(-0.5, 2)] [Row(295, -123)] public void AddThrowsOnNegativeOperands(double operand1, double operand2) {     ICalculator calculator = container.GetService<ICalculator>();       Assert.Throws<ArgumentException>(() => calculator.Add(operand1, operand2)); }   // in Calculator: public double Add(double operand1, double operand2) {     if (operand1 < 0.0)     {         throw new ArgumentException("Value must not be negative.", "operand1");     }     if (operand2 < 0.0)     {         throw new ArgumentException("Value must not be negative.", "operand2");     }     throw new NotImplementedException(); } So far, we have sheltered our method from unwanted input, and now we can safely operate on the parameters without further caring about their validity (this is my interpretation of the Fail Fast principle, which is regarded here in more detail). Now we can think about the method’s successful outcomes. First let’s write another test for that: [Test] [Row(1, 1, 2)] public void TestAdd(double operand1, double operand2, double expectedResult) {     ICalculator calculator = container.GetService<ICalculator>();       double result = calculator.Add(operand1, operand2);       Assert.AreEqual(expectedResult, result); } Again, I’m regularly using row based test methods for these kinds of unit tests. The above shown pattern proved to be extremely helpful for my development work, I call it the Defined-Input/Expected-Output test idiom: You define your input arguments together with the expected method result. There are two major benefits from that way of testing: In the course of refining a method, it’s very likely to come up with additional test cases. In our case, we might add tests for some edge cases like ‘one of the operands is zero’ or ‘the sum of the two operands causes an overflow’, or maybe there’s an external test protocol that has to be fulfilled (e.g. an ISO norm for medical software), and this results in the need of testing against additional values. In all these scenarios we only have to add another Row attribute to the test. Remember that the argument values are written to the test report, so as a side-effect this produces valuable documentation. (This can become especially important if the fulfillment of some sort of external requirements has to be proven). So your test method might look something like that in the end: [Test, Description("Arguments: operand1, operand2, expectedResult")] [Row(1, 1, 2)] [Row(0, 999999999, 999999999)] [Row(0, 0, 0)] [Row(0, double.MaxValue, double.MaxValue)] [Row(4, double.MaxValue - 2.5, double.MaxValue)] public void TestAdd(double operand1, double operand2, double expectedResult) {     ICalculator calculator = container.GetService<ICalculator>();       double result = calculator.Add(operand1, operand2);       Assert.AreEqual(expectedResult, result); } And this will produce the following HTML report (with Gallio):   Not bad for the amount of work we invested in it, huh? - There might be scenarios where reports like that can be useful for demonstration purposes during a Scrum sprint review… The last requirement to fulfill is that the LastResult property is expected to store the result of the last operation. I don’t show this here, it’s trivial enough and brings nothing new… And finally: Refactor (for the right reasons) To demonstrate my way of going through the refactoring portion of the red-green-refactor cycle, I added another method to our Calculator component, namely Subtract(). Here’s the code (tests and production): // CalculatorTest.cs:   [Test, Description("Arguments: operand1, operand2, expectedResult")] [Row(1, 1, 0)] [Row(0, 999999999, -999999999)] [Row(0, 0, 0)] [Row(0, double.MaxValue, -double.MaxValue)] [Row(4, double.MaxValue - 2.5, -double.MaxValue)] public void TestSubtract(double operand1, double operand2, double expectedResult) {     ICalculator calculator = container.GetService<ICalculator>();       double result = calculator.Subtract(operand1, operand2);       Assert.AreEqual(expectedResult, result); }   [Test, Description("Arguments: operand1, operand2, expectedResult")] [Row(1, 1, 0)] [Row(0, 999999999, -999999999)] [Row(0, 0, 0)] [Row(0, double.MaxValue, -double.MaxValue)] [Row(4, double.MaxValue - 2.5, -double.MaxValue)] public void TestSubtractGivesExpectedLastResult(double operand1, double operand2, double expectedResult) {     ICalculator calculator = container.GetService<ICalculator>();       calculator.Subtract(operand1, operand2);       Assert.AreEqual(expectedResult, calculator.LastResult); }   ...   // ICalculator.cs: /// <summary> /// Subtracts the specified operands. /// </summary> /// <param name="operand1">The operand1.</param> /// <param name="operand2">The operand2.</param> /// <returns>The result of the subtraction.</returns> /// <exception cref="ArgumentException"> /// Argument <paramref name="operand1"/> is &lt; 0.<br/> /// -- or --<br/> /// Argument <paramref name="operand2"/> is &lt; 0. /// </exception> double Subtract(double operand1, double operand2);   ...   // Calculator.cs:   public double Subtract(double operand1, double operand2) {     if (operand1 < 0.0)     {         throw new ArgumentException("Value must not be negative.", "operand1");     }       if (operand2 < 0.0)     {         throw new ArgumentException("Value must not be negative.", "operand2");     }       return (this.LastResult = operand1 - operand2).Value; }   Obviously, the argument validation stuff that was produced during the red-green part of our cycle duplicates the code from the previous Add() method. So, to avoid code duplication and minimize the number of code lines of the production code, we do an Extract Method refactoring. One more time, this is only a matter of a few mouse clicks (and giving the new method a name) with R#: Having done that, our production code finally looks like that: using System; using LinFu.IoC.Configuration;   namespace Calculator {     [Implements(typeof(ICalculator))]     internal class Calculator : ICalculator     {         #region ICalculator           public double? LastResult { get; private set; }           public double Add(double operand1, double operand2)         {             ThrowIfOneOperandIsInvalid(operand1, operand2);               return (this.LastResult = operand1 + operand2).Value;         }           public double Subtract(double operand1, double operand2)         {             ThrowIfOneOperandIsInvalid(operand1, operand2);               return (this.LastResult = operand1 - operand2).Value;         }           #endregion // ICalculator           #region Implementation (Helper)           private static void ThrowIfOneOperandIsInvalid(double operand1, double operand2)         {             if (operand1 < 0.0)             {                 throw new ArgumentException("Value must not be negative.", "operand1");             }               if (operand2 < 0.0)             {                 throw new ArgumentException("Value must not be negative.", "operand2");             }         }           #endregion // Implementation (Helper)       } // class Calculator   } // namespace Calculator But is the above worth the effort at all? It’s obviously trivial and not very impressive. All our tests were green (for the right reasons), and refactoring the code did not change anything. It’s not immediately clear how this refactoring work adds value to the project. Derick puts it like this: STOP! Hold on a second… before you go any further and before you even think about refactoring what you just wrote to make your test pass, you need to understand something: if your done with your requirements after making the test green, you are not required to refactor the code. I know… I’m speaking heresy, here. Toss me to the wolves, I’ve gone over to the dark side! Seriously, though… if your test is passing for the right reasons, and you do not need to write any test or any more code for you class at this point, what value does refactoring add? Derick immediately answers his own question: So why should you follow the refactor portion of red/green/refactor? When you have added code that makes the system less readable, less understandable, less expressive of the domain or concern’s intentions, less architecturally sound, less DRY, etc, then you should refactor it. I couldn’t state it more precise. From my personal perspective, I’d add the following: You have to keep in mind that real-world software systems are usually quite large and there are dozens or even hundreds of occasions where micro-refactorings like the above can be applied. It’s the sum of them all that counts. And to have a good overall quality of the system (e.g. in terms of the Code Duplication Percentage metric) you have to be pedantic on the individual, seemingly trivial cases. My job regularly requires the reading and understanding of ‘foreign’ code. So code quality/readability really makes a HUGE difference for me – sometimes it can be even the difference between project success and failure… Conclusions The above described development process emerged over the years, and there were mainly two things that guided its evolution (you might call it eternal principles, personal beliefs, or anything in between): Test-driven development is the normal, natural way of writing software, code-first is exceptional. So ‘doing TDD or not’ is not a question. And good, stable code can only reliably be produced by doing TDD (yes, I know: many will strongly disagree here again, but I’ve never seen high-quality code – and high-quality code is code that stood the test of time and causes low maintenance costs – that was produced code-first…) It’s the production code that pays our bills in the end. (Though I have seen customers these days who demand an acceptance test battery as part of the final delivery. Things seem to go into the right direction…). The test code serves ‘only’ to make the production code work. But it’s the number of delivered features which solely counts at the end of the day - no matter how much test code you wrote or how good it is. With these two things in mind, I tried to optimize my coding process for coding speed – or, in business terms: productivity - without sacrificing the principles of TDD (more than I’d do either way…).  As a result, I consider a ratio of about 3-5/1 for test code vs. production code as normal and desirable. In other words: roughly 60-80% of my code is test code (This might sound heavy, but that is mainly due to the fact that software development standards only begin to evolve. The entire software development profession is very young, historically seen; only at the very beginning, and there are no viable standards yet. If you think about software development as a kind of casting process, where the test code is the mold and the resulting production code is the final product, then the above ratio sounds no longer extraordinary…) Although the above might look like very much unnecessary work at first sight, it’s not. With the aid of the mentioned add-ins, doing all the above is a matter of minutes, sometimes seconds (while writing this post took hours and days…). The most important thing is to have the right tools at hand. Slow developer machines or the lack of a tool or something like that - for ‘saving’ a few 100 bucks -  is just not acceptable and a very bad decision in business terms (though I quite some times have seen and heard that…). Production of high-quality products needs the usage of high-quality tools. This is a platitude that every craftsman knows… The here described round-trip will take me about five to ten minutes in my real-world development practice. I guess it’s about 30% more time compared to developing the ‘traditional’ (code-first) way. But the so manufactured ‘product’ is of much higher quality and massively reduces maintenance costs, which is by far the single biggest cost factor, as I showed in this previous post: It's the maintenance, stupid! (or: Something is rotten in developerland.). In the end, this is a highly cost-effective way of software development… But on the other hand, there clearly is a trade-off here: coding speed vs. code quality/later maintenance costs. The here described development method might be a perfect fit for the overwhelming majority of software projects, but there certainly are some scenarios where it’s not - e.g. if time-to-market is crucial for a software project. So this is a business decision in the end. It’s just that you have to know what you’re doing and what consequences this might have… Some last words First, I’d like to thank Derick Bailey again. His two aforementioned posts (which I strongly recommend for reading) inspired me to think deeply about my own personal way of doing TDD and to clarify my thoughts about it. I wouldn’t have done that without this inspiration. I really enjoy that kind of discussions… I agree with him in all respects. But I don’t know (yet?) how to bring his insights into the described production process without slowing things down. The above described method proved to be very “good enough” in my practical experience. But of course, I’m open to suggestions here… My rationale for now is: If the test is initially red during the red-green-refactor cycle, the ‘right reason’ is: it actually calls the right method, but this method is not yet operational. Later on, when the cycle is finished and the tests become part of the regular, automated Continuous Integration process, ‘red’ certainly must occur for the ‘right reason’: in this phase, ‘red’ MUST mean nothing but an unfulfilled assertion - Fail By Assertion, Not By Anything Else!

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  • Silverlight for Windows Embedded tutorial (step 4)

    - by Valter Minute
    I’m back with my Silverlight for Windows Embedded tutorial. Sorry for the long delay between step 3 and step 4, the MVP summit and some work related issue prevented me from working on the tutorial during the last weeks. In our first,  second and third tutorial steps we implemented some very simple applications, just to understand the basic structure of a Silverlight for Windows Embedded application, learn how to handle events and how to operate on images. In this third step our sample application will be slightly more complicated, to introduce two new topics: list boxes and custom control. We will also learn how to create controls at runtime. I choose to explain those topics together and provide a sample a bit more complicated than usual just to start to give the feeling of how a “real” Silverlight for Windows Embedded application is organized. As usual we can start using Expression Blend to define our main page. In this case we will have a listbox and a textblock. Here’s the XAML code: <UserControl xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" x:Class="ListDemo.Page" Width="640" Height="480" x:Name="ListPage" xmlns:ListDemo="clr-namespace:ListDemo">   <Grid x:Name="LayoutRoot" Background="White"> <ListBox Margin="19,57,19,66" x:Name="FileList" SelectionChanged="Filelist_SelectionChanged"/> <TextBlock Height="35" Margin="19,8,19,0" VerticalAlignment="Top" TextWrapping="Wrap" x:Name="CurrentDir" Text="TextBlock" FontSize="20"/> </Grid> </UserControl> In our listbox we will load a list of directories, starting from the filesystem root (there are no drives in Windows CE, the filesystem has a single root named “\”). When the user clicks on an item inside the list, the corresponding directory path will be displayed in the TextBlock object and the subdirectories of the selected branch will be shown inside the list. As you can see we declared an event handler for the SelectionChanged event of our listbox. We also used a different font size for the TextBlock, to make it more readable. XAML and Expression Blend allow you to customize your UI pretty heavily, experiment with the tools and discover how you can completely change the aspect of your application without changing a single line of code! Inside our ListBox we want to insert the directory presenting a nice icon and their name, just like you are used to see them inside Windows 7 file explorer, for example. To get this we will define a user control. This is a custom object that will behave like “regular” Silverlight for Windows Embedded objects inside our application. First of all we have to define the look of our custom control, named DirectoryItem, using XAML: <UserControl xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" xmlns:d="http://schemas.microsoft.com/expression/blend/2008" xmlns:mc="http://schemas.openxmlformats.org/markup-compatibility/2006" mc:Ignorable="d" x:Class="ListDemo.DirectoryItem" Width="500" Height="80">   <StackPanel x:Name="LayoutRoot" Orientation="Horizontal"> <Canvas Width="31.6667" Height="45.9583" Margin="10,10,10,10" RenderTransformOrigin="0.5,0.5"> <Canvas.RenderTransform> <TransformGroup> <ScaleTransform/> <SkewTransform/> <RotateTransform Angle="-31.27"/> <TranslateTransform/> </TransformGroup> </Canvas.RenderTransform> <Rectangle Width="31.6667" Height="45.8414" Canvas.Left="0" Canvas.Top="0.116943" Stretch="Fill"> <Rectangle.Fill> <LinearGradientBrush StartPoint="0.142631,0.75344" EndPoint="1.01886,0.75344"> <LinearGradientBrush.RelativeTransform> <TransformGroup> <SkewTransform CenterX="0.142631" CenterY="0.75344" AngleX="19.3128" AngleY="0"/> <RotateTransform CenterX="0.142631" CenterY="0.75344" Angle="-35.3436"/> </TransformGroup> </LinearGradientBrush.RelativeTransform> <LinearGradientBrush.GradientStops> <GradientStop Color="#FF7B6802" Offset="0"/> <GradientStop Color="#FFF3D42C" Offset="1"/> </LinearGradientBrush.GradientStops> </LinearGradientBrush> </Rectangle.Fill> </Rectangle> <Rectangle Width="29.8441" Height="43.1517" Canvas.Left="0.569519" Canvas.Top="1.05249" Stretch="Fill"> <Rectangle.Fill> <LinearGradientBrush StartPoint="0.142632,0.753441" EndPoint="1.01886,0.753441"> <LinearGradientBrush.RelativeTransform> <TransformGroup> <SkewTransform CenterX="0.142632" CenterY="0.753441" AngleX="19.3127" AngleY="0"/> <RotateTransform CenterX="0.142632" CenterY="0.753441" Angle="-35.3437"/> </TransformGroup> </LinearGradientBrush.RelativeTransform> <LinearGradientBrush.GradientStops> <GradientStop Color="#FFCDCDCD" Offset="0.0833333"/> <GradientStop Color="#FFFFFFFF" Offset="1"/> </LinearGradientBrush.GradientStops> </LinearGradientBrush> </Rectangle.Fill> </Rectangle> <Rectangle Width="29.8441" Height="43.1517" Canvas.Left="0.455627" Canvas.Top="2.28036" Stretch="Fill"> <Rectangle.Fill> <LinearGradientBrush StartPoint="0.142631,0.75344" EndPoint="1.01886,0.75344"> <LinearGradientBrush.RelativeTransform> <TransformGroup> <SkewTransform CenterX="0.142631" CenterY="0.75344" AngleX="19.3128" AngleY="0"/> <RotateTransform CenterX="0.142631" CenterY="0.75344" Angle="-35.3436"/> </TransformGroup> </LinearGradientBrush.RelativeTransform> <LinearGradientBrush.GradientStops> <GradientStop Color="#FFCDCDCD" Offset="0.0833333"/> <GradientStop Color="#FFFFFFFF" Offset="1"/> </LinearGradientBrush.GradientStops> </LinearGradientBrush> </Rectangle.Fill> </Rectangle> <Rectangle Width="29.8441" Height="43.1517" Canvas.Left="0.455627" Canvas.Top="1.34485" Stretch="Fill"> <Rectangle.Fill> <LinearGradientBrush StartPoint="0.142631,0.75344" EndPoint="1.01886,0.75344"> <LinearGradientBrush.RelativeTransform> <TransformGroup> <SkewTransform CenterX="0.142631" CenterY="0.75344" AngleX="19.3128" AngleY="0"/> <RotateTransform CenterX="0.142631" CenterY="0.75344" Angle="-35.3436"/> </TransformGroup> </LinearGradientBrush.RelativeTransform> <LinearGradientBrush.GradientStops> <GradientStop Color="#FFCDCDCD" Offset="0.0833333"/> <GradientStop Color="#FFFFFFFF" Offset="1"/> </LinearGradientBrush.GradientStops> </LinearGradientBrush> </Rectangle.Fill> </Rectangle> <Rectangle Width="26.4269" Height="45.8414" Canvas.Left="0.227798" Canvas.Top="0" Stretch="Fill"> <Rectangle.Fill> <LinearGradientBrush StartPoint="0.142631,0.75344" EndPoint="1.01886,0.75344"> <LinearGradientBrush.RelativeTransform> <TransformGroup> <SkewTransform CenterX="0.142631" CenterY="0.75344" AngleX="19.3127" AngleY="0"/> <RotateTransform CenterX="0.142631" CenterY="0.75344" Angle="-35.3436"/> </TransformGroup> </LinearGradientBrush.RelativeTransform> <LinearGradientBrush.GradientStops> <GradientStop Color="#FF7B6802" Offset="0"/> <GradientStop Color="#FFF3D42C" Offset="1"/> </LinearGradientBrush.GradientStops> </LinearGradientBrush> </Rectangle.Fill> </Rectangle> <Rectangle Width="1.25301" Height="45.8414" Canvas.Left="1.70862" Canvas.Top="0.116943" Stretch="Fill" Fill="#FFEBFF07"/> </Canvas> <TextBlock Height="80" x:Name="Name" Width="448" TextWrapping="Wrap" VerticalAlignment="Center" FontSize="24" Text="Directory"/> </StackPanel> </UserControl> As you can see, this XAML contains many graphic elements. Those elements are used to design the folder icon. The original drawing has been designed in Expression Design and then exported as XAML. In Silverlight for Windows Embedded you can use vector images. This means that your images will look good even when scaled or rotated. In our DirectoryItem custom control we have a TextBlock named Name, that will be used to display….(suspense)…. the directory name (I’m too lazy to invent fancy names for controls, and using “boring” intuitive names will make code more readable, I hope!). Now that we have some XAML code, we may execute XAML2CPP to generate part of the aplication code for us. We should then add references to our XAML2CPP generated resource file and include in our code and add a reference to the XAML runtime library to our sources file (you can follow the instruction of the first tutorial step to do that), To generate the code used in this tutorial you need XAML2CPP ver 1.0.1.0, that is downloadable here: http://geekswithblogs.net/WindowsEmbeddedCookbook/archive/2010/03/08/xaml2cpp-1.0.1.0.aspx We can now create our usual simple Win32 application inside Platform Builder, using the same step described in the first chapter of this tutorial (http://geekswithblogs.net/WindowsEmbeddedCookbook/archive/2009/10/01/silverlight-for-embedded-tutorial.aspx). We can declare a class for our main page, deriving it from the template that XAML2CPP generated for us: class ListPage : public TListPage<ListPage> { ... } We will see the ListPage class code in a short time, but before we will see the code of our DirectoryItem user control. This object will be used to populate our list, one item for each directory. To declare a user control things are a bit more complicated (but also in this case XAML2CPP will write most of the “boilerplate” code for use. To interact with a user control you should declare an interface. An interface defines the functions of a user control that can be called inside the application code. Our custom control is currently quite simple and we just need some member functions to store and retrieve a full pathname inside our control. The control will display just the last part of the path inside the control. An interface is declared as a C++ class that has only abstract virtual members. It should also have an UUID associated with it. UUID means Universal Unique IDentifier and it’s a 128 bit number that will identify our interface without the need of specifying its fully qualified name. UUIDs are used to identify COM interfaces and, as we discovered in chapter one, Silverlight for Windows Embedded is based on COM or, at least, provides a COM-like Application Programming Interface (API). Here’s the declaration of the DirectoryItem interface: class __declspec(novtable,uuid("{D38C66E5-2725-4111-B422-D75B32AA8702}")) IDirectoryItem : public IXRCustomUserControl { public:   virtual HRESULT SetFullPath(BSTR fullpath) = 0; virtual HRESULT GetFullPath(BSTR* retval) = 0; }; The interface is derived from IXRCustomControl, this will allow us to add our object to a XAML tree. It declares the two functions needed to set and get the full path, but don’t implement them. Implementation will be done inside the control class. The interface only defines the functions of our control class that are accessible from the outside. It’s a sort of “contract” between our control and the applications that will use it. We must support what’s inside the contract and the application code should know nothing else about our own control. To reference our interface we will use the UUID, to make code more readable we can declare a #define in this way: #define IID_IDirectoryItem __uuidof(IDirectoryItem) Silverlight for Windows Embedded objects (like COM objects) use a reference counting mechanism to handle object destruction. Every time you store a pointer to an object you should call its AddRef function and every time you no longer need that pointer you should call Release. The object keeps an internal counter, incremented for each AddRef and decremented on Release. When the counter reaches 0, the object is destroyed. Managing reference counting in our code can be quite complicated and, since we are lazy (I am, at least!), we will use a great feature of Silverlight for Windows Embedded: smart pointers.A smart pointer can be connected to a Silverlight for Windows Embedded object and manages its reference counting. To declare a smart pointer we must use the XRPtr template: typedef XRPtr<IDirectoryItem> IDirectoryItemPtr; Now that we have defined our interface, it’s time to implement our user control class. XAML2CPP has implemented a class for us, and we have only to derive our class from it, defining the main class and interface of our new custom control: class DirectoryItem : public DirectoryItemUserControlRegister<DirectoryItem,IDirectoryItem> { ... } XAML2CPP has generated some code for us to support the user control, we don’t have to mind too much about that code, since it will be generated (or written by hand, if you like) always in the same way, for every user control. But knowing how does this works “under the hood” is still useful to understand the architecture of Silverlight for Windows Embedded. Our base class declaration is a bit more complex than the one we used for a simple page in the previous chapters: template <class A,class B> class DirectoryItemUserControlRegister : public XRCustomUserControlImpl<A,B>,public TDirectoryItem<A,XAML2CPPUserControl> { ... } This class derives from the XAML2CPP generated template class, like the ListPage class, but it uses XAML2CPPUserControl for the implementation of some features. This class shares the same ancestor of XAML2CPPPage (base class for “regular” XAML pages), XAML2CPPBase, implements binding of member variables and event handlers but, instead of loading and creating its own XAML tree, it attaches to an existing one. The XAML tree (and UI) of our custom control is created and loaded by the XRCustomUserControlImpl class. This class is part of the Silverlight for Windows Embedded framework and implements most of the functions needed to build-up a custom control in Silverlight (the guys that developed Silverlight for Windows Embedded seem to care about lazy programmers!). We have just to initialize it, providing our class (DirectoryItem) and interface (IDirectoryItem). Our user control class has also a static member: protected:   static HINSTANCE hInstance; This is used to store the HINSTANCE of the modules that contain our user control class. I don’t like this implementation, but I can’t find a better one, so if somebody has good ideas about how to handle the HINSTANCE object, I’ll be happy to hear suggestions! It also implements two static members required by XRCustomUserControlImpl. The first one is used to load the XAML UI of our custom control: static HRESULT GetXamlSource(XRXamlSource* pXamlSource) { pXamlSource->SetResource(hInstance,TEXT("XAML"),IDR_XAML_DirectoryItem); return S_OK; }   It initializes a XRXamlSource object, connecting it to the XAML resource that XAML2CPP has included in our resource script. The other method is used to register our custom control, allowing Silverlight for Windows Embedded to create it when it load some XAML or when an application creates a new control at runtime (more about this later): static HRESULT Register() { return XRCustomUserControlImpl<A,B>::Register(__uuidof(B), L"DirectoryItem", L"clr-namespace:DirectoryItemNamespace"); } To register our control we should provide its interface UUID, the name of the corresponding element in the XAML tree and its current namespace (namespaces compatible with Silverlight must use the “clr-namespace” prefix. We may also register additional properties for our objects, allowing them to be loaded and saved inside XAML. In this case we have no permanent properties and the Register method will just register our control. An additional static method is implemented to allow easy registration of our custom control inside our application WinMain function: static HRESULT RegisterUserControl(HINSTANCE hInstance) { DirectoryItemUserControlRegister::hInstance=hInstance; return DirectoryItemUserControlRegister<A,B>::Register(); } Now our control is registered and we will be able to create it using the Silverlight for Windows Embedded runtime functions. But we need to bind our members and event handlers to have them available like we are used to do for other XAML2CPP generated objects. To bind events and members we need to implement the On_Loaded function: virtual HRESULT OnLoaded(__in IXRDependencyObject* pRoot) { HRESULT retcode; IXRApplicationPtr app; if (FAILED(retcode=GetXRApplicationInstance(&app))) return retcode; return ((A*)this)->Init(pRoot,hInstance,app); } This function will call the XAML2CPPUserControl::Init member that will connect the “root” member with the XAML sub tree that has been created for our control and then calls BindObjects and BindEvents to bind members and events to our code. Now we can go back to our application code (the code that you’ll have to actually write) to see the contents of our DirectoryItem class: class DirectoryItem : public DirectoryItemUserControlRegister<DirectoryItem,IDirectoryItem> { protected:   WCHAR fullpath[_MAX_PATH+1];   public:   DirectoryItem() { *fullpath=0; }   virtual HRESULT SetFullPath(BSTR fullpath) { wcscpy_s(this->fullpath,fullpath);   WCHAR* p=fullpath;   for(WCHAR*q=wcsstr(p,L"\\");q;p=q+1,q=wcsstr(p,L"\\")) ;   Name->SetText(p); return S_OK; }   virtual HRESULT GetFullPath(BSTR* retval) { *retval=SysAllocString(fullpath); return S_OK; } }; It’s pretty easy and contains a fullpath member (used to store that path of the directory connected with the user control) and the implementation of the two interface members that can be used to set and retrieve the path. The SetFullPath member parses the full path and displays just the last branch directory name inside the “Name” TextBlock object. As you can see, implementing a user control in Silverlight for Windows Embedded is not too complex and using XAML also for the UI of the control allows us to re-use the same mechanisms that we learnt and used in the previous steps of our tutorial. Now let’s see how the main page is managed by the ListPage class. class ListPage : public TListPage<ListPage> { protected:   // current path TCHAR curpath[_MAX_PATH+1]; It has a member named “curpath” that is used to store the current directory. It’s initialized inside the constructor: ListPage() { *curpath=0; } And it’s value is displayed inside the “CurrentDir” TextBlock inside the initialization function: virtual HRESULT Init(HINSTANCE hInstance,IXRApplication* app) { HRESULT retcode;   if (FAILED(retcode=TListPage<ListPage>::Init(hInstance,app))) return retcode;   CurrentDir->SetText(L"\\"); return S_OK; } The FillFileList function is used to enumerate subdirectories of the current dir and add entries for each one inside the list box that fills most of the client area of our main page: HRESULT FillFileList() { HRESULT retcode; IXRItemCollectionPtr items; IXRApplicationPtr app;   if (FAILED(retcode=GetXRApplicationInstance(&app))) return retcode; // retrieves the items contained in the listbox if (FAILED(retcode=FileList->GetItems(&items))) return retcode;   // clears the list if (FAILED(retcode=items->Clear())) return retcode;   // enumerates files and directory in the current path WCHAR filemask[_MAX_PATH+1];   wcscpy_s(filemask,curpath); wcscat_s(filemask,L"\\*.*");   WIN32_FIND_DATA finddata; HANDLE findhandle;   findhandle=FindFirstFile(filemask,&finddata);   // the directory is empty? if (findhandle==INVALID_HANDLE_VALUE) return S_OK;   do { if (finddata.dwFileAttributes&=FILE_ATTRIBUTE_DIRECTORY) { IXRListBoxItemPtr listboxitem;   // add a new item to the listbox if (FAILED(retcode=app->CreateObject(IID_IXRListBoxItem,&listboxitem))) { FindClose(findhandle); return retcode; }   if (FAILED(retcode=items->Add(listboxitem,NULL))) { FindClose(findhandle); return retcode; }   IDirectoryItemPtr directoryitem;   if (FAILED(retcode=app->CreateObject(IID_IDirectoryItem,&directoryitem))) { FindClose(findhandle); return retcode; }   WCHAR fullpath[_MAX_PATH+1];   wcscpy_s(fullpath,curpath); wcscat_s(fullpath,L"\\"); wcscat_s(fullpath,finddata.cFileName);   if (FAILED(retcode=directoryitem->SetFullPath(fullpath))) { FindClose(findhandle); return retcode; }   XAML2CPPXRValue value((IXRDependencyObject*)directoryitem);   if (FAILED(retcode=listboxitem->SetContent(&value))) { FindClose(findhandle); return retcode; } } } while (FindNextFile(findhandle,&finddata));   FindClose(findhandle); return S_OK; } This functions retrieve a pointer to the collection of the items contained in the directory listbox. The IXRItemCollection interface is used by listboxes and comboboxes and allow you to clear the list (using Clear(), as our function does at the beginning) and change its contents by adding and removing elements. This function uses the FindFirstFile/FindNextFile functions to enumerate all the objects inside our current directory and for each subdirectory creates a IXRListBoxItem object. You can insert any kind of control inside a list box, you don’t need a IXRListBoxItem, but using it will allow you to handle the selected state of an item, highlighting it inside the list. The function creates a list box item using the CreateObject function of XRApplication. The same function is then used to create an instance of our custom control. The function returns a pointer to the control IDirectoryItem interface and we can use it to store the directory full path inside the object and add it as content of the IXRListBox item object, adding it to the listbox contents. The listbox generates an event (SelectionChanged) each time the user clicks on one of the items contained in the listbox. We implement an event handler for that event and use it to change our current directory and repopulate the listbox. The current directory full path will be displayed in the TextBlock: HRESULT Filelist_SelectionChanged(IXRDependencyObject* source,XRSelectionChangedEventArgs* args) { HRESULT retcode;   IXRListBoxItemPtr listboxitem;   if (!args->pAddedItem) return S_OK;   if (FAILED(retcode=args->pAddedItem->QueryInterface(IID_IXRListBoxItem,(void**)&listboxitem))) return retcode;   XRValue content; if (FAILED(retcode=listboxitem->GetContent(&content))) return retcode;   if (content.vType!=VTYPE_OBJECT) return E_FAIL;   IDirectoryItemPtr directoryitem;   if (FAILED(retcode=content.pObjectVal->QueryInterface(IID_IDirectoryItem,(void**)&directoryitem))) return retcode;   content.pObjectVal->Release(); content.pObjectVal=NULL;   BSTR fullpath=NULL;   if (FAILED(retcode=directoryitem->GetFullPath(&fullpath))) return retcode;   CurrentDir->SetText(fullpath);   wcscpy_s(curpath,fullpath); FillFileList(); SysFreeString(fullpath);     return S_OK; } }; The function uses the pAddedItem member of the XRSelectionChangedEventArgs object to retrieve the currently selected item, converts it to a IXRListBoxItem interface using QueryInterface, and then retrives its contents (IDirectoryItem object). Using the GetFullPath method we can get the full path of our selected directory and assing it to the curdir member. A call to FillFileList will update the listbox contents, displaying the list of subdirectories of the selected folder. To build our sample we just need to add code to our WinMain function: int WINAPI WinMain(HINSTANCE hInstance, HINSTANCE hPrevInstance, LPTSTR lpCmdLine, int nCmdShow) { if (!XamlRuntimeInitialize()) return -1;   HRESULT retcode;   IXRApplicationPtr app; if (FAILED(retcode=GetXRApplicationInstance(&app))) return -1;   if (FAILED(retcode=DirectoryItem::RegisterUserControl(hInstance))) return retcode;   ListPage page;   if (FAILED(page.Init(hInstance,app))) return -1;   page.FillFileList();   UINT exitcode;   if (FAILED(page.GetVisualHost()->StartDialog(&exitcode))) return -1;   return 0; } This code is very similar to the one of the WinMains of our previous samples. The main differences are that we register our custom control (you should do that as soon as you have initialized the XAML runtime) and call FillFileList after the initialization of our ListPage object to load the contents of the root folder of our device inside the listbox. As usual you can download the full sample source code from here: http://cid-9b7b0aefe3514dc5.skydrive.live.com/self.aspx/.Public/ListBoxTest.zip

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  • What&rsquo;s New in ASP.NET 4.0 Part Two: WebForms and Visual Studio Enhancements

    - by Rick Strahl
    In the last installment I talked about the core changes in the ASP.NET runtime that I’ve been taking advantage of. In this column, I’ll cover the changes to the Web Forms engine and some of the cool improvements in Visual Studio that make Web and general development easier. WebForms The WebForms engine is the area that has received most significant changes in ASP.NET 4.0. Probably the most widely anticipated features are related to managing page client ids and of ViewState on WebForm pages. Take Control of Your ClientIDs Unique ClientID generation in ASP.NET has been one of the most complained about “features” in ASP.NET. Although there’s a very good technical reason for these unique generated ids - they guarantee unique ids for each and every server control on a page - these unique and generated ids often get in the way of client-side JavaScript development and CSS styling as it’s often inconvenient and fragile to work with the long, generated ClientIDs. In ASP.NET 4.0 you can now specify an explicit client id mode on each control or each naming container parent control to control how client ids are generated. By default, ASP.NET generates mangled client ids for any control contained in a naming container (like a Master Page, or a User Control for example). The key to ClientID management in ASP.NET 4.0 are the new ClientIDMode and ClientIDRowSuffix properties. ClientIDMode supports four different ClientID generation settings shown below. For the following examples, imagine that you have a Textbox control named txtName inside of a master page control container on a WebForms page. <%@Page Language="C#"      MasterPageFile="~/Site.Master"     CodeBehind="WebForm2.aspx.cs"     Inherits="WebApplication1.WebForm2"  %> <asp:Content ID="content"  ContentPlaceHolderID="content"               runat="server"               ClientIDMode="Static" >       <asp:TextBox runat="server" ID="txtName" /> </asp:Content> The four available ClientIDMode values are: AutoID This is the existing behavior in ASP.NET 1.x-3.x where full naming container munging takes place. <input name="ctl00$content$txtName" type="text"        id="ctl00_content_txtName" /> This should be familiar to any ASP.NET developer and results in fairly unpredictable client ids that can easily change if the containership hierarchy changes. For example, removing the master page changes the name in this case, so if you were to move a block of script code that works against the control to a non-Master page, the script code immediately breaks. Static This option is the most deterministic setting that forces the control’s ClientID to use its ID value directly. No naming container naming at all is applied and you end up with clean client ids: <input name="ctl00$content$txtName"         type="text" id="txtName" /> Note that the name property which is used for postback variables to the server still is munged, but the ClientID property is displayed simply as the ID value that you have assigned to the control. This option is what most of us want to use, but you have to be clear on that because it can potentially cause conflicts with other controls on the page. If there are several instances of the same naming container (several instances of the same user control for example) there can easily be a client id naming conflict. Note that if you assign Static to a data-bound control, like a list child control in templates, you do not get unique ids either, so for list controls where you rely on unique id for child controls, you’ll probably want to use Predictable rather than Static. I’ll write more on this a little later when I discuss ClientIDRowSuffix. Predictable The previous two values are pretty self-explanatory. Predictable however, requires some explanation. To me at least it’s not in the least bit predictable. MSDN defines this value as follows: This algorithm is used for controls that are in data-bound controls. The ClientID value is generated by concatenating the ClientID value of the parent naming container with the ID value of the control. If the control is a data-bound control that generates multiple rows, the value of the data field specified in the ClientIDRowSuffix property is added at the end. For the GridView control, multiple data fields can be specified. If the ClientIDRowSuffix property is blank, a sequential number is added at the end instead of a data-field value. Each segment is separated by an underscore character (_). The key that makes this value a bit confusing is that it relies on the parent NamingContainer’s ClientID to build its own ClientID value. This effectively means that the value is not predictable at all but rather very tightly coupled to the parent naming container’s ClientIDMode setting. For my simple textbox example, if the ClientIDMode property of the parent naming container (Page in this case) is set to “Predictable” you’ll get this: <input name="ctl00$content$txtName" type="text"         id="content_txtName" /> which gives an id that based on walking up to the currently active naming container (the MasterPage content container) and starting the id formatting from there downward. Think of this as a semi unique name that’s guaranteed unique only for the naming container. If, on the other hand, the Page is set to “AutoID” you get the following with Predictable on txtName: <input name="ctl00$content$txtName" type="text"         id="ctl00_content_txtName" /> The latter is effectively the same as if you specified AutoID because it inherits the AutoID naming from the Page and Content Master Page control of the page. But again - predictable behavior always depends on the parent naming container and how it generates its id, so the id may not always be exactly the same as the AutoID generated value because somewhere in the NamingContainer chain the ClientIDMode setting may be set to a different value. For example, if you had another naming container in the middle that was set to Static you’d end up effectively with an id that starts with the NamingContainers id rather than the whole ctl000_content munging. The most common use for Predictable is likely to be for data-bound controls, which results in each data bound item getting a unique ClientID. Unfortunately, even here the behavior can be very unpredictable depending on which data-bound control you use - I found significant differences in how template controls in a GridView behave from those that are used in a ListView control. For example, GridView creates clean child ClientIDs, while ListView still has a naming container in the ClientID, presumably because of the template container on which you can’t set ClientIDMode. Predictable is useful, but only if all naming containers down the chain use this setting. Otherwise you’re right back to the munged ids that are pretty unpredictable. Another property, ClientIDRowSuffix, can be used in combination with ClientIDMode of Predictable to force a suffix onto list client controls. For example: <asp:GridView runat="server" ID="gvItems"              AutoGenerateColumns="false"             ClientIDMode="Static"              ClientIDRowSuffix="Id">     <Columns>     <asp:TemplateField>         <ItemTemplate>             <asp:Label runat="server" id="txtName"                        Text='<%# Eval("Name") %>'                   ClientIDMode="Predictable"/>         </ItemTemplate>     </asp:TemplateField>     <asp:TemplateField>         <ItemTemplate>         <asp:Label runat="server" id="txtId"                     Text='<%# Eval("Id") %>'                     ClientIDMode="Predictable" />         </ItemTemplate>     </asp:TemplateField>     </Columns>  </asp:GridView> generates client Ids inside of a column in the master page described earlier: <td>     <span id="txtName_0">Rick</span> </td> where the value after the underscore is the ClientIDRowSuffix field - in this case “Id” of the item data bound to the control. Note that all of the child controls require ClientIDMode=”Predictable” in order for the ClientIDRowSuffix to be applied, and the parent GridView controls need to be set to Static either explicitly or via Naming Container inheritance to give these simple names. It’s a bummer that ClientIDRowSuffix doesn’t work with Static to produce this automatically. Another real problem is that other controls process the ClientIDMode differently. For example, a ListView control processes the Predictable ClientIDMode differently and produces the following with the Static ListView and Predictable child controls: <span id="ctrl0_txtName_0">Rick</span> I couldn’t even figure out a way using ClientIDMode to get a simple ID that also uses a suffix short of falling back to manually generated ids using <%= %> expressions instead. Given the inconsistencies inside of list controls using <%= %>, ids for the ListView might not be a bad idea anyway. Inherit The final setting is Inherit, which is the default for all controls except Page. This means that controls by default inherit the parent naming container’s ClientIDMode setting. For more detailed information on ClientID behavior and different scenarios you can check out a blog post of mine on this subject: http://www.west-wind.com/weblog/posts/54760.aspx. ClientID Enhancements Summary The ClientIDMode property is a welcome addition to ASP.NET 4.0. To me this is probably the most useful WebForms feature as it allows me to generate clean IDs simply by setting ClientIDMode="Static" on either the page or inside of Web.config (in the Pages section) which applies the setting down to the entire page which is my 95% scenario. For the few cases when it matters - for list controls and inside of multi-use user controls or custom server controls) - I can use Predictable or even AutoID to force controls to unique names. For application-level page development, this is easy to accomplish and provides maximum usability for working with client script code against page controls. ViewStateMode Another area of large criticism for WebForms is ViewState. ViewState is used internally by ASP.NET to persist page-level changes to non-postback properties on controls as pages post back to the server. It’s a useful mechanism that works great for the overall mechanics of WebForms, but it can also cause all sorts of overhead for page operation as ViewState can very quickly get out of control and consume huge amounts of bandwidth in your page content. ViewState can also wreak havoc with client-side scripting applications that modify control properties that are tracked by ViewState, which can produce very unpredictable results on a Postback after client-side updates. Over the years in my own development, I’ve often turned off ViewState on pages to reduce overhead. Yes, you lose some functionality, but you can easily implement most of the common functionality in non-ViewState workarounds. Relying less on heavy ViewState controls and sticking with simpler controls or raw HTML constructs avoids getting around ViewState problems. In ASP.NET 3.x and prior, it wasn’t easy to control ViewState - you could turn it on or off and if you turned it off at the page or web.config level, you couldn’t turn it back on for specific controls. In short, it was an all or nothing approach. With ASP.NET 4.0, the new ViewStateMode property gives you more control. It allows you to disable ViewState globally either on the page or web.config level and then turn it back on for specific controls that might need it. ViewStateMode only works when EnableViewState="true" on the page or web.config level (which is the default). You can then use ViewStateMode of Disabled, Enabled or Inherit to control the ViewState settings on the page. If you’re shooting for minimal ViewState usage, the ideal situation is to set ViewStateMode to disabled on the Page or web.config level and only turn it back on particular controls: <%@Page Language="C#"      CodeBehind="WebForm2.aspx.cs"     Inherits="Westwind.WebStore.WebForm2"        ClientIDMode="Static"                ViewStateMode="Disabled"     EnableViewState="true"  %> <!-- this control has viewstate  --> <asp:TextBox runat="server" ID="txtName"  ViewStateMode="Enabled" />       <!-- this control has no viewstate - it inherits  from parent container --> <asp:TextBox runat="server" ID="txtAddress" /> Note that the EnableViewState="true" at the Page level isn’t required since it’s the default, but it’s important that the value is true. ViewStateMode has no effect if EnableViewState="false" at the page level. The main benefit of ViewStateMode is that it allows you to more easily turn off ViewState for most of the page and enable only a few key controls that might need it. For me personally, this is a perfect combination as most of my WebForm apps can get away without any ViewState at all. But some controls - especially third party controls - often don’t work well without ViewState enabled, and now it’s much easier to selectively enable controls rather than the old way, which required you to pretty much turn off ViewState for all controls that you didn’t want ViewState on. Inline HTML Encoding HTML encoding is an important feature to prevent cross-site scripting attacks in data entered by users on your site. In order to make it easier to create HTML encoded content, ASP.NET 4.0 introduces a new Expression syntax using <%: %> to encode string values. The encoding expression syntax looks like this: <%: "<script type='text/javascript'>" +     "alert('Really?');</script>" %> which produces properly encoded HTML: &lt;script type=&#39;text/javascript&#39; &gt;alert(&#39;Really?&#39;);&lt;/script&gt; Effectively this is a shortcut to: <%= HttpUtility.HtmlEncode( "<script type='text/javascript'>" + "alert('Really?');</script>") %> Of course the <%: %> syntax can also evaluate expressions just like <%= %> so the more common scenario applies this expression syntax against data your application is displaying. Here’s an example displaying some data model values: <%: Model.Address.Street %> This snippet shows displaying data from your application’s data store or more importantly, from data entered by users. Anything that makes it easier and less verbose to HtmlEncode text is a welcome addition to avoid potential cross-site scripting attacks. Although I listed Inline HTML Encoding here under WebForms, anything that uses the WebForms rendering engine including ASP.NET MVC, benefits from this feature. ScriptManager Enhancements The ASP.NET ScriptManager control in the past has introduced some nice ways to take programmatic and markup control over script loading, but there were a number of shortcomings in this control. The ASP.NET 4.0 ScriptManager has a number of improvements that make it easier to control script loading and addresses a few of the shortcomings that have often kept me from using the control in favor of manual script loading. The first is the AjaxFrameworkMode property which finally lets you suppress loading the ASP.NET AJAX runtime. Disabled doesn’t load any ASP.NET AJAX libraries, but there’s also an Explicit mode that lets you pick and choose the library pieces individually and reduce the footprint of ASP.NET AJAX script included if you are using the library. There’s also a new EnableCdn property that forces any script that has a new WebResource attribute CdnPath property set to a CDN supplied URL. If the script has this Attribute property set to a non-null/empty value and EnableCdn is enabled on the ScriptManager, that script will be served from the specified CdnPath. [assembly: WebResource(    "Westwind.Web.Resources.ww.jquery.js",    "application/x-javascript",    CdnPath =  "http://mysite.com/scripts/ww.jquery.min.js")] Cool, but a little too static for my taste since this value can’t be changed at runtime to point at a debug script as needed, for example. Assembly names for loading scripts from resources can now be simple names rather than fully qualified assembly names, which make it less verbose to reference scripts from assemblies loaded from your bin folder or the assembly reference area in web.config: <asp:ScriptManager runat="server" id="Id"          EnableCdn="true"         AjaxFrameworkMode="disabled">     <Scripts>         <asp:ScriptReference          Name="Westwind.Web.Resources.ww.jquery.js"         Assembly="Westwind.Web" />     </Scripts>        </asp:ScriptManager> The ScriptManager in 4.0 also supports script combining via the CompositeScript tag, which allows you to very easily combine scripts into a single script resource served via ASP.NET. Even nicer: You can specify the URL that the combined script is served with. Check out the following script manager markup that combines several static file scripts and a script resource into a single ASP.NET served resource from a static URL (allscripts.js): <asp:ScriptManager runat="server" id="Id"          EnableCdn="true"         AjaxFrameworkMode="disabled">     <CompositeScript          Path="~/scripts/allscripts.js">         <Scripts>             <asp:ScriptReference                    Path="~/scripts/jquery.js" />             <asp:ScriptReference                    Path="~/scripts/ww.jquery.js" />             <asp:ScriptReference            Name="Westwind.Web.Resources.editors.js"                 Assembly="Westwind.Web" />         </Scripts>     </CompositeScript> </asp:ScriptManager> When you render this into HTML, you’ll see a single script reference in the page: <script src="scripts/allscripts.debug.js"          type="text/javascript"></script> All you need to do to make this work is ensure that allscripts.js and allscripts.debug.js exist in the scripts folder of your application - they can be empty but the file has to be there. This is pretty cool, but you want to be real careful that you use unique URLs for each combination of scripts you combine or else browser and server caching will easily screw you up royally. The script manager also allows you to override native ASP.NET AJAX scripts now as any script references defined in the Scripts section of the ScriptManager trump internal references. So if you want custom behavior or you want to fix a possible bug in the core libraries that normally are loaded from resources, you can now do this simply by referencing the script resource name in the Name property and pointing at System.Web for the assembly. Not a common scenario, but when you need it, it can come in real handy. Still, there are a number of shortcomings in this control. For one, the ScriptManager and ClientScript APIs still have no common entry point so control developers are still faced with having to check and support both APIs to load scripts so that controls can work on pages that do or don’t have a ScriptManager on the page. The CdnUrl is static and compiled in, which is very restrictive. And finally, there’s still no control over where scripts get loaded on the page - ScriptManager still injects scripts into the middle of the HTML markup rather than in the header or optionally the footer. This, in turn, means there is little control over script loading order, which can be problematic for control developers. MetaDescription, MetaKeywords Page Properties There are also a number of additional Page properties that correspond to some of the other features discussed in this column: ClientIDMode, ClientTarget and ViewStateMode. Another minor but useful feature is that you can now directly access the MetaDescription and MetaKeywords properties on the Page object to set the corresponding meta tags programmatically. Updating these values programmatically previously required either <%= %> expressions in the page markup or dynamic insertion of literal controls into the page. You can now just set these properties programmatically on the Page object in any Control derived class on the page or the Page itself: Page.MetaKeywords = "ASP.NET,4.0,New Features"; Page.MetaDescription = "This article discusses the new features in ASP.NET 4.0"; Note, that there’s no corresponding ASP.NET tag for the HTML Meta element, so the only way to specify these values in markup and access them is via the @Page tag: <%@Page Language="C#"      CodeBehind="WebForm2.aspx.cs"     Inherits="Westwind.WebStore.WebForm2"      ClientIDMode="Static"                MetaDescription="Article that discusses what's                      new in ASP.NET 4.0"     MetaKeywords="ASP.NET,4.0,New Features" %> Nothing earth shattering but quite convenient. Visual Studio 2010 Enhancements for Web Development For Web development there are also a host of editor enhancements in Visual Studio 2010. Some of these are not Web specific but they are useful for Web developers in general. Text Editors Throughout Visual Studio 2010, the text editors have all been updated to a new core engine based on WPF which provides some interesting new features for various code editors including the nice ability to zoom in and out with Ctrl-MouseWheel to quickly change the size of text. There are many more API options to control the editor and although Visual Studio 2010 doesn’t yet use many of these features, we can look forward to enhancements in add-ins and future editor updates from the various language teams that take advantage of the visual richness that WPF provides to editing. On the negative side, I’ve noticed that occasionally the code editor and especially the HTML and JavaScript editors will lose the ability to use various navigation keys like arrows, back and delete keys, which requires closing and reopening the documents at times. This issue seems to be well documented so I suspect this will be addressed soon with a hotfix or within the first service pack. Overall though, the code editors work very well, especially given that they were re-written completely using WPF, which was one of my big worries when I first heard about the complete redesign of the editors. Multi-Targeting Visual Studio now targets all versions of the .NET framework from 2.0 forward. You can use Visual Studio 2010 to work on your ASP.NET 2, 3.0 and 3.5 applications which is a nice way to get your feet wet with the new development environment without having to make changes to existing applications. It’s nice to have one tool to work in for all the different versions. Multi-Monitor Support One cool feature of Visual Studio 2010 is the ability to drag windows out of the Visual Studio environment and out onto the desktop including onto another monitor easily. Since Web development often involves working with a host of designers at the same time - visual designer, HTML markup window, code behind and JavaScript editor - it’s really nice to be able to have a little more screen real estate to work on each of these editors. Microsoft made a welcome change in the environment. IntelliSense Snippets for HTML and JavaScript Editors The HTML and JavaScript editors now finally support IntelliSense scripts to create macro-based template expansions that have been in the core C# and Visual Basic code editors since Visual Studio 2005. Snippets allow you to create short XML-based template definitions that can act as static macros or real templates that can have replaceable values that can be embedded into the expanded text. The XML syntax for these snippets is straight forward and it’s pretty easy to create custom snippets manually. You can easily create snippets using XML and store them in your custom snippets folder (C:\Users\rstrahl\Documents\Visual Studio 2010\Code Snippets\Visual Web Developer\My HTML Snippets and My JScript Snippets), but it helps to use one of the third-party tools that exist to simplify the process for you. I use SnippetEditor, by Bill McCarthy, which makes short work of creating snippets interactively (http://snippeteditor.codeplex.com/). Note: You may have to manually add the Visual Studio 2010 User specific Snippet folders to this tool to see existing ones you’ve created. Code snippets are some of the biggest time savers and HTML editing more than anything deals with lots of repetitive tasks that lend themselves to text expansion. Visual Studio 2010 includes a slew of built-in snippets (that you can also customize!) and you can create your own very easily. If you haven’t done so already, I encourage you to spend a little time examining your coding patterns and find the repetitive code that you write and convert it into snippets. I’ve been using CodeRush for this for years, but now you can do much of the basic expansion natively for HTML and JavaScript snippets. jQuery Integration Is Now Native jQuery is a popular JavaScript library and recently Microsoft has recently stated that it will become the primary client-side scripting technology to drive higher level script functionality in various ASP.NET Web projects that Microsoft provides. In Visual Studio 2010, the default full project template includes jQuery as part of a new project including the support files that provide IntelliSense (-vsdoc files). IntelliSense support for jQuery is now also baked into Visual Studio 2010, so unlike Visual Studio 2008 which required a separate download, no further installs are required for a rich IntelliSense experience with jQuery. Summary ASP.NET 4.0 brings many useful improvements to the platform, but thankfully most of the changes are incremental changes that don’t compromise backwards compatibility and they allow developers to ease into the new features one feature at a time. None of the changes in ASP.NET 4.0 or Visual Studio 2010 are monumental or game changers. The bigger features are language and .NET Framework changes that are also optional. This ASP.NET and tools release feels more like fine tuning and getting some long-standing kinks worked out of the platform. It shows that the ASP.NET team is dedicated to paying attention to community feedback and responding with changes to the platform and development environment based on this feedback. If you haven’t gotten your feet wet with ASP.NET 4.0 and Visual Studio 2010, there’s no reason not to give it a shot now - the ASP.NET 4.0 platform is solid and Visual Studio 2010 works very well for a brand new release. Check it out. © Rick Strahl, West Wind Technologies, 2005-2010Posted in ASP.NET  

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  • Powershell script to change screen Orientation

    - by user161964
    I wrote a script to change Primary screen orientation to portrait. my screen is 1920X1200 It runs and no error reported. But the screen does not rotated as i expected. The code was modified from Set-ScreenResolution (Andy Schneider) Does anybody can help me take a look? some reference site: 1.set-screenresolution http://gallery.technet.microsoft.com/ScriptCenter/2a631d72-206d-4036-a3f2-2e150f297515/ 2.C code for change oridentation (MSDN) Changing Screen Orientation Programmatically http://msdn.microsoft.com/en-us/library/ms812499.aspx my code as below: Function Set-ScreenOrientation { $pinvokeCode = @" using System; using System.Runtime.InteropServices; namespace Resolution { [StructLayout(LayoutKind.Sequential)] public struct DEVMODE1 { [MarshalAs(UnmanagedType.ByValTStr, SizeConst = 32)] public string dmDeviceName; public short dmSpecVersion; public short dmDriverVersion; public short dmSize; public short dmDriverExtra; public int dmFields; public short dmOrientation; public short dmPaperSize; public short dmPaperLength; public short dmPaperWidth; public short dmScale; public short dmCopies; public short dmDefaultSource; public short dmPrintQuality; public short dmColor; public short dmDuplex; public short dmYResolution; public short dmTTOption; public short dmCollate; [MarshalAs(UnmanagedType.ByValTStr, SizeConst = 32)] public string dmFormName; [MarshalAs(UnmanagedType.U4)] public short dmDisplayOrientation public short dmLogPixels; public short dmBitsPerPel; public int dmPelsWidth; public int dmPelsHeight; public int dmDisplayFlags; public int dmDisplayFrequency; public int dmICMMethod; public int dmICMIntent; public int dmMediaType; public int dmDitherType; public int dmReserved1; public int dmReserved2; public int dmPanningWidth; public int dmPanningHeight; }; class User_32 { [DllImport("user32.dll")] public static extern int EnumDisplaySettings(string deviceName, int modeNum, ref DEVMODE1 devMode); [DllImport("user32.dll")] public static extern int ChangeDisplaySettings(ref DEVMODE1 devMode, int flags); public const int ENUM_CURRENT_SETTINGS = -1; public const int CDS_UPDATEREGISTRY = 0x01; public const int CDS_TEST = 0x02; public const int DISP_CHANGE_SUCCESSFUL = 0; public const int DISP_CHANGE_RESTART = 1; public const int DISP_CHANGE_FAILED = -1; } public class PrmaryScreenOrientation { static public string ChangeOrientation() { DEVMODE1 dm = GetDevMode1(); if (0 != User_32.EnumDisplaySettings(null, User_32.ENUM_CURRENT_SETTINGS, ref dm)) { dm.dmDisplayOrientation = DMDO_90 dm.dmPelsWidth = 1200; dm.dmPelsHeight = 1920; int iRet = User_32.ChangeDisplaySettings(ref dm, User_32.CDS_TEST); if (iRet == User_32.DISP_CHANGE_FAILED) { return "Unable To Process Your Request. Sorry For This Inconvenience."; } else { iRet = User_32.ChangeDisplaySettings(ref dm, User_32.CDS_UPDATEREGISTRY); switch (iRet) { case User_32.DISP_CHANGE_SUCCESSFUL: { return "Success"; } case User_32.DISP_CHANGE_RESTART: { return "You Need To Reboot For The Change To Happen.\n If You Feel Any Problem After Rebooting Your Machine\nThen Try To Change Resolution In Safe Mode."; } default: { return "Failed"; } } } } else { return "Failed To Change."; } } private static DEVMODE1 GetDevMode1() { DEVMODE1 dm = new DEVMODE1(); dm.dmDeviceName = new String(new char[32]); dm.dmFormName = new String(new char[32]); dm.dmSize = (short)Marshal.SizeOf(dm); return dm; } } } "@ Add-Type $pinvokeCode -ErrorAction SilentlyContinue [Resolution.PrmaryScreenOrientation]::ChangeOrientation() }

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  • How to use Bonjour?

    - by Roman
    First, what exactly Bonjour does (pleas read my guesses written bellow)? Here I found out that Bonjour enables automatic discovery of computers, devices, and services on IP networks. But I thought that it not only "discovers devices on IP network" it also creates an IP network by assigning IP addresses to devices where Bonjour is running. Am I right? And I still miss the essence. Does it work in the following way? First I connect devices (for example laptops) physically so that they potentially can communicate with each other. Then, let say, on some laptops I have Bonjour running and then, as a consequence, these laptops assign IP addresses to them self in automatic way. So, laptops (where Bonjour is running) build an IP network. Does it work in this way? Or may be a computer running Bonjour is not considered as a service and it does not broadcast itself just because Bonjour is running on this computer. I mean that the applications running on the computers need to use Bonjour to broadcast themself. So, it is applications that broadcast themself (not computers) and it is not done automatically (application needs to broadcast themself explicitly). Is it right? How exactly my application can broadcast itself? Can I use command line to register an service (so that all applications using Bonjour knows that a new service appeared)? Further, I would like to have an application which use the IP network created by Bonjour. For that my application needs to know which devices/services are present in the network. In more details, my application needs to have a list of services. Each service in the list should have a name, the IP address where it is running and the port which is used by the application. Can Bonjour provide this information in some way? If it is the case, how exactly it works. How my program can get this information from Bonjour? Can my program read some file created by Bonjour and containing the above mentioned information? Can I use some commands in command line to retrieve this information? I have a special interest in accessing the information about services from files, environment variables or commands in command line. These options seems to me to be the simplest! Since in these case I do not need to use any additional libraries to communicate with Bonjour from a particular programming language. P.S. Pleas ask questions if something is not clear in my question. I will try to formulate my question in a more clear way. P.P.S. I use Windows 7. ADDED: I plan to write my applications in PHP. Every computer should run a Apache web server. And I want to use Bonjour to help computer discover each other (computers are working in a local network).

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  • How can I automatically update Flash Player whenever a new version is released? [closed]

    - by user219950
    Summary: Flash Player Update Service doesn't run on a reliable schedule, and doesn't automatically download and apply updates when it does run. Given the importance of having an up-to-date version of Flash Player installed (for those of us who don't use Chrome with its built-in player), I would like to find a way to ensure that new updates are promptly detected and installed. What follows are the details of my efforts to solve this problem on my own... Appendix A: Flash Player Update Service OK, way back in Flash Player 11.2 (or so?) Adobe added the Flash Player Update Service (FlashPlayerUpdateService.exe), it was supposed to keep the Flash Player updated... Upon installation, FPUS is configured to run as a Windows Service, with Start Type set to Manual. A Scheduled Task (Adobe Flash Player Updater.job) is added to start this service every hour. So far, so good - this set-up avoids having a constantly-running service, but makes sure that the checks are run often enough to catch any updates quickly. Google's software updater is configured in a similar fashion, and that works just fine... ...And yet, when I checked the version of my installed Flash Player, I found it was 11.6.602.180, which, based on looking at the timestamps of the files in C:\Windows\System32\Macromed\Flash was last updated (or installed) on Tue, Mar 12, 2013 --- 3/12/13, 5:00:08pm. I made this observation on Thu, Apr 25, 2013 --- 4/25/13, 7:00:00pm, and upon checking Adobe's website found that the current version of Flash Player was 11.7.700.169. That's over a month since the last update, with a new one clearly available on the website but with no indication that the hourly check running on my machine has noticed it or has any intention of downloading it. Appendix B: running the Flash Player updater manually Once upon a time, running FlashUtil32_<version_Plugin.exe -update plugin would give you a window with an Install button; pressing it would download the installer for the current version (automatically, without opening a browser) and run it, then you'd click thru that installer & be done. It was manual, but it worked! Finding my current installation out of date (see Appendix A), I first tried this manual update process. However... Running FlashUtil32_<version_ActiveX.exe -update activex (in my case, that's FlashUtil32_11_6_602_180_ActiveX.exe -update activex) ...only presents a window with a Download button, clicking that Download button opens my browser to the URL https://get3.adobe.com/flashplayer/update/activex. Running FlashUtil32_<version_Plugin.exe -update plugin (in my case, that's FlashUtil32_11_6_602_180_Plugin.exe -update plugin) ...only presents a window with a Download button, clicking that Download button opens my browser to the URL https://get3.adobe.com/flashplayer/update/plugin. I could continue with the Download page it sent me to, uncheck the foistware box ("Free! McAfee Security Scan Plus"), download that installer (ActiveX, no foistware: install_flashplayer11x32axau_mssd_aih.exe, Plugin, no foistware: install_flashplayer11x32au_mssd_aih.exe) & probably have an updated Flash...but then, what is the point of the Flash Player Update Service if I have to manually download & run another exe? Epilogue I've since come to suspect that the update service is intentionally hobbled to drive early adopters to the manual download page. If this is true, there's probably no solution to this short of writing my own updater; hopefully I am wrong.

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  • nginx: How can I set proxy_* directives only for matching URIs?

    - by Artem Russakovskii
    I've been at this for hours and I can't figure out a clean solution. Basically, I have an nginx proxy setup, which works really well, but I'd like to handle a few urls more manually. Specifically, there are 2-3 locations for which I'd like to set proxy_ignore_headers to Set-Cookie to force nginx to cache them (nginx doesn't cache responses with Set-Cookie as per http://wiki.nginx.org/HttpProxyModule#proxy_ignore_headers). So for these locations, all I'd like to do is set proxy_ignore_headers Set-Cookie; I've tried everything I could think of outside of setting up and duplicating every config value, but nothing works. I tried: Nesting location directives, hoping the inner location which matches on my files would just set this value and inherit the rest, but that wasn't the case - it seemed to ignore anything set in the outer location, most notably proxy_pass and I end up with a 404). Specifying the proxy_cache_valid directive in an if block that matches on $request_uri, but nginx complains that it's not allowed ("proxy_cache_valid" directive is not allowed here). Specifying a variable equal to "Set-Cookie" in an if block, and then trying to set proxy_cache_valid to that variable later, but nginx isn't allowing variables for this case and throws up. It should be so simple - modifying/appending a single directive for some requests, and yet I haven't been able to make nginx do that. What am I missing here? Is there at least a way to wrap common directives in a reusable block and have multiple location blocks refer to it, after adding their own unique bits? Thank you. Just for reference, the main location / block is included below, together with my failed proxy_ignore_headers directive for a specific URI. location / { # Setup var defaults set $no_cache ""; # If non GET/HEAD, don't cache & mark user as uncacheable for 1 second via cookie if ($request_method !~ ^(GET|HEAD)$) { set $no_cache "1"; } if ($http_user_agent ~* '(iphone|ipod|ipad|aspen|incognito|webmate|android|dream|cupcake|froyo|blackberry|webos|s8000|bada)') { set $mobile_request '1'; set $no_cache "1"; } # feed crawlers, don't want these to get stuck with a cached version, especially if it caches a 302 back to themselves (infinite loop) if ($http_user_agent ~* '(FeedBurner|FeedValidator|MediafedMetrics)') { set $no_cache "1"; } # Drop no cache cookie if need be # (for some reason, add_header fails if included in prior if-block) if ($no_cache = "1") { add_header Set-Cookie "_mcnc=1; Max-Age=17; Path=/"; add_header X-Microcachable "0"; } # Bypass cache if no-cache cookie is set, these are absolutely critical for Wordpress installations that don't use JS comments if ($http_cookie ~* "(_mcnc|comment_author_|wordpress_(?!test_cookie)|wp-postpass_)") { set $no_cache "1"; } if ($request_uri ~* wpsf-(img|js)\.php) { proxy_ignore_headers Set-Cookie; } # Bypass cache if flag is set proxy_no_cache $no_cache; proxy_cache_bypass $no_cache; # under no circumstances should there ever be a retry of a POST request, or any other request for that matter proxy_next_upstream off; proxy_read_timeout 86400s; # Point nginx to the real app/web server proxy_pass http://localhost; # Set cache zone proxy_cache microcache; # Set cache key to include identifying components proxy_cache_key $scheme$host$request_method$request_uri$mobile_request; # Only cache valid HTTP 200 responses for this long proxy_cache_valid 200 15s; #proxy_cache_min_uses 3; # Serve from cache if currently refreshing proxy_cache_use_stale updating timeout; # Send appropriate headers through proxy_set_header Host $host; # no need for this proxy_set_header X-Real-IP $remote_addr; # no need for this proxy_set_header X-Forwarded-For $proxy_add_x_forwarded_for; # Set files larger than 1M to stream rather than cache proxy_max_temp_file_size 1M; access_log /var/log/nginx/androidpolice-microcache.log custom; }

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  • Windows: what is the difference between DEP always on and DEP opt-out with no exceptions?

    - by Peter Mortensen
    What is the difference between DEP always on ("/NoExecute=AlwaysOn" in boot.ini) and DEP opt-out ( "/NoExecute=OptOut" in boot.ini) with no exceptions? "no exceptions" = empty list of programs for which DEP does not apply. DEP = Data Execution Prevention (hardware). One would expect it to work the same way, but it makes a difference for some applications. E.g. for all versions of UltraEdit 14 (14.2). It crashes at startup for DEP always on, at least on Microsoft Windows XP Professional Edition x64 edition. (2010-03-11: this problem has been fixed with UltraEdit 15.2 and later.) Update 1: I think this difference is caused by the backdoors that Microsoft has put into hardware DEP for OptOut, according to Fabrice Roux (see below). In the case of IrfanView, for which Steve Gibson observed the same difference as I did for UltraEdit (see below), the difference is caused by a non-DEP aware EXE packer (ASPack) that Microsoft coded a backdoor for. Is there a difference between Windows XP, Windows Vista and Windows 7 ? Is there a difference between 32 bit and 64 bit versions of Windows ? Sources: From [http://blog.fabriceroux.com/index.php/2007/02/26/hardware_dep_has_a_backdoor?blog=1], "Hardware DEP has a backdoor" by Fabrice Roux. 2007-02-26. "IrfanView was not using any trick to evade DEP ... Microsoft just coded a backdoor used only in OPTOUT. Bascially Microsoft checks the executable header for a section matching one of the 3 strings. If one these strings is found, DEP will be turned OFF for this application by windows. ... 'aspack', 'pcle', 'sforce'" From [http://www.grc.com/sn/sn-078.htm], by Steve Gibson. "I can’t find any documentation on Microsoft’s site anywhere, because we’re seeing a difference between always-on and opt-out. That is, you would imagine that always-on mode would be the same as opting out if you weren’t having any opt-out programs. It turns out it’s not the case. For example ... the IrfanView file viewer ... runs fine in opt-out mode, even if it has not been opted out. But it won’t launch, Windows blocks it from launching ... in always-on mode." From [http://www.grc.com/sn/sn-083.htm], by Steve Gibson. "... IrfanView ... won’t run with DEP turned on. It’s because it uses an EXE packer, an executable compression program called ASPack. And it makes sense that it wouldn’t because naturally an executable compressor has got to decompress the executable, so it allocates a bunch of data memory into which it decompresses the compressed executable, and then it runs it. Well, it’s running a data allocation, which is exactly what DEP is designed to stop. On the other hand, UPX, which is actually the leading and most popular EXE compressor, it’s DEP- compatible because those guys realized, hey, when we allocate this memory, we should mark the pages as executable."

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  • DHCP and DNS services configuration for VOIP system, windows domain, etc

    - by Stemen
    My company has numerous physical offices (for purposes of this discussion, 15 buildings). Some of them are well-connected to our primary data center via fiber. Others will be connected to the data center by P2P T1. We are in the beginning stages of implementing an Avaya VOIP telephone system, and we will be replacing a significant portion of our network infrastructure in the process. In tandem with the phone system implementation, we are going to be re-addressing some of our networks, and consolidating most of our Windows domains into one (not all domains, just most). We currently have quite a few Windows domains, and they of course each have their own DNS zones. A few of those networks currently use DHCP, but the majority use static IP assignments for every device. I'm tired of managing static assignments -- I want to use DHCP configuration on everything except servers. Printers and etc will have DHCP reservations. The new IP phones will need to get IP addresses from DHCP, though they need to be in a separate VLAN from the computers/printers/etc. The computers and printers need to be registered in DNS. That's currently handled by the Windows DHCP servers on each of the respective domains. We need to place a priority on DHCP and DNS being available on a per-site basis (in case something were to interrupt the WAN connection) for computers and (primarily) phones. Smaller locations (which will have IP phones but not be a member of any Windows domain) will not have any Windows DNS/DHCP server(s) available. We also are looking for the easiest way to replace a part if it were to fail. That is to say, if a server/appliance/router hosting DHCP were to crash hard, and we couldn't extremely quickly recover the DHCP reservations and leases (and subsequently restore them onto a cold spare), we anticipate that bad things could happen. What is the best idea for how to re-implement DNS and DHCP keeping all of the above in mind? Some thoughts that have been raised (by myself or my coworkers): Use Windows DNS and DHCP servers, where they exist, and use IP helpers to route DHCP requests to some other Windows server if necessary. May not be acceptable if the WAN goes down and clients don't get a DHCP response. Use Windows DNS (everywhere, over WAN in some cases) and a mix of Windows DHCP and DHCP provided by Cisco routers. Every site would be covered for DHCP, but from what I've read, Cisco routers can't handle dynamic registration of DHCP clients to Windows DNS servers, which might create a problem where Cisco routers are used for DHCP. Use Windows DNS (everywhere, over WAN in some cases) and a mix of Windows DHCP and DHCP provided by some service running on an extremely low-price linux server. Is there any such software that would allow DHCP leases granted by these linux boxes to be dynamically registered on the Windows DNS servers? Come up with a Linux solution for both DNS and DHCP, and deploy low-price linux servers to every site. Requirements would be that the DNS zone be multi-master (like Windows DNS integrated with Active Directory), that DHCP be able to make dynamic DNS registrations in that zone, for every lease (where a hostname is provided and is thus possible), and that multiple servers be either authoritative for the same DHCP scope or at least receiving a real-time copy / replication / sync of the leases table so that if one server dies, we still know which MAC has what address. Purchase dedicated DNS/DHCP appliances, deploying to all sites. From what I read/see, this solves all of our technical problems. Then come the financial problems... I don't have a ton of money to spend on this. Or, some other solution that we've thus far overlooked and will consider upon recommendation. Can Cisco routers or Windows servers sync DHCP lease tables so that multiple servers can be authoritative (or active/passive for all I care) for the same scope, in case one of the partners were to fail? I've read online (repeatedly) that ISC's DHCP is able to maintain the same lease table across multiple servers, in order to solve this problem. Does anyone have any experience or advice to regarding that?

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  • Windows 7 .NET 3.5.1 - 2.0 Slightly Corrupted, How to Repair?

    - by Quinxy von Besiex
    My Windows 7 included .NET installation (3.5 to 2.0) appears very slightly and particularly corrupted and I am trying to fix it without reinstalling Windows or trying to revert to backups. Everything was working and then my hard drive started corrupting a few files and checkdisk found bad clusters so I imaged the drive to a new one. As soon as I booted on the new drive everything worked except programs which call the System.Net.NetworkInformation methods within .NET 3.5 to 2.0 (like Ping() and IsNetworkAvailable()), which immediately crash the app in which the calls are (those calls in .NET 4.0 works fine). Those methods are found inside System.dll, and I assume call native methods which I believe are inside winnsi.dll or iphlpapi.dll or something else (I've not found this yet); I assume it calls native methods because the exception which causes the crash is Fatal Execution Engine Error which people mention is usually related to calling native methods and marshaling data between them. A huge clue about the culprit is likely found in the fact that when I launch the exact same crashing application through a code profiler (which executes the exe and captures stats on which methods took the longest) the app works fine, no crash at all! How could running it within the profiler work and running it outside not work? That seems the key to the mystery. I've used procmon to catch all the registry, filesystem, and network events from the crashing execution and the profiler-run successful execution and compared the two outputs but didn't learn much (I see the moment at which the non-profiled app crashes, but up until then they behave the same, loaded the same modules, ). The only big difference seems to be that at the moment before the app crash the profiler-executed code creates 4-6 new threads and the directly executed code only creates 1-2. I have diffed the files/directories which seemed most relevant (the .NET stuff under Windows and Program Files) pre- and post- disk trouble and seen no changes where I didn't expect any (no obvious file corruption). I have diffed the software and system registry hives pre- and post- disk trouble and seen no changes which seemed relevant. I have created a new user account and cleaned up any environment variables in case environment was related. No change. I did "sfc /scannow" and it found no integrity problems. I tried "ngen update" to regenerate pre-compiled code in case I missed something that might be damaged and nothing changed. I assume I need to repair my .NET installation but because Windows 7 included .NET 3.5 - 2.0 you can't just re-run a .NET installer to redo it. I do not have access to the Windows disks to try to re-install Windows over itself (the computer has a recovery partition but it is unusable); also the drive uses a whole-disk encryption solution and re-installing would be difficult. I absolutely do not want to start from scratch here and install a fresh Windows, reinstall dozens of software packages, try and remember dozens of development-related customizations/etc. Given all that... does anyone have any helpful advice? I need .NET 3.5 - 2.0 working as I am a developer and need to build and test against it. Thanks! Quinxy

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  • Why can I resolve this hostname but not a cname to this hostname?

    - by Joe Hopfgartner
    if a run dig against a hostname, i get the according cname, however i get an NXDOMAIN error (non existent domain). if i run dig against the cname i got, I can resolve it to an IP address successfully. It is reproduceable. On the system I am currently on it is always the case, on other systems it sometimes works and sometimes not, and on other systems it seems to work all the time. If i run using a nameserver i specify (for example googles public nameserver) i can successfully resolve the hostname. i would just blame the local system, but it seems i am not having the only one problems the 2nd domain (unrestrict.me) is hosted on amazon route 53 nameservers. the 1st one on another dns server which has proofen to be fully functional and reliable over the years. i once switchted with the other domain to amazon dns as well, everything seemed to work, also various dns health check tests reported fine, however i recieved a lot of support tickets that dns resolution would not work. is amazon just "bad" or am i doing something wrong? i did not tamper with the domain in any way on the local system (in case of caching or making a custom dns view or whatever...) joe@joe:~$ dig scorpion.premiumize.me ; <<>> DiG 9.8.1-P1 <<>> scorpion.premiumize.me ;; global options: +cmd ;; Got answer: ;; ->>HEADER<<- opcode: QUERY, status: NXDOMAIN, id: 10222 ;; flags: qr rd ra; QUERY: 1, ANSWER: 1, AUTHORITY: 0, ADDITIONAL: 0 ;; QUESTION SECTION: ;scorpion.premiumize.me. IN A ;; ANSWER SECTION: scorpion.premiumize.me. 180 IN CNAME alpha.nue.scorpion.unrestrict.me. ;; Query time: 28 msec ;; SERVER: 127.0.0.1#53(127.0.0.1) ;; WHEN: Mon Jun 18 10:28:39 2012 ;; MSG SIZE rcvd: 84 joe@joe:~$ dig alpha.nue.scorpion.unrestrict.me ; <<>> DiG 9.8.1-P1 <<>> alpha.nue.scorpion.unrestrict.me ;; global options: +cmd ;; Got answer: ;; ->>HEADER<<- opcode: QUERY, status: NOERROR, id: 25381 ;; flags: qr rd ra; QUERY: 1, ANSWER: 1, AUTHORITY: 0, ADDITIONAL: 0 ;; QUESTION SECTION: ;alpha.nue.scorpion.unrestrict.me. IN A ;; ANSWER SECTION: alpha.nue.scorpion.unrestrict.me. 300 IN A 78.46.25.130 ;; Query time: 48 msec ;; SERVER: 127.0.0.1#53(127.0.0.1) ;; WHEN: Mon Jun 18 10:28:47 2012 ;; MSG SIZE rcvd: 66 joe@joe:~$

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  • Excel: conditionally format a cell using the format of another, content-matching cell

    - by Eric A. Meyer
    I have an Excel spreadsheet where I’d like to be able to create a “key” of formatted cells with unique values, and then in another sheet format cells using the key formatting. So for example, my key is as follows, with one value per cell and the visual formatting indicated in parentheses: A (red background) B (green background) C (blue background) So that’s on one sheet (or in a remote corner of the current sheet—whichever is better). Then, in an area that I mark for conditional formatting, I can type one of those three letters and have the cell where I typed it visually formatted according to the key. So if I type a “B” into one of the conditionally formatted cells, it gets a green background. (Note that I’m using backgrounds here solely for ease of explanation: ideally I want to have all visual formatting copied over, whether it’s foreground color, background color, font weight, borders, or whatever. But I’ll take what I can get, obviously.) And—just to make it extra-tricky—if I change the formatting in the key, that change should be reflected in cells that reference the key. Thus, if I change the “B” formatting in the key from a green background to a purple background, any “B” in the main sheet should switch to the new color. Similarly, it should be possible to add or remove values from the key and have those changes applied to the main data set. I’m okay with the formatting-update-on-key-change being triggered by clicking a button or something. I suspect that if any of this is possible it will require VBA, but I’ve never used it so I’ve no idea where to start if that’s the case. I’m hoping it’s possible without VBA. I know it’s possible to just use multiple conditional formats, but my use case here is that I’m trying to create the above-described capability for someone who isn’t conversant with conditional formatting. I’d like to let them be able to define a key, update it if necessary, and keep on truckin’ without me having to rewrite the spreadsheet’s formatting rules for them. --- UPDATE --- So I think I was a bit unclear about my original request. Let me try again with an image. The image shows the “key” on the left, where values and styles are defined using keyboard and mouse input. On the right, you see the data that should be formatted to match the key. Thus if I type a “C” into a cell in the Data area, it should be blue-backed. Furthermore, if I change the formatting of “C” in the Key to have a purple background, all the “C” cells should switch from blue to purple. For further craziness, if I add more to the Key (say, “D” with a yellow background) then any “D” cells will be styled to match; if I remove a Key entry, then matching values in the Data area should revert to default styling. So. Is that more clear? Is it possible, in whole or in part? I don’t have to use conditional formatting for this; in fact, at this point I suspect I probably shouldn’t. But I’m open to any approach!

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  • Why the server is not responding?

    - by par
    Hello! Our server occasionally refuses to serve a simple HTML page. This is happening during a relatively high number of requests. However, the processor is not heavy loaded and there are a lot of free memory. The error seems to occure 1 out of 50 requests in average, depending on the server load. I need to find the source of the problem and take the appropriate actions to eliminate it. I have a suspicion that the problem source is a huge number of incoming network packets. There are 5000 packets per second on average. Traffic - 2 MBits/sec Can this be the cause of the error? There is an interesting thing, in case the server fails to respond, the request string is not logged to access.log by Apache. The error is repeatable from several client computers. DNS is not involved, since I have accessed the server by the IP. I have profiled the problem case with tcpdump utility. These are the good and bad sessions traced by tcpdump. The request is the same in both experiments. Good - server returns response. Bad - no response, time-out error. ---- Bad ---- 12:23:36.366292 IP 123.45.67.890.61749 > myserver.superbservers.com.www: S 2125316338:2125316338(0) win 8192 <mss 1460,nop,wscale 2,nop,nop,sackOK> 12:23:39.362394 IP 123.45.67.890.61749 > myserver.superbservers.com.www: S 2125316338:2125316338(0) win 8192 <mss 1460,nop,wscale 2,nop,nop,sackOK> 12:23:45.365567 IP 123.45.67.890.61749 > myserver.superbservers.com.www: S 2125316338:2125316338(0) win 8192 <mss 1460,nop,nop,sackOK> -------- ---- Good ---- 12:27:07.632229 IP 123.45.67.890.63914 > myserver.superbservers.com.www: S 3581365570:3581365570(0) win 8192 <mss 1460,nop,wscale 2,nop,nop,sackOK> 12:27:10.620946 IP 123.45.67.890.63914 > myserver.superbservers.com.www: S 3581365570:3581365570(0) win 8192 <mss 1460,nop,wscale 2,nop,nop,sackOK> 12:27:10.620969 IP myserver.superbservers.com.www > 123.45.67.890.63914: S 2654770980:2654770980(0) ack 3581365571 win 5840 <mss 1460,nop,nop,sackOK,nop,wscale 6> 12:27:10.838747 IP 123.45.67.890.63914 > myserver.superbservers.com.www: . ack 1 win 4380 12:27:10.957143 IP 123.45.67.890.63914 > myserver.superbservers.com.www: P 1:213(212) ack 1 win 4380 12:27:10.957152 IP myserver.superbservers.com.www > 123.45.67.890.63914: . ack 213 win 108 12:27:10.965543 IP myserver.superbservers.com.www > 123.45.67.890.63914: P 1:630(629) ack 213 win 108 12:27:10.965621 IP myserver.superbservers.com.www > 123.45.67.890.63914: F 630:630(0) ack 213 win 108 12:27:11.183540 IP 123.45.67.890.63914 > myserver.superbservers.com.www: . ack 631 win 4222 12:27:11.185657 IP 123.45.67.890.63914 > myserver.superbservers.com.www: F 213:213(0) ack 631 win 4222 12:27:11.185663 IP myserver.superbservers.com.www > 123.45.67.890.63914: . ack 214 win 108 -------- Hoster: SuperbHosting OS: Ubuntu Server parameters: E6300 CONROE 1.86GHZ 2 X 1MB CACHE 1066 1GB DDR2 667MHZ This is a link to apache configuration file we use http://repkin5.snow.prohosting.com/apache.txt This is server-status report taken right after time-out error. http://repkin5.snow.prohosting.com/server-status.htm There are only 10 Child Servers running out of 120, so enough space for new requests. VMSTAT procs -----------memory---------- ---swap-- -----io---- -system-- ----cpu---- r b swpd free buff cache si so bi bo in cs us sy id wa 0 0 8900 725900 8468 65684 0 0 5 18 11 33 4 3 92 1

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  • Multiple Homed Windows 2008 Server / Windows 7 Client

    - by Daniel Scott
    I have a small Windows 2008 network, with some Windows 7 clients. The clients are both laptops with docking stations and I would like them to communicate with the Windows 2008 server (for filesharing) through the wired network whilst they're docked. Internet connectivity for all machines (clients and server) is via a Wireless LAN, so the wireless adapter in the Windows 7 clients stays active while they're docked. When the laptops are un-docked, it would be nice to still be able to contact the windows 2008 server for print sharing (and slower file sharing) - hence the server also being on the wireless LAN. The windows 2008 server is running Active Directory, DHCP and DNS. It controls DHCP leases on the wired network and holds the DNS records for "myserver.mycompany.local", which is what the filesharing clients connect to. Ideally I'd like the DNS records to return the wired IP first so that this is the address that the laptops will attempt initially - but there doesn't seem to be a way to do that? At present the server's IP on the wireless LAN comes out of an nslookup above the wired Lan IP. The multi-homing works perfectly - but in the wrong order! Switch on the wireless lan and ping myserver and it goes to the wireless IP. Disable the wireless on the client and do the same ping again and after a couple of seconds it starts pinging the wired address. Does anyone have any suggestions on how to make this work in a predictable order? - or even if it can work. Alternative 1? If it can't work, then would this work: Remove the wireless adapter from the server, put a wireless router/bridge on the wired network (set up to route to/from the wireless LAN's subnet), then configure the clients with two routes to the (now) single IP of the server with metrics favouring direct communication over the wired LAN first? Alternative 2? Should I instead single-home the laptops so all of their connectivity is via the wired-LAN while they're docked? (and route via the windows 2008 server - or a dedicated wireless bridge/router)? My concern here is that I'd like undocking to be seamless - and if the clients are in the middle of downloading something from the internet I wouldn't want whatever they're doing interupted as they switch IP addresses onto the Wireless network. Perhaps this isn't the case and I'm concerned over nothing? Any thoughts? :) UPDATE I seem to have cracked it (at least DNS entries come out in the order I hope for - and pinging the server with various combinations of wired, wireless and both interfaces enabled uses the IP I want) ... I set the binding order of the NICs on the Server (which is acting as Domain Controller, DHCP and DNS server) so that the Wired NIC is before the Wireless adapter. (Start -- type "Network Interfaces" -- Select "View Network Connections" -- Press Alt to show classic dropdown menus -- Advanced -- Advanced Settings) Now, an nslookup (from the client) of the server's hostname returns the Wired IP first, followed by the Wireless IP. The wired IP now seems to be used whenever it's contactable. Incidentally, the metrics on the wired and wireless routes (on the client) also favour the wired LAN (based on Windows' automatically assigned metrics) - but this was always the case, even when I was having trouble getting the wired IP to be "favoured". I'm not entirely sure if this is coincidence - or if a DNS server running on Windows, handing back IP addresses for itself does actually take the binding order of it's own network interfaces into account? It would be interesting to hear from someone who can confirm or deny that (or confirm that the binding order on the server plays a role for some other reason?)

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  • Magento hosting on a budget

    - by spa
    I have to do a setup for Magento. My constraint is primarily ease of setup and fault tolerance/fail over. Furthermore costs are an issue. I have three identical physical servers to get the job done. Each server node has an i7 quad core, 16GB RAM, and 2x3TB HD in a software RAID 1 configuration. Each node runs Ubuntu 12.04. right now. I have an additional IP address which can be routed to any of these nodes. The Magento shop has max. 1000 products, 50% of it are bundle products. I would estimate that max. 100 users are active at once. This leads me to the conclusion, that performance is not top priority here. My first setup idea One node (lb) runs nginx as a load balancer. The additional IP is used with domain name and routed to this node by default. Nginx distributes the load equally to the other two nodes (shop1, shop2). Shop1 and shop2 are configured equally: each server runs Apache2 and MySQL. The Mysqls are configured with master/slave replication. My failover strategy: Lb fails = Route IP to shop1 (MySQL master), continue. Shop1 fails = Lb will handle that automatically, promote MySQL slave on shop2 to master, reconfigure Magento to use shop2 for writes, continue. Shop2 fails = Lb will handle that automatically, continue. Is this a sane strategy? Has anyone done a similar setup with Magento? My second setup idea Another way to do it would be to use drbd for storing the MySQL data files on shop1 and shop2. I understand that in this scenario only one node/MySQL instance can be active and the other is used as hot standby. So in case shop1 fails, I would start up MySQL on shop2, route the IP to shop2, and continue. I like that as the MySQL setup is easier and the nodes can be configured 99% identical. So in this case the load balancer becomes useless and I have a spare server. My third setup idea The third way might be master-master replication of MySQL databases. However, in my optinion this might be tricky, as Magento isn't build for this scenario (e.g. conflicting ids for new rows). I would not do that until I have heard of a working example. Could you give me an advice which route to follow? There seems not one "good" way to do it. E.g. I read blog posts which describe a MySQL master/slave setup for Magento, but elsewhere I read, that data might get duplicated when the slave lags behind the master (e.g. when an order is placed, a customer might get created twice). I'm kind of lost here.

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  • Duplicate DNS Zones (Error 4515 in Event Log )

    - by Campo
    I am getting these two error in the DNS Event log (errors at end of question). I have confirmed I do have duplicate zones. I am wondering which ones to delete. The DomainDNSZone contains all of our DNS records but it does not have the _msdcs zone.... that is in the ForestDNSZone with the duplicates that are not in use. here is a picture of that 3 Questions. I understand the advantages of having DNS in the ForestDNSZone. so... Why is DNS using the DomainDNSZone and is that acceptable considering _msdcs... is in the ForestDNSZone? If so, should I just delete the DC=1.168.192.in-addr.arpa and DC=supernova.local from the ForestDNSZone? Or should I try to get those to be the ones in use? What are those steps? I understand how to delete. That is simple but if i must move zones some info would be appreaciated there. Just to confirm. from my understanding. I can delete the two duplicates in the ForestDNSZone and leave the _msdcs.supernova.local as thats required there. This will resolve the erros I see. Just fyi when I look in those folders from the ForestDNSZone they have just 2 and 1 entries respectively. So obviously not in use compared to the others. I am pretty sure I understand the steps to complete this. But if you would like to provide that info, bonus points! Event Type: Warning Event Source: DNS Event Category: None Event ID: 4515 Date: 1/4/2011 Time: 2:14:18 PM User: N/A Computer: STANLEY Description: The zone 1.168.192.in-addr.arpa was previously loaded from the directory partition DomainDnsZones.supernova.local but another copy of the zone has been found in directory partition ForestDnsZones.supernova.local. The DNS Server will ignore this new copy of the zone. Please resolve this conflict as soon as possible. If an administrator has moved this zone from one directory partition to another this may be a harmless transient condition. In this case, no action is necessary. The deletion of the original copy of the zone should soon replicate to this server. If there are two copies of this zone in two different directory partitions but this is not a transient caused by a zone move operation then one of these copies should be deleted as soon as possible to resolve this conflict. To change the replication scope of an application directory partition containing DNS zones and for more details on storing DNS zones in the application directory partitions, please see Help and Support. For more information, see Help and Support Center at http://go.microsoft.com/fwlink/events.asp. Data: 0000: 89 25 00 00 %.. AND Event Type: Warning Event Source: DNS Event Category: None Event ID: 4515 Date: 1/4/2011 Time: 2:14:18 PM User: N/A Computer: STANLEY Description: The zone supernova.local was previously loaded from the directory partition DomainDnsZones.supernova.local but another copy of the zone has been found in directory partition ForestDnsZones.supernova.local. The DNS Server will ignore this new copy of the zone. Please resolve this conflict as soon as possible. If an administrator has moved this zone from one directory partition to another this may be a harmless transient condition. In this case, no action is necessary. The deletion of the original copy of the zone should soon replicate to this server. If there are two copies of this zone in two different directory partitions but this is not a transient caused by a zone move operation then one of these copies should be deleted as soon as possible to resolve this conflict. To change the replication scope of an application directory partition containing DNS zones and for more details on storing DNS zones in the application directory partitions, please see Help and Support. For more information, see Help and Support Center at http://go.microsoft.com/fwlink/events.asp. Data: 0000: 89 25 00 00 %..

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  • Cheapest way to connect 20-24 Sata II HDDs in a budget storage server?

    - by Joe Hopfgartner
    I need to assemble a high density storage server for as cheap as possible. It's been a while for me and the last systems I integrated didn't even have Sata yet... During my Research I of course stumbled about Nexsan SATA Beast, the BackBlaze storage Pods as well as some ridiculously overpriced HP Proliant or Dell storage solutions. Finally I choose Norco cases as the way to go. My eye is set on the RPC-4020, which is a 4U 19" Rackmount case with 20 Hot Swap 3.5" SATA/SAS Hdd trays (Backplanes included) and room for two 2.5" OS drives as well as a Slim Line CD-Rom. The backplanes connect with a single SATA port for each drive, so there are 20 internal SATA ports to to be connected. They also have redundant power ports which I think is quite nice. The cheapest price I have found is 290$ + 40$ shipping. In europe the cheapest unfortunately is 370€ (500$) + 40 € shipping... A nice alternative would be the RPC-4224 which has SFF-8087 Mini SAS connectors that bundle 4 SATA trays each. But it doesn't seem to be available in Europe (where i am) anywhere. So here comes my problem: What Mainboard/Controller to choose to connect them for as cheap as possible while still having nice data rates? I have to say that the server is intended as a Storage server with 1gps connectivity and the data transfer will be distributed very evenly across all drives. I also don't require any raid functionality. This is all done at application level, I just need JBOD. So for example if I go for the RPC 4020 Model I need to connect 20 Storage + 1 OS + 1 CDROM Sata ports. I searched a bit and stumbled across this very low priced controller: http://www.intel.com/products/server/raid-controllers/SASWT4I/SASWT4I-overview.htm They sell it for 115 € here and the specs say it can control up to 122 hard discs and has 4 Mini SAS connectors. So I would use 4 Mini SAS 36pin - 4 SATA 7pin cables to connect 4 SATA drives to each port and choose a Mainboard taht has 6 SATA on board (for example this one) and hurray, I can connect my 22 SATA devices for as low as about ~ 220 EUR (cpu, ram, psu, case not counted) Question: WOULD THAT WORK? And if not, why? 2nd Question: If I go for the 4220 or 4224 Model, I have internal Mini SAS connectors. Am I right in assuming that the backplane than acts as a "SAS Expander"? And can I just plug these SAS connectors into any SAS port I can find on my controller / mainboard or are there certain requirements? I know that SATA port multipliers only work with controllers that are ready for that. But isn't this expansion already implemented in the SAS standard? I am sorry that this is a very broad question, but I really spent the last week reading up and it seems to be not so clear! Especially all the controlling hardware specifications! 3rd Question: A lot of hardware specs feature "internal channels" and "internal connectors". The connecors are the physical numbers of places where I can plug a cable in. I got that. But are the "internal channels" always the maximum numbers of physical drives that can be used in the end? Or can I enhance this further by Expanders/Fanouts? 4th and last question: What do you think about the setup so far? Do you know any good alternatives? Maby I am completely going the wrong way and some DAS would be way better? Are there any comparable chassis available in europe? Please feel free to say whatever you think is relevant to the subject!

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  • Varnish VCL - Regular Expression Evaluation

    - by Hugues ALARY
    I have been struggling for the past few days with this problem: Basically, I want to send to a client browser a cookie of the form foo[sha1oftheurl]=[randomvalue] if and only if the cookie has not already been set. e.g. If a client browser requests "/page.html", the HTTP response will be like: resp.http.Set-Cookie = "foo4c9ae249e9e061dd6e30893e03dc10a58cc40ee6=ABCD;" then, if the same client request "/index.html", the HTTP response will contain a header: resp.http.Set-Cookie = "foo14fe4559026d4c5b5eb530ee70300c52d99e70d7=QWERTY;" In the end, the client browser will have 2 cookies: foo4c9ae249e9e061dd6e30893e03dc10a58cc40ee6=ABCD foo14fe4559026d4c5b5eb530ee70300c52d99e70d7=QWERTY Now, that, is not complicated in itself. The following code does it: import digest; import random; ##This vmod does not exist, it's just for the example. sub vcl_recv() { ## We compute the sha1 of the requested URL and store it in req.http.Url-Sha1 set req.http.Url-Sha1 = digest.hash_sha1(req.url); set req.http.random-value = random.get_rand(); } sub vcl_deliver() { ## We create a cookie on the client browser by creating a "Set-Cookie" header ## In our case the cookie we create is of the form foo[sha1]=[randomvalue] ## e.g for a URL "/page.html" the cookie will be foo4c9ae249e9e061dd6e30893e03dc10a58cc40ee6=[randomvalue] set resp.http.Set-Cookie = {""} + resp.http.Set-Cookie + "foo"+req.http.Url-Sha1+"="+req.http.random-value; } However, this code does not take into account the case where the Cookie already exists. I need to check that the Cookie does not exists before generating a random value. So I thought about this code: import digest; import random; sub vcl_recv() { ## We compute the sha1 of the requested URL and store it in req.http.Url-Sha1 set req.http.Url-Sha1 = digest.hash_sha1(req.url); set req.http.random-value = random.get_rand(); set req.http.regex = "abtest"+req.http.Url-Sha1; if(!req.http.Cookie ~ req.http.regex) { set req.http.random-value = random.get_rand(); } } The problem is that Varnish does not compute Regular expression at run time. Which leads to this error when I try to compile: Message from VCC-compiler: Expected CSTR got 'req.http.regex' (program line 940), at ('input' Line 42 Pos 31) if(req.http.Cookie !~ req.http.regex) { ------------------------------##############--- Running VCC-compiler failed, exit 1 VCL compilation failed One could propose to solve my problem by matching on the "abtest" part of the cookie or even "abtest[a-fA-F0-9]{40}": if(!req.http.Cookie ~ "abtest[a-fA-F0-9]{40}") { set req.http.random-value = random.get_rand(); } But this code matches any cookie starting by 'abtest' and containing an hexadecimal string of 40 characters. Which means that if a client requests "/page.html" first, then "/index.html", the condition will evaluate to true even if the cookie for the "/index.html" has not been set. I found in bug report phk or someone else stating that computing regular expressions was extremely expensive which is why they are evaluated during compilation. Considering this, I believe that there is no way of achieving what I want the way I've been trying to. Is there any way of solving this problem, other than writting a vmod? Thanks for your help! -Hugues

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  • LDAP object class violation: attribute ou not allowed in suffix?

    - by Paramaeleon
    I am about to set up a LDAP directory. It is used as a tool to communicate user permissions from a web application to WebDav file system access, e.g. adding a user to the web platform shall allow login to the file system with the same credentials. There are no other usages intended. Following this German tutorial which encourages the use of the attributes c, o, ou etc. over dc, I configured the following suffix and root: suffix "ou=webtool,o=myOrg,c=de" rootdn "cn=ldapadmin,ou=webtool,o=myOrg,c=de" Server starts and I can connect to it by LDAP Admin, which reports “LDAP error: Object lacks”. Well, there aren’t any objects yet. I now want to create the root and admin elements from shell. I created an init.ldif file: dn: ou=webtool,o=myOrg,c=de objectclass: dcObject objectclass: organization dc: webtool o: webtool dn: cn=ldapadmin,ou=webtool,o=myOrg,c=de objectclass: organizationalRole cn: ldapadmin Trying to load the file runs into an error, telling me that ou is not allowed: server:~ # ldapadd -x -D "cn=ldapadmin,ou=webtool,o=myOrg,c=de" -W -f init.ldif Enter LDAP Password: adding new entry "ou=webtool,o=myOrg,c=de" ldap_add: Object class violation (65) additional info: attribute 'ou' not allowed I am not using ou anywhere except in the suffix, so the question: Isn’t it allowed here? What is allowed here? Here is my answer. I am not allowed to post it as answer for 8 hours, so don’t mind that it is part of the question by now. I will move it outside some day, if I don’t forget to do so. There are numberous dependencies for the creation of elements, and error messages are rather confusing if you don’t know of the concept. The objectclass isn’t necessarily dcObject for the databases’ root node, as it is likely to guess when you read several tutoriales. Instead, it must correspond to the object’s type: Here, for a name starting with ou=, it must be organizationalUnit. I found this piece of information in these tables [Link removed due to restriction: Oops! Your edit couldn't be submitted because: We're sorry, but as a spam prevention mechanism, new users can only post a maximum of two hyperlinks. Earn more than 10 reputation to post more hyperlinks. Link is below]. Further on, the object class dictates which properties must and can be added in the record. Here, organizationalUnit must have an ou: entry and must not have neither dc: nor o: entry. The healthy init.ldif file looks like that: dn: ou=webtool,o=myOrg,c=de objectclass: organizationalUnit ou: LDAP server for my webtool dn: cn=ldapadmin,ou=webtool,o=myOrg,c=de objectclass: organizationalRole cn: ldapadmin Note: The page also states: “While many objectClasses show no MUST attributes you must (ouch) follow any hierarchy […] to determine if this is the really case.” I thought that would mean my root record would have to provide the must fields for c= and o= (c: and o:, respectively) but this isn’t the case. Link in answer is (1): http :// www (dot) zytrax (dot) com/books/ldap/ape/ "Appendix E: LDAP - Object Classes and Attributes"

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  • Hardware recommendations / parts list for a modern, quiet ZFS NAS box - 2011-Feb edition [closed]

    - by dandv
    I want to build some really reliable storage for my data, and it seems that ZFS is the only filesystem at the moment that does live checksumming. That rules out DroboPro, so I'm looking to building a quiet ZFS NAS that would start with 4 2TB or larger hard drives. I'd like this system to be very reliable and relatively future-proof for 2-3 years, so I'm willing to invest some $$$ and buy higher end components. I did see questions here and on other forums about low-cost servers, but I'm not looking for those. I'd be super happy to go for an off-the-shelf solution, but I haven't found one that's quiet. I started doing the research (summarized on my wiki), but I realized that it just gets too complicated for what I know as a software dude, and I'm entering the analysis paralysis area. At this point, I'm basically looking for a parts list for a configuration that will work (and is modern), and I know there are folks around here who are way more competent than me. I've built computers and am comfortable assembling one and messing with *nix; I can follow guides; I just want to end the decision process for the hardware and software configuration. What I've researched so far (not that these are very modern components): Case: I think I've settled on the Antec Twelve Hundred case because it cools well, is quiet, and simply has 12 bays that allow elastic mounting. The SilverStone Raven is its counter-candidate, but I find its construction quite odd. For the PSU, I'm torn between Antec CP-850 and Nexus RX-8500, but I did this research more than a year ago. The Nexus has a very uniform power profile, and I'd rather not have the Antec spin up and down based on load. On the other hand, I'm not sure how often my file server will draw more than 400W under use. For the hard drives, I've read that WD Black drives are actually WD RE3 with a software setting changed. I'd also like to buy different drive types, not just 4 WDs. Recommendations? Right now I have a 2TB Hitachi Deskstar 7K300. For the motherboard, CPU and RAM I have no idea, other than the RAM must be ECC. I already asked a question here about ECC RAM, but I was misguided and was looking for a motherboard that would support USB 3.0 as well. I've learned to go with eSATA, or worry about USB later. Then there's the (liquid) cooling, Wi-Fi card, and FreeBSD vs. OpenSolaris Express. Lastly, I'm wondering if I can make this PC into a media server by adding a Blu-ray drive and a good sound card. But support for Blu-Ray is spotty on Linux, and I don't know if Windows 7 on VirtualBox would get sufficient hardware access to output HDMI or SPDIFF signals. (Running OpenSolaris virtualized is not an option because of the reliability risk.) Then there are HDCP concerns. Suggestions on that would be appreciated as well, but I don't want us to get sidetracked. A specific shopping list on the core components would be great, so I can start ordering, and in the meantime educate myself with regards to the other issues. Finally, I think this could become a great FAQ for those technically inclined to build their own ZFS server, but confused by the dizzying array of options out there, and I promise to compile the results and share my experience building and benchmarking the server.

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  • Why is syslog so much slower than file IO?

    - by ceving
    I wrote a simple test program to measure the performance of the syslog function. This are the results of my test system: (Debian 6.0.2 with Linux 2.6.32-5-amd64) Test Case Calls Payload Duration Thoughput [] [MB] [s] [MB/s] -------------------- ---------- ---------- ---------- ---------- syslog 200000 10.00 7.81 1.28 syslog %s 200000 10.00 9.94 1.01 write /dev/null 200000 10.00 0.03 343.93 printf %s 200000 10.00 0.13 76.29 The test program did 200000 system calls writing 50 Bytes of data during each call. Why is Syslog more than ten times slower than file IO? This is the program I used to perform the test: #include <fcntl.h> #include <stdio.h> #include <string.h> #include <sys/stat.h> #include <sys/time.h> #include <sys/types.h> #include <syslog.h> #include <unistd.h> const int iter = 200000; const char msg[] = "123456789 123456789 123456789 123456789 123456789"; struct timeval t0; struct timeval t1; void start () { gettimeofday (&t0, (void*)0); } void stop () { gettimeofday (&t1, (void*)0); } void report (char *action) { double dt = (double)t1.tv_sec - (double)t0.tv_sec + 1e-6 * ((double)t1.tv_usec - (double)t0.tv_usec); double mb = 1e-6 * sizeof (msg) * iter; if (action == NULL) printf ("Test Case Calls Payload Duration Thoughput \n" " [] [MB] [s] [MB/s] \n" "-------------------- ---------- ---------- ---------- ----------\n"); else { if (strlen (action) > 20) action[20] = 0; printf ("%-20s %-10d %-10.2f %-10.2f %-10.2f\n", action, iter, mb, dt, mb / dt); } } void test_syslog () { int i; openlog ("test_syslog", LOG_PID | LOG_NDELAY, LOG_LOCAL0); start (); for (i = 0; i < iter; i++) syslog (LOG_DEBUG, msg); stop (); closelog (); report ("syslog"); } void test_syslog_format () { int i; openlog ("test_syslog", LOG_PID | LOG_NDELAY, LOG_LOCAL0); start (); for (i = 0; i < iter; i++) syslog (LOG_DEBUG, "%s", msg); stop (); closelog (); report ("syslog %s"); } void test_write_devnull () { int i, fd; fd = open ("/dev/null", O_WRONLY); start (); for (i = 0; i < iter; i++) write (fd, msg, sizeof(msg)); stop (); close (fd); report ("write /dev/null"); } void test_printf () { int i; FILE *fp; fp = fopen ("/tmp/test_printf", "w"); start (); for (i = 0; i < iter; i++) fprintf (fp, "%s", msg); stop (); fclose (fp); report ("printf %s"); } int main (int argc, char **argv) { report (NULL); test_syslog (); test_syslog_format (); test_write_devnull (); test_printf (); }

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  • What NAS setup for two-way syncing over the internet?

    - by Jamse
    I have family living a few hours away and have a lot of files that I would like to share - especially lots of folders of digital photos, but also documents etc. - partially so they can see them, partially so I can have access when I visit them and partially for backup / redundancy purposes. My current hard drives on my main machine are getting pretty full anyway, and I have a MythTV box where my music is currently stored, so I was thinking of getting a NAS anyway. And at the other end my family have a few computers, so they would probably benefit from a NAS too. My general idea (though I'm willing to shift on this if there are any bright ideas about other ways of achieving my objectives) is to get a matching pair of NASs and have them sync over the internet. (To cut down on bandwidth use I would get them in sync locally to start with.) Having read around as best I can it seems that syncing over the internet is generally only a feature on quite high end units. However, I have seen that QNAP seem to feature this on their TS-110 and TS-210 units, which might work (they call it "remote replication"). They seem pretty reasonably priced for what they are, but of course with buying 2 of them and then adding the drives (say 1TB or 2TB each) I'd be looking at about £400 total. So, I'm looking for recommendations really. I don't want to spend more than the QNAPs would cost me, but any other ideas would be most appreciated. I am comfortable with technology and tinkering around, but I don't have as much time for that as I would like, so I guess I would favour solutions that require less tinkering rather than more (even though that's less fun!). Any thoughts would be welcome, as would any comments from people who have used the QNAP boxes for this. Thanks in advance. Some specifications: Two-way syncing. Changes made at either end should be synced to the other. There shouldn't be one unit that is effectively a read-only mirror of the other. Not real time. The syncing doesn't need to be real time - if it updated, say, daily overnight that would be fine. Set and forget. I would prefer minimal user interaction once set up - it would be great if syncs were scheduled and automatic. OS independence. I am running Windows XP plus an Ubuntu-based MythTV box. At the other end there are Windows 7 and Windows XP machines, plus a networked TV set top box which I think can play files off the network. Machine independence. I would favour a system that is self-contained, i.e. not reliant on any particular PC being switched on. If the system had enough else going for it I could perhaps work around it at this end, where I only have one PC that's used as such, but it would be harder at the other where there are at least two PCs that might be accessing the files. Notifications. I guess things like getting an email notification if the syncing fell over for any reason would be useful, though it's not a deal breaker. Update I've been digging some more and it looks like QNAP's Remote Replication function is actually just Rsync, so only really suitable for one-way syncing. I've posted on their forum to double check, but I think that's the case. In which case, I think the focus of my question is now either: do any reasonably-priced NASs support bidirectional syncing over the internet?, or has anyone had any luck installing onto NASs for this purpose? (Also, updated question to clarify that I'm after two-way syncing.)

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  • Determining the required depth and specifications for a server cabinet

    - by Bingu Bingme
    I'm trying to understand the considerations ("why") that go into determining the specifications ("what") for a rackmount server cabinet, in order to determine what sort of rack I should purchase for my home use. Since this is for home use, I won't be following certain best practices (eg. hot/cold aisle, not even air conditioning) and may be willing to sacrifice in various areas in order to reduce cost and footprint - but please advise if there are safety concerns or other considerations to note. The most basic specs for a server cabinet are the dimensions (external width x external depth x usable height). Width: commonly 600mm or 800mm (if the use case requires extra clearance around the sides, such as if there is lots of cabling). In my case and most common cases, I'm going to stick with 600mm. Height: Select a sufficiently tall rack to fit my equipment. But how much may I stuff into it? Eg, if there is a 15U rack, can I really populate it with 15U of servers, or should I leave 1U at top and bottom for air circulation? Depth: Racks commonly have external depth of 600mm (network equipment), 800mm, 1000mm, or even longer. I'm trying to see how to fit into the 800mm depth. With reference to http://www.server-racks.com/rack-mount-depth.html, I'm hoping to have the front and rear posts mounted ~ 28.5" (72cm) apart, which would leave only 8cm for front space and rear space. How much rear space (from rear posts to back of rack) do I really need? I won't use cable management arms, so can I mount a 72cm depth server since the power, KVM, network cables won't take up much depth? My most important equipment are all < 60cm depth (4U chassis) and should comfortably fit within the 800mm cabinet. The rest of the equipment are very old 1U servers that range from 65-72cm depth. I might still want to make further use of them, or I might discard them since they are so old. Even if the 72cm servers cannot be powered on in an 800mm rack, I should be able to use them as 1U shelves. But, what server depth can I expect to be able to operate? Or am I forced to upgrade to 1000mm depth racks in order to use any servers deeper than 60cm? With reference to best practices for HP racks, some other specs and installation considerations: There aren't any minimum recommendations for clearance on the sides of the rack. It is recommended to leave 48" front clearance. The 48" front clearance is based on 32" chassis depth, 13" to extend the rack rails and mate the inner/outer rails, and 3" for movement. If I don't use such rails (eg, use shelves instead), it should be sufficient to leave front clearance of chassis depth + 3". It is recommended to leave 30" rear clearance "to provide space for servicing the rack". I'm planning to back the rack into a corner of the room, and wheel it slightly out when I need to access the rear. If the wheeling plan is ok, I still need to know how much rear clearance is required for air circulation and ventilation purposes. Castor wheels and stabilising feet. Since I'm backing the rack into a corner of the room, I'll only be able to set the stabilising feet on the front corners. Thoughts on safety? The rack that I'm considering has front glass doors with side ventilation slits and fully perforated rear doors. I'm hoping this will be a good balance between temperature and noise (only ventilation slits facing out the front, while the rear is facing the walls). Or is the sound of high-rpm fans going to escape through the front slits anyway and destroy my sanity?

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  • Supermicro IPMI on MBD-X8DAH+-F-O motherboard. Keyboard and mouse do not work after booting Windows Server 2008 R2

    - by LDelgado
    Hell Everyone, I built a server with the mentioned motherboard. I installed Windows Server 2008 R2 Enterprise on this server. IPMI is integrated on the motherboard with its own dedicated NIC. I've got that NIC configured with its own IP address. I can remote into it using IPMI, and I can remotely control the server settings before booting the OS ( BIOS, RAID configuration, etc). When the OS boots, I lose the mouse and keyboard. I cannot use the keyboard or mouse when installing the OS either. So the Keyboard and Mouse only work when no OS is loaded. Once the OS loads I lose it - that is my problem. I've been doing some research and trying a few things, but I have not been successful in fixing this issue. I may be wrong, but based on the things I've found online, it seems that the problem could be caused by the way the OS handles USB. The server is headless. There is no keyboard, mouse, or monitor plugged into it. When I boot up the OS and remote into it, I cannot see a mouse or keyboard listed in the Device Manager. Based on what I've read, it seems that the OS should detect a mouse and a keyboard when connecting remotely via IPMI. The following are the solutions I've tried. Nothing has worked so far: I've updated the firmware of the IPMI component to the latest firmware - 1.33. I made sure that the mouse mode was set to Absolute (Windows OS). I've loaded the factory defaults several times. I've enabled Port64h/60h Emulation under the USB settings in the BIOS. I've disabled USB legacy support in the BIOS. I made sure the firewall wasn't blocking IPMI (disabled the firewall). And that's about it. I've found threads in some forums from people having the same issue as me, but they were not running the same OS. They were either running Linux or FreeBSD. Most of them fixed their problem by selecting the right mouse mode (Linux in their case). There was one other that solved the problem by disabling USB Mass Storage mode. He stated "When I set it to disable USB Mass Storage when no image is loaded, the ukbd came alive, and I'm typing this on the IPMI Console. " source: http://freebsd.1045724.n5.nabble.com/IPMI-Console-No-luck-once-OS-is-booted-td3967868.html I suspect the solution described in the previous paragraph is somehow related to my problem. I've found several threads on the internet with issues describing the same problem, but none of them were with Windows Server 2008 R2. Again, I may be wrong, but it seems like that could be the issue. I just don't know how I go about applying a solution in Windows Server 2008 R2. In any case, I could use your expertise. Maybe I am missing something, or maybe I'm on the right track. Your help is much appreciated. Thank you in advance,

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  • Pushing DNSSEC updates with offline keys

    - by eggyal
    In a non-professional capacity, I look after the DNS of some 18 domains: mostly personal/vanity domains for immediate family. I outsource the whole shebang to an inexpensive managed hosting provider with a web interface through which I manage the zones; since the provider also offers DNSSEC, I have successfully deployed that too. These domains are so unimportant that an attack targetted against them seems much less likely than a general compromise of my provider's systems, at which point the records of all their customers might be changed to misdirect traffic (perhaps with extremely long TTLs). DNSSEC could protect against such an attack, but only if the zone's private keys are not held by the hosting provider. So, I wonder: how can one keep DNSSEC private keys offline yet still transfer signed zones to an outsourced DNS host? The most obvious answer (to me, at least) is to run one's own shadow/hidden master (from which the provider can slave) and then copy offline-signed zonefiles to the master as required. The problem is that the only machine I (want to*) control is my personal laptop, which usually connects from a typical home ADSL (behind NAT over a dynamically-assigned IP address). Having them slave from that (e.g. with a very long Expiry time on the zone for periods when my laptop is offline/unavailable) would not only require a Dynamic DNS record from which they can slave (if indeed they can slave from a named host rather than a static IP address), but would also involve me running a DNS server on my laptop and opening both it and my home network up to the incoming zone transfer requests: not ideal. I would prefer a much more push-oriented design, whereby my laptop initiates transfer of offline-signed zonefiles/updates to the provider's servers. I looked into whether nsupdate could fit the bill: documentation is a little sketchy, but my testing (with BIND 9.7) suggests it can indeed update DNSSEC zones, but only where the server holds the keys to perform the zone signing; I have not found a way to have it take an update including the relevant RRSIG/NSEC/etc. records and have the server accept them. Is this a supported use-case? If not, I suspect the only solutions which could fit the bill will involve non-DNS-based transfer of the zone updates and would welcome recommendations that are supported by (hopefully inexpensive) hosting providers: SFTP/SCP? rsync? RDBMS replication? Proprietary API? Finally, what would be the practical implications of such a setup? Key rotation is jumping out at me as being an obvious difficulty, especially if my laptop is offline for extended periods. But the zones are extremely stable, so perhaps I could get away with long-lived ZSKs**...? * Whilst I could run a shadow/hidden master on e.g. an outsourced VPS, I dislike the overhead of having to secure / manage / monitor / maintain yet another system; not to mention the additional financial costs of so doing. ** Okay, this would enable a concerted attacker to replay outdated records—but the risk and impact of such are both tolerable in the case of these domains.

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