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  • why my home page not showing in google ?

    - by user298788
    hello members i have 1 site : http://www.magentocommerceexperts.com/ it is not showing cache in google toolabr also i am not getiing result properly also other pages has been cached and it showing and the site genrated in wordpress so will u show me whats a problem. all seo stuff done completd and i am facing problem since 2 months

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  • Download SVN for windows

    - by Gandalf StormCrow
    Every now and then I have a problem with SVN inside eclipse folder gets locked, I have to check out projects few times update and stuff like that . is there a SVN that I can use to commit files directly from console or windows folders?

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  • Error handling with try catch (AGAIN)

    - by Eatdoku
    Hi, just a general question, do you ALWAYS have to handle error? i was just having this debate with one of my coworker where in his code I see a lot places where stuff are wrapped around a try statement and in the catch statement there is nothing. I always thought it is a bad practice to not handling error or hide them from the user (except log them in the log file). just want to know what other people thinks thanks.

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  • If you knew then what you know now, what language(s) will you start learning first?

    - by John
    I will probably start with C and then Lisp. Although I started programming in C, I did not program in it as much as I want. It's fun working with the low level stuff. Learning C helped me a lot when I started working. I've only started learning Lisp now and I found it amazing that the concepts it supported from the very start are still very relevant today. Well I think it's better late than never.

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  • Eclipse/Aptana Subprojects

    - by TPorteus
    I've been using a eclipse with the aptana plugin and have all my projects neatly defined. However one project is a main corporate website lets say http://sun.com and it's set up nicely for FTP transfers. However i was wondering if there was a way to define directories of that as subprojects or projects in there own right without messing up the file transfer stuff. Any ideas?

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  • Better way to write this method with this pattern?

    - by Slorthe
    I have written a lot of methods and I want to time how long it takes for them to run. public void myMethod(){ startTiming(); doLotsOfStuff(); stopTiming(); } I am not only timing, I am also doing some other stuff before AND after the doLotsOfStuff() method. I was wondering if there is a better/smarter way to do this in C# to achieve lesser amount of lines/coding needed for this particular pattern.

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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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  • Apple Airport Express, Extreme and Time Capsules, BT Home Hub, Wireless Extenders confusion

    - by Jamie Hartnoll
    I post quite frequently in Stack Overflow, but use Superuser less frequently. Mainly as I don't change hardware often and rarely have software issues! I live in a small stone cottage, and have an office in a separate building across a yard. I have a BT Homehub which is located in the cottage and a series of Ethernet cables running across the yard to the office. This is fine for my wired stuff. My main office computers are PCs running Windows 7 Ultimate, and one on Win7 Home, all working fine. I also have an old laptop on Win XP which works fine wirelessly in the house for those evenings in front of the TV catching up on a bit of work. I also have an iPhone and an iPad. Recently, I have been trying to get WiFi in the office so I can use Adobe Shadow (or whatever it now is!) to improve mobile web development efficiency using my iPhone and iPad, so I bought this: http://www.ebuyer.com/393462-zyxel-wre2205-500mbps-powerline-wireless-n300-range-extender-wre2205-gb0101f Thinking that would be lovely just plugged into the socket by the door in the office, extending the perimeter of the WiFi from my Homehub. I can't get it to work properly! If I plug a laptop into its ethernet port I can get it to connect to the Homehub and give me a kinda of wired, wireless extender. If, however, I plug the ethernet port into my home hub, it then seems to extend the network, but only my iOs devices work, and all my wired stuff stops working, and seems to create an infinite loop where windows connects to my homehob, and then rather to the internet, it then connects back to the extender thing. Anyway... in the meantime, I took a fatal trip to the Apple Store, where I purchased an Airport Express... solely for the purpose of hooking my iOs devices up as wireless music players in the house. I knew it had WiFi, but didn't want to use that part as an extender, I didn't think it would work on a Homehub anyway. It doesn't work on a Homehub! I now have a new wireless network in the house, which, when anything connects to it cannot connect to the Internet, so it works ONLY as a wireless music player. I then borrowed some Powerline Adaptors from someone and realised that this whole thing was getting totally out of control! It seems all the technology is out there but it's so complicated to get the right series of devices. To further add to the confusion, I wouldn't mind a network hard drive. I bought one that broke and lost everything, so now we're on to looking at the Apple Time Capsules. So my question is... IF... I buy an Apple Time Capsule, can I: Hook that up to my Homehub, leaving the homehub connected to the Internet so my Hub phones still work, then disable wireless on the homehub Link up my Airport Express to the Time Capsule PROPERLY so it will connect to the Internet Do the above with an Apple TV box should I buy one in future Use the Time Capsule as a network hard drive to store video and music that can be viewed/listened to via my iOS devices/Apple TV/Aiport Express anywhere even with my main PC off (this currently stores all this data) Hope that the IOS devices like the WiFi from the TimeCapsule better than the Homehub and work without extension, or buy another Airport Express to get WiFI in the office. Or... should I buy an Airport Extreme and use a USB hard drive for the network drive?

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  • Redmine with Apache 2 + Passenger nightmare --- site is up and available, but Redmine doesn't execute

    - by CptSupermrkt
    I was determined to figure this out myself, but I've been at it for a total of more than 10 hours, and I just can't figure this out. First, let me detail my environment (which I cannot change): Server version: Apache/2.2.15 (Unix) Ruby version: ruby 1.9.3p448 Rails version: Rails 4.0.1 Passenger version: Phusion Passenger version 4.0.5 Redmine version: 2.3.3 I have followed the Redmine instructions all the way through the test webserver to check that installation was successful with this command: ruby script/rails server webrick -e production The roadblock which I cannot overcome is getting Apache and Passenger to interpret and properly serve Redmine. I have searched pretty much every possible link within the first 10 pages or so of Google results. Everywhere I go I come across conflicting/contradicting/outdated information. We have a "weird" setup with Apache (which I inherited and cannot change). Redmine needs to be served through SSL, but Apache already has another website it's serving through SSL called Twiki. By "weird", what I mean is that our file structure is entirely different from all the tutorials out there on this version of Apache which have directories like "available-sites" and such. Here are the abbreviated versions of some of our config files: /etc/httpd/conf/httpd.conf (the global configuration file --- note that NO VirtualHost is defined here): ServerRoot "/etc/httpd" ... LoadModule passenger_module /usr/local/pkg/ruby/1.9.3-p448/lib/ruby/gems/1.9.1/gems/passenger-4.0.5/libout/apache2/mod_passenger.so PassengerRoot /usr/local/pkg/ruby/1.9.3-p448/lib/ruby/gems/1.9.1/gems/passenger-4.0.5 PassengerDefaultRuby /usr/local/pkg/ruby/1.9.3-p448/bin/ruby Include conf.d/*.conf ... User apache Group apache ... DocumentRoot "/var/www/html" So just to clarify, the above httpd.conf file does NOT have a VirtualHost section. /etc/httpd/conf.d/ssl.conf (defines the VirtualHost for ssl): Listen 443 <VirtualHost _default_:443> SSLEngine on ... SSLCertificateFile /etc/pki/tls/certs/localhost.crt </VirtualHost> /etc/httpd/conf.d/twiki.conf (this works just fine --- note this does NOT define a VirtualHost): ScriptAlias /twiki/bin/ "/var/www/twiki/bin/" Alias /twiki/ "/var/www/twiki/" <Directory "/var/www/twiki/bin"> AllowOverride None Order Deny,Allow Deny from all AuthType Basic AuthName "our team" AuthBasicProvider ldap ...a lot of ldap and authorization stuff Options ExecCGI FollowSymLinks SetHandler cgi-script </Directory> /etc/httpd/conf.d/redmine.conf: Alias /redmine/ "/var/www/redmine/public/" <Directory "/var/www/redmine/public"> Options Indexes ExecCGI FollowSymLinks Order allow,deny Allow from all AllowOverride all </Directory> The amazing thing is that this doesn't completely NOT work: I can successfully open up https://someserver/redmine/ with SSL and the https://someserver/twiki/ site remains unaffected. This tells me that it IS possible to have two separate sites up with one SSL configuration, so I don't think that's the problem. The problem is is that it opens up to the file index. I can navigate around my Redmine file structure, but no code ever gets executed. For example, there is a file included with Redmine called dispatch.fcgi in the public folder. https://someserver/redmine/dispatch.fcgi opens, but just as plain text code in the browser. As I understand it, in the case of using Passenger, CGI and FastCGI stuff is irrelevant/unused.

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  • Process not Listed by PS or in /proc/

    - by Hammer Bro.
    I'm trying to figure out how to operate a rather large Java program, 'prog'. If I go to its /bin/ dir and configure its setenv.sh and prog.sh to use local directories and my current user account. Then I try to run it via "./prog.sh start". Here are all the relevant bits of prog.sh: USER=(my current account) _CMD="/opt/jdk/bin/java -server -Xmx768m -classpath "${CLASSPATH}" -jar "${DIR}/prog.jar"" case "${ACTION}" in start) nohup su ${USER} -c "exec ${_CMD} >>${_LOGFILE} 2>&1" >/dev/null & echo $! >${_PID} echo "Prog running. PID="`cat ${_PID}` ;; stop) PID=`cat ${_PID} 2>/dev/null` echo "Shutting down prog: ${PID} kill -QUIT ${PID} 2>/dev/null kill ${PID} 2>/dev/null kill -KILL ${PID} 2>/dev/null rm -f ${_PID} echo "STOPPED `date`" >>${_LOGFILE} ;; When I actually do ./prog.sh start, it starts. But I can't find it at all on the process list. Nor can I kill it manually, using the same command the shell script uses. But I can tell it's running, because if I do ./prog.sh stop, it stops (and some temporary files elsewhere clean themselves out). ./prog.sh start Prog running. PID=1234 ps eaux | grep 1234 ps eaux | grep -i prog.jar ps eaux >> pslist.txt (It's not there either by PID or any clear name I can find: prog, java or jar.) cd /proc/1234/ -bash: cd: /proc/1234/: No such file or directory kill -QUIT 1234 kill 1234 kill -KILL 1234 -bash: kill: (1234) - No such process ./prog.sh stop Shutting down prog: 1234 As far as I can tell, the process is running yet not in any way listed by the system. I can't find it in ps or /proc/, nor can I kill it. But the shell script can still stop it properly. So my question is, how can something like this happen? Is the process supremely hidden, actually unlisted, or am I just missing it in some fashion? I'm trying to figure out what makes this program tick, and I can barely prove that it's ticking! Edit: ps eu | grep prog.sh (after having restarted; so random PID) 50038 19381 0.0 0.0 4412 632 pts/3 S+ 16:09 0:00 grep prog.sh HOSTNAME=machine.server.com TERM=vt100 SHELL=/bin/bash HISTSIZE=1000 SSH_CLIENT=::[STUFF] 1754 22 CVSROOT=:[DIR] SSH_TTY=/dev/pts/3 ANT_HOME=/opt/apache-ant-1.7.1 USER=[USER] LS_COLORS=[COLORS] SSH_AUTH_SOCK=[DIR] KDEDIR=/usr MAIL=[DIR] PATH=[DIRS] INPUTRC=/etc/inputrc PWD=[PWD] JAVA_HOME=/opt/jdk1.6.0_21 LANG=en_US.UTF-8 SSH_ASKPASS=/usr/libexec/openssh/gnome-ssh-askpass M2_HOME=/opt/apache-maven-2.2.1 SHLVL=1 HOME=[~] LOGNAME=[USER] SSH_CONNECTION=::[STUFF] LESSOPEN=|/usr/bin/lesspipe.sh %s G_BROKEN_FILENAMES=1 _=/bin/grep OLDPWD=[DIR] I just realized that the stop) part of prog.sh isn't actually a guarantee that the process it claims to be stopping is running -- it just tries to kill the PID and suppresses all output then deletes the temporary file and manually inserts STOPPED into the log file. So I'm no longer so certain that the process is always running when I ps for it, although the code sample above indicates that it at least runs erratically. I'll continue looking into this undocumented behemoth when I return to work tomorrow.

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  • Single-Signon options for Exchange 2010

    - by freiheit
    We're working on a project to migrate employee email from Unix/open-source (courier IMAP, exim, squirrelmail, etc) to Exchange 2010, and trying to figure out options for single-signon for Outlook Web Access. So far all the options I've found are very ugly and "unsupportable", and may simply not work with Forefront. We already have JA-SIG CAS for token-based single-signon and Shibboleth for SAML. Users are directed to a simple in-house portal (a Perl CGI, really) that they use to sign in to most stuff. We have an HA OpenLDAP cluster that's already synchronized against another AD domain and will be synchronized with the AD domain Exchange will be using. CAS authenticates against LDAP. The portal authenticates against CAS. Shibboleth authenticates with CAS but pulls additional data from LDAP. We're moving in the direction of having web services authenticate against CAS or Shibboleth. (Students are already on SAML/Shibboleth authenticated Google Apps for Education) With Squirrelmail we have a horrible hack linked to from that portal page that authenticates against CAS, gets your original plaintext password (yes, I know, evil), and gives you an HTTP form pre-filled with all the necessary squirrelmail login details with javaScript onLoad stuff to immediately submit the form. Trying to find out exactly what is possible with Exchange/OWA seems to be difficult. "CAS" is both the acronym for our single-signon server and an Exchange component. From what I've been able to tell there's an addon for Exchange that does SAML, but only for federating things like free/busy calendar info, not authenticating users. Plus it costs additional money so there's no way to experiment with it to see if it can be coaxed into doing what we want. Our plans for the Exchange cluster involve Forefront Threat Management Gateway (the new ISA) in the DMZ front-ending the CAS servers. So, the real question: Has anybody managed to make Exchange authenticate with CAS (token-based single-signon) or SAML, or with something I can reasonably likely make authenticate with one of those (such as anything that will accept apache's authentication)? With Forefront? Failing that, anybody have some tips on convincing OWA Forms Based Authentication (FBA) into letting us somehow "pre-login" the user? (log in as them and pass back cookies to the user, or giving the user a pre-filled form that autosubmits like we do with squirrelmail). This is the least-favorite option for a number of reasons, but it would (just barely) satisfy our requirements. From what I hear from the guy implementing Forefront, we may have to set OWA to basic authentication and do forms in Forefront for authentication, so it's possible this isn't even possible. I did find CasOwa, but it only mentions Exchange 2007, looks kinda scary, and as near as I can tell is mostly the same OWA FBA hack I was considering slightly more integrated with the CAS server. It also didn't look like many people had had much success with it. And it may not work with Forefront. There's also "CASifying Outlook Web Access 2", but that one scares me, too, and involves setting up a complex proxy config, which seems more likely to break. And, again, doesn't look like it would work with Forefront. Am I missing something with Exchange SAML (OWA Federated whatchamacallit) where it is possible to configure to do user authentication and not just free/busy access authorization?

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  • synchronization of file locations between two machines

    - by intuited
    Although similar threads have been asked on this site and its siblings before, I've not managed to glean the answer to this persistent question. Any help is much appreciated. The situation: I've got two laptops; both contain a ton of music. Sometimes I move these music files to different locations, or change the metadata in them, or convert them to a different format. I might do any of these things on either machine. I rarely do all of them at once — ie it's unlikely that I'll convert a file's format and move it to a different location all in one go. I'd like to be able to synchronize these changes without having to sift through everything that was renamed or moved. I'm familiar with rsync but I find it inadequate, because although it can compute checksums, it doesn't have any way to store them. So if a file differs, it can't figure out which side it changed on. This also means that it can't attempt to match a missing file to a new one with the same checksum (ie a move) if the filesize and date are the same, it , so it takes an epoch to do a sync on a large repository. I would like to only check the checksum if the files even if you turn on checksumming, it still doesn't use it intelligently: ie it checksums files even if the sizes differ. IIRC. it's not able to use file metadata as a means of file comparison. this is sort of a wishlist item but it seems doable. I've also looked into rsnapshot, but its requirement to create a full backup is impractical in this situation. I don't need a backup, I just need a record of what file with each hash was where when. Unison seems like it might be able to do something vaguely along these lines, but I'm loathe to spend hours wading through its details only to discover that it's sadly lacking. Plus, it's fun asking questions on here. What I'd like is a tool that does something along these lines: keeps track of file checksums or of actual renames, possibly using inotify to greatly reduce resource consumption/latency stores a database containing this info, along with other pertinencies like the file format and metadata, the actual inode, the filename history, etc. uses this info to provide more-intelligent synchronization with a counterpart on the other side. So for example: if a file has been converted from flac to ogg, but kept the same base filename, or the same metadata, it should be able to send the new version over, and the other side should delete the original. Probably it should actually sequester it somewhere in case they or you screwed up, but that's a detail. And then when the transaction is done, the state is logged so that the next time the two interact they can work out their differences. Maybe all this metadata stuff is a fancy pipe dream. I would actually be pretty happy if there was something out there that could just use checksums in an intelligent way. This would be sort of like having the intelligence of something like git, minus the need to duplicate data in an index/backup/etc (and branching, and checkouts, and all the other great stuff that RCSs do. basically just fast forward commit pushes are all I want, with maybe the option to roll back.) So is there something out there that can do this? If not, can someone suggest a good way to start making it?

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  • Centos 7. Freeradius fails to start on boot

    - by Alex
    I was messing around with FreeRADIUS and MySQL (MariaDB) and it seems FreeRADIUS service can't start properly on startup. But it starts fine using root user or in debug mode (radiusd -X) and works just fine! Debug mode shows no errors. systemctl command shows that radiusd.service has failed to start. /var/log/messages output: Aug 21 15:52:29 nexus-test systemd: Starting The Apache HTTP Server... Aug 21 15:52:29 nexus-test systemd: Starting MariaDB database server... Aug 21 15:52:29 nexus-test systemd: Starting FreeRADIUS high performance RADIUS server.... Aug 21 15:52:29 nexus-test systemd: Started OpenSSH server daemon. Aug 21 15:52:29 nexus-test mysqld_safe: 140821 15:52:29 mysqld_safe Logging to '/var/log/mariadb/mariadb.log'. Aug 21 15:52:29 nexus-test mysqld_safe: 140821 15:52:29 mysqld_safe Starting mysqld daemon with databases from /var/lib/mysql Aug 21 15:52:30 nexus-test systemd: Started Postfix Mail Transport Agent. Aug 21 15:52:30 nexus-test avahi-daemon[604]: Registering new address record for fe80::250:56ff:fe85:e4af on eth0.*. Aug 21 15:52:30 nexus-test systemd: radiusd.service: control process exited, code=exited status=1 Aug 21 15:52:30 nexus-test systemd: Failed to start FreeRADIUS high performance RADIUS server.. Aug 21 15:52:30 nexus-test systemd: Unit radiusd.service entered failed state. Aug 21 15:52:31 nexus-test kdumpctl: kexec: loaded kdump kernel Aug 21 15:52:31 nexus-test kdumpctl: Starting kdump: [OK] Aug 21 15:52:31 nexus-test systemd: Started Crash recovery kernel arming. Aug 21 15:52:31 nexus-test systemd: Started The Apache HTTP Server. Aug 21 15:52:31 nexus-test systemd: Started MariaDB database server. /var/log/radius/radius.log output: Thu Aug 21 15:24:16 2014 : Info: rlm_sql (sql): Driver rlm_sql_mysql (module rlm_sql_mysql) loaded and linked Thu Aug 21 15:24:16 2014 : Info: rlm_sql (sql): Attempting to connect to database "radius" Thu Aug 21 15:24:16 2014 : Info: rlm_sql (sql): Opening additional connection (0) Thu Aug 21 15:24:16 2014 : Error: rlm_sql_mysql: Couldn't connect socket to MySQL server radius@localhost:radius Thu Aug 21 15:24:16 2014 : Error: rlm_sql_mysql: Mysql error 'Can't connect to local MySQL server through socket '/var/lib/mysql/mysql.sock' (2)' Thu Aug 21 15:24:16 2014 : Error: rlm_sql (sql): Opening connection failed (0) Thu Aug 21 15:24:16 2014 : Error: /etc/raddb/mods-enabled/sql[47]: Instantiation failed for module "sql" After seeing this I tried to replicate the problem, killed mariadb.service and started to run debug mode again. And it spits out the same problem as in the radius.log. I tried disabling iptables and firewalld and rebooting, but no luck: systemctl disable iptables systemctl disable firewalld So maybe the problem is in the process startup order or delay of some kind. Maybe FreeRADIUS's SQL module can't connect to not yet started MariaDB? If it, how can I fix this? In earlier versions of RHEL/CENTOS I know you easily see service start order in like rc.d or stuff, now IDK. I am new to this fancy "systemd", "systemctl", "firewalld" stuff Centos 7 introduced so sorry I'm a little bit confused. Also this new FreeRADIUS 3 structure... PS. MariaDB is enabled on startup, credentials in FR DB configuration are correct A little update: cat /etc/systemd/system/multi-user.target.wants/radiusd.service output: [Unit] Description=FreeRADIUS high performance RADIUS server. After=syslog.target network.target [Service] Type=forking PIDFile=/var/run/radiusd/radiusd.pid ExecStartPre=-/bin/chown -R radiusd.radiusd /var/run/radiusd ExecStartPre=/usr/sbin/radiusd -C ExecStart=/usr/sbin/radiusd -d /etc/raddb ExecReload=/usr/sbin/radiusd -C ExecReload=/bin/kill -HUP $MAINPID [Install] WantedBy=multi-user.target

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  • "Can't create table" when having to many partitions

    - by Chris
    I am currently having a problem I dont understand. Wherever I look it says mySQL (5.5) / InnoDB doesnt have a table limit. I wanted to test the InnoDB compression and was about to create an empty copy of an existing table and ran into the following problem. this one works: CREATE TABLE `hsc` ( LOTS OF STUFF ) ENGINE=InnoDB CHARSET=utf8 PARTITION BY RANGE (pid) SUBPARTITION BY HASH (cons) SUBPARTITIONS 2 (PARTITION hsc_p0 VALUES LESS THAN (10000) , PARTITION hsc_p1 VALUES LESS THAN (20000) , PARTITION hsc_p2 VALUES LESS THAN (30000) , PARTITION hsc_p3 VALUES LESS THAN (40000) , PARTITION hsc_p4 VALUES LESS THAN (50000) , PARTITION hsc_p40 VALUES LESS THAN (4000000) ); this one doesn't: CREATE TABLE `hsc` ( LOTS OF STUFF ) ENGINE=InnoDB CHARSET=utf8 PARTITION BY RANGE (pid) SUBPARTITION BY HASH (cons) SUBPARTITIONS 2 (PARTITION hsc_p0 VALUES LESS THAN (10000) , PARTITION hsc_p1 VALUES LESS THAN (20000) , PARTITION hsc_p2 VALUES LESS THAN (30000) , PARTITION hsc_p3 VALUES LESS THAN (40000) , PARTITION hsc_p4 VALUES LESS THAN (50000) , PARTITION hsc_p5 VALUES LESS THAN (75000) , PARTITION hsc_p6 VALUES LESS THAN (100000) , PARTITION hsc_p7 VALUES LESS THAN (125000) , PARTITION hsc_p8 VALUES LESS THAN (150000) , PARTITION hsc_p9 VALUES LESS THAN (175000) , PARTITION hsc_p40 VALUES LESS THAN (4000000) ); ERROR 1005 (HY000): Can't create table 'hsc' (errno: 1) Its reproducable by removing the number of partitions and adding them again. it does not have to do anything with the name of the table as i tried various names. there is also enough empty space on the HDD. /dev/simfs 230G 26G 192G 12% /var/lib/mysql.mnt There should be no limit on the partitions http://dev.mysql.com/doc/refman/5.5/en/partitioning-limitations.html Maximum number of partitions. The maximum possible number of partitions for a given table (that does not use the NDB storage engine) is 1024. This number includes subpartitions. i have increased both open_files show variables where variable_name LIKE '%open_files%'; +-------------------+-------+ | Variable_name | Value | +-------------------+-------+ | innodb_open_files | 512 | | open_files_limit | 1536 | +-------------------+-------+ No change. Any clues where should I start looking? UPDATE: the whole thing is running in an openvz environment. i saw in users_beancounters that the numflock was a problem, so i increased it. but the problem still persists. maybe this helps: ulimit -a core file size (blocks, -c) 0 data seg size (kbytes, -d) unlimited scheduling priority (-e) 0 file size (blocks, -f) unlimited pending signals (-i) 515011 max locked memory (kbytes, -l) 64 max memory size (kbytes, -m) unlimited open files (-n) 1024 pipe size (512 bytes, -p) 8 POSIX message queues (bytes, -q) 819200 real-time priority (-r) 0 stack size (kbytes, -s) 10240 cpu time (seconds, -t) unlimited max user processes (-u) 515011 virtual memory (kbytes, -v) unlimited file locks (-x) unlimited cat /proc/user_beancounters Version: 2.5 uid resource held maxheld barrier limit failcnt 200: kmemsize 9309653 13357056 14372700 14790164 0 lockedpages 0 1008 2048 2048 0 privvmpages 675424 686528 1048576 1572864 0 shmpages 33 673 21504 21504 0 dummy 0 0 9223372036854775807 9223372036854775807 0 numproc 49 90 240 240 0 physpages 243761 246945 0 9223372036854775807 0 vmguarpages 0 0 1048576 1048576 0 oomguarpages 81672 83305 1048576 1048576 0 numtcpsock 6 8 360 360 0 numflock 175 188 512 512 8 numpty 1 9 16 16 0 numsiginfo 0 48 256 256 0 tcpsndbuf 104640 263912 1720320 2703360 0 tcprcvbuf 98304 131072 1720320 2703360 0 othersockbuf 32368 89304 1126080 2097152 0 dgramrcvbuf 0 2312 262144 262144 0 numothersock 19 28 360 360 0 dcachesize 2285052 3624426 3409920 3624960 0 numfile 616 870 9312 9312 0 dummy 0 0 9223372036854775807 9223372036854775807 0 dummy 0 0 9223372036854775807 9223372036854775807 0 dummy 0 0 9223372036854775807 9223372036854775807 0 numiptent 24 24 128 128 0

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  • Apache config that uses two document roots based on whether the requested resource exists in the first

    - by mattalexx
    Background I have a client site that consists of a CakePHP installation and a Magento installation: /web/example.com/ /web/example.com/app/ <== CakePHP /web/example.com/app/webroot/ <== DocumentRoot /web/example.com/app/webroot/store/ <== Magento /web/example.com/config/ <== Site-wide config /web/example.com/vendors/ <== Site-wide libraries The server runs Apache 2.2.3. The problem The whole company has FTP access and got used to clogging up the /web/example.com/, /web/example.com/app/webroot/, and /web/example.com/app/webroot/store/ directories with their own files. Sometimes these files need HTTP access and sometimes they don't. In any case, this mess makes my job harder when it comes to maintaining the site. Code merges, tarring the live code, etc, is very complicated and usually requires a bunch of filters. Abandoned solution At first, I thought I would set up a new subdomain on the same server, move all of their files there, and change their FTP chroot. But that wouldn't work for these reasons: Firstly, I have no idea (and neither do they remember) what marketing materials they've sent out that contain URLs to certain resources they've uploaded to the server, using the main domain, and also using abstract subdomains that use the main virtual host because it has ServerAlias *.example.com. So suddenly having them only use static.example.com isn't feasible. Secondly, The PHP scripts in their projects are potentially very non-portable. I want their files to stay in as similar an environment as they were built as I can. Also, I do not want to debug their code to make it portable. Half-baked solution After some thought, I decided to find a way to section off the actual website files into another directory that they would not touch. The company's uploaded files would stay where they were. This would ensure that I didn't break any of their projects that needed HTTP access. It would look something like this: /web/example.com/ <== A bunch of their files are in here /web/example.com/app/webroot/ <== 1st DocumentRoot; A bunch of their files are in here /web/example.com/app/webroot/store/ <== Some more are in here /web/example.com/site/ <== New dir; Contains only site files /web/example.com/site/app/ <== CakePHP /web/example.com/site/app/webroot/ <== 2nd DocumentRoot /web/example.com/site/app/webroot/store/ <== Magento /web/example.com/site/config/ <== Site-wide config /web/example.com/site/vendors/ <== Site-wide libraries After I made this change, I would not need to pay attention to anything except for the stuff within /web/example.com/site/ and my job would be a lot easier. I would be the only one changing stuff in there. So here's where the Apache magic would happen: I need an HTTP request to http://www.example.com/ to first use /web/example.com/app/webroot/ as the document root. If nothing is found (no miscellaneous uploaded company projects are found), try finding something within /web/example.com/site/app/webroot/. Another thing to keep in mind is, the site might have some problems if the $_SERVER['DOCUMENT_ROOT'] variable reads /web/example.com/app/webroot/ but the actual files are within /web/example.com/site/app/webroot/. It would be better if the DOCUMENT_ROOT environment variable could be /web/example.com/site/app/webroot/ for anything within the /web/example.com/site/app/webroot/ directory. Conclusion Is my half-baked solution possible with Apache 2.2.3? Is there a better way to solve this problem?

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  • Apache sends plain-text response when accessing SSL-enabled site without HTTPS

    - by animuson
    I've never encountered something such as this before. I was attempting to simply redirect the page to the HTTPS version if it determined that HTTPS was off, but instead it's displaying an HTML page rather than actually redirecting; and even odder, it's displaying it as text/plain! The VirtualHost Declaration (Sort of): ServerAdmin [email protected] DocumentRoot "/path/to/files" ServerName example.com SSLEngine On SSLCertificateFile /etc/ssh/certify/example.com.crt SSLCertificateKeyFile /etc/ssh/certify/example.com.key SSLCertificateChainFile /etc/ssh/certify/sub.class1.server.ca.pem <Directory "/path/to/files/"> AllowOverride All Options +FollowSymLinks DirectoryIndex index.php Order allow,deny Allow from all </Directory> RewriteEngine On RewriteCond %{HTTPS} off RewriteRule .* https://example.com:6161 [R=301] The Page Output: <!DOCTYPE HTML PUBLIC "-//IETF//DTD HTML 2.0//EN"> <html><head> <title>301 Moved Permanently</title> </head><body> <h1>Moved Permanently</h1> <p>The document has moved <a href="https://example.com:6161">here</a>.</p> <hr> <address>Apache/2.2.21 (Unix) mod_ssl/2.2.21 OpenSSL/1.0.0e DAV/2 Server at example.com Port 443</address> </body></html> I've tried moving the Rewrite stuff up above the SSL stuff hoping it'd do something and nothing happens. If I view the page with via HTTPS, it displays fine like it should. It's obviously detecting that I'm trying to rewrite the path, but it's not acting. The Apache error log does not indicate anything to me that might have gone wrong. When I remove the RewriteRules: <!DOCTYPE HTML PUBLIC "-//IETF//DTD HTML 2.0//EN"> <html><head> <title>400 Bad Request</title> </head><body> <h1>Bad Request</h1> <p>Your browser sent a request that this server could not understand.<br /> Reason: You're speaking plain HTTP to an SSL-enabled server port.<br /> Instead use the HTTPS scheme to access this URL, please.<br /> <blockquote>Hint: <a href="https://example.com/"><b>https://example.com/</b></a></blockquote></p> <p>Additionally, a 404 Not Found error was encountered while trying to use an ErrorDocument to handle the request.</p> <hr> <address>Apache/2.2.21 (Unix) mod_ssl/2.2.21 OpenSSL/1.0.0e DAV/2 Server at example.com Port 443</address> </body></html> I get the standard "you can't do this because you're not using SSL" response, which is also provided in text/plain rather than being rendered as HTML. This would make sense, it should only work for HTTPS-enabled connections, but I still want to redirect them to the HTTPS connection when it determines that it is not enabled. Thinking I could circumvent the system: I tried adding a ErrorDocument 400 https://example.com:6161 to the config file instead of using RewriteRules, and that just gave me a new message, still no cheese. <!DOCTYPE HTML PUBLIC "-//IETF//DTD HTML 2.0//EN"> <html><head> <title>302 Found</title> </head><body> <h1>Found</h1> <p>The document has moved <a href="https://example.com:6161">here</a>.</p> <hr> <address>Apache/2.2.21 (Unix) mod_ssl/2.2.21 OpenSSL/1.0.0e DAV/2 Server at example.com Port 443</address> </body></html> How can I force Apache to actually redirect rather than displaying a "301" page that shows HTML in plain-text format?

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  • In Linux, what's the best way to delegate administration responsibilities, like for Apache, a database, or some other application?

    - by Andrew Banks
    In Linux, what's the best way to delegate administration responsibilities for Apache and other "applications"? File permissions? Sudo? A mix of both? Something else? At work we have two tiers of "administrators" Operating system administrators. These are your run-of-the-mill "server administrators." They are responsible for just the operating system. Application administrators. The people who build the web site. This includes not only writing the SQL, PHP, and HTML, but also setting up and running Apache and PostgreSQL or MySQL. The aforementioned OS admins will install this stuff, but it's mainly up to the app admins to edit all the config files, start and stop processes when needed, and so on. I am one of the app admins. This is different than what I am used to. I used to just write code. The sysadmin took care not only of the OS but also installing, setting up, and keeping up the server software. But he left. Now I'm in charge of setting up Apache and the database. The new sysadmins say they just handle the operating system. It's no problem. I welcome learning new stuff. But there is a learning curve, even for the OS admins. Apache, by default, seems to be set up for administration by root directly. All the config files and scripts are 644 and owned by root:root. I'm not given the root password, naturally, so the OS admins must somehow give my ordinary OS user account all the rights necessary to edit Apache's config files, start and stop it, read its log files, and so on. Right now they're using a mix of: (1) giving me certain sudo rights, (2) adding me to certain groups, and (3) changing the file permissions of various directories, to make them writable by one of the groups I'm in. This never goes smoothly. There's always a back-and-forth between me and the sysadmins. They say it's ready. Then I try certain things, and half of them I still can't do. So they make some more changes. Then finally I seem to be independent and can administer Apache and the database without pestering them anymore. It's the sheer complication and amount of changes that make me uncomfortable. Even though it finally works, more or less, it seems hackneyed. I feel like we're doing it wrong. It seems like the makers of the software would have anticipated this scenario (someone other than root administering it) and have a clean two- or three-step program to delegate responsibility to me. But it feels like we are really chewing up the filesystem and making it far and away from the default set-up. Any suggestions? Are we doing it the recommended way? P.S. For PostgreSQL it seems a little better. Its files are owned by a system user named postgres. So giving me the right to run sudo su - postgres gives me just about everything. I'm just now getting into MySQL, but it seems to be set up similarly. But it seems a little weird doing all my work as another user.

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  • My PC suddenly doesn't detect the primary drive (SSD)

    - by smoth190
    My computer has been working fine for months, and it worked today, but tonight I went to start it up to find that my OCZ Vertex 2 isn't being found. When I turn on my computer, the loading screen gets stuck at "Detecting IDE drives...". After a while, it keeps going and lists the drives it finds. The first one in the list should be my Vertex 2, but it just says "None". The computer proceeds to get stuck on "Loading operating system...", which is understandable because the drive with the OS is "gone". My first thought was drive failure, but every time drives have crashed on me, they're still detected--they just don't work. This drive is an SSD, it's pretty new, and I had no problems beforehand. I find it hard to believe it failed. I'm sure it's possible, but I hope this isn't the case. There has been nothing strange going on at all with my PC, it's been running perfect until now. I was just about to do my monthly dskchk and defrag today. I popped in my Windows 7 Home Premium disk and booted from it. When I launched the repair tool, it didn't list any operating systems (because the drive is 100% missing...). When I've had disks crash before, it still listed the OS, you just couldn't do anything with it. I tried to restore from an image, but I don't have any of those, either. I opened the command console and listed the drivers with wmic logicaldisk get name. Only C: and D: came up. C: was my 1TB storage driver (luckily, all my stuff is here--only the OS is on the SSD!) and D: was the disk driver. So I still had an MIA drive... The SSD didn't come with any driver disks, so I can't install drivers. If there's a way to do this from a CD I can burn with my other PC, please let me know. What the heck do I do? Although only the OS is on my SSD, a new SSD is expensive. I'll probably also have to buy a new copy of Windows (an upgrade would be nice, though...) because I've found it eats my registration key when my PC crashes (and my thousands of dollars of Adobe programs, I'll be on the phone with tech support for a week to get those keys back). And I'll lose my registry, all my settings, all sorts of other stuff that I'll spend weeks restoring. My computer is a pain in the butt to take out and open up, so if I can't fix it, I'll try fiddling with the plug or putting it into a new computer, but not right now. Any help is greatly appreciated! The day when they make crash-less drives will be the day I live without worry.

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  • Log transport and aggregation at scale

    - by markdrayton
    How're you analysing log files from UNIX/Linux machines? We run several hundred servers which all generate their own log files, either directly or through syslog. I'm looking for a decent solution to aggregate these and pick out important events. This problem breaks down into 3 components: 1) Message transport The classic way is to use syslog to log messages to a remote host. This works fine for applications that log into syslog but less useful for apps that write to a local file. Solutions for this might include having the application log into a FIFO connected to a program to send the message using syslog, or by writing something that will grep the local files and send the output to the central syslog host. However, if we go to the trouble of writing tools to get messages into syslog would we be better replacing the whole lot with something like Facebook's Scribe which offers more flexibility and reliability than syslog? 2) Message aggregation Log entries seem to fall into one of two types: per-host and per-service. Per-host messages are those which occur on one machine; think disk failures or suspicious logins. Per-service messages occur on most or all of the hosts running a service. For instance, we want to know when Apache finds an SSI error but we don't want the same error from 100 machines. In all cases we only want to see one of each type of message: we don't want 10 messages saying the same disk has failed, and we don't want a message each time a broken SSI is hit. One approach to solving this is to aggregate multiple messages of the same type into one on each host, send the messages to a central server and then aggregate messages of the same kind into one overall event. SER can do this but it's awkward to use. Even after a couple of days of fiddling I had only rudimentary aggregations working and had to constantly look up the logic SER uses to correlate events. It's powerful but tricky stuff: I need something which my colleagues can pick up and use in the shortest possible time. SER rules don't meet that requirement. 3) Generating alerts How do we tell our admins when something interesting happens? Mail the group inbox? Inject into Nagios? So, how're you solving this problem? I don't expect an answer on a plate; I can work out the details myself but some high-level discussion on what is surely a common problem would be great. At the moment we're using a mishmash of cron jobs, syslog and who knows what else to find events. This isn't extensible, maintainable or flexible and as such we miss a lot of stuff we shouldn't. Updated: we're already using Nagios for monitoring which is great for detected down hosts/testing services/etc but less useful for scraping log files. I know there are log plugins for Nagios but I'm interested in something more scalable and hierarchical than per-host alerts.

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  • A failed disk (Pay for professional service or SpinRite?)(new edit)

    - by huggie
    EDIT: After much negotiating and begging and seeing through promotion smoke screen, thanks to the nice representative who took my case, I now know that the engineer has already fixed my NTFS partition (I guess it might be a bad block in the partition table?). She told me that the problem was considered minor, and I should be able to boot normally and just copy stuff out. Whew..I'm glad I didn't agree to the NTD $16,000 deal. New question (should this be in a new thread?): is it safer to use the linux "dd" command or is it better to boot normally into Windows XP and just copy stuff out? EDIT2: Thanks to all the help. I give the best answer to Console as it's most directed related to my question. But many suggestion are helpful and informational. ---- ORIGINAL POST BELOW --- Hi, in my previous post (You don't need to read but it's at http://superuser.com/questions/48838/windows-xp-a-disk-read-error-occurred), I said that my hard disk was not booting and is showing "a disk read error occurred". I took it to a recovery professional. A representative responded today told me that the NTFS partitions have a "NTFS partition system crash". I have no idea what that means. The engineer handling my drive will not be available for contact till tomorrow. Now the company charges me NTD (New Taiwan Dollar) $16,000 to recover lost data, that's kind of a lot considering that my graduate student monthly stipend is currently NTD $32,000 (max. allowed by regulation, may be lower, may change depend on funding). Now I'm weighting in between the options. Option A: let the professional recovers it with the half of my monthly stipend. If file/directories I designated are not recovered I don't pay a penny. (other than the initial examination fee of NTD $1000 which I've already paid.) Option B: let me try SpinRite, if failed, back to Option A. I spoke to the representative at the company they recommended me not to handle it on my own (yeah of course that's what they all want to say, right?), and at the price tag the disk error is probably relatively minor and data recoverable. But the representative really did not have detailed information of the disk failure so I didn't take her recommendation readily. Though one thing I heed was that she said that what they would do is to duplicate the disk before attempting discovery, so there would be no data loss (Is this true? can't duplicating invoke further data loss?). That sounds very good to me. Or maybe a third option: Option C: Negotiate with them to pay them to duplicate the disk hopefully for a much smaller price tag. Let me try SpinRite, if failed, back to Option A. This is a difficult decision. Ultimately I want my data back, but if a cheaper way is available to achieve the same thing... Can operating with SpinRite also corrupt data in someway? I've no idea what happened to my drive. I'll attempt to contact the engineer and hope to get it clarified and make an edit here.

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  • how to remove obsolete device and network entries? Device manager "uninstall" option has no effect

    - by Gizmo
    I am trying to remove a few "obsolete" things which annoy me (because I like to have everything cleen, working and not interferring with each other, fresh, etc..). I tried looking for solutions without any help, so here I am to ask. My first part is about removing obsolete networks, let me explain by showing the ipconfig output: C:\windows\system32>ipconfig Windows IP Configuration Wireless LAN adapter Local Area Connection* 11: Media State . . . . . . . . . . . : Media disconnected Connection-specific DNS Suffix . : Wireless LAN adapter Local Area Connection* 9: Media State . . . . . . . . . . . : Media disconnected Connection-specific DNS Suffix . : Ethernet adapter LAN: Media State . . . . . . . . . . . : Media disconnected Connection-specific DNS Suffix . : Wireless LAN adapter Wi-Fi: Connection-specific DNS Suffix . : home Link-local IPv6 Address . . . . . : fe80::c129:8d57:bbd1:3564%10 IPv4 Address. . . . . . . . . . . : 192.168.2.1 Subnet Mask . . . . . . . . . . . : 255.255.255.0 Default Gateway . . . . . . . . . : 192.168.2.254 Tunnel adapter isatap.home: Media State . . . . . . . . . . . : Media disconnected Connection-specific DNS Suffix . : home C:\windows\system32> Specificalyy the first two adapter entries annoy me because the adapters are not visible in the network connection menu (invisible folder / file visibility set to "show"): And here is the second problem altogether with the first one: No matter what I click/do, Uninstall option has no effect on the multiplexor driver. (bridging stuff, right?) I really want to remove the Wireless LAN adapter Local Area Connection entries and the adapter multiplexor stuff but it's impossible? Why is this? How can I remove them anyway?

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