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  • Setting useLegacyV2RuntimeActivationPolicy At Runtime

    - by Reed
    Version 4.0 of the .NET Framework included a new CLR which is almost entirely backwards compatible with the 2.0 version of the CLR.  However, by default, mixed-mode assemblies targeting .NET 3.5sp1 and earlier will fail to load in a .NET 4 application.  Fixing this requires setting useLegacyV2RuntimeActivationPolicy in your app.Config for the application.  While there are many good reasons for this decision, there are times when this is extremely frustrating, especially when writing a library.  As such, there are (rare) times when it would be beneficial to set this in code, at runtime, as well as verify that it’s running correctly prior to receiving a FileLoadException. Typically, loading a pre-.NET 4 mixed mode assembly is handled simply by changing your app.Config file, and including the relevant attribute in the startup element: <?xml version="1.0" encoding="utf-8" ?> <configuration> <startup useLegacyV2RuntimeActivationPolicy="true"> <supportedRuntime version="v4.0"/> </startup> </configuration> .csharpcode { background-color: #ffffff; font-family: consolas, "Courier New", courier, monospace; color: black; font-size: small } .csharpcode pre { background-color: #ffffff; font-family: consolas, "Courier New", courier, monospace; color: black; font-size: small } .csharpcode pre { margin: 0em } .csharpcode .rem { color: #008000 } .csharpcode .kwrd { color: #0000ff } .csharpcode .str { color: #006080 } .csharpcode .op { color: #0000c0 } .csharpcode .preproc { color: #cc6633 } .csharpcode .asp { background-color: #ffff00 } .csharpcode .html { color: #800000 } .csharpcode .attr { color: #ff0000 } .csharpcode .alt { background-color: #f4f4f4; margin: 0em; width: 100% } .csharpcode .lnum { color: #606060 } This causes your application to run correctly, and load the older, mixed-mode assembly without issues. For full details on what’s happening here and why, I recommend reading Mark Miller’s detailed explanation of this attribute and the reasoning behind it. Before I show any code, let me say: I strongly recommend using the official approach of using app.config to set this policy. That being said, there are (rare) times when, for one reason or another, changing the application configuration file is less than ideal. While this is the supported approach to handling this issue, the CLR Hosting API includes a means of setting this programmatically via the ICLRRuntimeInfo interface.  Normally, this is used if you’re hosting the CLR in a native application in order to set this, at runtime, prior to loading the assemblies.  However, the F# Samples include a nice trick showing how to load this API and bind this policy, at runtime.  This was required in order to host the Managed DirectX API, which is built against an older version of the CLR. This is fairly easy to port to C#.  Instead of a direct port, I also added a little addition – by trapping the COM exception received if unable to bind (which will occur if the 2.0 CLR is already bound), I also allow a runtime check of whether this property was setup properly: public static class RuntimePolicyHelper { public static bool LegacyV2RuntimeEnabledSuccessfully { get; private set; } static RuntimePolicyHelper() { ICLRRuntimeInfo clrRuntimeInfo = (ICLRRuntimeInfo)RuntimeEnvironment.GetRuntimeInterfaceAsObject( Guid.Empty, typeof(ICLRRuntimeInfo).GUID); try { clrRuntimeInfo.BindAsLegacyV2Runtime(); LegacyV2RuntimeEnabledSuccessfully = true; } catch (COMException) { // This occurs with an HRESULT meaning // "A different runtime was already bound to the legacy CLR version 2 activation policy." LegacyV2RuntimeEnabledSuccessfully = false; } } [ComImport] [InterfaceType(ComInterfaceType.InterfaceIsIUnknown)] [Guid("BD39D1D2-BA2F-486A-89B0-B4B0CB466891")] private interface ICLRRuntimeInfo { void xGetVersionString(); void xGetRuntimeDirectory(); void xIsLoaded(); void xIsLoadable(); void xLoadErrorString(); void xLoadLibrary(); void xGetProcAddress(); void xGetInterface(); void xSetDefaultStartupFlags(); void xGetDefaultStartupFlags(); [MethodImpl(MethodImplOptions.InternalCall, MethodCodeType = MethodCodeType.Runtime)] void BindAsLegacyV2Runtime(); } } Using this, it’s possible to not only set this at runtime, but also verify, prior to loading your mixed mode assembly, whether this will succeed. In my case, this was quite useful – I am working on a library purely for internal use which uses a numerical package that is supplied with both a completely managed as well as a native solver.  The native solver uses a CLR 2 mixed-mode assembly, but is dramatically faster than the pure managed approach.  By checking RuntimePolicyHelper.LegacyV2RuntimeEnabledSuccessfully at runtime, I can decide whether to enable the native solver, and only do so if I successfully bound this policy. There are some tricks required here – To enable this sort of fallback behavior, you must make these checks in a type that doesn’t cause the mixed mode assembly to be loaded.  In my case, this forced me to encapsulate the library I was using entirely in a separate class, perform the check, then pass through the required calls to that class.  Otherwise, the library will load before the hosting process gets enabled, which in turn will fail. This code will also, of course, try to enable the runtime policy before the first time you use this class – which typically means just before the first time you check the boolean value.  As a result, checking this early on in the application is more likely to allow it to work. Finally, if you’re using a library, this has to be called prior to the 2.0 CLR loading.  This will cause it to fail if you try to use it to enable this policy in a plugin for most third party applications that don’t have their app.config setup properly, as they will likely have already loaded the 2.0 runtime. As an example, take a simple audio player.  The code below shows how this can be used to properly, at runtime, only use the “native” API if this will succeed, and fallback (or raise a nicer exception) if this will fail: public class AudioPlayer { private IAudioEngine audioEngine; public AudioPlayer() { if (RuntimePolicyHelper.LegacyV2RuntimeEnabledSuccessfully) { // This will load a CLR 2 mixed mode assembly this.audioEngine = new AudioEngineNative(); } else { this.audioEngine = new AudioEngineManaged(); } } public void Play(string filename) { this.audioEngine.Play(filename); } } Now – the warning: This approach works, but I would be very hesitant to use it in public facing production code, especially for anything other than initializing your own application.  While this should work in a library, using it has a very nasty side effect: you change the runtime policy of the executing application in a way that is very hidden and non-obvious.

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  • Testing Workflows &ndash; Test-After

    - by Timothy Klenke
    Originally posted on: http://geekswithblogs.net/TimothyK/archive/2014/05/30/testing-workflows-ndash-test-after.aspxIn this post I’m going to outline a few common methods that can be used to increase the coverage of of your test suite.  This won’t be yet another post on why you should be doing testing; there are plenty of those types of posts already out there.  Assuming you know you should be testing, then comes the problem of how do I actual fit that into my day job.  When the opportunity to automate testing comes do you take it, or do you even recognize it? There are a lot of ways (workflows) to go about creating automated tests, just like there are many workflows to writing a program.  When writing a program you can do it from a top-down approach where you write the main skeleton of the algorithm and call out to dummy stub functions, or a bottom-up approach where the low level functionality is fully implement before it is quickly wired together at the end.  Both approaches are perfectly valid under certain contexts. Each approach you are skilled at applying is another tool in your tool belt.  The more vectors of attack you have on a problem – the better.  So here is a short, incomplete list of some of the workflows that can be applied to increasing the amount of automation in your testing and level of quality in general.  Think of each workflow as an opportunity that is available for you to take. Test workflows basically fall into 2 categories:  test first or test after.  Test first is the best approach.  However, this post isn’t about the one and only best approach.  I want to focus more on the lesser known, less ideal approaches that still provide an opportunity for adding tests.  In this post I’ll enumerate some test-after workflows.  In my next post I’ll cover test-first. Bug Reporting When someone calls you up or forwards you a email with a vague description of a bug its usually standard procedure to create or verify a reproduction plan for the bug via manual testing and log that in a bug tracking system.  This can be problematic.  Often reproduction plans when written down might skip a step that seemed obvious to the tester at the time or they might be missing some crucial environment setting. Instead of data entry into a bug tracking system, try opening up the test project and adding a failing unit test to prove the bug.  The test project guarantees that all aspects of the environment are setup properly and no steps are missing.  The language in the test project is much more precise than the English that goes into a bug tracking system. This workflow can easily be extended for Enhancement Requests as well as Bug Reporting. Exploratory Testing Exploratory testing comes in when you aren’t sure how the system will behave in a new scenario.  The scenario wasn’t planned for in the initial system requirements and there isn’t an existing test for it.  By definition the system behaviour is “undefined”. So write a new unit test to define that behaviour.  Add assertions to the tests to confirm your assumptions.  The new test becomes part of the living system specification that is kept up to date with the test suite. Examples This workflow is especially good when developing APIs.  When you are finally done your production API then comes the job of writing documentation on how to consume the API.  Good documentation will also include code examples.  Don’t let these code examples merely exist in some accompanying manual; implement them in a test suite. Example tests and documentation do not have to be created after the production API is complete.  It is best to write the example code (tests) as you go just before the production code. Smoke Tests Every system has a typical use case.  This represents the basic, core functionality of the system.  If this fails after an upgrade the end users will be hosed and they will be scratching their heads as to how it could be possible that an update got released with this core functionality broken. The tests for this core functionality are referred to as “smoke tests”.  It is a good idea to have them automated and run with each build in order to avoid extreme embarrassment and angry customers. Coverage Analysis Code coverage analysis is a tool that reports how much of the production code base is exercised by the test suite.  In Visual Studio this can be found under the Test main menu item. The tool will report a total number for the code coverage, which can be anywhere between 0 and 100%.  Coverage Analysis shouldn’t be used strictly for numbers reporting.  Companies shouldn’t set minimum coverage targets that mandate that all projects must have at least 80% or 100% test coverage.  These arbitrary requirements just invite gaming of the coverage analysis, which makes the numbers useless. The analysis tool will break down the coverage by the various classes and methods in projects.  Instead of focusing on the total number, drill down into this view and see which classes have high or low coverage.  It you are surprised by a low number on a class this is an opportunity to add tests. When drilling through the classes there will be generally two types of reaction to a surprising low test coverage number.  The first reaction type is a recognition that there is low hanging fruit to be picked.  There may be some classes or methods that aren’t being tested, which could easy be.  The other reaction type is “OMG”.  This were you find a critical piece of code that isn’t under test.  In both cases, go and add the missing tests. Test Refactoring The general theme of this post up to this point has been how to add more and more tests to a test suite.  I’ll step back from that a bit and remind that every line of code is a liability.  Each line of code has to be read and maintained, which costs money.  This is true regardless whether the code is production code or test code. Remember that the primary goal of the test suite is that it be easy to read so that people can easily determine the specifications of the system.  Make sure that adding more and more tests doesn’t interfere with this primary goal. Perform code reviews on the test suite as often as on production code.  Hold the test code up to the same high readability standards as the production code.  If the tests are hard to read then change them.  Look to remove duplication.  Duplicate setup code between two or more test methods that can be moved to a shared function.  Entire test methods can be removed if it is found that the scenario it tests is covered by other tests.  Its OK to delete a test that isn’t pulling its own weight anymore. Remember to only start refactoring when all the test are green.  Don’t refactor the tests and the production code at the same time.  An automated test suite can be thought of as a double entry book keeping system.  The unchanging, passing production code serves as the tests for the test suite while refactoring the tests. As with all refactoring, it is best to fit this into your regular work rather than asking for time later to get it done.  Fit this into the standard red-green-refactor cycle.  The refactor step no only applies to production code but also the tests, but not at the same time.  Perhaps the cycle should be called red-green-refactor production-refactor tests (not quite as catchy).   That about covers most of the test-after workflows I can think of.  In my next post I’ll get into test-first workflows.

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  • Creating a Training Lab on Windows Azure

    - by Michael Stephenson
    Originally posted on: http://geekswithblogs.net/michaelstephenson/archive/2013/06/17/153149.aspxThis week we are preparing for a training course that Alan Smith will be running for the support teams at one of my customers around Windows Azure. In order to facilitate the training lab we have a few prerequisites we need to handle. One of the biggest ones is that although the support team all have MSDN accounts the local desktops they work on are not ideal for running most of the labs as we want to give them some additional developer background training around Azure. Some recent Azure announcements really help us in this area: MSDN software can now be used on Azure VM You don't pay for Azure VM's when they are no longer used  Since the support team only have limited experience of Windows Azure and the organisation also have an Enterprise Agreement we decided it would be best value for money to spin up a training lab in a subscription on the EA and then we can turn the machines off when we are done. At the same time we would be able to spin them back up when the users need to do some additional lab work once the training course is completed. In order to achieve this I wanted to create a powershell script which would setup my training lab. The aim was to create 18 VM's which would be based on a prebuilt template with Visual Studio and the Azure development tools. The script I used is described below The Start & Variables The below text will setup the powershell environment and some variables which I will use elsewhere in the script. It will also import the Azure Powershell cmdlets. You can see below that I will need to download my publisher settings file and know some details from my Azure account. At this point I will assume you have a basic understanding of Azure & Powershell so already know how to do this. Set-ExecutionPolicy Unrestrictedcls $startTime = get-dateImport-Module "C:\Program Files (x86)\Microsoft SDKs\Windows Azure\PowerShell\Azure\Azure.psd1"# Azure Publisher Settings $azurePublisherSettings = '<Your settings file>.publishsettings'  # Subscription Details $subscriptionName = "<Your subscription name>" $defaultStorageAccount = "<Your default storage account>"  # Affinity Group Details $affinityGroup = '<Your affinity group>' $dataCenter = 'West Europe' # From Get-AzureLocation  # VM Details $baseVMName = 'TRN' $adminUserName = '<Your admin username>' $password = '<Your admin password>' $size = 'Medium' $vmTemplate = '<The name of your VM template image>' $rdpFilePath = '<File path to save RDP files to>' $machineSettingsPath = '<File path to save machine info to>'    Functions In the next section of the script I have some functions which are used to perform certain actions. The first is called CreateVM. This will do the following actions: If the VM already exists it will be deleted Create the cloud service Create the VM from the template I have created Add an endpoint so we can RDP to them all over the same port Download the RDP file so there is a short cut the trainees can easily access the machine via Write settings for the machine to a log file  function CreateVM($machineNo) { # Specify a name for the new VM $machineName = "$baseVMName-$machineNo" Write-Host "Creating VM: $machineName"       # Get the Azure VM Image      $myImage = Get-AzureVMImage $vmTemplate   #If the VM already exists delete and re-create it $existingVm = Get-AzureVM -Name $machineName -ServiceName $serviceName if($existingVm -ne $null) { Write-Host "VM already exists so deleting it" Remove-AzureVM -Name $machineName -ServiceName $serviceName }   "Creating Service" $serviceName = "bupa-azure-train-$machineName" Remove-AzureService -Force -ServiceName $serviceName New-AzureService -Location $dataCenter -ServiceName $serviceName   Write-Host "Creating VM: $machineName" New-AzureQuickVM -Windows -name $machineName -ServiceName $serviceName -ImageName $myImage.ImageName -InstanceSize $size -AdminUsername $adminUserName -Password $password  Write-Host "Updating the RDP endpoint for $machineName" Get-AzureVM -name $machineName -ServiceName $serviceName ` | Add-AzureEndpoint -Name RDP -Protocol TCP -LocalPort 3389 -PublicPort 550 ` | Update-AzureVM    Write-Host "Get the RDP File for machine $machineName" $machineRDPFilePath = "$rdpFilePath\$machineName.rdp" Get-AzureRemoteDesktopFile -name $machineName -ServiceName $serviceName -LocalPath "$machineRDPFilePath"   WriteMachineSettings "$machineName" "$serviceName" }    The delete machine settings function is used to delete the log file before we start re-running the process.  function DeleteMachineSettings() { Write-Host "Deleting the machine settings output file" [System.IO.File]::Delete("$machineSettingsPath"); }    The write machine settings function will get the VM and then record its details to the log file. The importance of the log file is that I can easily provide the information for all of the VM's to our infrastructure team to be able to configure access to all of the VM's    function WriteMachineSettings([string]$vmName, [string]$vmServiceName) { Write-Host "Writing to the machine settings output file"   $vm = Get-AzureVM -name $vmName -ServiceName $vmServiceName $vmEndpoint = Get-AzureEndpoint -VM $vm -Name RDP   $sb = new-object System.Text.StringBuilder $sb.Append("Service Name: "); $sb.Append($vm.ServiceName); $sb.Append(", "); $sb.Append("VM: "); $sb.Append($vm.Name); $sb.Append(", "); $sb.Append("RDP Public Port: "); $sb.Append($vmEndpoint.Port); $sb.Append(", "); $sb.Append("Public DNS: "); $sb.Append($vmEndpoint.Vip); $sb.AppendLine(""); [System.IO.File]::AppendAllText($machineSettingsPath, $sb.ToString());  } # end functions    Rest of Script In the rest of the script it is really just the bit that orchestrates the actions we want to happen. It will load the publisher settings, select the Azure subscription and then loop around the CreateVM function and create 16 VM's  Import-AzurePublishSettingsFile $azurePublisherSettings Set-AzureSubscription -SubscriptionName $subscriptionName -CurrentStorageAccount $defaultStorageAccount Select-AzureSubscription -SubscriptionName $subscriptionName  DeleteMachineSettings    "Starting creating Bupa International Azure Training Lab" $numberOfVMs = 16  for ($index=1; $index -le $numberOfVMs; $index++) { $vmNo = "$index" CreateVM($vmNo); }    "Finished creating Bupa International Azure Training Lab" # Give it a Minute Start-Sleep -s 60  $endTime = get-date "Script run time " + ($endTime - $startTime)    Conclusion As you can see there is nothing too fancy about this script but in our case of creating a small isolated training lab which is not connected to our corporate network then we can easily use this to provision the lab. Im sure if this is of use to anyone you can easily modify it to do other things with the lab environment too. A couple of points to note are that there are some soft limits in Azure about the number of cores and services your subscription can use. You may need to contact the Azure support team to be able to increase this limit. In terms of the real business value of this approach, it was not possible to use the existing desktops to do the training on, and getting some internal virtual machines would have been relatively expensive and time consuming for our ops team to do. With the Azure option we are able to spin these machines up for a temporary period during the training course and then throw them away when we are done. We expect the costing of this test lab to be very small, especially considering we have EA pricing. As a ball park I think my 18 lab VM training environment will cost in the region of $80 per day on our EA. This is a fraction of the cost of the creation of a single VM on premise.

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  • Building an Infrastructure Cloud with Oracle VM for x86 + Enterprise Manager 12c

    - by Richard Rotter
    Cloud Computing? Everyone is talking about Cloud these days. Everyone is explaining how the cloud will help you to bring your service up and running very fast, secure and with little effort. You can find these kinds of presentations at almost every event around the globe. But what is really behind all this stuff? Is it really so simple? And the answer is: Yes it is! With the Oracle SW Stack it is! In this post, I will try to bring this down to earth, demonstrating how easy it could be to build a cloud infrastructure with Oracle's solution for cloud computing.But let me cover some basics first: How fast can you build a cloud?How elastic is your cloud so you can provide new services on demand? How much effort does it take to monitor and operate your Cloud Infrastructure in order to meet your SLAs?How easy is it to chargeback for your services provided? These are the critical success factors of Cloud Computing. And Oracle has an answer to all those questions. By using Oracle VM for X86 in combination with Enterprise Manager 12c you can build and control your cloud environment very fast and easy. What are the fundamental building blocks for your cloud? Oracle Cloud Building Blocks #1 Hardware Surprise, surprise. Even the cloud needs to run somewhere, hence you will need hardware. This HW normally consists of servers, storage and networking. But Oracles goes beyond that. There are Optimized Solutions available for your cloud infrastructure. This is a cookbook to build your HW cloud platform. For example, building your cloud infrastructure with blades and our network infrastructure will reduce complexity in your datacenter (Blades with switch network modules, splitter cables to reduce the amount of cables, TOR (Top Of the Rack) switches which are building the interface to your infrastructure environment. Reducing complexity even in the cabling will help you to manage your environment more efficient and with less risk. Of course, our engineered systems fit into the cloud perfectly too. Although they are considered as a PaaS themselves, having the database SW (for Exadata) and the application development environment (for Exalogic) already deployed on them, in general they are ideal systems to enable you building your own cloud and PaaS infrastructure. #2 Virtualization The next missing link in the cloud setup is virtualization. For me personally, it's one of the most hidden "secret", that oracle can provide you with a complete virtualization stack in terms of a hypervisor on both architectures: X86 and Sparc CPUs. There is Oracle VM for X86 and Oracle VM for Sparc available at no additional  license costs if your are running this virtualization stack on top of Oracle HW (and with Oracle Premier Support for HW). This completes the virtualization portfolio together with Solaris Zones introduced already with Solaris 10 a few years ago. Let me explain how Oracle VM for X86 works: Oracle VM for x86 consists of two main parts: - The Oracle VM Server: Oracle VM Server is installed on bare metal and it is the hypervisor which is able to run virtual machines. It has a very small footprint. The ISO-Image of Oracle VM Server is only 200MB large. It is very small but efficient. You can install a OVM-Server in less than 5 mins by booting the Server with the ISO-Image assigned and providing the necessary configuration parameters (like installing an Linux distribution). After the installation, the OVM-Server is ready to use. That's all. - The Oracle VM-Manager: OVM-Manager is the central management tool where you can control your OVM-Servers. OVM-Manager provides the graphical user interface, which is an Application Development Framework (ADF) application, with a familiar web-browser based interface, to manage Oracle VM Servers, virtual machines, and resources. The Oracle VM Manager has the following capabilities: Create virtual machines Create server pools Power on and off virtual machines Manage networks and storage Import virtual machines, ISO files, and templates Manage high availability of Oracle VM Servers, server pools, and virtual machines Perform live migration of virtual machines I want to highlight one of the goodies which you can use if you are running Oracle VM for X86: Preconfigured, downloadable Virtual Machine Templates form edelivery With these templates, you can download completely preconfigured Virtual Machines in your environment, boot them up, configure them at first time boot and use it. There are templates for almost all Oracle SW and Applications (like Fusion Middleware, Database, Siebel, etc.) available. #3) Cloud Management The management of your cloud infrastructure is key. This is a day-to-day job. Acquiring HW, installing a virtualization layer on top of it is done just at the beginning and if you want to expand your infrastructure. But managing your cloud, keeping it up and running, deploying new services, changing your chargeback model, etc, these are the daily jobs. These jobs must be simple, secure and easy to manage. The Enterprise Manager 12c Cloud provides this functionality from one management cockpit. Enterprise Manager 12c uses Oracle VM Manager to control OVM Serverpools. Once you registered your OVM-Managers in Enterprise Manager, then you are able to setup your cloud infrastructure and manage everything from Enterprise Manager. What you need to do in EM12c is: ">Register your OVM Manager in Enterprise ManagerAfter Registering your OVM Manager, all the functionality of Oracle VM for X86 is also available in Enterprise Manager. Enterprise Manager works as a "Manger" of the Manager. You can register as many OVM-Managers you want and control your complete virtualization environment Create Roles and Users for your Self Service Portal in Enterprise ManagerWith this step you allow users to logon on the Enterprise Manager Self Service Portal. Users can request Virtual Machines in this portal. Setup the Cloud InfrastructureSetup the Quotas for your self service users. How many VMs can they request? How much of your resources ( cpu, memory, storage, network, etc. etc.)? Which SW components (templates, assemblys) can your self service users request? In this step, you basically set up the complete cloud infrastructure. Setup ChargebackOnce your cloud is set up, you need to configure your chargeback mechanism. The Enterprise Manager collects the resources metrics, which are used in a very deep level. Almost all collected Metrics could be used in the chargeback module. You can define chargeback plans based on configurations (charge for the amount of cpu, memory, storage is assigned to a machine, or for a specific OS which is installed) or chargeback on resource consumption (% of cpu used, storage used, etc). Or you can also define a combination of configuration and consumption chargeback plans. The chargeback module is very flexible. Here is a overview of the workflow how to handle infrastructure cloud in EM: Summary As you can see, setting up an Infrastructure Cloud Service with Oracle VM for X86 and Enterprise Manager 12c is really simple. I personally configured a complete cloud environment with three X86 servers and a small JBOD san box in less than 3 hours. There is no magic in it, it is all straightforward. Of course, you have to have some experience with Oracle VM and Enterprise Manager. Experience in setting up Linux environments helps as well. I plan to publish a technical cookbook in the next few weeks. I hope you found this post useful and will see you again here on our blog. Any hints, comments are welcome!

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

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

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  • Possible SWITCH Optimization in DAX – #powerpivot #dax #tabular

    - by Marco Russo (SQLBI)
    In one of the Advanced DAX Workshop I taught this year, I had an interesting discussion about how to optimize a SWITCH statement (which could be frequently used checking a slicer, like in the Parameter Table pattern). Let’s start with the problem. What happen when you have such a statement? Sales :=     SWITCH (         VALUES ( Period[Period] ),         "Current", [Internet Total Sales],         "MTD", [MTD Sales],         "QTD", [QTD Sales],         "YTD", [YTD Sales],          BLANK ()     ) The SWITCH statement is in reality just syntax sugar for a nested IF statement. When you place such a measure in a pivot table, for every cell of the pivot table the IF options are evaluated. In order to optimize performance, the DAX engine usually does not compute cell-by-cell, but tries to compute the values in bulk-mode. However, if a measure contains an IF statement, every cell might have a different execution path, so the current implementation might evaluate all the possible IF branches in bulk-mode, so that for every cell the result from one of the branches will be already available in a pre-calculated dataset. The price for that could be high. If you consider the previous Sales measure, the YTD Sales measure could be evaluated for all the cells where it’s not required, and also when YTD is not selected at all in a Pivot Table. The actual optimization made by the DAX engine could be different in every build, and I expect newer builds of Tabular and Power Pivot to be better than older ones. However, we still don’t live in an ideal world, so it could be better trying to help the engine finding a better execution plan. One student (Niek de Wit) proposed this approach: Selection := IF (     HASONEVALUE ( Period[Period] ),     VALUES ( Period[Period] ) ) Sales := CALCULATE (     [Internet Total Sales],     FILTER (         VALUES ( 'Internet Sales'[Order Quantity] ),         'Internet Sales'[Order Quantity]             = IF (                 [Selection] = "Current",                 'Internet Sales'[Order Quantity],                 -1             )     ) )     + CALCULATE (         [MTD Sales],         FILTER (             VALUES ( 'Internet Sales'[Order Quantity] ),             'Internet Sales'[Order Quantity]                 = IF (                     [Selection] = "MTD",                     'Internet Sales'[Order Quantity],                     -1                 )         )     )     + CALCULATE (         [QTD Sales],         FILTER (             VALUES ( 'Internet Sales'[Order Quantity] ),             'Internet Sales'[Order Quantity]                 = IF (                     [Selection] = "QTD",                     'Internet Sales'[Order Quantity],                     -1                 )         )     )     + CALCULATE (         [YTD Sales],         FILTER (             VALUES ( 'Internet Sales'[Order Quantity] ),             'Internet Sales'[Order Quantity]                 = IF (                     [Selection] = "YTD",                     'Internet Sales'[Order Quantity],                     -1                 )         )     ) At first sight, you might think it’s impossible that this approach could be faster. However, if you examine with the profiler what happens, there is a different story. Every original IF’s execution branch is now a separate CALCULATE statement, which applies a filter that does not execute the required measure calculation if the result of the FILTER is empty. I used the ‘Internet Sales’[Order Quantity] column in this example just because in Adventure Works it has only one value (every row has 1): in the real world, you should use a column that has a very low number of distinct values, or use a column that has always the same value for every row (so it will be compressed very well!). Because the value –1 is never used in this column, the IF comparison in the filter discharge all the values iterated in the filter if the selection does not match with the desired value. I hope to have time in the future to write a longer article about this optimization technique, but in the meantime I’ve seen this optimization has been useful in many other implementations. Please write your feedback if you find scenarios (in both Power Pivot and Tabular) where you obtain performance improvements using this technique!

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  • Resolving data redundancy up front

    - by okeofs
    Introduction As all of us do when confronted with a problem, the resource of choice is to ‘Google it’. This is where the plot thickens. Recently I was asked to stage data from numerous databases which were to be loaded into a data warehouse. To make a long story short, I was looking for a manner in which to obtain the table names from each database, to ascertain potential overlap.   As the source data comes from a SQL database created from dumps of a third party product,  one could say that there were +/- 95 tables for each database.   Yes I know that first instinct is to use the system stored procedure “exec sp_msforeachdb 'select "?" AS db, * from [?].sys.tables'”. However, if one stops to think about this, it would be nice to have all the results in a temporary or disc based  table; which in itself , implies additional labour. This said,  I decided to ‘re-invent’ the wheel. The full code sample may be found at the bottom of this article.   Define a few temporary tables and variables   declare @SQL varchar(max); declare @databasename varchar(75) /* drop table ##rawdata3 drop table #rawdata1 drop table #rawdata11 */ -- A temp table to hold the names of my databases CREATE TABLE #rawdata1 (    database_name varchar(50) ,    database_size varchar(50),    remarks Varchar(50) )     --A temp table with the same database names as above, HOWEVER using an --Identity number (recNO) as a loop variable. --You will note below that I loop through until I reach 25 (see below) as at --that point the system databases, the reporting server database etc begin. --1- 24 are user databases. These are really what I was looking for. --Whilst NOT the best solution,it works and the code was meant as a quick --and dirty. CREATE TABLE #rawdata11 (    recNo int identity(1,1),    database_name varchar(50) ,    database_size varchar(50),    remarks Varchar(50) )   --My output table showing the database name and table name CREATE TABLE ##rawdata3 (    database_name varchar(75) ,    table_name varchar(75), )   Insert the database names into a temporary table I pull the database names using the system stored procedure sp_databases   INSERT INTO #rawdata1 EXEC sp_databases Go   Insert the results from #rawdata1 into a table containing a record number  #rawdata11 so that I can LOOP through the extract   INSERT into #rawdata11 select * from  #rawdata1   We now declare 3 more variables:  @kounter is used to keep track of our position within the loop. @databasename is used to keep track of the’ current ‘ database name being used in the current pass of the loop;  as inorder to obtain the tables for that database we  need to issue a ‘USE’ statement, an insert command and other related code parts. This is the challenging part. @sql is a varchar(max) variable used to contain the ‘USE’ statement PLUS the’ insert ‘ code statements. We now initalize @kounter to 1 .   declare @kounter int; declare @databasename varchar(75); declare @sql varchar(max); set @kounter = 1   The Loop The astute reader will remember that the temporary table #rawdata11 contains our  database names  and each ‘database row’ has a record number (recNo). I am only interested in record numbers under 25. I now set the value of the temporary variable @DatabaseName (see below) .Note that I used the row number as a part of the predicate. Now, knowing the database name, I can create dynamic T-SQL to be executed using the sp_sqlexec stored procedure (see the code in red below). Finally, after all the tables for that given database have been placed in temporary table ##rawdata3, I increment the counter and continue on. Note that I used a global temporary table to ensure that the result set persists after the termination of the run. At some stage, I plan to redo this part of the code, as global temporary tables are not really an ideal solution.    WHILE (@kounter < 25)  BEGIN  select @DatabaseName = database_name from #rawdata11 where recNo = @kounter  set @SQL = 'Use ' + @DatabaseName + ' Insert into ##rawdata3 ' + + ' SELECT table_catalog,Table_name FROM information_schema.tables' exec sp_sqlexec  @Sql  SET @kounter  = @kounter + 1  END   The full code extract   Here is the full code sample.   declare @SQL varchar(max); declare @databasename varchar(75) /* drop table ##rawdata3 drop table #rawdata1 drop table #rawdata11 */ CREATE TABLE #rawdata1 (    database_name varchar(50) ,    database_size varchar(50),    remarks Varchar(50) ) CREATE TABLE #rawdata11 (    recNo int identity(1,1),    database_name varchar(50) ,    database_size varchar(50),    remarks Varchar(50) ) CREATE TABLE ##rawdata3 (    database_name varchar(75) ,    table_name varchar(75), )   INSERT INTO #rawdata1 EXEC sp_databases go INSERT into #rawdata11 select * from  #rawdata1 declare @kounter int; declare @databasename varchar(75); declare @sql varchar(max); set @kounter = 1 WHILE (@kounter < 25)  BEGIN  select @databasename = database_name from #rawdata11 where recNo = @kounter  set @SQL = 'Use ' + @DatabaseName + ' Insert into ##rawdata3 ' + + ' SELECT table_catalog,Table_name FROM information_schema.tables' exec sp_sqlexec  @Sql  SET @kounter  = @kounter + 1  END    select * from ##rawdata3  where table_name like '%SalesOrderHeader%'

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  • Three ways to upload/post/convert iMovie to YouTube

    - by user44251
    For Mac users, iMovie is probably a convenient tool for making, editing their own home movies so as to upload to YouTube for sharing with more people. However, uploading iMovie files to YouTube can't be always a smooth run, I did notice many people complaining about it. This article is delivered for guiding those who are haunted by the nightmare by providing three common ways to upload iMovie files to YouTube. YouTube and iMovie YouTube is the most popular video sharing website for users to upload, share and view videos. It empowers anyone with an Internet connection the ability to upload video clips and share them with friends, family and the world. Users are invited to leave comments, pick favourites, send messages to each other and watch videos sorted into subjects and channels. YouTube accepts videos uploaded in most container formats, including WMV (Windows Media Video), 3GP (Cell Phones), AVI (Windows), MOV (Mac), MP4 (iPod/PSP), FLV (Adobe Flash), MKV (H.264). These include video codecs such as MP4, MPEG and WMV. iMovie is a common video editing software application comes with every Mac for users to edit their own home movies. It imports video footage to the Mac using either the Firewire interface on most MiniDV format digital video cameras, the USB port, or by importing the files from a hard drive where users can edit the video clips, add titles, and add music. Since 1999, eight versions of iMovie have been released by Apple, each with its own functions and characteristic, and each of them deal with videos in a way more or less different. But the most common formats handled with iMovie if specialty discarded as far as to my research are MOV, DV, HDV, MPEG-4. Three ways for successful upload iMovie files to YouTube Solution one and solution two suitable for those who are 100 certainty with their iMovie files which are fully compatible with YouTube. For smooth uploading, you are required to get a YouTube account first. Solution 1: Directly upload iMovie to YouTube Step 1: Launch iMovie, select the project you want to upload in YouTube. Step 2: Go to the file menu, click Share, select Export Movie Step 3: Specify the output file name and directory and then type the video type and video size. Solution 2: Post iMovie to YouTube straightly Step 1: Launch iMovie, choose the project you want to post in YouTube Step 2: From the Share menu, choose YouTube Step 3: In the pop-up YouTube windows, specify the name of your YouTube account, the password, choose the Category and fill in the description and tags of the project. Tick Make this movie more private on the bottom of the window, if possible, to limit those who can view the project. Click Next, and then click Publish. iMovie will automatically export and upload the movie to YouTube. Step 4: Click Tell a Friend to email friends and your family about your film. You are also allowed to copy the URL from Tell a Friend window and paste it into an email you created in your favourite email application if you like. Anyone you send to email to will be able to follow the URL directly to your movie. Note: Videos uploaded to YouTube are limited to ten minutes in length and a file size of 2GB. Solution 3: Upload to iMovie after conversion If neither of the above mentioned method works, there is still a third way to turn to. Sometimes, your iMovie files may not be recognized by YouTube due to the versions of iMovie (settings and functions may varies among versions), video itself (video format difference because of file extension, resolution, video size and length), compatibility (videos that are completely incompatible with YouTube). In this circumstance, the best and reliable method is to convert your iMovie files to YouTube accepted files, iMovie to YouTube converter will be inevitably the ideal choice. iMovie to YouTube converter is an elaborately designed tool for convert iMovie files to YouTube workable WMV, 3GP, AVI, MOV, MP4, FLV, MKV for smooth uploading with hard-to-believe conversion speed and second to none output quality. It can also convert between almost all popular popular file formats like AVI, WMV, MPG, MOV, VOB, DV, MP4, FLV, 3GP, RM, ASF, SWF, MP3, AAC, AC3, AIFF, AMR, WAV, WMA etc so as to put on various portable devices, import to video editing software or play on vast amount video players. iMovie to YouTube converter can also served as an excellent video editing tool to meet your specific program requirements. For example, you can cut your video files to a certain length, or split your video files to smaller ones and select the proper resolution suitable for demands of YouTube by Clip or Settings separately. Crop allows you to cut off unwanted black edges from your videos. Besides, you can also have a good command of the whole process or snapshot your favourite pictures from the preview window. More can be expected if you have a try.

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  • Strategy for using snapshots to back up Ubuntu Linux server?

    - by MountainX
    I need some backup advice for my home file server. Here are the mount points, volume groups, logical volumes and used/total space of all the volumes on my Ubuntu 8.10 home file server. / vgA/lvRoot [7.5G/50G] /tmp vgB/lvTmp [195M/30G] /var vgB/lvVar [780M/30G] swap vgB/lvSwap [16.00 GB] /media1 vgC/lvMedia1 [400G/975G] /media2 vgC/lvMedia2 [75G/295G] /boot partition (no volume group) [95M/200M] /video partition (no volume group) [450G/950G] /backups vgD/lvBackupTarget [800G/925G] /home vgE/lvHome [85G/200G] I have just added a 2.0 TB external USB drive that I would like to use to backup everything. (It will be a close fit to get it all on one 2.0 TB drive. I actually have a 2nd external USB drive if needed.) I'd like to backup "/", var, /media1, media2 and /home. I'll deal with /boot and /video separately since they are not logical volumes. For all the logical volumes I'm anticipating taking snapshots and then copying those snapshots to the 2.0 TB external USB drive. I have never done a task like that before. If I do that, I could use the tutorial I found here: http://www.howtoforge.com/linux_lvm_snapshots My questions are: What is the best overall strategy? Is it LVM snapshots, as I'm assuming? How should I prepare, subdivide and mount the 2.0 TB external USB drive? 2.a. Should I create one or more regular partitions or should I create a physical volume with one or more logical volumes? 2.b. Would it be advisable to extactly mirror the source pv/lv layout on the external drive, and if so, is this a good strategy? What's the best way to get the snapshots onto the external drive? dd? Even though this is a strategy question, feedback with actual commands is appreciated. I need step-by-step cookbook-style help because I don't do much server admin work. (Background: This is a home file server that I have rarely had to touch in about 2 years. It has done its job without much intervention. The really old PC that I used to back everything up recently failed, so I'm replacing that with the external USB drive(s) and I'd like to upgrade my backup strategy at the same time. Previously, I just copied stuff from /backups over to the other computer and that would not have made things very easy in a real restore situation. The /backups mount point contains backup copies of "most" of the important data on a file by file basis, but it does not contain copies of /boot, etc. BTW, the actual internal HDD that holds /backups is separate from the other storage devices.) EDIT: I'll propose a strategy... The idea came from a comment here: LVM mirroring VS RAID1 "LVM mirrors are for replication of a logical volume to a different physical volume. It's essentially meant to "move the data to a different disk". The mirror is then broken..." That would fit my requirements well. Here is an ideal situation: establish the LV mirror on the external drive break the link with the mirror create a (persistent) snapshot on the mirror after a week, resync the mirror with the original source and update the mirror break the link and create another snapshot on the mirror. Obviously, the mirror will be like a weekly full backup. And the snapshots on the mirror will represent earlier points in time. If this would work and if it would be time efficient, it would give a nice full & differential type backup on the external drive based on LVM. I have not heard of a strategy like this before. Will it work? Could it be scripted? Thoughts? EDIT 2: Creating Portable DiskSafes With LoopbackFS And LVM Snapshots This article seems intriguing: http://www.howtoforge.com/creating-portable-disksafes-with-loopbackfs-and-lvm-snapshots Unfortunately, I don't understand exactly how to map those ideas to the strategy I'm proposing above. I'm going to ask this last bit as a separate question. I will leave my original question in place because I still desire feedback on the overall best strategy. At this moment I'm assuming it is LVM mirroring in the style of "Creating Portable DiskSafes with LVM Snapshots" but that might be wrong.

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  • Three ways to upload/post/convert iMovie to YouTube [closed]

    - by alexyu2010
    For Mac users, iMovie is probably a convenient tool for making, editing their own home movies so as to upload to YouTube for sharing with more people. However, uploading iMovie files to YouTube can't be always a smooth run, I did notice many people complaining about it. This article is delivered for guiding those who are haunted by the nightmare by providing three common ways to upload iMovie files to YouTube. YouTube and iMovie YouTube is the most popular video sharing website for users to upload, share and view videos. It empowers anyone with an Internet connection the ability to upload video clips and share them with friends, family and the world. Users are invited to leave comments, pick favourites, send messages to each other and watch videos sorted into subjects and channels. YouTube accepts videos uploaded in most container formats, including WMV (Windows Media Video), 3GP (Cell Phones), AVI (Windows), MOV (Mac), MP4 (iPod/PSP), FLV (Adobe Flash), MKV (H.264). These include video codecs such as MP4, MPEG and WMV. iMovie is a common video editing software application comes with every Mac for users to edit their own home movies. It imports video footage to the Mac using either the Firewire interface on most MiniDV format digital video cameras, the USB port, or by importing the files from a hard drive where users can edit the video clips, add titles, and add music. Since 1999, eight versions of iMovie have been released by Apple, each with its own functions and characteristic, and each of them deal with videos in a way more or less different. But the most common formats handled with iMovie if specialty discarded as far as to my research are MOV, DV, HDV, MPEG-4. Three ways for successful upload iMovie files to YouTube Solution one and solution two suitable for those who are 100 certainty with their iMovie files which are fully compatible with YouTube. For smooth uploading, you are required to get a YouTube account first. Solution 1: Directly upload iMovie to YouTube Step 1: Launch iMovie, select the project you want to upload in YouTube. Step 2: Go to the file menu, click Share, select Export Movie Step 3: Specify the output file name and directory and then type the video type and video size. Solution 2: Post iMovie to YouTube straightly Step 1: Launch iMovie, choose the project you want to post in YouTube Step 2: From the Share menu, choose YouTube Step 3: In the pop-up YouTube windows, specify the name of your YouTube account, the password, choose the Category and fill in the description and tags of the project. Tick Make this movie more private on the bottom of the window, if possible, to limit those who can view the project. Click Next, and then click Publish. iMovie will automatically export and upload the movie to YouTube. Step 4: Click Tell a Friend to email friends and your family about your film. You are also allowed to copy the URL from Tell a Friend window and paste it into an email you created in your favourite email application if you like. Anyone you send to email to will be able to follow the URL directly to your movie. Note: Videos uploaded to YouTube are limited to ten minutes in length and a file size of 2GB. Solution 3: Upload to iMovie after conversion If neither of the above mentioned method works, there is still a third way to turn to. Sometimes, your iMovie files may not be recognized by YouTube due to the versions of iMovie (settings and functions may varies among versions), video itself (video format difference because of file extension, resolution, video size and length), compatibility (videos that are completely incompatible with YouTube). In this circumstance, the best and reliable method is to convert your iMovie files to YouTube accepted files, iMovie to YouTube converter will be inevitably the ideal choice. iMovie to YouTube converter is an elaborately designed tool for convert iMovie files to YouTube workable WMV, 3GP, AVI, MOV, MP4, FLV, MKV for smooth uploading with hard-to-believe conversion speed and second to none output quality. It can also convert between almost all popular popular file formats like AVI, WMV, MPG, MOV, VOB, DV, MP4, FLV, 3GP, RM, ASF, SWF, MP3, AAC, AC3, AIFF, AMR, WAV, WMA etc so as to put on various portable devices, import to video editing software or play on vast amount video players. iMovie to YouTube converter can also served as an excellent video editing tool to meet your specific program requirements. For example, you can cut your video files to a certain length, or split your video files to smaller ones and select the proper resolution suitable for demands of YouTube by Clip or Settings separately. Crop allows you to cut off unwanted black edges from your videos. Besides, you can also have a good command of the whole process or snapshot your favourite pictures from the preview window. More can be expected if you have a try.

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  • Creating Persistent Drive Labels With UDEV Using /dev/disk/by-path

    - by Matt
    I have a new BackBlaze Pod (BackBlaze Pod 2.0). It has 45 3TB drives and they when I first set it up they were labeled /dev/sda through /dev/sdz and /dev/sdaa through /dev/sdas. I used mdadm to setup three really big 15 drive RAID6 arrays. However, since first setup a few weeks ago I had a couple of the hard drives fail on me. I've replaced them but now the arrays are complaining because they can't find the missing drives. When I list the the disks... ls -l /dev/sd* I see that /dev/sda /dev/sdf /dev/sdk /dev/sdp no longer appear and now there are 4 new ones... /dev/sdau /dev/sdav /dev/sdaw /dev/sdax I also just found that I can do this... ls -l /dev/disk/by-path/ total 0 lrwxrwxrwx 1 root root 10 Sep 19 18:08 pci-0000:02:04.0-scsi-0:0:0:0 -> ../../sdau lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-0:1:0:0 -> ../../sdb lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-0:2:0:0 -> ../../sdc lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-0:3:0:0 -> ../../sdd lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-0:4:0:0 -> ../../sde lrwxrwxrwx 1 root root 10 Sep 19 18:08 pci-0000:02:04.0-scsi-2:0:0:0 -> ../../sdae lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-2:1:0:0 -> ../../sdg lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-2:2:0:0 -> ../../sdh lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-2:3:0:0 -> ../../sdi lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-2:4:0:0 -> ../../sdj lrwxrwxrwx 1 root root 10 Sep 19 18:08 pci-0000:02:04.0-scsi-3:0:0:0 -> ../../sdav lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-3:1:0:0 -> ../../sdl lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-3:2:0:0 -> ../../sdm lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-3:3:0:0 -> ../../sdn lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:02:04.0-scsi-3:4:0:0 -> ../../sdo lrwxrwxrwx 1 root root 10 Sep 19 18:08 pci-0000:04:04.0-scsi-0:0:0:0 -> ../../sdax lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:04:04.0-scsi-0:1:0:0 -> ../../sdq lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:04:04.0-scsi-0:2:0:0 -> ../../sdr lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:04:04.0-scsi-0:3:0:0 -> ../../sds lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:04:04.0-scsi-0:4:0:0 -> ../../sdt lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:04:04.0-scsi-2:0:0:0 -> ../../sdu lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:04:04.0-scsi-2:1:0:0 -> ../../sdv lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:04:04.0-scsi-2:2:0:0 -> ../../sdw lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:04:04.0-scsi-2:3:0:0 -> ../../sdx lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:04:04.0-scsi-2:4:0:0 -> ../../sdy lrwxrwxrwx 1 root root 9 Sep 19 18:08 pci-0000:04:04.0-scsi-3:0:0:0 -> ../../sdz I didn't list them all....you can see the problem above. They're sorted by scsi id here but sda is missing...replaced by sdau...etc... So obviously the arrays are complaining. Is it possible to get Linux to reread the drive labels in the correct order or am I screwed? My initial design with 15 drive arrays is not ideal. With 3TB drives the rebuild times were taking 3 or 4 days....maybe more. I'm scrapping the whole design and I think I am going to go with 6 x 7 RAID5 disk arrays and 3 hot spares to make the arrays a bit easier to manage and shorten the rebuild times. But I'd like to clean up the drive labels so they aren't out of order. I haven't figured out how to do this yet. Does anyone know how to get this straightened out? Thanks, Matt

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  • Weighted round robins via TTL - possible?

    - by Joe Hopfgartner
    I currently use DNS round robin for load balancing, which works great. The records look like this (I have a ttl of 120 seconds) ;; ANSWER SECTION: orion.2x.to. 116 IN A 80.237.201.41 orion.2x.to. 116 IN A 87.230.54.12 orion.2x.to. 116 IN A 87.230.100.10 orion.2x.to. 116 IN A 87.230.51.65 I learned that not every ISP / device treats such a response the same way. For example some DNS servers rotate the addresses randomly or always cycle them through. Some just propagate the first entry, others try to determine which is best (regionally near) by looking at the ip address. However if the userbase is big enough (spreads over multiple ISPs etc) it balances pretty well. The discrepancies from highest to lowest loaded server hardly every exceeds 15%. However now I have the problem that I am introducing more servers into the systems, that not all have the same capacities. I currently only have 1gbps servers, but I want to work with 100mbit and also 10gbps servers too. So what I want is I want to introduce a server with 10 GBps with a weight of 100, a 1 gbps server with a weight of 10 and a 100 mbit server with a weight of 1. I used to add servers twice to bring more traffic to them (which worked nice. the bandwidth doubled almost.) But adding a 10gbit server 100 times to DNS is a bit rediculous. So I thought about using the TTL. If I give server A 240 seconds ttl and server B only 120 seconds (which is about about the minimum to use for round robin, as a lot of dns servers set to 120 if a lower ttl is specified.. so i have heard) I think something like this should occour in an ideal scenario: first 120 seconds 50% of requests get server A -> keep it for 240 seconds. 50% of requests get server B -> keep it for 120 seconds second 120 seconds 50% of requests still have server A cached -> keep it for another 120 seconds. 25% of requests get server A -> keep it for 240 seconds 25% of requests get server B -> keep it for 120 seconds third 120 seconds 25% will get server A (from the 50% of Server A that now expired) -> cache 240 sec 25% will get server B (from the 50% of Server A that now expired) -> cache 120 sec 25% will have server A cached for another 120 seconds 12.5% will get server B (from the 25% of server B that now expired) -> cache 120sec 12.5% will get server A (from the 25% of server B that now expired) -> cache 240 sec fourth 120 seconds 25% will have server A cached -> cache for another 120 secs 12.5% will get server A (from the 25% of b that now expired) -> cache 240 secs 12.5% will get server B (from the 25% of b that now expired) -> cache 120 secs 12.5% will get server A (from the 25% of a that now expired) -> cache 240 secs 12.5% will get server B (from the 25% of a that now expired) -> cache 120 secs 6.25% will get server A (from the 12.5% of b that now expired) -> cache 240 secs 6.25% will get server B (from the 12.5% of b that now expired) -> cache 120 secs 12.5% will have server A cached -> cache another 120 secs ... i think i lost something at this point but i think you get the idea.... As you can see this gets pretty complicated to predict and it will for sure not work out like this in practice. But it should definitely have an effect on the distribution! I know that weighted round robin exists and is just controlled by the root server. It just cycles through dns records when responding and returns dns records with a set propability that corresponds to the weighting. My DNS server does not support this, and my requirements are not that precise. If it doesnt weight perfectly its okay, but it should go into the right direction. I think using the TTL field could be a more elegant and easier solution - and it deosnt require a dns server that controls this dynamically, which saves resources - which is in my opinion the whole point of dns load balancing vs hardware load balancers. My question now is... are there any best prectices / methos / rules of thumb to weight round robin distribution using the TTL attribute of DNS records? Edit: The system is a forward proxy server system. The amount of Bandwidth (not requests) exceeds what one single server with ethernet can handle. So I need a balancing solution that distributes the bandwidth to several servers. Are there any alternative methods than using DNS? Of course I can use a load balancer with fibre channel etc, but the costs are rediciulous and it also increases only the width of the bottleneck and does not eliminate it. The only thing i can think of are anycast (is it anycast or multicast?) ip addresses, but I don't have the means to set up such a system.

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  • OpenVPN Configuration - Windows 7 client & debian server

    - by Guillaume
    I recently formatted my Windows 7 computer and lost my client's config files for OpenVPN. I recovered the certificates and default config that were left on the server but I haven't managed to make the whole thing work again. I assume the server's config and routing table are OK because it was working before (although quite some time ago). Would any of you experts be able to help? server.conf # Serveur TCP/666 mode server proto udp port 666 dev tun # Cles et certificats ca ca.crt cert server.crt key server.key dh dh1024.pem tls-auth ta.key 0 cipher AES-256-CBC # Reseau server 10.8.0.0 255.255.255.0 #push "redirect-gateway def1 bypass-dhcp" push "dhcp-option DNS 208.67.222.222" push "dhcp-option DNS 208.67.220.220" push "redirect-gateway def1" keepalive 10 120 # Securite user nobody group nogroup chroot /etc/openvpn/jail persist-key persist-tun comp-lzo # Log verb 3 mute 20 status openvpn-status.log log-append /var/log/openvpn.log client.conf # Client client dev tun proto udp remote *my server's ip address*:666 cipher AES-256-CBC # Cles ca ca.crt cert client1.crt key client1.key tls-auth ta.key 1 # Securite nobind persist-key persist-tun comp-lzo verb 3 Routing table on debian server when OpenVPN server is running: Destination Gateway Genmask Indic Metric Ref Use Iface 10.8.0.2 * 255.255.255.255 UH 0 0 0 tun0 10.8.0.0 10.8.0.2 255.255.255.0 UG 0 0 0 tun0 my server's ip * 255.255.255.0 U 0 0 0 eth0 default 72815.trg.dedic 0.0.0.0 UG 0 0 0 eth0 Routing table on Windows 7 client (OpenVPN not working) =========================================================================== Interface List 19...00 f0 8a 1b 6e 5c ......TAP-Win32 Adapter V9 12...90 2e 34 33 84 7b ......Atheros AR8151 PCI-E Gigabit Ethernet Controller ( NDIS 6.20) 1...........................Software Loopback Interface 1 12...00 00 00 00 00 00 00 e0 Microsoft ISATAP Adapter 13...00 00 00 00 00 00 00 e0 Teredo Tunneling Pseudo-Interface 16...00 00 00 00 00 00 00 e0 Microsoft ISATAP Adapter #2 =========================================================================== IPv4 Route Table =========================================================================== Active Routes: Network Destination Netmask Gateway Interface Metric 0.0.0.0 0.0.0.0 192.168.1.1 192.168.1.11 20 127.0.0.0 255.0.0.0 On-link 127.0.0.1 306 127.0.0.1 255.255.255.255 On-link 127.0.0.1 306 127.255.255.255 255.255.255.255 On-link 127.0.0.1 306 192.168.1.0 255.255.255.0 On-link 192.168.1.11 276 192.168.1.11 255.255.255.255 On-link 192.168.1.11 276 192.168.1.255 255.255.255.255 On-link 192.168.1.11 276 224.0.0.0 240.0.0.0 On-link 127.0.0.1 306 224.0.0.0 240.0.0.0 On-link 192.168.1.11 276 255.255.255.255 255.255.255.255 On-link 127.0.0.1 306 255.255.255.255 255.255.255.255 On-link 192.168.1.11 276 =========================================================================== Persistent Routes: None IPv6 Route Table =========================================================================== Active Routes: [...] =========================================================================== Persistent Routes: None And when the link is established between my client and the server: The server's routing table stays the same. The client's becomes: =========================================================================== Interface List 19...00 f0 8a 1b 6e 5c ......TAP-Win32 Adapter V9 12...90 2e 34 33 84 7b ......Atheros AR8151 PCI-E Gigabit Ethernet Controller ( NDIS 6.20) 1...........................Software Loopback Interface 1 12...00 00 00 00 00 00 00 e0 Microsoft ISATAP Adapter 13...00 00 00 00 00 00 00 e0 Teredo Tunneling Pseudo-Interface 16...00 00 00 00 00 00 00 e0 Microsoft ISATAP Adapter #2 =========================================================================== IPv4 Route Table =========================================================================== Active Routes: Network Destination Netmask Gateway Interface Metric 0.0.0.0 0.0.0.0 192.168.1.1 192.168.1.11 20 0.0.0.0 128.0.0.0 10.8.0.5 10.8.0.6 30 10.8.0.1 255.255.255.255 10.8.0.5 10.8.0.6 30 10.8.0.4 255.255.255.252 On-link 10.8.0.6 286 10.8.0.6 255.255.255.255 On-link 10.8.0.6 286 10.8.0.7 255.255.255.255 On-link 10.8.0.6 286 my server's ip 255.255.255.255 192.168.1.1 192.168.1.11 20 127.0.0.0 255.0.0.0 On-link 127.0.0.1 306 127.0.0.1 255.255.255.255 On-link 127.0.0.1 306 127.255.255.255 255.255.255.255 On-link 127.0.0.1 306 128.0.0.0 128.0.0.0 10.8.0.5 10.8.0.6 30 192.168.1.0 255.255.255.0 On-link 192.168.1.11 276 192.168.1.11 255.255.255.255 On-link 192.168.1.11 276 192.168.1.255 255.255.255.255 On-link 192.168.1.11 276 224.0.0.0 240.0.0.0 On-link 127.0.0.1 306 224.0.0.0 240.0.0.0 On-link 192.168.1.11 276 224.0.0.0 240.0.0.0 On-link 10.8.0.6 286 255.255.255.255 255.255.255.255 On-link 127.0.0.1 306 255.255.255.255 255.255.255.255 On-link 192.168.1.11 276 255.255.255.255 255.255.255.255 On-link 10.8.0.6 286 =========================================================================== Persistent Routes: None What's working: Server and client do connect to each other, SSL certificates are OK. The client gets an IP (10.8.0.6) from the server OpenVPN client is started as an administrator. But: I cannot ping the other one on either side. 'Gateway' value is empty on client's side (in the adapter's "status" window). Client has got no internet access when the link is up. Ideal configuration: I only want the client to be able to use the server's Internet access and access its resources (MySQL server in particular). I do not need or want the server to access the client's local network. The client needs to be able to access it's local network, although all Internet traffic should be redirected to the VPN link. I spent a considerable amount of time on this but it's still not working, any help would be much appreciated. Thanks :)

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  • Silverlight 4 + MVVM + KeyDown event

    - by jturn
    I'm trying to build a sample game in Silverlight 4 using the MVVM design pattern to broaden my knowledge. I'm using Laurent Bugnion's MvvmLight toolkit as well (found here: http://mvvmlight.codeplex.com/ ). All I want to do right now is move a shape around within a Canvas by pressing specific keys. My solution contains a Player.xaml (just a rectangle; this will be moved around) and MainPage.xaml (the Canvas and an instance of the Player control). To my understanding, Silverlight doesn't support tunneling routed events, only bubbling. My big problem is that Player.xaml never recognizes the KeyDown event. It's always intercepted by MainPage.xaml first and it never reaches any child controls because it bubbles upward. I'd prefer that the logic to move the Player be in the PlayerViewModel class, but I don't think the Player can know about any KeyDown events firing without me explicitly passing them on down from the MainPage. I ended up adding the handler logic to the MainPageViewModel class. Now my problem is that the MainPageViewModel has no knowledge of Player.xaml so it cannot move this object when handling KeyDown events. I guess this is expected, as ViewModels should not have any knowledge of their associated Views. In not so many words...is there a way this Player user control within my MainPage.xaml can directly accept and handle KeyDown events? If not, what's the ideal method for my MainPageViewModel to communicate with its View's child controls? I'm trying to keep code out of the code-behind files as much as possible. Seems like it's best to put logic in the ViewModels for ease of testing and to decouple UI from logic. (MainPage.xaml) <UserControl x:Class="MvvmSampleGame.MainPage" 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" xmlns:game="clr-namespace:MvvmSampleGame" xmlns:i="clr-namespace:System.Windows.Interactivity;assembly=System.Windows.Interactivity" xmlns:cmd="clr-namespace:GalaSoft.MvvmLight.Command;assembly=GalaSoft.MvvmLight.Extras.SL4" mc:Ignorable="d" Height="300" Width="300" DataContext="{Binding Main, Source={StaticResource Locator}}"> <i:Interaction.Triggers> <i:EventTrigger EventName="KeyDown"> <cmd:EventToCommand Command="{Binding KeyPressCommand}" PassEventArgsToCommand="True" /> </i:EventTrigger> </i:Interaction.Triggers> <Canvas x:Name="LayoutRoot"> <game:Player x:Name="Player1"></game:Player> </Canvas> (MainViewModel.cs) public MainViewModel() { KeyPressCommand = new RelayCommand<KeyEventArgs>(KeyPressed); } public RelayCommand<KeyEventArgs> KeyPressCommand { get; private set; } private void KeyPressed(KeyEventArgs e) { if (e.Key == Key.Up || e.Key == Key.W) { // move player up } else if (e.Key == Key.Left || e.Key == Key.A) { // move player left } else if (e.Key == Key.Down || e.Key == Key.S) { // move player down } else if (e.Key == Key.Right || e.Key == Key.D) { // move player right } } Thanks in advance, Jeremy

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  • Animating the offset of the scrollView in a UICollectionView/UITableView causes prematurely disappearing cells

    - by radutzan
    We have a UICollectionView with a custom layout very similar to UITableView (it scrolls vertically). The UICollectionView displays only 3 cells simultaneously, with one of them being the currently active cell: [ 1 ] [*2*] [ 3 ] (The active cell here is #2.) The cells are roughly 280 points high, so only the active cell is fully visible on the screen. The user doesn't directly scroll the view to navigate, instead, she swipes the active cell horizontally to advance to the next cell. We then do some fancy animations and scroll the UICollectionView so the next cell is in the "active" position, thus making it the active one, moving the old one away and bringing up the next cell in the queue: [ 2 ] [*3*] [ 4 ] The problem here is setting the UICollectionView's offset. We currently set it in a UIView animation block (self.collectionView.contentOffset = targetOffset;) along with three other animating properties, which mostly works great, but causes the first cell (the previously active one, in the latter case, #2) to vanish as soon as the animation starts running, even before the delay interval completes. This is definitely not ideal. I've thought of some solutions, but can't figure out the best one: Absurdly enlarge the UICollectionView's frame to fit five cells instead of three, thus forcing it to keep the cells in memory even if they are offscreen. I've tried this and it works, but it sounds like an awfully dirty hack. Take a snapshot of the rendered content of the vanishing cell, put it in a UIImageView, add the UIImageView as a subview of the scrollView just before the cell goes away in the exact same position of the old cell, removing it once the animation ends. Sounds less sucky than the previous option (memory-wise, at least), but still kinda hacky. I also don't know the best way to accomplish this, please point me in the right direction. Switch to UIScrollView's setContentOffset:animated:. We actually used to have this, and it fixed the disappearing cell issue, but running this in parallel with the other UIView animations apparently competes for the attention of the main thread, thus creating a terribly choppy animation on single-core devices (iPhone 3GS/4). It also doesn't allow us to change the duration or easing of the animation, so it feels out of sync with the rest. Still an option if we can find a way to make it work in harmony with the UIView block animations. Switch to UICollectionView's scrollToItemAtIndexPath:atScrollPosition:animated:. Haven't tried this, but it has a big downside: it only takes 3 possible constants (that apply to this case, at least) for the target scroll position: UICollectionViewScrollPositionTop, UICollectionViewScrollPositionCenteredVertically and UICollectionViewScrollPositionBottom. The active cell could vary its height, but it always has to be 35 points from the top of the window, and these options don't provide enough control to accomplish the design. It could also potentially be just as problematic as 3.1. Still an option because there might be a way to go around the scroll position thing that I don't know of, and it might not have the same issue with the main thread, which seems unlikely. Any help will be greatly appreciated. Please ask if you need clarification. Thanks a lot!

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  • NSURLSession and amazon S3 uploads

    - by George Green
    I have an app which is currently uploading images to amazon S3. I have been trying to switch it from using NSURLConnection to NSURLSession so that the uploads can continue while the app is in the background! I seem to be hitting a bit of an issue. The NSURLRequest is created and passed to the NSURLSession but amazon sends back a 403 - forbidden response, if I pass the same request to a NSURLConnection it uploads the file perfectly. Here is the code that creates the response: NSString *requestURLString = [NSString stringWithFormat:@"http://%@.%@/%@/%@", BUCKET_NAME, AWS_HOST, DIRECTORY_NAME, filename]; NSURL *requestURL = [NSURL URLWithString:requestURLString]; NSMutableURLRequest *request = [NSMutableURLRequest requestWithURL:requestURL cachePolicy:NSURLRequestReloadIgnoringLocalAndRemoteCacheData timeoutInterval:60.0]; // Configure request [request setHTTPMethod:@"PUT"]; [request setValue:[NSString stringWithFormat:@"%@.%@", BUCKET_NAME, AWS_HOST] forHTTPHeaderField:@"Host"]; [request setValue:[self formattedDateString] forHTTPHeaderField:@"Date"]; [request setValue:@"public-read" forHTTPHeaderField:@"x-amz-acl"]; [request setHTTPBody:imageData]; And then this signs the response (I think this came from another SO answer): NSString *contentMd5 = [request valueForHTTPHeaderField:@"Content-MD5"]; NSString *contentType = [request valueForHTTPHeaderField:@"Content-Type"]; NSString *timestamp = [request valueForHTTPHeaderField:@"Date"]; if (nil == contentMd5) contentMd5 = @""; if (nil == contentType) contentType = @""; NSMutableString *canonicalizedAmzHeaders = [NSMutableString string]; NSArray *sortedHeaders = [[[request allHTTPHeaderFields] allKeys] sortedArrayUsingSelector:@selector(caseInsensitiveCompare:)]; for (id key in sortedHeaders) { NSString *keyName = [(NSString *)key lowercaseString]; if ([keyName hasPrefix:@"x-amz-"]){ [canonicalizedAmzHeaders appendFormat:@"%@:%@\n", keyName, [request valueForHTTPHeaderField:(NSString *)key]]; } } NSString *bucket = @""; NSString *path = request.URL.path; NSString *query = request.URL.query; NSString *host = [request valueForHTTPHeaderField:@"Host"]; if (![host isEqualToString:@"s3.amazonaws.com"]) { bucket = [host substringToIndex:[host rangeOfString:@".s3.amazonaws.com"].location]; } NSString* canonicalizedResource; if (nil == path || path.length < 1) { if ( nil == bucket || bucket.length < 1 ) { canonicalizedResource = @"/"; } else { canonicalizedResource = [NSString stringWithFormat:@"/%@/", bucket]; } } else { canonicalizedResource = [NSString stringWithFormat:@"/%@%@", bucket, path]; } if (query != nil && [query length] > 0) { canonicalizedResource = [canonicalizedResource stringByAppendingFormat:@"?%@", query]; } NSString* stringToSign = [NSString stringWithFormat:@"%@\n%@\n%@\n%@\n%@%@", [request HTTPMethod], contentMd5, contentType, timestamp, canonicalizedAmzHeaders, canonicalizedResource]; NSString *signature = [self signatureForString:stringToSign]; [request setValue:[NSString stringWithFormat:@"AWS %@:%@", self.S3AccessKey, signature] forHTTPHeaderField:@"Authorization"]; Then if I use this line of code: [NSURLConnection connectionWithRequest:request delegate:self]; It works and uploads the file, but if I use: NSURLSessionUploadTask *task = [self.session uploadTaskWithRequest:request fromFile:[NSURL fileURLWithPath:filePath]]; [task resume]; I get the forbidden error..!? Has anyone tried uploading to S3 with this and hit similar issues? I wonder if it is to do with the way the session pauses and resumes uploads, or it is doing something funny to the request..? One possible solution would be to upload the file to an interim server that I control and have that forward it to S3 when it is complete... but this is clearly not an ideal solution! Any help is much appreciated!! Thanks!

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  • IE6 iframe anchor links moves iframe up

    - by WastedSpace
    Hi, Having a real head scratching moment... I have a site where there is a footer div that always sits at the bottom of the screen (26px high), and above that I have an iFrame which sizes to 100% of the remaining height. This works well. Even clicking on anchor links inside the iframe works as it should in all browsers (apart from IE6). Unfortunately I still have to support IE6. What is happening in IE6 is that the footer jumps up the page with the iframe still above it when I click on an anchor link. The top part of the iframe is cut off. Even the iframe's scroll bars disappear under the top of the browser. I have created some screen shots to show you what I mean. I have blurred out the actual data for now. How it should look (and does look) in other browsers: http://img100.imageshack.us/i/screen1om.jpg/ How it looks in IE6 before clicking on an anchor link: http://img532.imageshack.us/i/screen2e.jpg/ (I had to make the iframe's height 95%, because if I set it to 100% height weirdly it wouldn't show anything...) How it looks in IE6 after clicking on an anchor link: http://img214.imageshack.us/i/screen3g.jpg/ It's hard for me to show the fool code I am using, as there are lots of other things going on (of which I'm confident doesn't affect the layout), so will try to summarise: The html code (simplified): <div id="ifra"><iframe src="home.php" frameborder="0" name="content_pane" id="content_pane" marginheight="10" marginwidth="10"></iframe></div> <h1 class="toolbar"><a id="footerlink">Site Name</a></h1> The CSS (simplified): html, body { overflow: hidden; } html, body, iframe { height: 100%; } body { padding: 0; margin: 0; } #ifra, iframe { position: absolute; width: 100%; left: 0; } #ifra { top: 0px; bottom: 26px; } iframe { border: 0 none; } .toolbar { height: 26px; background-color: #C2C7C9; position: fixed; bottom: 0; width: 100%; background-image: url(bg.png); background-repeat: repeat-x; background-position: left top; } IE8 specific CSS: #ifra, .toolbar { position: fixed; } IE7 specific CSS: html { padding: 0px; } #ifra, iframe { position: absolute; } #ifra { top: 0px; bottom: 26px; } * html body { padding /**/: 100px 0 50px 0; overflow-y /**/: hidden; } IE6 specific CSS: .toolbar { position: fixed; } * html { overflow-y: hidden; } * html body { overflow-y: auto; height: 100%; } * html .toolbar { position: absolute; } iframe { height: 95%; } #ifra { height: 100%; } I know it's not ideal not seeing the full code, but just wondering if there is anything jumping out at anyone from these lines of code? By the way I did consider dropping the div surrounding the iframe, but for some reason the scroll bars would disappear under the footer in all browsers... Thanks for looking! Ali.

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  • Running multiple image manipulations in parallel causing OutOfMemory exception

    - by Tom
    I am working on a site where I need to be able to split and image around 4000x6000 into 4 parts (amongst many other tasks) and I need this to be as quick as possible for multiple users. My current code for doing this is var bitmaps = new RenderTargetBitmap[elements.Length]; using (var stream = blobService.Stream(key)) { BitmapImage bi = new BitmapImage(); bi.BeginInit(); bi.StreamSource = stream; bi.EndInit(); for (var i = 0; i < elements.Length; i++) { var element = elements[i]; TransformGroup transformGroup = new TransformGroup(); TranslateTransform translateTransform = new TranslateTransform(); translateTransform.X = -element.Left; translateTransform.Y = -element.Top; transformGroup.Children.Add(translateTransform); DrawingVisual vis = new DrawingVisual(); DrawingContext cont = vis.RenderOpen(); cont.PushTransform(transformGroup); cont.DrawImage(bi, new Rect(new Size(bi.PixelWidth, bi.PixelHeight))); cont.Close(); RenderTargetBitmap rtb = new RenderTargetBitmap(element.Width, element.Height, 96d, 96d, PixelFormats.Default); rtb.Render(vis); bitmaps[i] = rtb; } } for (var i = 0; i < bitmaps.Length; i++) { using (MemoryStream ms = new MemoryStream()) { PngBitmapEncoder encoder = new PngBitmapEncoder(); encoder.Frames.Add(BitmapFrame.Create(bitmaps[i])); encoder.Save(ms); var regionKey = WebPath.Variant(key, elements[i].Id); saveBlobService.Save("image/png", regionKey, ms); } } I am running multiple threads which take jobs off a queue. I am finding that if this part of code is hit by 4 threads at once I get an OutOfMemory exception. I can stop this happening by wrapping all the code above in a lock(obj) but this isn't ideal. I have tried wrapping just the first using block (where the file is read from disk and split) but I still get the out of memory exceptions (this part of the code executes quite quickly). I this normal considering the amount of memory this should be taking up? Are there any optimisations I could make? Can I increase the memory available? UPDATE: My new code as per Moozhe's help public static void GenerateRegions(this IBlobService blobService, string key, Element[] elements) { using (var stream = blobService.Stream(key)) { foreach (var element in elements) { stream.Position = 0; BitmapImage bi = new BitmapImage(); bi.BeginInit(); bi.SourceRect = new Int32Rect(element.Left, element.Top, element.Width, element.Height); bi.StreamSource = stream; bi.EndInit(); DrawingVisual vis = new DrawingVisual(); DrawingContext cont = vis.RenderOpen(); cont.DrawImage(bi, new Rect(new Size(element.Width, element.Height))); cont.Close(); RenderTargetBitmap rtb = new RenderTargetBitmap(element.Width, element.Height, 96d, 96d, PixelFormats.Default); rtb.Render(vis); using (MemoryStream ms = new MemoryStream()) { PngBitmapEncoder encoder = new PngBitmapEncoder(); encoder.Frames.Add(BitmapFrame.Create(rtb)); encoder.Save(ms); var regionKey = WebPath.Variant(key, element.Id); blobService.Save("image/png", regionKey, ms); } } } }

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  • Visitor and templated virtual methods

    - by Thomas Matthews
    In a typical implementation of the Visitor pattern, the class must account for all variations (descendants) of the base class. There are many instances where the same method content in the visitor is applied to the different methods. A templated virtual method would be ideal in this case, but for now, this is not allowed. So, can templated methods be used to resolve virtual methods of the parent class? Given (the foundation): struct Visitor_Base; // Forward declaration. struct Base { virtual accept_visitor(Visitor_Base& visitor) = 0; }; // More forward declarations struct Base_Int; struct Base_Long; struct Base_Short; struct Base_UInt; struct Base_ULong; struct Base_UShort; struct Visitor_Base { virtual void operator()(Base_Int& b) = 0; virtual void operator()(Base_Long& b) = 0; virtual void operator()(Base_Short& b) = 0; virtual void operator()(Base_UInt& b) = 0; virtual void operator()(Base_ULong& b) = 0; virtual void operator()(Base_UShort& b) = 0; }; struct Base_Int : public Base { void accept_visitor(Visitor_Base& visitor) { visitor(*this); } }; struct Base_Long : public Base { void accept_visitor(Visitor_Base& visitor) { visitor(*this); } }; struct Base_Short : public Base { void accept_visitor(Visitor_Base& visitor) { visitor(*this); } }; struct Base_UInt : public Base { void accept_visitor(Visitor_Base& visitor) { visitor(*this); } }; struct Base_ULong : public Base { void accept_visitor(Visitor_Base& visitor) { visitor(*this); } }; struct Base_UShort : public Base { void accept_visitor(Visitor_Base& visitor) { visitor(*this); } }; Now that the foundation is laid, here is where the kicker comes in (templated methods): struct Visitor_Cout : public Visitor { template <class Receiver> void operator() (Receiver& r) { std::cout << "Visitor_Cout method not implemented.\n"; } }; Intentionally, Visitor_Cout does not contain the keyword virtual in the method declaration. All the other attributes of the method signatures match the parent declaration (or perhaps specification). In the big picture, this design allows developers to implement common visitation functionality that differs only by the type of the target object (the object receiving the visit). The implementation above is my suggestion for alerts when the derived visitor implementation hasn't implement an optional method. Is this legal by the C++ specification? (I don't trust when some says that it works with compiler XXX. This is a question against the general language.)

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  • How to do inclusive range queries when only half-open range is supported (ala SortedMap.subMap)

    - by polygenelubricants
    On SortedMap.subMap This is the API for SortedMap<K,V>.subMap: SortedMap<K,V> subMap(K fromKey, K toKey) : Returns a view of the portion of this map whose keys range from fromKey, inclusive, to toKey, exclusive. This inclusive lower bound, exclusive upper bound combo ("half-open range") is something that is prevalent in Java, and while it does have its benefits, it also has its quirks, as we shall soon see. The following snippet illustrates a simple usage of subMap: static <K,V> SortedMap<K,V> someSortOfSortedMap() { return Collections.synchronizedSortedMap(new TreeMap<K,V>()); } //... SortedMap<Integer,String> map = someSortOfSortedMap(); map.put(1, "One"); map.put(3, "Three"); map.put(5, "Five"); map.put(7, "Seven"); map.put(9, "Nine"); System.out.println(map.subMap(0, 4)); // prints "{1=One, 3=Three}" System.out.println(map.subMap(3, 7)); // prints "{3=Three, 5=Five}" The last line is important: 7=Seven is excluded, due to the exclusive upper bound nature of subMap. Now suppose that we actually need an inclusive upper bound, then we could try to write a utility method like this: static <V> SortedMap<Integer,V> subMapInclusive(SortedMap<Integer,V> map, int from, int to) { return (to == Integer.MAX_VALUE) ? map.tailMap(from) : map.subMap(from, to + 1); } Then, continuing on with the above snippet, we get: System.out.println(subMapInclusive(map, 3, 7)); // prints "{3=Three, 5=Five, 7=Seven}" map.put(Integer.MAX_VALUE, "Infinity"); System.out.println(subMapInclusive(map, 5, Integer.MAX_VALUE)); // {5=Five, 7=Seven, 9=Nine, 2147483647=Infinity} A couple of key observations need to be made: The good news is that we don't care about the type of the values, but... subMapInclusive assumes Integer keys for to + 1 to work. A generic version that also takes e.g. Long keys is not possible (see related questions) Not to mention that for Long, we need to compare against Long.MAX_VALUE instead Overloads for the numeric primitive boxed types Byte, Character, etc, as keys, must all be written individually A special check need to be made for toInclusive == Integer.MAX_VALUE, because +1 would overflow, and subMap would throw IllegalArgumentException: fromKey > toKey This, generally speaking, is an overly ugly and overly specific solution What about String keys? Or some unknown type that may not even be Comparable<?>? So the question is: is it possible to write a general subMapInclusive method that takes a SortedMap<K,V>, and K fromKey, K toKey, and perform an inclusive-range subMap queries? Related questions Are upper bounds of indexed ranges always assumed to be exclusive? Is it possible to write a generic +1 method for numeric box types in Java? On NavigableMap It should be mentioned that there's a NavigableMap.subMap overload that takes two additional boolean variables to signify whether the bounds are inclusive or exclusive. Had this been made available in SortedMap, then none of the above would've even been asked. So working with a NavigableMap<K,V> for inclusive range queries would've been ideal, but while Collections provides utility methods for SortedMap (among other things), we aren't afforded the same luxury with NavigableMap. Related questions Writing a synchronized thread-safety wrapper for NavigableMap On API providing only exclusive upper bound range queries Does this highlight a problem with exclusive upper bound range queries? How were inclusive range queries done in the past when exclusive upper bound is the only available functionality?

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  • Google Maps: remember id of marker with open info window

    - by AP257
    I have a Google map that is showing a number of markers. When the user moves the map, the markers are redrawn for the new boundaries, using the code below: GEvent.addListener(map, "moveend", function() { var newBounds = map.getBounds(); for(var i = 0; i < places_json.places.length ; i++) { // if marker is within the new bounds then do... var latlng = new GLatLng(places_json.places[i].lat, places_json.places[i].lon); var html = "blah"; var marker = createMarker(latlng, html); map.addOverlay(marker); } }); My question is simple. If the user has clicked on a marker so that it is showing an open info window, currently when the boundaries are redrawn the info window is closed, because the marker is added again from scratch. How can I prevent this? It is not ideal, because often the boundaries are redrawn when the user clicks on a marker and the map moves to display the info window - so the info window appears and then disappears again :) I guess there are a couple of possible ways: remember which marker has an open info window, and open it again when the markers are redrawn don't actually re-add the marker with an open info window, just leave it there However, both require the marker with an open window to have some kind of ID number, and I don't know that this is actually the case in the Google Maps API. Anyone? ----------UPDATE------------------ I've tried doing it by loading the markers into an initial array, as suggested. This loads OK, but the page crashes after the map is dragged. <script type="text/javascript" src="{{ MEDIA_URL }}js/markerclusterer.js"></script> <script type='text/javascript'> function createMarker(point,html, hideMarker) { //alert('createMarker'); var icon = new GIcon(G_DEFAULT_ICON); icon.image = "http://chart.apis.google.com/chart?cht=mm&chs=24x32&chco=FFFFFF,008CFF,000000&ext=.png"; var tmpMarker = new GMarker(point, {icon: icon, hide: hideMarker}); GEvent.addListener(tmpMarker, "click", function() { tmpMarker.openInfoWindowHtml(html); }); return tmpMarker; } var map = new GMap2(document.getElementById("map_canvas")); map.addControl(new GSmallMapControl()); var mapLatLng = new GLatLng({{ place.lat }}, {{ place.lon }}); map.setCenter(mapLatLng, 12); map.addOverlay(new GMarker(mapLatLng)); // load initial markers from json array var markers = []; var initialBounds = map.getBounds(); for(var i = 0; i < places_json.places.length ; i++) { var latlng = new GLatLng(places_json.places[i].lat, places_json.places[i].lon); var html = "<strong><a href='/place/" + places_json.places[i].placesidx + "/" + places_json.places[i].area + "'>" + places_json.places[i].area + "</a></strong><br/>" + places_json.places[i].county; var hideMarker = true; if((initialBounds.getSouthWest().lat() < places_json.places[i].lat) && (places_json.places[i].lat < initialBounds.getNorthEast().lat()) && (initialBounds.getSouthWest().lng() < places_json.places[i].lon) && (places_json.places[i].lon < initialBounds.getNorthEast().lng()) && (places_json.places[i].placesidx != {{ place.placesidx }})) { hideMarker = false; } var marker = createMarker(latlng, html, hideMarker); markers.push(marker); } var markerCluster = new MarkerClusterer(map, markers, {maxZoom: 11}); </script>

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  • Rendering a view to a string in MVC, then redirecting -- workarounds?

    - by James S
    Hi -- I can't render a view to a string and then redirect, despite this answer from Feb (after version 1.0, I think) that claims it's possible. I thought I was doing something wrong, and then I read this answer from Haack in July that claims it's not possible. If somebody has it working and can help me get it working, that's great (and I'll post code, errors). However, I'm now at the point of needing workarounds. There are a few, but nothing ideal. Has anybody solved this, or have any comments on my ideas? This is to render email. While I can surely send the email outside of the web request (store info in a db and get it later), there are many types of emails and I don't want to store the template data (user object, a few other LINQ objects) in a db to let it get rendered later. I could create a simpler, serializable POCO and save that in the db, but why? ... I just want rendered text! I can create a new RedirectToAction object that checks if the headers have been sent (can't figure out how to do this -- try/catch?) and, if so, builds out a simple page with a meta redirect, a javascript redirect, and also a "click here" link. Within my controller, I can remember if I've rendered an email and, if so, manually do #2 by displaying a view. I can manually send the redirect headers before any potential email rendering. Then, rather than using the MVC infrastructure to redirecttoaction, I just call result.end. This seems easiest, but really messy. Anything else? EDIT: I've tried Dan's code (very similar to the code from Jan/Feb that I've already tried) and I'm still getting the same error. The only substantial difference I can see is that his example uses a view while I use a partial view. I'll try testing this later with a view. Here's what I've got: Controller public ActionResult Certifications(string email_intro) { //a lot of stuff ViewData["users"] = users; if (isPost()) { //create the viewmodel var view_model = new ViewModels.Emails.Certifications.Open(userContext) { emailIntro = email_intro }; //i've tried stopping this after just one iteration, in case the problem is due to calling it multiple times foreach (var user in users) { if (user.Email_Address.IsValidEmailAddress()) { //add more stuff to the view model specific to this user view_model.user = user; view_model.certification302Summary.subProcessesOwner = new SubProcess_Certifications(RecordUpdating.Role.Owner, null, null, user.User_ID, repository); //more here.... //if i comment out the next line, everything works ok SendEmail(view_model, this.ControllerContext); } } return RedirectToAction("Certifications"); } return View(); } SendEmail() public static void SendEmail(ViewModels.Emails.Certifications.Open model, ControllerContext context) { var vd = context.Controller.ViewData; vd["model"] = model; var renderer = new CustomRenderers(); //i fixed an error in your code here var text = renderer.RenderViewToString3(context, "~/Views/Emails/Certifications/Open.ascx", "", vd, null); var a = text; } CustomRenderers public class CustomRenderers { public virtual string RenderViewToString3(ControllerContext controllerContext, string viewPath, string masterPath, ViewDataDictionary viewData, TempDataDictionary tempData) { //copy/paste of dan's code } } Error [HttpException (0x80004005): Cannot redirect after HTTP headers have been sent.] System.Web.HttpResponse.Redirect(String url, Boolean endResponse) +8707691 Thanks, James

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  • Select number of rows for each group where two column values makes one group

    - by Fábio Antunes
    I have a two select statements joined by UNION ALL. In the first statement a where clause gathers only rows that have been shown previously to the user. The second statement gathers all rows that haven't been shown to the user, therefore I end up with the viewed results first and non-viewed results after. Of course this could simply be achieved with the same select statement using a simple ORDER BY, however the reason for two separate selects is simple after you realize what I hope to accomplish. Consider the following structure and data. +----+------+-----+--------+------+ | id | from | to | viewed | data | +----+------+-----+--------+------+ | 1 | 1 | 10 | true | .... | | 2 | 10 | 1 | true | .... | | 3 | 1 | 10 | true | .... | | 4 | 6 | 8 | true | .... | | 5 | 1 | 10 | true | .... | | 6 | 10 | 1 | true | .... | | 7 | 8 | 6 | true | .... | | 8 | 10 | 1 | true | .... | | 9 | 6 | 8 | true | .... | | 10 | 2 | 3 | true | .... | | 11 | 1 | 10 | true | .... | | 12 | 8 | 6 | true | .... | | 13 | 10 | 1 | false | .... | | 14 | 1 | 10 | false | .... | | 15 | 6 | 8 | false | .... | | 16 | 10 | 1 | false | .... | | 17 | 8 | 6 | false | .... | | 18 | 3 | 2 | false | .... | +----+------+-----+--------+------+ Basically I wish all non viewed rows to be selected by the statement, that is accomplished by checking weather the viewed column is true or false, pretty simple and straightforward, nothing to worry here. However when it comes to the rows already viewed, meaning the column viewed is TRUE, for those records I only want 3 rows to be returned for each group. The appropriate result in this instance should be the 3 most recent rows of each group. +----+------+-----+--------+------+ | id | from | to | viewed | data | +----+------+-----+--------+------+ | 6 | 10 | 1 | true | .... | | 7 | 8 | 6 | true | .... | | 8 | 10 | 1 | true | .... | | 9 | 6 | 8 | true | .... | | 10 | 2 | 3 | true | .... | | 11 | 1 | 10 | true | .... | | 12 | 8 | 6 | true | .... | +----+------+-----+--------+------+ As you see from the ideal result set we have three groups. Therefore the desired query for the viewed results should show a maximum of 3 rows for each grouping it finds. In this case these groupings were 10 with 1 and 8 with 6, both which had three rows to be shown, while the other group 2 with 3 only had one row to be shown. Please note that where from = x and to = y, makes the same grouping as if it was from = y and to = x. Therefore considering the first grouping (10 with 1), from = 10 and to = 1 is the same group if it was from = 1 and to = 10. However there are plenty of groups in the whole table that I only wish the 3 most recent of each to be returned in the select statement, and thats my problem, I not sure how that can be accomplished in the most efficient way possible considering the table will have hundreds if not thousands of records at some point. Thanks for your help. Note: The columns id, from, to and viewed are indexed, that should help with performance. PS: I'm unsure on how to name this question exactly, if you have a better idea, be my guest and edit the title.

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  • UIImage rounded corners

    - by catlan
    I try to get rounded corners on a UIImage, what I read so far, the easiest way is to use a mask images. For this I used code from TheElements iPhone Example and some image resize code I found. My problem is that resizedImage is always nil and I don't find the error... - (UIImage *)imageByScalingProportionallyToSize:(CGSize)targetSize { CGSize imageSize = [self size]; float width = imageSize.width; float height = imageSize.height; // scaleFactor will be the fraction that we'll // use to adjust the size. For example, if we shrink // an image by half, scaleFactor will be 0.5. the // scaledWidth and scaledHeight will be the original, // multiplied by the scaleFactor. // // IMPORTANT: the "targetHeight" is the size of the space // we're drawing into. The "scaledHeight" is the height that // the image actually is drawn at, once we take into // account the ideal of maintaining proportions float scaleFactor = 0.0; float scaledWidth = targetSize.width; float scaledHeight = targetSize.height; CGPoint thumbnailPoint = CGPointMake(0,0); // since not all images are square, we want to scale // proportionately. To do this, we find the longest // edge and use that as a guide. if ( CGSizeEqualToSize(imageSize, targetSize) == NO ) { // use the longeset edge as a guide. if the // image is wider than tall, we'll figure out // the scale factor by dividing it by the // intended width. Otherwise, we'll use the // height. float widthFactor = targetSize.width / width; float heightFactor = targetSize.height / height; if ( widthFactor < heightFactor ) scaleFactor = widthFactor; else scaleFactor = heightFactor; // ex: 500 * 0.5 = 250 (newWidth) scaledWidth = width * scaleFactor; scaledHeight = height * scaleFactor; // center the thumbnail in the frame. if // wider than tall, we need to adjust the // vertical drawing point (y axis) if ( widthFactor < heightFactor ) thumbnailPoint.y = (targetSize.height - scaledHeight) * 0.5; else if ( widthFactor > heightFactor ) thumbnailPoint.x = (targetSize.width - scaledWidth) * 0.5; } CGContextRef mainViewContentContext; CGColorSpaceRef colorSpace; colorSpace = CGColorSpaceCreateDeviceRGB(); // create a bitmap graphics context the size of the image mainViewContentContext = CGBitmapContextCreate (NULL, targetSize.width, targetSize.height, 8, 0, colorSpace, kCGImageAlphaPremultipliedLast); // free the rgb colorspace CGColorSpaceRelease(colorSpace); if (mainViewContentContext==NULL) return NULL; //CGContextSetFillColorWithColor(mainViewContentContext, [[UIColor whiteColor] CGColor]); //CGContextFillRect(mainViewContentContext, CGRectMake(0, 0, targetSize.width, targetSize.height)); CGContextDrawImage(mainViewContentContext, CGRectMake(thumbnailPoint.x, thumbnailPoint.y, scaledWidth, scaledHeight), self.CGImage); // Create CGImageRef of the main view bitmap content, and then // release that bitmap context CGImageRef mainViewContentBitmapContext = CGBitmapContextCreateImage(mainViewContentContext); CGContextRelease(mainViewContentContext); CGImageRef maskImage = [[UIImage imageNamed:@"Mask.png"] CGImage]; CGImageRef resizedImage = CGImageCreateWithMask(mainViewContentBitmapContext, maskImage); CGImageRelease(mainViewContentBitmapContext); // convert the finished resized image to a UIImage UIImage *theImage = [UIImage imageWithCGImage:resizedImage]; // image is retained by the property setting above, so we can // release the original CGImageRelease(resizedImage); // return the image return theImage; }

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  • Paging, sorting and filtering in a stored procedure (SQL Server)

    - by Fruitbat
    I was looking at different ways of writing a stored procedure to return a "page" of data. This was for use with the asp ObjectDataSource, but it could be considered a more general problem. The requirement is to return a subset of the data based on the usual paging paremeters, startPageIndex and maximumRows, but also a sortBy parameter to allow the data to be sorted. Also there are some parameters passed in to filter the data on various conditions. One common way to do this seems to be something like this: [Method 1] ;WITH stuff AS ( SELECT CASE WHEN @SortBy = 'Name' THEN ROW_NUMBER() OVER (ORDER BY Name) WHEN @SortBy = 'Name DESC' THEN ROW_NUMBER() OVER (ORDER BY Name DESC) WHEN @SortBy = ... ELSE ROW_NUMBER() OVER (ORDER BY whatever) END AS Row, ., ., ., FROM Table1 INNER JOIN Table2 ... LEFT JOIN Table3 ... WHERE ... (lots of things to check) ) SELECT * FROM stuff WHERE (Row > @startRowIndex) AND (Row <= @startRowIndex + @maximumRows OR @maximumRows <= 0) ORDER BY Row One problem with this is that it doesn't give the total count and generally we need another stored procedure for that. This second stored procedure has to replicate the parameter list and the complex WHERE clause. Not nice. One solution is to append an extra column to the final select list, (SELECT COUNT(*) FROM stuff) AS TotalRows. This gives us the total but repeats it for every row in the result set, which is not ideal. [Method 2] An interesting alternative is given here (http://www.4guysfromrolla.com/articles/032206-1.aspx) using dynamic SQL. He reckons that the performance is better because the CASE statement in the first solution drags things down. Fair enough, and this solution makes it easy to get the totalRows and slap it into an output parameter. But I hate coding dynamic SQL. All that 'bit of SQL ' + STR(@parm1) +' bit more SQL' gubbins. [Method 3] The only way I can find to get what I want, without repeating code which would have to be synchronised, and keeping things reasonably readable is to go back to the "old way" of using a table variable: DECLARE @stuff TABLE (Row INT, ...) INSERT INTO @stuff SELECT CASE WHEN @SortBy = 'Name' THEN ROW_NUMBER() OVER (ORDER BY Name) WHEN @SortBy = 'Name DESC' THEN ROW_NUMBER() OVER (ORDER BY Name DESC) WHEN @SortBy = ... ELSE ROW_NUMBER() OVER (ORDER BY whatever) END AS Row, ., ., ., FROM Table1 INNER JOIN Table2 ... LEFT JOIN Table3 ... WHERE ... (lots of things to check) SELECT * FROM stuff WHERE (Row > @startRowIndex) AND (Row <= @startRowIndex + @maximumRows OR @maximumRows <= 0) ORDER BY Row (Or a similar method using an IDENTITY column on the table variable). Here I can just add a SELECT COUNT on the table variable to get the totalRows and put it into an output parameter. I did some tests and with a fairly simple version of the query (no sortBy and no filter), method 1 seems to come up on top (almost twice as quick as the other 2). Then I decided to test probably I needed the complexity and I needed the SQL to be in stored procedures. With this I get method 1 taking nearly twice as long as the other 2 methods. Which seems strange. Is there any good reason why I shouldn't spurn CTEs and stick with method 3? UPDATE - 15 March 2012 I tried adapting Method 1 to dump the page from the CTE into a temporary table so that I could extract the TotalRows and then select just the relevant columns for the resultset. This seemed to add significantly to the time (more than I expected). I should add that I'm running this on a laptop with SQL Server Express 2008 (all that I have available) but still the comparison should be valid. I looked again at the dynamic SQL method. It turns out I wasn't really doing it properly (just concatenating strings together). I set it up as in the documentation for sp_executesql (with a parameter description string and parameter list) and it's much more readable. Also this method runs fastest in my environment. Why that should be still baffles me, but I guess the answer is hinted at in Hogan's comment.

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