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  • How to Make Ubuntu Play MP3 Files

    - by Trevor Bekolay
    Because of licensing issues, Ubuntu is unable to play MP3s out of the box. We’ll show you how to play MP3s and other restricted file formats in about four mouse clicks. The philosophy behind Ubuntu is that software should be free and accessible to all. Whether MP3 and other file formats are free is unclear in many countries, so Ubuntu does not include software to read these file formats by default. Fortunately, it does include a package that installs the most commonly used file formats all at once, including a Flash plugin for Firefox. Note: These instructions are for Ubuntu 10.04. There are small differences for earlier versions of Ubuntu. Play MP3 Files Open the Ubuntu Software Center, found in the Applications menu.   Click on View and ensure that All Software is selected. Type “restricted extras” into the search box at the top-right. Find the Ubuntu restricted extras package and click Install. Enter your password when prompted. Once the install is complete, close out of Ubuntu Software Center, and you’ll be able to play MP3 files! To confirm this, we’ll open up Rhythmbox, found in the Sound & Video section of the Applications menu. Our test MP3 plays with no problems! Note: If Rhythmbox tells you that MP3 plugins are not installed, close Rhythmbox and reopen it. You should not have to install anything extra through Rhythmbox.   Despite this extra step, playing the most common audio and video file formats – including Flash videos on the internet – is simple. All the software comes installed, you just have to teach them how to read your files. Similar Articles Productive Geek Tips How to Play .OGM Video Files in Windows VistaView Hidden Files and Folders in Ubuntu File BrowserMake Ubuntu Automatically Save Changes to Your SessionInstalling PHP4 and Apache on UbuntuInstalling PHP5 and Apache on Ubuntu TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Xobni Plus for Outlook All My Movies 5.9 CloudBerry Online Backup 1.5 for Windows Home Server Snagit 10 How to Forecast Weather, without Gadgets Outlook Tools, one stop tweaking for any Outlook version Zoofs, find the most popular tweeted YouTube videos Video preview of new Windows Live Essentials 21 Cursor Packs for XP, Vista & 7 Map the Stars with Stellarium

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  • Make Windows Position Your Dual Monitors Correctly

    - by Mysticgeek
    If you have a dual monitor setup and each monitor is a different size or height, it can be annoying trying to move the mouse pointer between them. Here is a quick tip that will help make the process easier. Align Monitors In our example, we’re using Windows 7, but the process is essentially the same in all versions, but getting to Display Settings is different. In Windows 7 open the Start menu and type display settings into the search box and hit Enter. In Vista right-click the desktop and click Personalize. Then from the Personalize appearance and sounds menu click on Display Settings. In XP right-click on the desktop and select Properties then in Display Properties click the Settings tab. Now here is where you can change the appearance of your monitors. In this example we have a larger 22” LCD and a smaller 19” and it can be annoying getting the mouse pointer from one to another depending where you are on each monitor. So what you want to do is simply move each display around to a particular height so it’s easier to get the pointer over. For example with this setting we know we’ll have no problem moving the pointer to the other screen at the top of each display.   Of course here you can flip your monitors around, change the display resolution, orientation, etc. If you have dual monitors where one might be larger or set up higher than the other, then this is a great way to get them finely tuned. You will have to play around with the settings a bit to settle on what works best for you. Similar Articles Productive Geek Tips GeekNewb: Get to Know These Windows 7 HotkeysDual Monitors: Use a Different Wallpaper on Each DesktopSet Windows as Default OS when Dual Booting UbuntuEasily Set Default OS in a Windows 7 / Vista and XP Dual-boot SetupSet XP as the Default OS in a Windows Vista Dual-Boot Setup TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows PC Tools Internet Security Suite 2010 Download Wallpapers From National Geographic Site Spyware Blaster v4.3 Yes, it’s Patch Tuesday Generate Stunning Tag Clouds With Tagxedo Install, Remove and HIDE Fonts in Windows 7 Need Help with Your Home Network?

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  • Using MAC Authentication for simple Web API’s consumption

    - by cibrax
    For simple scenarios of Web API consumption where identity delegation is not required, traditional http authentication schemas such as basic, certificates or digest are the most used nowadays. All these schemas rely on sending the caller credentials or some representation of it in every request message as part of the Authorization header, so they are prone to suffer phishing attacks if they are not correctly secured at transport level with https. In addition, most client applications typically authenticate two different things, the caller application and the user consuming the API on behalf of that application. For most cases, the schema is simplified by using a single set of username and password for authenticating both, making necessary to store those credentials temporally somewhere in memory. The true is that you can use two different identities, one for the user running the application, which you might authenticate just once during the first call when the application is initialized, and another identity for the application itself that you use on every call. Some cloud vendors like Windows Azure or Amazon Web Services have adopted an schema to authenticate the caller application based on a Message Authentication Code (MAC) generated with a symmetric algorithm using a key known by the two parties, the caller and the Web API. The caller must include a MAC as part of the Authorization header created from different pieces of information in the request message such as the address, the host, and some other headers. The Web API can authenticate the caller by using the key associated to it and validating the attached MAC in the request message. In that way, no credentials are sent as part of the request message, so there is no way an attacker to intercept the message and get access to those credentials. Anyways, this schema also suffers from some deficiencies that can generate attacks. For example, brute force can be still used to infer the key used for generating the MAC, and impersonate the original caller. This can be mitigated by renewing keys in a relative short period of time. This schema as any other can be complemented with transport security. Eran Rammer, one of the brains behind OAuth, has recently published an specification of a protocol based on MAC for Http authentication called Hawk. The initial version of the spec is available here. A curious fact is that the specification per se does not exist, and the specification itself is the code that Eran initially wrote using node.js. In that implementation, you can associate a key to an user, so once the MAC has been verified on the Web API, the user can be inferred from that key. Also a timestamp is used to avoid replay attacks. As a pet project, I decided to port that code to .NET using ASP.NET Web API, which is available also in github under https://github.com/pcibraro/hawknet Enjoy!.

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  • Changing the default installation path to a newly installed hard disk

    - by mgj
    Hi, I am currently working on a dual-booted PC. I am using Windows XP and Ubuntu 10.04 Lucid Lynx released in April 2010. The allocated partition to Ubuntu that I am making use of has almost exhausted. Current memory allocations on the PC wrt Ubuntu OS looks like this: bodhgaya@pc146724-desktop:~$ df -h Filesystem Size Used Avail Use% Mounted on /dev/sda2 8.6G 8.0G 113M 99% / none 998M 268K 998M 1% /dev none 1002M 580K 1002M 1% /dev/shm none 1002M 100K 1002M 1% /var/run none 1002M 0 1002M 0% /var/lock none 1002M 0 1002M 0% /lib/init/rw /dev/sda1 25G 16G 9.8G 62% /media/C /dev/sdb1 37G 214M 35G 1% /media/ubuntulinuxstore bodhgaya@pc146724-desktop:~$ cd /tmp I am trying to mount a 40GB(/dev/sdb1 - given below) new hard disk along with my existing Ubuntu system to overcome with hard disk space related issues. I referred to the following tutorial to mount a new hard disk onto the system:- http://www.smorgasbord.net/how-to-in...untu-linux%20/ I was able to successfully mount this hard disk for Ubuntu 0S. I have this new hard disk setup in /media/ubuntulinuxstore directory. The current partition in my system looks like this: bodhgaya@pc146724-desktop:/media/ubuntulinuxstore$ sudo fdisk -l [sudo] password for bodhgaya: Disk /dev/sda: 40.0 GB, 40000000000 bytes 255 heads, 63 sectors/track, 4863 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x446eceb5 Device Boot Start End Blocks Id System /dev/sda1 * 2 3264 26210047+ 7 HPFS/NTFS /dev/sda2 3265 4385 9004432+ 83 Linux /dev/sda3 4386 4863 3839535 82 Linux swap / Solaris Disk /dev/sdb: 40.0 GB, 40000000000 bytes 255 heads, 63 sectors/track, 4863 cylinders Units = cylinders of 16065 * 512 = 8225280 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0xfa8afa8a Device Boot Start End Blocks Id System /dev/sdb1 1 4862 39053983+ 7 HPFS/NTFS bodhgaya@pc146724-desktop:/media/ubuntulinuxstore$ Now, I have a concern wrt the "location" where the new softwares will be installed. Generally softwares are installed via the terminal and by default a fixed path is used to where the post installation set up files can be found (I am talking in context of the drive). This is like the typical case of Windows, where softwares by default are installed in the C: drive. These days people customize their installations to a drive which they find apt to serve their purpose (generally based on availability of hard disk space). I am trying to figure out how to customize the same for Ubuntu. As we all know the most softwares are installed via commands given from the Terminal. My road block is how do I redirect the default path set on the terminal where files get installed to this new hard disk. This if done will help me overcome space constraints I am currently facing wrt the partition on which my Ubuntu is initially installed. I would also by this, save time on not formatting my system and reinstalling Ubuntu and other softwares all over again. Please help me with this, your suggestions are much appreciated.

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  • Requesting Delegation (ActAs) Tokens using WSTrustChannel (as opposed to Configuration Madness)

    - by Your DisplayName here!
    Delegation using the ActAs approach has some interesting security features A security token service can make authorization and validation checks before issuing the ActAs token. Combined with proof keys you get non-repudiation features. The ultimate receiver sees the original caller as direct caller and can optionally traverse the delegation chain. Encryption and audience restriction can be tied down Most samples out there (including the SDK sample) use the CreateChannelActingAs extension method from WIF to request ActAs tokens. This method builds on top of the WCF binding configuration which may not always be suitable for your situation. You can also use the WSTrustChannel to request ActAs tokens. This allows direct and programmatic control over bindings and configuration and is my preferred approach. The below method requests an ActAs token based on a bootstrap token. The returned token can then directly be used with the CreateChannelWithIssued token extension method. private SecurityToken GetActAsToken(SecurityToken bootstrapToken) {     var factory = new WSTrustChannelFactory(         new UserNameWSTrustBinding(SecurityMode.TransportWithMessageCredential),         new EndpointAddress(_stsAddress));     factory.TrustVersion = TrustVersion.WSTrust13;     factory.Credentials.UserName.UserName = "middletier";     factory.Credentials.UserName.Password = "abc!123";     var rst = new RequestSecurityToken     {         AppliesTo = new EndpointAddress(_serviceAddress),         RequestType = RequestTypes.Issue,         KeyType = KeyTypes.Symmetric,         ActAs = new SecurityTokenElement(bootstrapToken)     };     var channel = factory.CreateChannel();     var delegationToken = channel.Issue(rst);     return delegationToken; }   HTH

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  • Fix: SqlDeploy Task Fails with NullReferenceException at ExtractPassword

    Still working on getting a TeamCity build working (see my last post).  Latest exception is: C:\Program Files\MSBuild\Microsoft\VisualStudio\v9.0\TeamData\Microsoft.Data.Schema.SqlTasks.targets(120, 5): error MSB4018: The "SqlDeployTask" task failed unexpectedly. System.NullReferenceException: Object reference not set to an instance of an object. at Microsoft.Data.Schema.Common.ConnectionStringPersistence.ExtractPassword(String partialConnection, String dbProvider) at Microsoft.Data.Schema.Common.ConnectionStringPersistence.RetrieveFullConnection(String partialConnection, String provider, Boolean presentUI, String password) at Microsoft.Data.Schema.Sql.Build.SqlDeployment.ConfigureConnectionString(String connectionString, String databaseName) at Microsoft.Data.Schema.Sql.Build.SqlDeployment.OnBuildConnectionString(String partialConnectionString, String databaseName) at Microsoft.Data.Schema.Build.Deployment.FinishInitialize(String targetConnectionString) at Microsoft.Data.Schema.Build.Deployment.Initialize(FileInfo sourceDbSchemaFile, ErrorManager errors, String targetConnectionString) at Microsoft.Data.Schema.Build.DeploymentConstructor.ConstructServiceImplementation() at Microsoft.Data.Schema.Extensibility.ServiceConstructor'1.ConstructService() at Microsoft.Data.Schema.Tasks.DBDeployTask.Execute() at Microsoft.Build.BuildEngine.TaskEngine.ExecuteInstantiatedTask(EngineProxy engineProxy, ItemBucket bucket, TaskExecutionMode howToExecuteTask, ITask task, Boolean& taskResult)   This time searching yielded some good stuff, including this thread that talks about how to resolve this via permissions.  The short answer is that the account that your build server runs under needs to have the necessary permissions in SQL Server.  Youll need to create a Login and then ensure at least the minimum rights are configured as described here: Required Permissions in Database Edition Alternately, you can just make your buildserver account an admin on the database (which is probably running on the same machine anyway) and at that point it should be able to do whatever it needs to. If youre certain the account has the necessary permissions, but youre still getting the error, the problem may be that the account has never logged into the build server.  In this case, there wont be any entry in the HKCU hive in the registry, which the system is checking for permissions (see this thread).  The solution in this case is quite simple: log into the machine (once is enough) with the build server account.  Then, open Visual Studio (thanks Brendan for the answer in this thread). Summary Make sure the build service account has the necessary database permissions Make sure the account has logged into the server so it has the necessary registry hive info Make sure the account has run Visual Studio at least once so its settings are established In my case I went through all 3 of these steps before I resolved the problem. Did you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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  • F# and ArcObjects, Part 3

    - by Marko Apfel
    Today i played a little bit with IFeature-sequences and piping data. The result was a calculator of the bounding box around all features in a feature class. Maybe a little bit dirty, but for learning was it OK. ;-) open System;; #I "C:\Program Files\ArcGIS\DotNet";; #r "ESRI.ArcGIS.System.dll";; #r "ESRI.ArcGIS.DataSourcesGDB.dll";; #r "ESRI.ArcGIS.Geodatabase.dll";; #r "ESRI.ArcGIS.Geometry.dll";; open ESRI.ArcGIS.esriSystem;; open ESRI.ArcGIS.DataSourcesGDB;; open ESRI.ArcGIS.Geodatabase;; open ESRI.ArcGIS.Geometry; let aoInitialize = new AoInitializeClass();; let status = aoInitialize.Initialize(esriLicenseProductCode.esriLicenseProductCodeArcEditor);; let workspacefactory = new SdeWorkspaceFactoryClass();; let connection = "SERVER=okul;DATABASE=p;VERSION=sde.default;INSTANCE=sde:sqlserver:okul;USER=s;PASSWORD=g";; let workspace = workspacefactory.OpenFromString(connection, 0);; let featureWorkspace = (box workspace) :?> IFeatureWorkspace;; let featureClass = featureWorkspace.OpenFeatureClass("Praxair.SFG.BP_L_ROHR");; let queryFilter = new QueryFilterClass();; let featureCursor = featureClass.Search(queryFilter, true);; let featureCursorSeq (featureCursor : IFeatureCursor) = let actualFeature = ref (featureCursor.NextFeature()) seq { while (!actualFeature) <> null do yield actualFeature do actualFeature := featureCursor.NextFeature() };; let min x y = if x < y then x else y;; let max x y = if x > y then x else y;; let info s (x : IEnvelope) = printfn "%s xMin:{%f} xMax: {%f} yMin:{%f} yMax: {%f}" s x.XMin x.XMax x.YMin x.YMax;; let con (env1 : IEnvelope) (env2 : IEnvelope) = let env = (new EnvelopeClass()) :> IEnvelope env.XMin <- min env1.XMin env2.XMin env.XMax <- max env1.XMax env2.XMax env.YMin <- min env1.YMin env2.YMin env.YMax <- max env1.YMax env2.YMax info "Intermediate" env env;; let feature = featureClass.GetFeature(100);; let ext = feature.Extent;; let BoundingBox featureClassName = let featureClass = featureWorkspace.OpenFeatureClass(featureClassName) let queryFilter = new QueryFilterClass() let featureCursor = featureClass.Search(queryFilter, true) let featureCursorSeq (featureCursor : IFeatureCursor) = let actualFeature = ref (featureCursor.NextFeature()) seq { while (!actualFeature) <> null do yield actualFeature do actualFeature := featureCursor.NextFeature() } featureCursorSeq featureCursor |> Seq.map (fun feature -> (!feature).Extent) |> Seq.fold (fun (acc : IEnvelope) a -> info "Intermediate" acc (con acc a)) ext ;; let boundingBox = BoundingBox "Praxair.SFG.BP_L_ROHR";; info "Ende-Info:" boundingBox;;

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  • How to Assign a Static IP to an Ubuntu 10.04 Desktop Computer

    - by Mysticgeek
    If you have a home network with several computers, assigning them static IP addresses can make troubleshooting easier. Today we take a look at switching from DHCP to a static IP in Ubuntu. Assign a Static IP Using Static IPs prevents address conflicts between machines and can allow easier access to them. If you have a small home network and are satisfied with the machines getting their IP address automatically via DHCP, there won’t be anything gained by using static addresses. Using Static IPs isn’t necessarily for the average user, but if you’re a geek who wants to know the address assigned to each machine, it can allow for faster troubleshooting.  To change your Ubuntu machine to a Static IP go to System \ Preferences \ Network Connections. In our example, we’re on a wired system so click on the Wired tab, then select Auto eth0 and click on Edit. Select the IPv4 settings tab, change Method to Manual, click the Add button. Then type in the Static IP Address, Subnet Mask, DNS Servers, and Default Gateway. Then click Apply when you’re finished. Make sure to hit Enter after typing in the Default Gateway otherwise it will revert back to 0.0.0.0 You’ll need to enter in your admin password before the changes go into affect. To verify the changes have been made successfully launch a Terminal session and type in ifconfig at the command prompt, or follow these directions. You also might want to ping the address from another machine to make sure everything is communicating. If you want to assign a Static IP to your Windows machines, check out our article on how to assign a Static IP on Windows systems (make sure to browse the comments as our readers have some good suggestions).  Whether you have a small office or home network set up with a server and several machines, using a Static IP on each device can help you manage them easily. Again, it isn’t for everyone as it really depends on how your network is setup and the way you use it. Similar Articles Productive Geek Tips Change Ubuntu Desktop from DHCP to a Static IP AddressAllow Remote Control To Your Desktop On UbuntuAssign Custom Shortcut Keys on Ubuntu LinuxKeyboard Ninja: 21 Keyboard Shortcut ArticlesChange Ubuntu Server from DHCP to a Static IP Address TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips HippoRemote Pro 2.2 Xobni Plus for Outlook All My Movies 5.9 CloudBerry Online Backup 1.5 for Windows Home Server TubeSort: YouTube Playlist Organizer XPS file format & XPS Viewer Explained Microsoft Office Web Apps Guide Know if Someone Accessed Your Facebook Account Shop for Music with Windows Media Player 12 Access Free Documentaries at BBC Documentaries

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  • mysql completely removing

    - by Dmitry Teplyakov
    I broke my mysql and now I want to completely reinstall it. I tried: $ sudo apt-get install --reinstall mysql-server $ sudo apt-get remove --purge mysql-client mysql-server But always I see popup with proposal to change root password, I change it and got an error that I can change it.. $ sudo apt-get remove --purge mysql-client mysql-server Reading package lists... Done Building dependency tree Reading state information... Done Package mysql-client is not installed, so not removed Package mysql-server is not installed, so not removed The following packages were automatically installed and are no longer required: libmygpo-qt1 libqtscript4-network libqtscript4-gui libtag-extras1 libqtscript4-sql libqtscript4-xml amarok-utils amarok-common libqtscript4-uitools liblastfm0 libloudmouth1-0 libqtscript4-core Use 'apt-get autoremove' to remove them. 0 upgraded, 0 newly installed, 0 to remove and 0 not upgraded. 1 not fully installed or removed. After this operation, 0 B of additional disk space will be used. Setting up mysql-server-5.5 (5.5.28-0ubuntu0.12.04.2) ... 121114 19:04:03 [Note] Plugin 'FEDERATED' is disabled. 121114 19:04:03 InnoDB: The InnoDB memory heap is disabled 121114 19:04:03 InnoDB: Mutexes and rw_locks use GCC atomic builtins 121114 19:04:03 InnoDB: Compressed tables use zlib 1.2.3.4 121114 19:04:03 InnoDB: Initializing buffer pool, size = 128.0M 121114 19:04:03 InnoDB: Completed initialization of buffer pool InnoDB: Error: auto-extending data file ./ibdata1 is of a different size InnoDB: 0 pages (rounded down to MB) than specified in the .cnf file: InnoDB: initial 640 pages, max 0 (relevant if non-zero) pages! 121114 19:04:03 InnoDB: Could not open or create data files. 121114 19:04:03 InnoDB: If you tried to add new data files, and it failed here, 121114 19:04:03 InnoDB: you should now edit innodb_data_file_path in my.cnf back 121114 19:04:03 InnoDB: to what it was, and remove the new ibdata files InnoDB created 121114 19:04:03 InnoDB: in this failed attempt. InnoDB only wrote those files full of 121114 19:04:03 InnoDB: zeros, but did not yet use them in any way. But be careful: do not 121114 19:04:03 InnoDB: remove old data files which contain your precious data! 121114 19:04:03 [ERROR] Plugin 'InnoDB' init function returned error. 121114 19:04:03 [ERROR] Plugin 'InnoDB' registration as a STORAGE ENGINE failed. 121114 19:04:03 [ERROR] Unknown/unsupported storage engine: InnoDB 121114 19:04:03 [ERROR] Aborting 121114 19:04:03 [Note] /usr/sbin/mysqld: Shutdown complete start: Job failed to start invoke-rc.d: initscript mysql, action "start" failed. dpkg: error processing mysql-server-5.5 (--configure): subprocess installed post-installation script returned error exit status 1 Errors were encountered while processing: mysql-server-5.5 E: Sub-process /usr/bin/dpkg returned an error code (1) It is good for me that I have not any important databases, but..

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  • ASP.NET Localization: Enabling resource expressions with an external resource assembly

    - by Brian Schroer
    I have several related projects that need the same localized text, so my global resources files are in a shared assembly that’s referenced by each of those projects. It took an embarrassingly long time to figure out how to have my .resx files generate “public” properties instead of “internal” so I could have a shared resources assembly (apparently it was pretty tricky pre-VS2008, and my “googling” bogged me down some out-of-date instructions). It’s easy though – Just change the “Custom Tool” to “PublicResXFileCodeGenerator”:    …which can be done via the “Access Modifier” dropdown of the resource file designer window:   A reference to my shared resources DLL gives me the ability to use the resources in code, but by default, the ASP.NET resource expression syntax: <asp:Button ID="BeerButton" runat="server" Text="<%$ Resources:MyResources, Beer %>" />   …assumes that your resources are in your web site project.   To make resource expressions work with my shared resources assembly, I added two classes to the resources assembly: 1) a custom IResourceProvider implementation:   1: using System; 2: using System.Web.Compilation; 3: using System.Globalization; 4:   5: namespace DuffBeer 6: { 7: public class CustomResourceProvider : IResourceProvider 8: { 9: public object GetObject(string resourceKey, CultureInfo culture) 10: { 11: return MyResources.ResourceManager.GetObject(resourceKey, culture); 12: } 13:   14: public System.Resources.IResourceReader ResourceReader 15: { 16: get { throw new NotSupportedException(); } 17: } 18: } 19: }   2) and a custom factory class inheriting from the ResourceProviderFactory base class:   1: using System; 2: using System.Web.Compilation; 3:   4: namespace DuffBeer 5: { 6: public class CustomResourceProviderFactory : ResourceProviderFactory 7: { 8: public override IResourceProvider CreateGlobalResourceProvider(string classKey) 9: { 10: return new CustomResourceProvider(); 11: } 12:   13: public override IResourceProvider CreateLocalResourceProvider(string virtualPath) 14: { 15: throw new NotSupportedException(String.Format( 16: "{0} does not support local resources.", 17: this.GetType().Name)); 18: } 19: } 20: }   In the “system.web / globalization” section of my web.config file, I point the “resourceProviderFactoryType" property to my custom factory:   <system.web> <globalization culture="auto:en-US" uiCulture="auto:en-US" resourceProviderFactoryType="DuffBeer.CustomResourceProviderFactory, DuffBeer" />   This simple approach met my needs for these projects , but if you want to create reusable resource provider and factory classes that allow you to specify the assembly in the resource expression, the instructions are here.

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  • Is inline SQL still classed as bad practice now that we have Micro ORMs?

    - by Grofit
    This is a bit of an open ended question but I wanted some opinions, as I grew up in a world where inline SQL scripts were the norm, then we were all made very aware of SQL injection based issues, and how fragile the sql was when doing string manipulations all over the place. Then came the dawn of the ORM where you were explaining the query to the ORM and letting it generate its own SQL, which in a lot of cases was not optimal but was safe and easy. Another good thing about ORMs or database abstraction layers were that the SQL was generated with its database engine in mind, so I could use Hibernate/Nhibernate with MSSQL, MYSQL and my code never changed it was just a configuration detail. Now fast forward to current day, where Micro ORMs seem to be winning over more developers I was wondering why we have seemingly taken a U-Turn on the whole in-line sql subject. I must admit I do like the idea of no ORM config files and being able to write my query in a more optimal manner but it feels like I am opening myself back up to the old vulnerabilities such as SQL injection and I am also tying myself to one database engine so if I want my software to support multiple database engines I would need to do some more string hackery which seems to then start to make code unreadable and more fragile. (Just before someone mentions it I know you can use parameter based arguments with most micro orms which offers protection in most cases from sql injection) So what are peoples opinions on this sort of thing? I am using Dapper as my Micro ORM in this instance and NHibernate as my regular ORM in this scenario, however most in each field are quite similar. What I term as inline sql is SQL strings within source code. There used to be design debates over SQL strings in source code detracting from the fundamental intent of the logic, which is why statically typed linq style queries became so popular its still just 1 language, but with lets say C# and Sql in one page you have 2 languages intermingled in your raw source code now. Just to clarify, the SQL injection is just one of the known issues with using sql strings, I already mention you can stop this from happening with parameter based queries, however I highlight other issues with having SQL queries ingrained in your source code, such as the lack of DB Vendor abstraction as well as losing any level of compile time error capturing on string based queries, these are all issues which we managed to side step with the dawn of ORMs with their higher level querying functionality, such as HQL or LINQ etc (not all of the issues but most of them). So I am less focused on the individual highlighted issues and more the bigger picture of is it now becoming more acceptable to have SQL strings directly in your source code again, as most Micro ORMs use this mechanism. Here is a similar question which has a few different view points, although is more about the inline sql without the micro orm context: http://stackoverflow.com/questions/5303746/is-inline-sql-hard-coding

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  • How do I Integrate Production Database Hot Fixes into Shared Database Development model?

    - by TetonSig
    We are using SQL Source Control 3, SQL Compare, SQL Data Compare from RedGate, Mercurial repositories, TeamCity and a set of 4 environments including production. I am working on getting us to a dedicated environment per developer, but for at least the next 6 months we are stuck with a shared model. To summarize our current system, we have a DEV SQL server where developers first make changes/additions. They commit their changes through SQL Source Control to a local hgdev repository. When they execute an hg push to the main repository, TeamCity listens for that and then (among other things) pushes hgdev repository to hgrc. Another TeamCity process listens for that and does a pull from hgrc and deploys the latest to a QA SQL Server where regression and integration tests are run. When those are passed a push from hgrc to hgprod occurs. We do a compare of hgprod to our PREPROD SQL Server and generate deployment/rollback scripts for our production release. Separate from the above we have database Hot Fixes that will need to be applied in between releases. The process there is for our Operations team make changes on the PreProd database, and then after testing, to use SQL Source Control to commit their hot fix changes to hgprod from the PREPROD database, and then do a compare from hgprod to PRODUCTION, create deployment scripts and run them on PRODUCTION. If we were in a dedicated database per developer model, we could simply automatically push hgprod back to hgdev and merge in the hot fix change (through TeamCity monitoring for hgprod checkins) and then developers would pick it up and merge it to their local repository and database periodically. However, given that with a shared model the DEV database itself is the source of all changes, this won't work. Pushing hotfixes back to hgdev will show up in SQL Source Control as being different than DEV SQL Server and therefore we need to overwrite the reposistory with the "change" from the DEV SQL Server. My only workaround so far is to just have OPS assign a developer the hotfix ticket with a script attached and then we run their hotfixes against DEV ourselves to merge them back in. I'm not happy with that solution. Other than working faster to get to dedicated environment, are they other ways to keep this loop going automatically?

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  • Oracle SOA Security for OUAF Web Services

    - by Anthony Shorten
    With the ability to use Oracle SOA Suite 11g with the Oracle Utilities Application Framework based products, an additional consideration needs to be configured to ensure correct integration. That additional consideration is security. By default, SOA Suite propagates any credentials from the calling application through to the interfacing applications. In most cases, this behavior is not appropriate as the calling application may use different credential stores and also some interfaces are “disconnected” from a calling application (for example, a file based load using the File Adapter). These situations require that the Web Service calls to the Oracle Utilities Application Framework based products have their own valid credentials. To do this the credentials must be attached at design time or at run time to provide the necessary credentials for the call. There are a number of techniques that can be used to do this: At design time, when integrating a Web Service from an Oracle Utilities Application Framework based product you can attach the security policy “oracle/wss_username_token_client_policy” in the composite.xml view. In this view select the Web Service you want to attach the policy to and right click to display the context menu and select “Configure WS Policies” and select the above policy from the list. If you are using SSL then you can use “oracle/wss_username_token_over_ssl_client_policy” instead. At design time, you can also specify the credential key (csf-key) associated with the above policy by selecting the policy and clicking “Edit Config Override Properties”. You name the key appropriately. Everytime the SOA components are deployed the credential configuration is also sent. You can also do this after deployment, or what I call at “runtime”, by specifying the policy and credential key in the Fusion Middleware Control. Refer to the Fusion Middleware Control documentation on how to do this. To complete the configuration you need to add a map and the key specified earlier to the credential store in the Oracle WebLogic instance used for Oracle SOA Suite. From Fusion Middleware Control, you do this by selecting the domain the SOA Suite is installed in a select “Credentials” from the context menu. You now need to add the credentials by adding the map “oracle.wsm.security” (the name is IMPORTANT) and creating a key with the necessary valid credentials. The example below added a key called “mdm.key”. The name I used is for example only. You can name the key anything you like as long as it corresponds to the key you specified in the design time component. Note: I used SYSUSER as an example credentials in the example, in real life you would use another credential as SYSUSER is not appropriate for production use. This key can be reused for other Oracle Utilities Application Framework Web Service integrations or you can use other keys for individual Web Service calls. Once the key is created and the SOA Suite components deployed the transactions should be able to be called as necessary. If you need to change the password for the credentials it can be done using the Fusion Middleware Control functionality.

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  • Fluid VS Responsive Website Development Questions

    - by Aditya P
    As I understand these form the basis for targeting a wide array of devices based on the browser size, given it would be a time consuming to generate different layouts targeting different/specific devices and their resolutions. Questions: Firstly right to the jargon, is there any actual difference between the two or do they mean the same? Is it safe to classify the current development mainly a html5/css3 based one? What popular frameworks are available to easily implement this? What testing methods used in this regard? What are the most common compatibility issues in terms of different browser types? I understand there are methods like this http://css-tricks.com/resolution-specific-stylesheets/ which does this come under?. Are there any external browser detection methods besides the API calls specific to the browser that are employed in this regard? Points of interest [Prior Research before asking these questions] Why shouldn't "responsive" web design be a consideration? Responsive Web Design Tips, Best Practices and Dynamic Image Scaling Techniques A recent list of tutorials 30 Responsive Web Design and Development Tutorials by Eric Shafer on May 14, 2012 Update Ive been reading that the basic point of designing content for different layouts to facilitate a responsive web design is to present the most relevant information. now obviously between the smallest screen width and the highest we are missing out on design elements. I gather from here http://flashsolver.com/2012/03/24/5-top-commercial-responsive-web-designs/ The top of the line design layouts (widths) are desktop layout (980px) tablet layout (768px) smartphone layout – landscape (480px) smartphone layout – portrait (320px) Also we have a popular responsive website testing site http://resizemybrowser.com/ which lists different screen resolutions. I've also come across this while trying to find out the optimal highest layout size to account for http://stackoverflow.com/questions/10538599/default-web-page-width-1024px-or-980px which brings to light seemingly that 1366x768 is a popular web resolution. Is it safe to assume that just accounting for proper scaling from width 980px onwards to the maximum size would be sufficient to accommodate this? given we aren't presenting any new information for the new size. Does it make sense to have additional information ( which conflicts with purpose of responsive web design) to utilize the top size and beyond?

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  • WIF, ADFS 2 and WCF&ndash;Part 3: ADFS Setup

    - by Your DisplayName here!
    In part 1 of this series I briefly gave an overview of the ADFS / WS-Trust infrastructure. In part 2 we created a basic WCF service that uses ADFS for authentication. This part will walk you through the steps to register the service in ADFS 2. I could provide screenshots for all the wizard pages here – but since this is really easy – I just go through the necessary steps in textual form. Step 1 – Select Data Source Here you can decide if you want to import a federation metadata file that describes the service you want to register. In that case all necessary information is inside the metadata document and you are done. FedUtil (a tool that ships with WIF) can generate such metadata for the most simple cases. Another tool to create metadata can be found here. We choose ‘Manual’ here. Step 2 – Specify Display Name I guess that’s self explaining. Step 3 – Choose Profile Choose ‘ADFS 2 Profile’ here. Step 4 – Configure Certificate Remember that we specified a certificate (or rather a private key) to be used to decrypting incoming tokens in the previous post. Here you specify the corresponding public key that ADFS 2 should use for encrypting the token. Step 5 – Configure URL This page is used to configure WS-Federation and SAML 2.0p support. Since we are using WS-Trust you can leave both boxes unchecked. Step 6 – Configure Identifier Here you specify the identifier (aka the realm, aka the appliesTo) that will be used to request tokens for the service. This value will be used in the token request and is used by ADFS 2 to make a connection to the relying party configuration and claim rules. Step 7 – Configure Issuance Authorization Rules Here you can configure who is allowed to request token for the service. I won’t go into details here how these rules exactly work – that’s for a separate blog post. For now simply use the “Permit all users” option. OK – that’s it. The service is now registered at ADFS 2. In the next part we will finally look at the service client. Stay tuned…

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  • Manage Your Favorite Social Accounts in Chrome and Iron with Seesmic

    - by Asian Angel
    Are you looking for a way to manage your Twitter, Facebook, Google Buzz, LinkedIn, and Foursquare accounts all in one place? Using the Seesmic Web App for Chrome and Iron you can access your favorite accounts and manage them in a single, simple-to-use interface. A feature that we loved from the start was the ability to access Twitter without creating a special Seesmic account. And in these days of multiple accounts who needs another one to complicate things up? All that you need to do is to sign in with your user name/e-mail along with your password. You do have to authorize access for Seesmic to connect with your account but the whole process (login & authorization) is handled in a single window instance. Now on to a quick look at some of the UI features… The sidebar allows you to add additional columns to the main interface, set your favorite location for Trends, and tie in additional social services as desired. You can also access additional options and controls in the upper right corner. When you are ready to start tweeting click in the blank at the top and enter your text, etc. in the convenient drop-down window that appears. Another nice perk is the ability to switch to a black and grey theme if the white is too bright for your needs. The Seesmic web app provides a simple-to-use, highly efficient way to manage your Twitter account and other favorite social services in a single tab interface. Seesmic [Chrome Web Store] Latest Features How-To Geek ETC Should You Delete Windows 7 Service Pack Backup Files to Save Space? What Can Super Mario Teach Us About Graphics Technology? Windows 7 Service Pack 1 is Released: But Should You Install It? How To Make Hundreds of Complex Photo Edits in Seconds With Photoshop Actions How to Enable User-Specific Wireless Networks in Windows 7 How to Use Google Chrome as Your Default PDF Reader (the Easy Way) Manage Your Favorite Social Accounts in Chrome and Iron with Seesmic E.T. II – Extinction [Fake Movie Sequel Video] Remastered King’s Quest Games Offer Classic Gaming on Modern Machines Compare Your Internet Cost and Speed to Global Averages [Infographic] Orbital Battle for Terra Wallpaper WizMouse Enables Mouse Over Scrolling on Any Window

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  • How to repair an external harddrive?

    - by dodohjk
    I would like to reformat my hard disk, and if possible recover the (somewhat unimportant) contents if possible. I have a Western Digital 1TB hard drive which had a NTFS partition. I unplugged the drive without safely removing it first. At first a pop up was asking me to use a Windows OS to run the chkdsk /f command, however, in the effort to keep using a Linux OS I used the ntfsfix command on the ubuntu terminal Now, when I try to access the hard drive, it doesn't show up anymore in Nautilus. I tried reformatting it using Disk Utility, but it gives me an error message, and Gparted would hang on the "Scanning devices" step infinitely. Please comment any output that you would like to see and I will add it to my question. EDIT disk utility tells me is on /dev/sdb the command sudo fdisk -l gives dodohjk@DodosPC:~$ sudo fdisk -l [sudo] password for dodohjk: Disk /dev/sda: 250.1 GB, 250059350016 bytes 255 heads, 63 sectors/track, 30401 cylinders, total 488397168 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x0006fa8c Device Boot Start End Blocks Id System /dev/sda1 * 4094 482344959 241170433 5 Extended /dev/sda2 482344960 488396799 3025920 82 Linux swap / Solaris /dev/sda5 4096 31461127 15728516 83 Linux /dev/sda6 31463424 52434943 10485760 83 Linux /dev/sda7 52436992 62923320 5243164+ 83 Linux /dev/sda8 62924800 482344959 209710080 83 Linux Disk /dev/sdb: 1000.2 GB, 1000202043392 bytes 255 heads, 63 sectors/track, 121600 cylinders, total 1953519616 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 512 bytes I/O size (minimum/optimal): 512 bytes / 512 bytes Disk identifier: 0x6e697373 This doesn't look like a partition table Probably you selected the wrong device. Device Boot Start End Blocks Id System /dev/sdb1 ? 1936269394 3772285809 918008208 4f QNX4.x 3rd part /dev/sdb2 ? 1917848077 2462285169 272218546+ 73 Unknown /dev/sdb3 ? 1818575915 2362751050 272087568 2b Unknown /dev/sdb4 ? 2844524554 2844579527 27487 61 SpeedStor Partition table entries are not in disk order I wrote something wrong here, however here the output of fsck /dev/sbd is dodohjk@DodosPC:~$ sudo fsck /dev/sdb fsck from util-linux 2.20.1 e2fsck 1.42.5 (29-Jul-2012) ext2fs_open2: Bad magic number in super-block fsck.ext2: Superblock invalid, trying backup blocks... fsck.ext2: Bad magic number in super-block while trying to open /dev/sdb The superblock could not be read or does not describe a correct ext2 filesystem. If the device is valid and it really contains an ext2 filesystem (and not swap or ufs or something else), then the superblock is corrupt, and you might try running e2fsck with an alternate superblock: e2fsck -b 8193 <device&gt;

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  • Can a Printer Print White?

    - by Jason Fitzpatrick
    The vast majority of the time we all print on white media: white paper, white cardstock, and other neutral white surfaces. But what about printing white? Can modern printers print white and if not, why not? Read on as we explore color theory, printer design choices, and why white is the foundation of the printing process. Today’s Question & Answer session comes to us courtesy of SuperUser—a subdivision of Stack Exchange, a community-driven grouping of Q&A web sites. Image by Coiote O.; available as wallpaper here. The Question SuperUser reader Curious_Kid is well, curious, about printers. He writes: I was reading about different color models, when this question hit my mind. Can the CMYK color model generate white color? Printers use CMYK color mode. What will happen if I try to print a white colored image (rabbit) on a black paper with my printer? Will I get any image on the paper? Does the CMYK color model have room for white? The Answer SuperUser contributor Darth Android offers some insight into the CMYK process: You will not get anything on the paper with a basic CMYK inkjet or laser printer. The CMYK color mixing is subtractive, meaning that it requires the base that is being colored to have all colors (i.e., White) So that it can create color variation through subtraction: White - Cyan - Yellow = Green White - Yellow - Magenta = Red White - Cyan - Magenta = Blue White is represented as 0 cyan, 0 yellow, 0 magenta, and 0 black – effectively, 0 ink for a printer that simply has those four cartridges. This works great when you have white media, as “printing no ink” simply leaves the white exposed, but as you can imagine, this doesn’t work for non-white media. If you don’t have a base color to subtract from (i.e., Black), then it doesn’t matter what you subtract from it, you still have the color Black. [But], as others are pointing out, there are special printers which can operate in the CMYW color space, or otherwise have a white ink or toner. These can be used to print light colors on top of dark or otherwise non-white media. You might also find my answer to a different question about color spaces helpful or informative. Given that the majority of printer media in the world is white and printing pure white on non-white colors is a specialty process, it’s no surprise that home and (most) commercial printers alike have no provision for it. Have something to add to the explanation? Sound off in the the comments. Want to read more answers from other tech-savvy Stack Exchange users? Check out the full discussion thread here.     

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  • C#/.NET Little Wonders: The ConcurrentDictionary

    - by James Michael Hare
    Once again we consider some of the lesser known classes and keywords of C#.  In this series of posts, we will discuss how the concurrent collections have been developed to help alleviate these multi-threading concerns.  Last week’s post began with a general introduction and discussed the ConcurrentStack<T> and ConcurrentQueue<T>.  Today's post discusses the ConcurrentDictionary<T> (originally I had intended to discuss ConcurrentBag this week as well, but ConcurrentDictionary had enough information to create a very full post on its own!).  Finally next week, we shall close with a discussion of the ConcurrentBag<T> and BlockingCollection<T>. For more of the "Little Wonders" posts, see the index here. Recap As you'll recall from the previous post, the original collections were object-based containers that accomplished synchronization through a Synchronized member.  While these were convenient because you didn't have to worry about writing your own synchronization logic, they were a bit too finely grained and if you needed to perform multiple operations under one lock, the automatic synchronization didn't buy much. With the advent of .NET 2.0, the original collections were succeeded by the generic collections which are fully type-safe, but eschew automatic synchronization.  This cuts both ways in that you have a lot more control as a developer over when and how fine-grained you want to synchronize, but on the other hand if you just want simple synchronization it creates more work. With .NET 4.0, we get the best of both worlds in generic collections.  A new breed of collections was born called the concurrent collections in the System.Collections.Concurrent namespace.  These amazing collections are fine-tuned to have best overall performance for situations requiring concurrent access.  They are not meant to replace the generic collections, but to simply be an alternative to creating your own locking mechanisms. Among those concurrent collections were the ConcurrentStack<T> and ConcurrentQueue<T> which provide classic LIFO and FIFO collections with a concurrent twist.  As we saw, some of the traditional methods that required calls to be made in a certain order (like checking for not IsEmpty before calling Pop()) were replaced in favor of an umbrella operation that combined both under one lock (like TryPop()). Now, let's take a look at the next in our series of concurrent collections!For some excellent information on the performance of the concurrent collections and how they perform compared to a traditional brute-force locking strategy, see this wonderful whitepaper by the Microsoft Parallel Computing Platform team here. ConcurrentDictionary – the fully thread-safe dictionary The ConcurrentDictionary<TKey,TValue> is the thread-safe counterpart to the generic Dictionary<TKey, TValue> collection.  Obviously, both are designed for quick – O(1) – lookups of data based on a key.  If you think of algorithms where you need lightning fast lookups of data and don’t care whether the data is maintained in any particular ordering or not, the unsorted dictionaries are generally the best way to go. Note: as a side note, there are sorted implementations of IDictionary, namely SortedDictionary and SortedList which are stored as an ordered tree and a ordered list respectively.  While these are not as fast as the non-sorted dictionaries – they are O(log2 n) – they are a great combination of both speed and ordering -- and still greatly outperform a linear search. Now, once again keep in mind that if all you need to do is load a collection once and then allow multi-threaded reading you do not need any locking.  Examples of this tend to be situations where you load a lookup or translation table once at program start, then keep it in memory for read-only reference.  In such cases locking is completely non-productive. However, most of the time when we need a concurrent dictionary we are interleaving both reads and updates.  This is where the ConcurrentDictionary really shines!  It achieves its thread-safety with no common lock to improve efficiency.  It actually uses a series of locks to provide concurrent updates, and has lockless reads!  This means that the ConcurrentDictionary gets even more efficient the higher the ratio of reads-to-writes you have. ConcurrentDictionary and Dictionary differences For the most part, the ConcurrentDictionary<TKey,TValue> behaves like it’s Dictionary<TKey,TValue> counterpart with a few differences.  Some notable examples of which are: Add() does not exist in the concurrent dictionary. This means you must use TryAdd(), AddOrUpdate(), or GetOrAdd().  It also means that you can’t use a collection initializer with the concurrent dictionary. TryAdd() replaced Add() to attempt atomic, safe adds. Because Add() only succeeds if the item doesn’t already exist, we need an atomic operation to check if the item exists, and if not add it while still under an atomic lock. TryUpdate() was added to attempt atomic, safe updates. If we want to update an item, we must make sure it exists first and that the original value is what we expected it to be.  If all these are true, we can update the item under one atomic step. TryRemove() was added to attempt atomic, safe removes. To safely attempt to remove a value we need to see if the key exists first, this checks for existence and removes under an atomic lock. AddOrUpdate() was added to attempt an thread-safe “upsert”. There are many times where you want to insert into a dictionary if the key doesn’t exist, or update the value if it does.  This allows you to make a thread-safe add-or-update. GetOrAdd() was added to attempt an thread-safe query/insert. Sometimes, you want to query for whether an item exists in the cache, and if it doesn’t insert a starting value for it.  This allows you to get the value if it exists and insert if not. Count, Keys, Values properties take a snapshot of the dictionary. Accessing these properties may interfere with add and update performance and should be used with caution. ToArray() returns a static snapshot of the dictionary. That is, the dictionary is locked, and then copied to an array as a O(n) operation.  GetEnumerator() is thread-safe and efficient, but allows dirty reads. Because reads require no locking, you can safely iterate over the contents of the dictionary.  The only downside is that, depending on timing, you may get dirty reads. Dirty reads during iteration The last point on GetEnumerator() bears some explanation.  Picture a scenario in which you call GetEnumerator() (or iterate using a foreach, etc.) and then, during that iteration the dictionary gets updated.  This may not sound like a big deal, but it can lead to inconsistent results if used incorrectly.  The problem is that items you already iterated over that are updated a split second after don’t show the update, but items that you iterate over that were updated a split second before do show the update.  Thus you may get a combination of items that are “stale” because you iterated before the update, and “fresh” because they were updated after GetEnumerator() but before the iteration reached them. Let’s illustrate with an example, let’s say you load up a concurrent dictionary like this: 1: // load up a dictionary. 2: var dictionary = new ConcurrentDictionary<string, int>(); 3:  4: dictionary["A"] = 1; 5: dictionary["B"] = 2; 6: dictionary["C"] = 3; 7: dictionary["D"] = 4; 8: dictionary["E"] = 5; 9: dictionary["F"] = 6; Then you have one task (using the wonderful TPL!) to iterate using dirty reads: 1: // attempt iteration in a separate thread 2: var iterationTask = new Task(() => 3: { 4: // iterates using a dirty read 5: foreach (var pair in dictionary) 6: { 7: Console.WriteLine(pair.Key + ":" + pair.Value); 8: } 9: }); And one task to attempt updates in a separate thread (probably): 1: // attempt updates in a separate thread 2: var updateTask = new Task(() => 3: { 4: // iterates, and updates the value by one 5: foreach (var pair in dictionary) 6: { 7: dictionary[pair.Key] = pair.Value + 1; 8: } 9: }); Now that we’ve done this, we can fire up both tasks and wait for them to complete: 1: // start both tasks 2: updateTask.Start(); 3: iterationTask.Start(); 4:  5: // wait for both to complete. 6: Task.WaitAll(updateTask, iterationTask); Now, if I you didn’t know about the dirty reads, you may have expected to see the iteration before the updates (such as A:1, B:2, C:3, D:4, E:5, F:6).  However, because the reads are dirty, we will quite possibly get a combination of some updated, some original.  My own run netted this result: 1: F:6 2: E:6 3: D:5 4: C:4 5: B:3 6: A:2 Note that, of course, iteration is not in order because ConcurrentDictionary, like Dictionary, is unordered.  Also note that both E and F show the value 6.  This is because the output task reached F before the update, but the updates for the rest of the items occurred before their output (probably because console output is very slow, comparatively). If we want to always guarantee that we will get a consistent snapshot to iterate over (that is, at the point we ask for it we see precisely what is in the dictionary and no subsequent updates during iteration), we should iterate over a call to ToArray() instead: 1: // attempt iteration in a separate thread 2: var iterationTask = new Task(() => 3: { 4: // iterates using a dirty read 5: foreach (var pair in dictionary.ToArray()) 6: { 7: Console.WriteLine(pair.Key + ":" + pair.Value); 8: } 9: }); The atomic Try…() methods As you can imagine TryAdd() and TryRemove() have few surprises.  Both first check the existence of the item to determine if it can be added or removed based on whether or not the key currently exists in the dictionary: 1: // try add attempts an add and returns false if it already exists 2: if (dictionary.TryAdd("G", 7)) 3: Console.WriteLine("G did not exist, now inserted with 7"); 4: else 5: Console.WriteLine("G already existed, insert failed."); TryRemove() also has the virtue of returning the value portion of the removed entry matching the given key: 1: // attempt to remove the value, if it exists it is removed and the original is returned 2: int removedValue; 3: if (dictionary.TryRemove("C", out removedValue)) 4: Console.WriteLine("Removed C and its value was " + removedValue); 5: else 6: Console.WriteLine("C did not exist, remove failed."); Now TryUpdate() is an interesting creature.  You might think from it’s name that TryUpdate() first checks for an item’s existence, and then updates if the item exists, otherwise it returns false.  Well, note quite... It turns out when you call TryUpdate() on a concurrent dictionary, you pass it not only the new value you want it to have, but also the value you expected it to have before the update.  If the item exists in the dictionary, and it has the value you expected, it will update it to the new value atomically and return true.  If the item is not in the dictionary or does not have the value you expected, it is not modified and false is returned. 1: // attempt to update the value, if it exists and if it has the expected original value 2: if (dictionary.TryUpdate("G", 42, 7)) 3: Console.WriteLine("G existed and was 7, now it's 42."); 4: else 5: Console.WriteLine("G either didn't exist, or wasn't 7."); The composite Add methods The ConcurrentDictionary also has composite add methods that can be used to perform updates and gets, with an add if the item is not existing at the time of the update or get. The first of these, AddOrUpdate(), allows you to add a new item to the dictionary if it doesn’t exist, or update the existing item if it does.  For example, let’s say you are creating a dictionary of counts of stock ticker symbols you’ve subscribed to from a market data feed: 1: public sealed class SubscriptionManager 2: { 3: private readonly ConcurrentDictionary<string, int> _subscriptions = new ConcurrentDictionary<string, int>(); 4:  5: // adds a new subscription, or increments the count of the existing one. 6: public void AddSubscription(string tickerKey) 7: { 8: // add a new subscription with count of 1, or update existing count by 1 if exists 9: var resultCount = _subscriptions.AddOrUpdate(tickerKey, 1, (symbol, count) => count + 1); 10:  11: // now check the result to see if we just incremented the count, or inserted first count 12: if (resultCount == 1) 13: { 14: // subscribe to symbol... 15: } 16: } 17: } Notice the update value factory Func delegate.  If the key does not exist in the dictionary, the add value is used (in this case 1 representing the first subscription for this symbol), but if the key already exists, it passes the key and current value to the update delegate which computes the new value to be stored in the dictionary.  The return result of this operation is the value used (in our case: 1 if added, existing value + 1 if updated). Likewise, the GetOrAdd() allows you to attempt to retrieve a value from the dictionary, and if the value does not currently exist in the dictionary it will insert a value.  This can be handy in cases where perhaps you wish to cache data, and thus you would query the cache to see if the item exists, and if it doesn’t you would put the item into the cache for the first time: 1: public sealed class PriceCache 2: { 3: private readonly ConcurrentDictionary<string, double> _cache = new ConcurrentDictionary<string, double>(); 4:  5: // adds a new subscription, or increments the count of the existing one. 6: public double QueryPrice(string tickerKey) 7: { 8: // check for the price in the cache, if it doesn't exist it will call the delegate to create value. 9: return _cache.GetOrAdd(tickerKey, symbol => GetCurrentPrice(symbol)); 10: } 11:  12: private double GetCurrentPrice(string tickerKey) 13: { 14: // do code to calculate actual true price. 15: } 16: } There are other variations of these two methods which vary whether a value is provided or a factory delegate, but otherwise they work much the same. Oddities with the composite Add methods The AddOrUpdate() and GetOrAdd() methods are totally thread-safe, on this you may rely, but they are not atomic.  It is important to note that the methods that use delegates execute those delegates outside of the lock.  This was done intentionally so that a user delegate (of which the ConcurrentDictionary has no control of course) does not take too long and lock out other threads. This is not necessarily an issue, per se, but it is something you must consider in your design.  The main thing to consider is that your delegate may get called to generate an item, but that item may not be the one returned!  Consider this scenario: A calls GetOrAdd and sees that the key does not currently exist, so it calls the delegate.  Now thread B also calls GetOrAdd and also sees that the key does not currently exist, and for whatever reason in this race condition it’s delegate completes first and it adds its new value to the dictionary.  Now A is done and goes to get the lock, and now sees that the item now exists.  In this case even though it called the delegate to create the item, it will pitch it because an item arrived between the time it attempted to create one and it attempted to add it. Let’s illustrate, assume this totally contrived example program which has a dictionary of char to int.  And in this dictionary we want to store a char and it’s ordinal (that is, A = 1, B = 2, etc).  So for our value generator, we will simply increment the previous value in a thread-safe way (perhaps using Interlocked): 1: public static class Program 2: { 3: private static int _nextNumber = 0; 4:  5: // the holder of the char to ordinal 6: private static ConcurrentDictionary<char, int> _dictionary 7: = new ConcurrentDictionary<char, int>(); 8:  9: // get the next id value 10: public static int NextId 11: { 12: get { return Interlocked.Increment(ref _nextNumber); } 13: } Then, we add a method that will perform our insert: 1: public static void Inserter() 2: { 3: for (int i = 0; i < 26; i++) 4: { 5: _dictionary.GetOrAdd((char)('A' + i), key => NextId); 6: } 7: } Finally, we run our test by starting two tasks to do this work and get the results… 1: public static void Main() 2: { 3: // 3 tasks attempting to get/insert 4: var tasks = new List<Task> 5: { 6: new Task(Inserter), 7: new Task(Inserter) 8: }; 9:  10: tasks.ForEach(t => t.Start()); 11: Task.WaitAll(tasks.ToArray()); 12:  13: foreach (var pair in _dictionary.OrderBy(p => p.Key)) 14: { 15: Console.WriteLine(pair.Key + ":" + pair.Value); 16: } 17: } If you run this with only one task, you get the expected A:1, B:2, ..., Z:26.  But running this in parallel you will get something a bit more complex.  My run netted these results: 1: A:1 2: B:3 3: C:4 4: D:5 5: E:6 6: F:7 7: G:8 8: H:9 9: I:10 10: J:11 11: K:12 12: L:13 13: M:14 14: N:15 15: O:16 16: P:17 17: Q:18 18: R:19 19: S:20 20: T:21 21: U:22 22: V:23 23: W:24 24: X:25 25: Y:26 26: Z:27 Notice that B is 3?  This is most likely because both threads attempted to call GetOrAdd() at roughly the same time and both saw that B did not exist, thus they both called the generator and one thread got back 2 and the other got back 3.  However, only one of those threads can get the lock at a time for the actual insert, and thus the one that generated the 3 won and the 3 was inserted and the 2 got discarded.  This is why on these methods your factory delegates should be careful not to have any logic that would be unsafe if the value they generate will be pitched in favor of another item generated at roughly the same time.  As such, it is probably a good idea to keep those generators as stateless as possible. Summary The ConcurrentDictionary is a very efficient and thread-safe version of the Dictionary generic collection.  It has all the benefits of type-safety that it’s generic collection counterpart does, and in addition is extremely efficient especially when there are more reads than writes concurrently. Tweet Technorati Tags: C#, .NET, Concurrent Collections, Collections, Little Wonders, Black Rabbit Coder,James Michael Hare

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  • SQL SERVER – Difference Between CURRENT_TIMESTAMP and GETDATE() – CURRENT_TIMESTAMP Equivalent in SQL Server

    - by pinaldave
    A common question – I often get from Oracle/MySQL Professionals: “What is the Equivalent to CURRENT_TIMESTAMP in SQL Server?” Here is a common question I often get from SQL Server Professionals: “What are differences between Difference Between CURRENT_TIMESTAMP and GETDATE ()?” Very simple question but have showed up so frequently that I feel like to write about it. Well in SQL Server GETDATE() is Equivalent to CURRENT_TIMESTAMP. However, if you use CURRENT_TIMESTAMP in your select statement it will work fine. You can see in the above example – both of them returns the same value. Now let us go to next question regarding difference between GETDATE and CURRENT_TIMESTAMP. Well, the matter of the fact, there is no difference between them in SQL Server (Reference Link). CURRENT_TIMESTAMP is an ANSI SQL function, whereas GETDATE is T-SQL implementation of the same function. Both of them derive value from the operating system of the computer on which SQL Server instance is running. Above discussion prompts another question – in this case, what should one use GETDATE or CURRENT_TIMESTAMP? Well, this is indeed tricky and interesting question. I think I am very comfortable using the GETDATE () so I will go to use it but a matter of the fact there is no right or wrong answer. If you want to follow ancient saying “When in Rome, do as the Romans do”, I suggest using the GETDATE (), or continue using CURRENT_TIMESTAMP. With that said, there is one very important property we all need to keep in mind. If you use CURRENT_TIMESTAMP while creating an object, they are automatically converted to GETDATE() and stored internally. To illustrate what I am suggesting here is the example - Create a table using the following script CREATE TABLE [dbo].[TestTable]( [Cold2] [datetime] NULL ) ON [PRIMARY] GO ALTER TABLE [dbo].[TestTable] ADD DEFAULT (CURRENT_TIMESTAMP) FOR [Cold2] GO Now go to SSMS and generate the script for the table and you will notice following syntax. CREATE TABLE [dbo].[TestTable]( [Cold2] [datetime] NULL ) ON [PRIMARY] GO ALTER TABLE [dbo].[TestTable] ADD DEFAULT (GETDATE()) FOR [Cold2] GO You can notice that SQL Server have automatically converted CURRENT_TIMESTAMP to GETDATE(). I guess this gives us an idea how they behave. Now go ahead and make your choice! Do let me know which one will you use CURRENT_TIMESTAMP or GETDATE () in the comments area. Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: PostADay, SQL, SQL Authority, SQL DateTime, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • New Process For Receiving Oracle Certification Exam Results

    - by Brandye Barrington
    On November 15, 2012, Oracle Certification exam results will be available directly from Oracle's certification portal, CertView. After completing an exam at a testing center, you will login to CertView to access and print your exam scores by selecting the See My New Exam Results Now link or the Print My New Exam Results Now link from the homepage. This will provide access to all certification and exam history in one place through Oracle, providing tighter integration with other activities at Oracle. This change in policy will also increase security around data privacy. AUTHENTICATE YOUR CERTVIEW ACCOUNT NOW One very important step you must take is to authenticate your CertView account BEFORE taking your exam. This way, if there are any issues with authorization, you have time to get these sorted out before testing. Keep in mind that it can take up to 3 business days for a CertView account to be manually authenticated, so completing this process before testing is key! You will need to create a web account at PearsonVUE prior to registering for your exam and you will need to create an Oracle Web Account prior to authenticating your CertView account. The CertView account will be available for authentication within 30 minutes of creating a Pearson VUE web account at certview.oracle.com. GETTING YOUR EXAM RESULTS FROM ORACLE Before taking the scheduled exam, you should authenticate your account at certview.oracle.com using the email address and Oracle Testing ID in your Pearson VUE profile. You will be required to have an Oracle Web Account to authenticate your CertView account. After taking the exam, you will receive an email from Oracle indicating that your exam results are available at certview.oracle.com If you have previously authenticated your CertView account, you will simply click on the link in the email, which will take you to CertView, login and select See My New Exam Results Now. If you have not authenticated your CertView account before receiving this notification email, you will be required to authenticate your CertView account before accessing your exam results. Authentication requires an Oracle Web Account user name and password and the following information from your Pearson VUE profile: email address and Oracle Testing ID. Click on the link in the email to authenticate your CertView account You will be given the option to create an Oracle Web Account if you do no already have one.  After account authentication, you will be able to login to CertView and select See My New Exam Results Now to view your exam results or Print My New Exam Results Now to print your exam results. As always, if you need assistance with your CertView account, please contact Oracle Certification Support. YOUR QUESTIONS ANSWERED More Information FAQ: Receiving Exam Scores FAQ: How Do I Log Into CertView? FAQ: How To Get Exam Results FAQ: Accessing Exam Results in CertView FAQ: How Will I Know When My Exam Results Are Available? FAQ: What If I Don't Get An Exam Results Email Alert? FAQ: How To Download and Print Exam Score Reports FAQ: What If I Think My Exam Results Are Wrong In CertView? FAQ: Is Oracle Changing The Way That Exams Are Scored?

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  • Unable to update/ install any files [closed]

    - by Surya
    Possible Duplicate: “Problem with MergeList” error when trying to do an update Just now I installed ubuntu 12.04 on my Lenovo G570 laptop. First I got an error at the time of installation (don't know about it) and I restarted the system and next time, it went well. So, after installing problems started.. There was a error with "Language recognition" and I tried to fix it but didn't work. I tried to install powerTop to check the status of power management. at terminal: sudo apt-get install powertop This is the error I got surya@surya-Lenovo-G570:~$ sudo apt-get powertop install [sudo] password for surya: E: Invalid operation powertop surya@surya-Lenovo-G570:~$ sudo apt-get install powertop Reading package lists... Error! E: Encountered a section with no Package: header E: Problem with MergeList /var/lib/apt/lists/extras.ubuntu.com_ubuntu_dists_precise_main_binary-i386_Packages E: The package lists or status file could not be parsed or opened. surya@surya-Lenovo-G570:~$ ^C surya@surya-Lenovo-G570:~$ ^C surya@surya-Lenovo-G570:~$ ^C surya@surya-Lenovo-G570:~$ I downloaded Google Chrome .deb one and tried to install but its not working. Software center is opened and its not loading. There was a notification on the status bar which says: An error occurred please run the package manager from the right-click menu ... .... ... E: Encountered a section with no Package: header E: Problem with MergeList /var/lib/apt/lists/extras.ubuntu.com_ubuntu_dists_precise_main_binary-i386_Packages "Copy & Paste" from terminal is not really working... When I press Ctrl + C; its showing ^C on terminal but its not working.. The most important error: I am unable to see a "chip" icon on the status bar so as to install proprietary drivers for my ATI drivers... The interesting part is, powertop worked will on live cd and it even detected my ATI card. Update When I opened "Software Up to Date", this showed a error: Could not initialize the package information An unresolvable problem occurred while initializing the package information. Please report this bug against the 'update-manager' package and include the following error message: 'E:Encountered a section with no Package: header, E:Problem with MergeList /var/lib/apt/lists/extras.ubuntu.com_ubuntu_dists_precise_main_binary-i386_Packages, E:The package lists or status file could not be parsed or opened.' : My laptop details Lenovo G570; Intel 2nd Gen i5 processor 4GB DDR3 RAM Intel in-build graphics + AMD Radeon HD 6370M 1GB graphics. I need help ASAP.

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  • C# XNA Normals Question

    - by Wade
    Hello all! I have been working on some simple XNA proof of concept for a game idea I have as well as just to further my learning in XNA. However, i seem to be stuck on these dreaded normals, and using the BasicEffect with default lighting i can't seem to tell if my normals are being calculated correctly, hence the question. I'm mainly drawing cubes at the moment, I'm using a triangle list and a VertexBuffer to get the job done. The north face of my cube has two polygons and 6 vectors: Vector3 startPosition = new Vector3(0,0,0); corners[0] = startPosition; // This is the start position. Block size is 5. corners[1] = new Vector3(startPosition.X, startPosition.Y + BLOCK_SIZE, startPosition.Z); corners[2] = new Vector3(startPosition.X + BLOCK_SIZE, startPosition.Y, startPosition.Z); corners[3] = new Vector3(startPosition.X + BLOCK_SIZE, startPosition.Y + BLOCK_SIZE, startPosition.Z); verts[0] = new VertexPositionNormalTexture(corners[0], normals[0], textCoordBR); verts[1] = new VertexPositionNormalTexture(corners[1], normals[0], textCoordTR); verts[2] = new VertexPositionNormalTexture(corners[2], normals[0], textCoordBL); verts[3] = new VertexPositionNormalTexture(corners[3], normals[0], textCoordTL); verts[4] = new VertexPositionNormalTexture(corners[2], normals[0], textCoordBL); verts[5] = new VertexPositionNormalTexture(corners[1], normals[0], textCoordTR); Using those coordinates I want to generate the normal for the north face, I have no clue how to get the average of all those vectors and create a normal for the two polygons that it makes. Here is what i tried: normals[0] = Vector3.Cross(corners[1], corners[2]); normals[0].Normalize(); It seems like its correct, but then using the same thing for other sides of the cube the lighting effect seems weird, and not cohesive with where i think the light source is coming from, not really sure with the BasicEffect. Am I doing this right? Can anyone explain in lay mans terms how normals are calculated. Any help is much appreciated. Note: I tried going through Riemers and such to figure it out with no luck, it seems no one really goes over the math well enough. Thanks!

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  • Use of Business Parameters in BPM12c

    - by Abhishek Mittal-Oracle
    With the release of BPM12c, a new feature to use Business Parameters is introduced through which we can define a business parameter which will behave as a global variable which can be used within BPM project. Business Administrator can be the one responsible to modify the business parameters value dynamically at run-time which may bring change in BPM process flow where it is used.This feature was a part of BPM10g product and was extensively used. In BPM11g, this feature is not present currently.Business Parameters can be defined in 2 ways:1. Using Jdev to define business parameters, and 2. Using BPM workspace to define business parameters.It is important to note that business parameters need to be mapped with a valid organisation unit defined in a BPM project. If the same is not handled, exceptions like 'BPM-70702' will be thrown by BPM Engine. This is because business parameters work along with organisation defined in a BPM project.At the same time, we can use same business parameter across different organisation units with different values. Business Parameters in BPM12c has this capability to handle multiple values with different organisation units defined in a single BPM project. This enables business to re-use same business parameters defined in a BPM project across different organisations.Business parameters can be defined using the below data types:1. int2. string 3. boolean4. double While defining an business parameter, it is mandatory to provide a default value. Below are the steps to define a business parameter in Jdev: Step 1:  Open 'Organization' and click on 'Business Parameters' tab.Step 2:  Click on '+' button.Step 3: Add business parameter name, type and provide default value(mandatory).Step 4: Click on 'OK' button.Step 5: Business parameter is defined. Below are the steps to define a business parameter in BPM workspace: Step 1: Login to BPM workspace using admin-username and password.Step 2: Click on 'Administration' on the right top side of workspace.Step 3: Click on 'Business Parameters' in the left navigation panel under 'Organization'. Step 4:  Click on '+' button.Step 5: Add business parameter name, type and provide default value(mandatory).Step 6: Click on 'OK' button.Step 7: Business parameter is defined. Note: As told earlier in the blog, it is necessary to define and map a valid organization ID with predefined variable 'organizationalUnit' under data associations in an BPM process before the business parameter is used. I have created one sample PoC demonstrating the use of Business Parameters in BPM12c and it can be found here.

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  • Mysql 5.5 server not working

    - by rajesh
    I had Ubuntu 14.04 installed on my system. I recently updated ubuntu and now my mysql does not start and workbench says that mysql server has been stopped. And when i try to start it gives me the following error 2014-08-12 23:02:04 - Checking server status... 2014-08-12 23:02:04 - Trying to connect to MySQL... 2014-08-12 23:02:04 - Can't connect to MySQL server on '127.0.0.1' (111) (2003) 2014-08-12 23:02:04 - Assuming server is not running 2014-08-12 23:02:04 - Server start done. 2014-08-12 23:02:04 - Checking server status... 2014-08-12 23:02:04 - Trying to connect to MySQL... 2014-08-12 23:02:04 - Can't connect to MySQL server on '127.0.0.1' (111) (2003) 2014-08-12 23:02:04 - Assuming server is not running And also when i try to login using terminal (mysql -u root -p <password>) i get the following error: ERROR 2002 (HY000): Can't connect to local MySQL server through socket '/var/run/mysqld/mysqld.sock' (2) I have also tried to reinstall Ubuntu but i am unable to do so. Gives me the following error: Reading package lists... Done Building dependency tree Reading state information... Done mysql-server-5.5 is already the newest version. 0 upgraded, 0 newly installed, 0 to remove and 4 not upgraded. I have data which i have not taken backup of as i am unable to log into the server. I am a newbie please help me resolve this issue without losing my data. Awaiting for your earliest response. Below is the error message from cat /var/log/mysql/error.log 140813 21:22:50 [Warning] Using unique option prefix myisam-recover instead of myisam-recover-options is deprecated and will be removed in a future release. Please use the full name instead. 140813 21:22:50 [Note] Plugin 'FEDERATED' is disabled. 140813 21:22:50 InnoDB: The InnoDB memory heap is disabled 140813 21:22:50 InnoDB: Mutexes and rw_locks use GCC atomic builtins 140813 21:22:50 InnoDB: Compressed tables use zlib 1.2.8 140813 21:22:50 InnoDB: Using Linux native AIO 140813 21:22:50 InnoDB: Initializing buffer pool, size = 128.0M 140813 21:22:50 InnoDB: Completed initialization of buffer pool 140813 21:22:50 InnoDB: highest supported file format is Barracuda. 140813 21:22:50 InnoDB: Waiting for the background threads to start 140813 21:22:51 InnoDB: 5.5.38 started; log sequence number 80726593570 140813 21:22:51 [Note] Server hostname (bind-address): '127.0.0.1'; port: 3306 140813 21:22:51 [Note] - '127.0.0.1' resolves to '127.0.0.1'; 140813 21:22:51 [Note] Server socket created on IP: '127.0.0.1'. 140813 21:22:51 [ERROR] Fatal error: Can't open and lock privilege tables: Incorrect file format 'user'

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