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  • Remotely Schedule and Stream Recorded TV in Windows 7 Media Center

    - by DigitalGeekery
    Have you ever been away from home and suddenly realized you forgot to record your favorite program? Now Windows 7 Media Center, users can schedule recordings remotely from their phones or mobile devices with Remote Potato. How it Works Remote Potato installs server software on the host computer running Windows 7 Media Center. Once the software is installed, we’ll need to do some port forwarding on the router and setup an optional dynamic DNS address. When setup is completed, we will access the application through a web based interface. Silverlight is required for Streaming recorded TV, but scheduling recordings can be done through an HTML interface. Installing Remote Potato Download and install Remote Potato on the Media Center PC. (See download link below) If you plan to stream any Recorded TV, you’ll also want to install the streaming pack located on the same page. It isn’t required to stream all shows, only shows that require the AC3 audio codec. Click Yes to allow Remote Potato to add rules to the Windows Firewall for remote access. You’ll likely need to accept a few UAC prompts. When notified that the rules were added, click OK. Remote Potato will then prompt you to allow administrator privileges to reserve a URL for it’s web server. Click Yes. Remote Potato server will start. Click on the configuration button at the right to to reveal the settings tabs.   One the General tab, you’ll have the option to run Remote Potato on startup and minimized in the System Tray. If you’re running Media Center on a dedicated HTPC, you’ll probably want to enable both startup options. Forwarding Ports on Your Router You’ll need to forward a couple ports on your router. By default, these will be ports 9080 and 9081. In this example we’re using a Linksys WRT54GL router, however, the steps for port forwarding will vary from router to router. On the Linksys configuration page, click on the Applications & Gaming Tab, and then the Port Range Forward tab. Under Application, type in a name of your choosing. In both the Start and End boxes, type the port number 9080. Enter the local IP address of your Media Center computer in the IP address column. Click the check box under Enable. Repeat the process on the next line, but this time use port 9081. When finished, click the Save Settings button. Note: It’s highly recommended that you configure the home computer running Media Center & Remote Potato with a static IP address.   Find your IP Address You’ll need to find the IP address assigned to your router from your ISP. There are many ways to do this but a quick and easy way is to visit a site like checkip.dyndns.org (link available below) The current external IP address of your router will be displayed in the browser.   Dynamic DNS This is an optional step, but  it’s highly recommended. Many routers, such as the Linksys WRT54GL we are using, support Dynamic DNS (DDNS). What Dynamic DNS allows you to do is affiliate your home router’s external IP address to a domain name. Every time your home router is assigned a a new IP address by your ISP, the domain name is updated to point to your new IP address. Remote Potato’s user interface is accessed over the Internet is by connecting to your router’s IP address followed by a colon and the port number. (Ex: XXX.XXX.XXX.XXX:9080) Instead of constantly having to look up and remember an IP address, you can use DDNS along with a 3rd party provider like DynDNS.com, to sign up for a free domain name and configure it to be updated each time your router is assigned a new IP address. Go to the DynDNS.com website (See link at the end of the article) and sign up for a free Domain name. You’ll need to register and confirm by email.   Once you’ve signed in and selected your domain name click Activate Services. You’ll get a confirmation message that your domain name has been activated.    On the Linksys WRT54GL click on the Setup tab an then DDNS. Select DynDNS.org, or TZO.com if you prefer to use their service, from the drop down list.   With DynDNS, you’ll need to fill in your username and password you signed up with at the DynDNS website and the hostname you chose. Note: You can connect over your local network with the IP Address of the computer running Remote Potato followed by a colon and the port number. Ex: 192.168.1.2:9080 Logging in Remote Potato and Recording a Show Once you connect, you’ll see the start page. To view the TV listings, click on TV Guide. You’ll then see your guide listings. There are a few ways to navigate the listings. At the top left, you can click on any of the preset time buttons to jump to  the listings at that time of the day.  Click on the arrows to the right and left of the day and date at the top center to proceed to the previous or next day. Or, jump to a specific day with the date and date buttons at the top right.   To setup a recording, click on a program.   You can choose to record the individual show or the entire series by clicking on Record Show or Record Series.   Remote Potato on Mobile Devices Perhaps the coolest feature of Remote Potato is the ability to schedule recording from your phone or mobile device. Note: For any devices or computers without Silverlight, you will be prompted to view the HTML page. Select Browse Listings. Select your program to record. In the Program Details, select Record Show to record the single episode or Record Series to record all instances of the series. You will then see a red dot on the program listing to indicate that the show is scheduled for recording.   Streaming Recorded TV Click on Recorded TV from the home screen to access your previously recorded TV programs. Click on the selection you wish to stream. Click on Play. If you receive this error message, you’ll need to install the streaming pack for Remote Potato. This is found on the same download page as installation files. (See link below) The Begin from slider allows you to start playback from the start (by default) or a different time of the program by moving the slider. The Quality (bitrate) setting  allows you to choose the quality of the playback. We found the video quality on the Normal setting to be pretty lousy, and Low was just pointless. High was the best overall viewing experience as it provided smooth quality video playback. We experienced significant stuttering during playback using the Ultra High setting.   Click Start when you are ready to begin. When playback begins you’ll see a slider at the top right.   Move the slider left or right to increase or decrease the size of the video. There’s also a button to switch to full screen.   Media Center users who travel frequently or are always on the go will likely find Remote Potato to be a blessing. Since being released earlier this year, updates for Remote Potato have come fast and furious. The latest beta release includes support for streaming music and photos. If you like those nice network TV logos, check out our article on adding TV channel logos to Windows Media Center. Downloads and Links Download Remote Potato and Streaming Pack Find your IP address Sign Up for a Domain Name at DynDNS.com Similar Articles Productive Geek Tips Schedule Updates for Windows Media CenterUsing Netflix Watchnow in Windows Vista Media Center (Gmedia)Add a Sleep Timer to Windows 7 Media CenterStartup Customizations for Media Center in Windows 7Enable Media Streaming in Windows Home Server to Windows Media Player 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 FoxClocks adds World Times in your Statusbar (Firefox) Have Fun Editing Photo Editing with Citrify Outlook Connector Upgrade Error Gadfly is a cool Twitter/Silverlight app Enable DreamScene in Windows 7 Microsoft’s “How Do I ?” Videos

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  • First round playing with Memcached

    - by Shaun
    To be honest I have not been very interested in the caching before I’m going to a project which would be using the multi-site deployment and high connection and concurrency and very sensitive to the user experience. That means we must cache the output data for better performance. After looked for the Internet I finally focused on the Memcached. What’s the Memcached? I think the description on its main site gives us a very good and simple explanation. Free & open source, high-performance, distributed memory object caching system, generic in nature, but intended for use in speeding up dynamic web applications by alleviating database load. Memcached is an in-memory key-value store for small chunks of arbitrary data (strings, objects) from results of database calls, API calls, or page rendering. Memcached is simple yet powerful. Its simple design promotes quick deployment, ease of development, and solves many problems facing large data caches. Its API is available for most popular languages. The original Memcached was built on *nix system are is being widely used in the PHP world. Although it’s not a problem to use the Memcached installed on *nix system there are some windows version available fortunately. Since we are WISC (Windows – IIS – SQL Server – C#, which on the opposite of LAMP) it would be much easier for us to use the Memcached on Windows rather than *nix. I’m using the Memcached Win X64 version provided by NorthScale. There are also the x86 version and other operation system version.   Install Memcached Unpack the Memcached file to a folder on the machine you want it to be installed, we can see that there are only 3 files and the main file should be the “memcached.exe”. Memcached would be run on the server as a service. To install the service just open a command windows and navigate to the folder which contains the “memcached.exe”, let’s say “C:\Memcached\”, and then type “memcached.exe -d install”. If you are using Windows Vista and Windows 7 system please be execute the command through the administrator role. Right-click the command item in the start menu and use “Run as Administrator”, otherwise the Memcached would not be able to be installed successfully. Once installed successful we can type “memcached.exe -d start” to launch the service. Now it’s ready to be used. The default port of Memcached is 11211 but you can change it through the command argument. You can find the help by typing “memcached -h”.   Using Memcached Memcahed has many good and ready-to-use providers for vary program language. After compared and reviewed I chose the Memcached Providers. It’s built based on another 3rd party Memcached client named enyim.com Memcached Client. The Memcached Providers is very simple to set/get the cached objects through the Memcached servers and easy to be configured through the application configuration file (aka web.config and app.config). Let’s create a console application for the demonstration and add the 3 DLL files from the package of the Memcached Providers to the project reference. Then we need to add the configuration for the Memcached server. Create an App.config file and firstly add the section on top of it. Here we need three sections: the section for Memcached Providers, for enyim.com Memcached client and the log4net. 1: <configSections> 2: <section name="cacheProvider" 3: type="MemcachedProviders.Cache.CacheProviderSection, MemcachedProviders" 4: allowDefinition="MachineToApplication" 5: restartOnExternalChanges="true"/> 6: <sectionGroup name="enyim.com"> 7: <section name="memcached" 8: type="Enyim.Caching.Configuration.MemcachedClientSection, Enyim.Caching"/> 9: </sectionGroup> 10: <section name="log4net" 11: type="log4net.Config.Log4NetConfigurationSectionHandler,log4net"/> 12: </configSections> Then we will add the configuration for 3 of them in the App.config file. The Memcached server information would be defined under the enyim.com section since it will be responsible for connect to the Memcached server. Assuming I installed the Memcached on two servers with the default port, the configuration would be like this. 1: <enyim.com> 2: <memcached> 3: <servers> 4: <!-- put your own server(s) here--> 5: <add address="192.168.0.149" port="11211"/> 6: <add address="10.10.20.67" port="11211"/> 7: </servers> 8: <socketPool minPoolSize="10" maxPoolSize="100" connectionTimeout="00:00:10" deadTimeout="00:02:00"/> 9: </memcached> 10: </enyim.com> Memcached supports the multi-deployment which means you can install the Memcached on the servers as many as you need. The protocol of the Memcached responsible for routing the cached objects into the proper server. So it’s very easy to scale-out your system by Memcached. And then define the Memcached Providers configuration. The defaultExpireTime indicates how long the objected cached in the Memcached would be expired, the default value is 2000 ms. 1: <cacheProvider defaultProvider="MemcachedCacheProvider"> 2: <providers> 3: <add name="MemcachedCacheProvider" 4: type="MemcachedProviders.Cache.MemcachedCacheProvider, MemcachedProviders" 5: keySuffix="_MySuffix_" 6: defaultExpireTime="2000"/> 7: </providers> 8: </cacheProvider> The last configuration would be the log4net. 1: <log4net> 2: <!-- Define some output appenders --> 3: <appender name="ConsoleAppender" type="log4net.Appender.ConsoleAppender"> 4: <layout type="log4net.Layout.PatternLayout"> 5: <conversionPattern value="%date [%thread] %-5level %logger [%property{NDC}] - %message%newline"/> 6: </layout> 7: </appender> 8: <!--<threshold value="OFF" />--> 9: <!-- Setup the root category, add the appenders and set the default priority --> 10: <root> 11: <priority value="WARN"/> 12: <appender-ref ref="ConsoleAppender"> 13: <filter type="log4net.Filter.LevelRangeFilter"> 14: <levelMin value="WARN"/> 15: <levelMax value="FATAL"/> 16: </filter> 17: </appender-ref> 18: </root> 19: </log4net>   Get, Set and Remove the Cached Objects Once we finished the configuration it would be very simple to consume the Memcached servers. The Memcached Providers gives us a static class named DistCache that can be used to operate the Memcached servers. Get<T>: Retrieve the cached object from the Memcached servers. If failed it will return null or the default value. Add: Add an object with a unique key into the Memcached servers. Assuming that we have an operation that retrieve the email from the name which is time consuming. This is the operation that should be cached. The method would be like this. I utilized Thread.Sleep to simulate the long-time operation. 1: static string GetEmailByNameSlowly(string name) 2: { 3: Thread.Sleep(2000); 4: return name + "@ethos.com.cn"; 5: } Then in the real retrieving method we will firstly check whether the name, email information had been searched previously and cached. If yes we will just return them from the Memcached, otherwise we will invoke the slowly method to retrieve it and then cached. 1: static string GetEmailByName(string name) 2: { 3: var email = DistCache.Get<string>(name); 4: if (string.IsNullOrEmpty(email)) 5: { 6: Console.WriteLine("==> The name/email not be in memcached so need slow loading. (name = {0})==>", name); 7: email = GetEmailByNameSlowly(name); 8: DistCache.Add(name, email); 9: } 10: else 11: { 12: Console.WriteLine("==> The name/email had been in memcached. (name = {0})==>", name); 13: } 14: return email; 15: } Finally let’s finished the calling method and execute. 1: static void Main(string[] args) 2: { 3: var name = string.Empty; 4: while (name != "q") 5: { 6: Console.Write("==> Please enter the name to find the email: "); 7: name = Console.ReadLine(); 8:  9: var email = GetEmailByName(name); 10: Console.WriteLine("==> The email of {0} is {1}.", name, email); 11: } 12: } The first time I entered “ziyanxu” it takes about 2 seconds to get the email since there’s nothing cached. But the next time I entered “ziyanxu” it returned very quickly from the Memcached.   Summary In this post I explained a bit on why we need cache, what’s Memcached and how to use it through the C# application. The example is fairly simple but hopefully demonstrated on how to use it. Memcached is very easy and simple to be used since it gives you the full opportunity to consider what, when and how to cache the objects. And when using Memcached you don’t need to consider the cache servers. The Memcached would be like a huge object pool in front of you. The next step I’m thinking now are: What kind of data should be cached? And how to determined the key? How to implement the cache as a layer on top of the business layer so that the application will not notice that the cache is there. How to implement the cache by AOP so that the business logic no need to consider the cache. I will investigate on them in the future and will share my thoughts and results.   Hope this helps, Shaun All documents and related graphics, codes are provided "AS IS" without warranty of any kind. Copyright © Shaun Ziyan Xu. This work is licensed under the Creative Commons License.

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  • Add Free Windows Live Apps to Your Website or Blog

    - by Matthew Guay
    Would you like to use Hotmail, Office Web Apps, Messenger, and more on your website domain?  Here’s how you can add Windows Live to your website for free. Microsoft offers a popular suite of online communications products including Hotmail and Messenger.  Although Hotmail hasn’t been as popular in recent years as Gmail, it is getting a refresh this summer that might make it an even better email solution.  Additionally, the new Office Web Apps offer great compatibility with Office documents. While Skydrive offers 25Gb of free online file storage for all users, so Windows Live can make a great communications solution for your domain. Note: To signup for Windows Live for your domain, you will need to be able to add info to your WordPress.com blog or change Domain settings manually. Getting Started Open the Windows Live Custom Domains page (Link below) to get started adding Windows Live to your domain.  Your free Windows Live account will let you create up to 500 accounts, so it’s great for teams and groups that want to have customized email addresses in addition to those who just want an email account for their website. Enter your domain or subdomain you want to add to Windows Live in the box, and then select whether you want to setup Hotmail with this or now.  We want to add email to our domain, so select Set up Windows Live Hotmail for my domain and click Continue. You’ll need to sign in with a Windows Live ID to create the account, or choose to create a new Windows Live account associated with your domain.   Sign in with your Windows Live ID…this can be a Hotmail, Live Messenger, XBOX Live, Zune ID, or Microsoft.com account. Or, enter your information to create a new Windows Live ID if you selected the second option. Now, review your settings and make sure everything looks correct.  Click the I Accept button to setup your account.   Your account is now fully setup, but you’ll need to add or edit DNS information on your site.  The steps are slightly different depending if your site is hosted on WordPress.com, on your own server, or hosting service. We’ll show you how to do it on either one. First, though, note the information below this box.  You’ll see settings for your Mail setup…   Security settings…   And Messenger integration.  Make note of the settings, especially the circled ones, as we’ll need them in the next step. Integrate Windows Live with Your WordPress Blog If the domain you added to Windows Live is for your WordPress blog, login to your WordPress dashboard in a separate browser window or tab.  Click the arrow beside Upgrades, and select Domains from the menu. Click the Edit DNS link beside the domain name you’re adding to Windows Live. In the text box on this page, enter the following, replacing Your_info with your code from the Mail Setup box in your Windows Live Dashboard.  Note that this is the blurred section in our screenshots.  It should be a numerical code like 1234567890.pamx1.hotmail.com. MX 10 Your_info.pamx1.hotmail.com. TXT v=spf1 include:hotmail.com ~all CNAME Your_info domains.live.com. Click Save DNS records, and your settings are saved to WordPress.  Note that this will only integrate email with your WordPress account; you cannot integrate Messenger with a domain hosted on WordPress.com. Finally, return to your Windows Live Settings page and click Refresh.  If your settings are correct, you’ll now be ready to use Windows Live on your WordPress.com domain. Integrate Windows Live with Your Own Server If your website is hosted on your own server or hosting account, you’ll need to take a few more steps to add Windows Live to your domain.  This is fairly easy, but the steps may be different depending on your hosting company or registrar.  With some hosts, you may have to contact support to have them add the MX records for you.  Our site’s host uses the popular cPanel for website administration, so here’s how we added the MX Entries through cPanel. Login to your website’s cPanel, and select MX Entry under the Mail section. In the text box on this page, enter the following, replacing Your_info with your code from the Mail Setup box in your Windows Live Dashboard.  Note that this is the blurred section in our screenshots.  It should be a numerical code like 1234567890.pamx1.hotmail.com. MX 10 Your_info.pamx1.hotmail.com. Now, go back to your cPanel home, and select Advanced DNS Zone Editor under Domains. Here, add a TXT record with the following info: Name: yoursite.com. TTL: 3600 TXT Data: v=spf1 include:hotmail.com ~all Click Add Record and your Mail integration data is all configured. To integrate Messenger with your own domain, you’ll have to add an SRV entry to your DNS settings.  cPanel doesn’t have an option for this, so we had to contact our site’s hosting company and they added the entry for us.  Copy all of the information in the Messenger box and send it to your domain support, and they should be able to add this for you.  Alternately, if you don’t want or need Messenger, then you can simply skip this step. Once all of your settings are in place, return to your Windows Live Settings page and click Refresh.  If your settings are correct, you’ll now be ready to use Windows Live on your WordPress.com domain. Create a New Email Account On Your Domain Welcome to your new Windows Live admin page!  Now you can add email accounts so you and anyone else you want can access Hotmail and the other Windows Live apps with your domain.  Click Add to add an account. Enter an account name, which will be the email address of the account, e.g. [email protected].  Then enter the user’s name and a password for the account.  By default this will be a temporary password, and the user will have to change it on first log-in, but if you’re setting up this account for yourself, you can uncheck the box and keep this as your standard password. Now, go to www.mail.live.com, and sign in with your new email address and password.  Remember, your email address is your username previously entered followed by @yourdomain.com. To finish setting up the email account, enter your password, secret question and answer, alternate email, and location information.  Click I accept to finish setting up your new email account. Enter the characters in the Captcha to confirm you’re a human, and click Continue. Your new Hotmail inbox will now load, and you’ll have a welcome email in your inbox.  This works the same as normal Hotmail, except this time, your email address is with your own domain. You can now access any of the Windows Live services from the top-level menu. Here’s an Excel Spreadsheet open in the new Office Web Apps via SkyDrive on our new Windows Live account. If you setup Messenger access previously, you can now sign in to Windows Live Messenger using your new @yourdomain.com account as well. Important Links Accessing your Windows Live accounts is easy.  Simply go to any Windows Live site, such as www.hotmail.com or www.skydrive.com, and sign in with your new Windows Live ID from your domain as normal.  You don’t need a special address to access your account; it works just like the standard public Hotmail accounts. To administer your Windows Live for your domain, go to https://domains.live.com/ and sign in with the Windows Live ID you used to create the account.  Here you can add more users, change settings, and view usage details for the Windows Live accounts on your domain. Conclusion Windows Live is easy to add to your domain, and lets you create up to 500 email address for it.  With the upcoming updates to Hotmail and Office Web Apps coming this summer, this can be a nice way to make your domain even more useful.  And with 500 email accounts, you can easily let your team take advantage of your unique address as well. If you’d rather use Google’s online applications with your domain, check out our article on how to add free Google apps to your website or blog. Link Signup for Windows Live for Your Domain Similar Articles Productive Geek Tips Tools to Help Post Content On Your WordPress BlogBackup Your Windows Live Writer SettingsInstall Windows Live Essentials In Windows 7Add Your Gmail To Windows Live MailMysticgeek Blog: A Look at Internet Explorer 8 Beta 1 on Windows XP 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 Backup Drivers With Driver Magician 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

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  • Why should you choose Oracle WebLogic 12c instead of JBoss EAP 6?

    - by Ricardo Ferreira
    In this post, I will cover some technical differences between Oracle WebLogic 12c and JBoss EAP 6, which was released a couple days ago from Red Hat. This article claims to help you in the evaluation of key points that you should consider when choosing for an Java EE application server. In the following sections, I will present to you some important aspects that most customers ask us when they are seriously evaluating for an middleware infrastructure, specially if you are considering JBoss for some reason. I would suggest that you keep the following question in mind while you are reading the points: "Why should I choose JBoss instead of WebLogic?" 1) Multi Datacenter Deployment and Clustering - D/R ("Disaster & Recovery") architecture support is embedded on the WebLogic Server 12c product. JBoss EAP 6 on the other hand has no direct D/R support included, Red Hat relies on third-part tools with higher prices. When you consider a middleware solution to host your business critical application, you should worry with every architectural aspect that are related with the solution. Fail-over support is one little aspect of a truly reliable solution. If you do not worry about D/R, your solution will not be reliable. Having said that, with Red Hat and JBoss EAP 6, you have this extra cost that will increase considerably the total cost of ownership of the solution. As we commonly hear from analysts, open-source are not so cheaper when you start seeing the big picture. - WebLogic Server 12c supports advanced LAN clustering, detection of death servers and have a common alert framework. JBoss EAP 6 on the other hand has limited LAN clustering support with no server death detection. They do not generate any alerts when servers goes down (only if you buy JBoss ON which is a separated technology, but until now does not support JBoss EAP 6) and manual intervention are required when servers goes down. In most cases, admin people must rely on "kill -9", "tail -f someFile.log" and "ps ax | grep java" commands to manage failures and clustering anomalies. - WebLogic Server 12c supports the concept of Node Manager, which is a separated process that runs on the physical | virtual servers that allows extend the administration of the cluster to WebLogic managed servers that are often distributed across multiple machines and geographic locations. JBoss EAP 6 on the other hand has no equivalent technology. Whole server instances must be managed individually. - WebLogic Server 12c Node Manager supports Coherence to boost performance when managing servers. JBoss EAP 6 on the other hand has no similar technology. There is no way to coordinate JBoss and infiniband instances provided by JBoss using high throughput and low latency protocols like InfiniBand. The Node Manager feature also allows another very important feature that JBoss EAP lacks: secure the administration. When using WebLogic Node Manager, all the administration tasks are sent to the managed servers in a secure tunel protected by a certificate, which means that the transport layer that separates the WebLogic administration console from the managed servers are secured by SSL. - WebLogic Server 12c are now integrated with OTD ("Oracle Traffic Director") which is a web server technology derived from the former Sun iPlanet Web Server. This software complements the web server support offered by OHS ("Oracle HTTP Server"). Using OTD, WebLogic instances are load-balanced by a high powerful software that knows how to handle SDP ("Socket Direct Protocol") over InfiniBand, which boost performance when used with engineered systems technologies like Oracle Exalogic Elastic Cloud. JBoss EAP 6 on the other hand only offers support to Apache Web Server with custom modules created to deal with JBoss clusters, but only across standard TCP/IP networks.  2) Application and Runtime Diagnostics - WebLogic Server 12c have diagnostics capabilities embedded on the server called WLDF ("WebLogic Diagnostic Framework") so there is no need to rely on third-part tools. JBoss EAP 6 on the other hand has no diagnostics capabilities. Their only diagnostics tool is the log generated by the application server. Admin people are encouraged to analyse thousands of log lines to find out what is going on. - WebLogic Server 12c complement WLDF with JRockit MC ("Mission Control"), which provides to administrators and developers a complete insight about the JVM performance, behavior and possible bottlenecks. WebLogic Server 12c also have an classloader analysis tool embedded, and even a log analyzer tool that enables administrators and developers to view logs of multiple servers at the same time. JBoss EAP 6 on the other hand relies on third-part tools to do something similar. Again, only log searching are offered to find out whats going on. - WebLogic Server 12c offers end-to-end traceability and monitoring available through Oracle EM ("Enterprise Manager"), including monitoring of business transactions that flows through web servers, ESBs, application servers and database servers, all of this with high deep JVM analysis and diagnostics. JBoss EAP 6 on the other hand, even using JBoss ON ("Operations Network"), which is a separated technology, does not support those features. Red Hat relies on third-part tools to provide direct Oracle database traceability across JVMs. One of those tools are Oracle EM for non-Oracle middleware that manage JBoss, Tomcat, Websphere and IIS transparently. - WebLogic Server 12c with their JRockit support offers a tool called JRockit Flight Recorder, which can give developers a complete visibility of a certain period of application production monitoring with zero extra overhead. This automatic recording allows you to deep analyse threads latency, memory leaks, thread contention, resource utilization, stack overflow damages and GC ("Garbage Collection") cycles, to observe in real time stop-the-world phenomenons, generational, reference count and parallel collects and mutator threads analysis. JBoss EAP 6 don't even dream to support something similar, even because they don't have their own JVM. 3) Application Server Administration - WebLogic Server 12c offers a complete administration console complemented with scripting and macro-like recording capabilities. A single WebLogic console can managed up to hundreds of WebLogic servers belonging to the same domain. JBoss EAP 6 on the other hand has a limited console and provides a XML centric administration. JBoss, after ten years, started the development of a rudimentary centralized administration that still leave a lot of administration tasks aside, so admin people and developers must touch scripts and XML configuration files for most advanced and even simple administration tasks. This lead applications to error prone and risky deployments. Even using JBoss ON, JBoss EAP are not able to offer decent administration features for admin people which must be high skilled in JBoss internal architecture and its managing capabilities. - Oracle EM is available to manage multiple domains, databases, application servers, operating systems and virtualization, with a complete end-to-end visibility. JBoss ON does not provide management capabilities across the complete architecture, only basic monitoring. Even deployment must be done aside JBoss ON which does no integrate well with others softwares than JBoss. Until now, JBoss ON does not supports JBoss EAP 6, so even their minimal support for JBoss are not available for JBoss EAP 6 leaving customers uncovered and subject to high skilled JBoss admin people. - WebLogic Server 12c has the same administration model whatever is the topology selected by the customer. JBoss EAP 6 on the other hand differentiates between two operational models: standalone-mode and domain-mode, that are not consistent with each other. Depending on the mode used, the administration skill is different. - WebLogic Server 12c has no point-of-failures processes, and it does not need to define any specialized server. Domain model in WebLogic is available for years (at least ten years or more) and is production proven. JBoss EAP 6 on the other hand needs special processes to garantee JBoss integrity, the PC ("Process-Controller") and the HC ("Host-Controller"). Different from WebLogic, the domain model in JBoss is quite new (one year at tops) of maturity, and need to mature considerably until start doing things like WebLogic domain model does. - WebLogic Server 12c supports parallel deployment model which enables some artifacts being deployed at the same time. JBoss EAP 6 on the other hand does not have any similar feature. Every deployment are done atomically in the containers. This means that if you have a huge EAR (an EAR of 120 MB of size for instance) and deploy onto JBoss EAP 6, this EAR will take some minutes in order to starting accept thread requests. The same EAR deployed onto WebLogic Server 12c will reduce the deployment time at least in 2X compared to JBoss. 4) Support and Upgrades - WebLogic Server 12c has patch management available. JBoss EAP 6 on the other hand has no patch management available, each JBoss EAP instance should be patched manually. To achieve such feature, you need to buy a separated technology called JBoss ON ("Operations Network") that manage this type of stuff. But until now, JBoss ON does not support JBoss EAP 6 so, in practice, JBoss EAP 6 does not have this feature. - WebLogic Server 12c supports previuous WebLogic domains without any reconfiguration since its kernel is robust and mature since its creation in 1995. JBoss EAP 6 on the other hand has a proven lack of supportability between JBoss AS 4, 5, 6 and 7. Different kernels and messaging engines were implemented in JBoss stack in the last five years reveling their incapacity to create a well architected and proven middleware technology. - WebLogic Server 12c has patch prescription based on customer configuration. JBoss EAP 6 on the other hand has no such capability. People need to create ticket supports and have their installations revised by Red Hat support guys to gain some patch prescription from them. - Oracle WebLogic Server independent of the version has 8 years of support of new patches and has lifetime release of existing patches beyond that. JBoss EAP 6 on the other hand provides patches for a specific application server version up to 5 years after the release date. JBoss EAP 4 and previous versions had only 4 years. A good question that Red Hat will argue to answer is: "what happens when you find issues after year 5"?  5) RAC ("Real Application Clusters") Support - WebLogic Server 12c ships with a specific JDBC driver to leverage Oracle RAC clustering capabilities (Fast-Application-Notification, Transaction Affinity, Fast-Connection-Failover, etc). Oracle JDBC thin driver are also available. JBoss EAP 6 on the other hand ships only the standard Oracle JDBC thin driver. Load balancing with Oracle RAC are not supported. Manual intervention in case of planned or unplanned RAC downtime are necessary. In JBoss EAP 6, situation does not reestablish automatically after downtime. - WebLogic Server 12c has a feature called Active GridLink for Oracle RAC which provides up to 3X performance on OLTP applications. This seamless integration between WebLogic and Oracle database enable more value added to critical business applications leveraging their investments in Oracle database technology and Oracle middleware. JBoss EAP 6 on the other hand has no performance gains at all, even when admin people implement some kind of connection-pooling tuning. - WebLogic Server 12c also supports transaction and web session affinity to the Oracle RAC, which provides aditional gains of performance. This is particularly interesting if you are creating a reliable solution that are distributed not only in an LAN cluster, but into a different data center. JBoss EAP 6 on the other hand has no such support. 6) Standards and Technology Support - WebLogic Server 12c is fully Java EE 6 compatible and production ready since december of 2011. JBoss EAP 6 on the other hand became fully compatible with Java EE 6 only in the community version after three months, and production ready only in a few days considering that this article was written in June of 2012. Red Hat says that they are the masters of innovation and technology proliferation, but compared with Oracle and even other proprietary vendors like IBM, they historically speaking are lazy to deliver the most newest technologies and standards adherence. - Oracle is the steward of Java, driving innovation into the platform from commercial and open-source vendors. Red Hat on the other hand does not have its own JVM and relies on third-part JVMs to complete their application server offer. 95% of Red Hat customers are using Oracle HotSpot as JVM, which means that without Oracle involvement, their support are limited exclusively to the application server layer and we all know that most problems are happens in the JVM layer. - WebLogic Server 12c supports natively JDK 7, which empower developers to explore the maximum of the Java platform productivity when writing code. This feature differentiate WebLogic from others application servers (except GlassFish that are also managed by Oracle) because the usage of JDK 7 introduce such remarkable productivity features like the "try-with-resources" enhancement, catching multiple exceptions with one try block, Strings in the switch statements, JVM improvements in terms of JDBC, I/O, networking, security, concurrency and of course, the most important feature of Java 7: native support for multiple non-Java languages. More features regarding JDK 7 can be found here. JBoss EAP 6 on the other hand does not support JDK 7 officially, they comment in their community version that "Java SE 7 can be used with JBoss 7" which does not gives you any guarantees of enterprise support for JDK 7. - Oracle WebLogic Server 12c supports integration with Spring framework allowing Spring applications to use WebLogic special transaction manager, exposing bean interfaces to WebLogic MBeans to take advantage of all WebLogic monitoring and administration advantages. JBoss EAP 6 on the other hand has no special integration with Spring. In fact, Red Hat offers a suspicious package called "JBoss Web Platform" that in theory supports Spring, but in practice this package does not offers any special integration. It is just a facility for Red Hat customers to have support from both JBoss and Spring technology using the same customer support. 7) Lightweight Development - Oracle WebLogic Server 12c and Oracle GlassFish are completely integrated and can share applications without any modifications. Starting with the 12c version, WebLogic now understands natively GlassFish deployment descriptors and specific configurations in order to offer you a truly and reliable migration path from a community Java EE application server to a enterprise middleware product like WebLogic. JBoss EAP 6 on the other hand has no support to natively reuse an existing (or still in development) application from JBoss AS community server. Users of JBoss suffer of critical issues during deployment time that includes: changing the libraries and dependencies of the application, patching the DTD or XSD deployment descriptors, refactoring of the application layers due classloading issues and anomalies, rebuilding of persistence, business and web layers due issues with "usage of the certified version of an certain dependency" or "frameworks that Red Hat potentially does not recommend" etc. If you have the culture or enterprise IT directive of developing Java EE applications using community middleware to in a certain future, transition to enterprise (supported by a vendor) middleware, Oracle WebLogic plus Oracle GlassFish offers you a more sustainable solution. - WebLogic Server 12c has a very light ZIP distribution (less than 165 MB). JBoss EAP 6 ZIP size is around 130 MB, together with JBoss ON you have more 100 MB resulting in a higher download footprint. This is particularly interesting if you plan to use automated setup of application server instances (for example, to rapidly setup a development or staging environment) using Maven or Hudson. - WebLogic Server 12c has a complete integration with Maven allowing developers to setup WebLogic domains with few commands. Tasks like downloading WebLogic, installation, domain creation, data sources deployment are completely integrated. JBoss EAP 6 on the other hand has a limited offer integration with those tools.  - WebLogic Server 12c has a startup mode called WLX that turns-off EJB, JMS and JCA containers leaving enabled only the web container with Java EE 6 web profile. JBoss EAP 6 on the other hand has no such feature, you need to disable manually the containers that you do not want to use. - WebLogic Server 12c supports fastswap, which enables you to change classes without redeployment. This is particularly interesting if you are developing patches for the application that is already deployed and you do not want to redeploy the entire application. This is the same behavior that most application servers offers to JSP pages, but with WebLogic Server 12c, you have the same feature for Java classes in general. JBoss EAP 6 on the other hand has no such support. Even JBoss EAP 5 does not support this until now. 8) JMS and Messaging - WebLogic Server 12c has a proven and high scalable JMS implementation since its initial release in 1995. JBoss EAP 6 on the other hand has a still immature technology called HornetQ, which was introduced in JBoss EAP 5 replacing everything that was implemented in the previous versions. Red Hat loves to introduce new technologies across JBoss versions, playing around with customers and their investments. And when they are asked about why they have changed the implementation and caused such a mess, their answer is always: "the previous implementation was inadequate and not aligned with the community strategy so we are creating a new a improved one". This Red Hat practice leads to uncomfortable investments that in a near future (sometimes less than a year) will be affected in someway. - WebLogic Server 12c has troubleshooting and monitoring features included on the WebLogic console and WLDF. JBoss EAP 6 on the other hand has no direct monitoring on the console, activity is reflected only on the logs, no debug logs available in case of JMS issues. - WebLogic Server 12c has extremely good performance and scalability. JBoss EAP 6 on the other hand has a JMS storage mechanism relying on Oracle database or MySQL. This means that if an issue in production happens and Red Hat affirms that an performance issue is happening due to database problems, they will not support you on the performance issue. They will orient you to call Oracle instead. - WebLogic Server 12c supports messaging enterprise features like SAF ("Store and Forward"), Distributed Queues/Topics and Foreign JMS providers support that leverage JMS implementations without compromise developer code making things completely transparent. JBoss EAP 6 on the other hand do not even dream to support such features. 9) Caching and Grid - Coherence, which is the leading and most mature data grid technology from Oracle, is available since early 2000 and was integrated with WebLogic in 2009. Coherence and WebLogic clusters can be both managed from WebLogic administrative console. Even Node Manager supports Coherence. JBoss on the other hand discontinued JBoss Cache, which was their caching implementation just like they did with the messaging implementation (JBossMQ) which was a issue for long term customers. JBoss EAP 6 ships InfiniSpan version 1.0 which is immature and lack a proven record of successful cases and reliability. - WebLogic Server 12c has a feature called ActiveCache which uses Coherence to, without any code changes, replicate HTTP sessions from both WebLogic and other application servers like JBoss, Tomcat, Websphere, GlassFish and even Microsoft IIS. JBoss EAP 6 on the other hand does have such support and even when they do in the future, they probably will support only their own application server. - Coherence can be used to manage both L1 and L2 cache levels, providing support to Oracle TopLink and others JPA compliant implementations, even Hibernate. JBoss EAP 6 and Infinispan on the other hand supports only Hibernate. And most important of all: Infinispan does not have any successful case of L1 or L2 caching level support using Hibernate, which lead us to reflect about its viability. 10) Performance - WebLogic Server 12c is certified with Oracle Exalogic Elastic Cloud and can run unchanged applications at this engineered system. This approach can benefit customers from Exalogic optimization's of both kernel and JVM layers to boost performance in terms of 10X for web, OLTP, JMS and grid applications. JBoss EAP 6 on the other hand has no investment on engineered systems: customers do not have the choice to deploy on a Java ultra fast system if their project becomes relevant and performance issues are detected. - WebLogic Server 12c maintains a performance gain across each new release: starting on WebLogic 5.1, the overall performance gain has been close to 4X, which close to a 20% gain release by release. JBoss on the other hand does not provide SPECJAppServer or SPECJEnterprise performance benchmarks. Their so called "performance gains" remains hidden in their customer environments, which lead us to think if it is true or not since we will never get access to those environments. - WebLogic Server 12c has industry performance benchmarks with submissions across platforms and configurations leading SPECJ. Oracle WebLogic leads SPECJAppServer performance in multiple categories, fitting all customer topologies like: dual-node, single-node, multi-node and multi-node with RAC. JBoss... again, does not provide any SPECJAppServer performance benchmarks. - WebLogic Server 12c has a feature called work manager which allows your application to embrace new performance levels based on critical resource utilization of the CPUs usage. Work managers prioritizes work and allocates threads based on an execution model that takes into account administrator-defined parameters and actual run-time performance and throughput. JBoss EAP 6 on the other hand has no compared feature and probably they never will. Not supporting such feature like work managers, JBoss EAP 6 forces admin people and specially developers to uncover performance gains in a intrusive way, rewriting the code and doing performance refactorings. 11) Professional Services Support - WebLogic Server 12c and any other technology sold by Oracle give customers the possibility of hire OCS ("Oracle Consulting Services") to manage critical scenarios, deployment assistance of new applications, high skilled consultancy of architecture, best practices and people allocation together with customer teams. All OCS services are available without any restrictions, having the customer bought software from Oracle or just starting their implementation before any acquisition. JBoss EAP 6 or Red Hat to be more specifically, only offers professional services if you buy subscriptions from them. If you are developing a new critical application for your business and need the help of Red Hat for a serious issue or architecture decision, they will probably say: "OK... I can help you but after you buy subscriptions from me". Red Hat also does not allows their professional services consultants to manage environments that uses community based software. They will probably force you to first buy a subscription, download their "enterprise" version and them, optionally hire their consultants. - Oracle provides you our university to educate your team into our technologies, including of course specialized trainings of WebLogic application server. At any time and location, you can hire Oracle to train your team so you get trustful knowledge according to your specific needs. Certifications for the products are also available if your technical people desire to differentiate themselves as professionals. Red Hat on the other hand have a limited pool of resources to train your team in their technologies. Basically they are selling training and certification for RHEL ("Red Hat Enterprise Linux") but if you demand more specialized training in JBoss middleware, they will probably connect you to some "certified" partner localized training since they are apparently discontinuing their education center, at least here in Brazil. They were not able to reproduce their success with RHEL education to their middleware division since they need first sell the subscriptions to after gives you specialized training. And again, they only offer you specialized training based on their enterprise version (EAP in the case of JBoss) which means that the courses will be a quite outdated. There are reports of developers that took official training's from Red Hat at this year (2012) and in a certain JBoss advanced course, Red Hat supposedly covered JBossMQ as the messaging subsystem, and even the printed material provided was based on JBossMQ since the training was created for JBoss EAP 4.3. 12) Encouraging Transparency without Ulterior Motives - WebLogic Server 12c like any other software from Oracle can be downloaded any time from anywhere, you should only possess an OTN ("Oracle Technology Network") credential and you can download any enterprise software how many times you want. And is not some kind of "trial" version. It is the official binaries that will be running for ever in your data center. Oracle does not encourages the usage of "specific versions" of our software. The binaries you buy from Oracle are the same binaries anyone in the world could download and use for testing and personal education. JBoss EAP 6 on the other hand are not available for download unless you buy a subscription and get access to the Red Hat enterprise repositories. If you need to test, learn or just start creating your application using Red Hat's middleware software, you should download it from the community website. You are not allowed to download the enterprise version that, according to Red Hat are more secure, reliable and robust. But no one of us want to start the development of a software with an unsecured, unreliable and not scalable middleware right? So what you do? You are "invited" by Red Hat to buy subscriptions from them to get access to the "cool" version of the software. - WebLogic Server 12c prices are publicly available in the Oracle website. If you want to know right now how much WebLogic will cost to your organization, just click here and get access to our price list. In the case of WebLogic, check out the "US Oracle Technology Commercial Price List". Oracle also encourages you to get in touch with a sales representative to discuss discounts that would make possible the investment into our technology. But you are not required to do this, only if you are interested in buying our technology or maybe you want to discuss some discount scenarios. JBoss EAP 6 on the other hand does not have its cost publicly available in Red Hat's website or in any other media, at least is not so easy to get such information. The only link you will possibly find in their website is a "Contact a Sales Representative" link. This is not a very good relationship between an customer and an vendor. This is not an example of transparency, mainly when the software are sold as open. In this situations, customers expects to see the software prices publicly available, so they can have the chance to decide, based on the existing features of the software, if the cost is fair or not. Conclusion Oracle WebLogic is the most mature, secure, reliable and scalable Java EE application server of the market, and have a proven record of success around the globe to prove it's majority. Don't lose the chance to discover today how WebLogic could fit your needs and sustain your global IT middleware strategy, no matter if your strategy are completely based on the Cloud or not.

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  • Communication Between Your PC and Azure VM via Windows Azure Connect

    - by Shaun
    With the new release of the Windows Azure platform there are a lot of new features available. In my previous post I introduced a little bit about one of them, the remote desktop access to azure virtual machine. Now I would like to talk about another cool stuff – Windows Azure Connect.   What’s Windows Azure Connect I would like to quote the definition of the Windows Azure Connect in MSDN With Windows Azure Connect, you can use a simple user interface to configure IP-sec protected connections between computers or virtual machines (VMs) in your organization’s network, and roles running in Windows Azure. IP-sec protects communications over Internet Protocol (IP) networks through the use of cryptographic security services. There’s an image available at the MSDN as well that I would like to forward here As we can see, using the Windows Azure Connect the Worker Role 1 and Web Role 1 are connected with the development machines and database servers which some of them are inside the organization some are not. With the Windows Azure Connect, the roles deployed on the cloud could consume the resource which located inside our Intranet or anywhere in the world. That means the roles can connect to the local database, access the local shared resource such as share files, folders and printers, etc.   Difference between Windows Azure Connect and AppFabric It seems that the Windows Azure Connect are duplicated with the Windows Azure AppFabric. Both of them are aiming to solve the problem on how to communication between the resource in the cloud and inside the local network. The table below lists the differences in my understanding. Category Windows Azure Connect Windows Azure AppFabric Purpose An IP-sec connection between the local machines and azure roles. An application service running on the cloud. Connectivity IP-sec, Domain-joint Net Tcp, Http, Https Components Windows Azure Connect Driver Service Bus, Access Control, Caching Usage Azure roles connect to local database server Azure roles use local shared files,  folders and printers, etc. Azure roles join the local AD. Expose the local service to Internet. Move the authorization process to the cloud. Integrate with existing identities such as Live ID, Google ID, etc. with existing local services. Utilize the distributed cache.   And also some scenarios on which of them should be used. Scenario Connect AppFabric I have a service deployed in the Intranet and I want the people can use it from the Internet.   Y I have a website deployed on Azure and need to use a database which deployed inside the company. And I don’t want to expose the database to the Internet. Y   I have a service deployed in the Intranet and is using AD authorization. I have a website deployed on Azure which needs to use this service. Y   I have a service deployed in the Intranet and some people on the Internet can use it but need to be authorized and authenticated.   Y I have a service in Intranet, and a website deployed on Azure. This service can be used from Internet and that website should be able to use it as well by AD authorization for more functionalities. Y Y   How to Enable Windows Azure Connect OK we talked a lot information about the Windows Azure Connect and differences with the Windows Azure AppFabric. Now let’s see how to enable and use the Windows Azure Connect. First of all, since this feature is in CTP stage we should apply before use it. On the Windows Azure Portal we can see our CTP features status under Home, Beta Program page. You can send the apply to join the Beta Programs to Microsoft in this page. After a few days the Microsoft will send an email to you (the email of your Live ID) when it’s available. In my case we can see that the Windows Azure Connect had been activated by Microsoft and then we can click the Connect button on top, or we can click the Virtual Network item from the left navigation bar.   The first thing we need, if it’s our first time to enter the Connect page, is to enable the Windows Azure Connect. After that we can see our Windows Azure Connect information in this page.   Add a Local Machine to Azure Connect As we explained below the Windows Azure Connect can make an IP-sec connection between the local machines and azure role instances. So that we firstly add a local machine into our Azure Connect. To do this we will click the Install Local Endpoint button on top and then the portal will give us an URL. Copy this URL to the machine we want to add and it will download the software to us. This software will be installed in the local machines which we want to join the Connect. After installed there will be a tray-icon appeared to indicate this machine had been joint our Connect. The local application will be refreshed to the Windows Azure Platform every 5 minutes but we can click the Refresh button to let it retrieve the latest status at once. Currently my local machine is ready for connect and we can see my machine in the Windows Azure Portal if we switched back to the portal and selected back Activated Endpoints node.   Add a Windows Azure Role to Azure Connect Let’s create a very simple azure project with a basic ASP.NET web role inside. To make it available on Windows Azure Connect we will open the azure project property of this role from the solution explorer in the Visual Studio, and select the Virtual Network tab, check the Activate Windows Azure Connect. The next step is to get the activation token from the Windows Azure Portal. In the same page there is a button named Get Activation Token. Click this button then the portal will display the token to me. We copied this token and pasted to the box in the Visual Studio tab. Then we deployed this application to azure. After completed the deployment we can see the role instance was listed in the Windows Azure Portal - Virtual Connect section.   Establish the Connect Group The final task is to create a connect group which contains the machines and role instances need to be connected each other. This can be done in the portal very easy. The machines and instances will NOT be connected until we created the group for them. The machines and instances can be used in one or more groups. In the Virtual Connect section click the Groups and Roles node from the left side navigation bar and clicked the Create Group button on top. This will bring up a dialog to us. What we need to do is to specify a group name, description; and then we need to select the local computers and azure role instances into this group. After the Azure Fabric updated the group setting we can see the groups and the endpoints in the page. And if we switch back to the local machine we can see that the tray-icon have been changed and the status turned connected. The Windows Azure Connect will update the group information every 5 minutes. If you find the status was still in Disconnected please right-click the tray-icon and select the Refresh menu to retrieve the latest group policy to make it connected.   Test the Azure Connect between the Local Machine and the Azure Role Instance Now our local machine and azure role instance had been connected. This means each of them can communication to others in IP level. For example we can open the SQL Server port so that our azure role can connect to it by using the machine name or the IP address. The Windows Azure Connect uses IPv6 to connect between the local machines and role instances. You can get the IP address from the Windows Azure Portal Virtual Network section when select an endpoint. I don’t want to take a full example for how to use the Connect but would like to have two very simple tests. The first one would be PING.   When a local machine and role instance are connected through the Windows Azure Connect we can PING any of them if we opened the ICMP protocol in the Filewall setting. To do this we need to run a command line before test. Open the command window on the local machine and the role instance, execute the command as following netsh advfirewall firewall add rule name="ICMPv6" dir=in action=allow enable=yes protocol=icmpv6 Thanks to Jason Chen, Patriek van Dorp, Anton Staykov and Steve Marx, they helped me to enable  the ICMPv6 setting. For the full discussion we made please visit here. You can use the Remote Desktop Access feature to logon the azure role instance. Please refer my previous blog post to get to know how to use the Remote Desktop Access in Windows Azure. Then we can PING the machine or the role instance by specifying its name. Below is the screen I PING my local machine from my azure instance. We can use the IPv6 address to PING each other as well. Like the image following I PING to my role instance from my local machine thought the IPv6 address.   Another example I would like to demonstrate here is folder sharing. I shared a folder in my local machine and then if we logged on the role instance we can see the folder content from the file explorer window.   Summary In this blog post I introduced about another new feature – Windows Azure Connect. With this feature our local resources and role instances (virtual machines) can be connected to each other. In this way we can make our azure application using our local stuff such as database servers, printers, etc. without expose them to Internet.   Hope this helps, Shaun All documents and related graphics, codes are provided "AS IS" without warranty of any kind. Copyright © Shaun Ziyan Xu. This work is licensed under the Creative Commons License.

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  • Run WordPress & Other Web Apps with Windows Web Platform

    - by Matthew Guay
    Would you like to run WordPress or other web apps on your PC so you can easily test and design websites?  Here we’ll look at how you can get the latest web apps on your computer in only a few quick steps. Many web apps today, such as WordPress, MediaWiki, and more, are open source and can be run for free from any computer with even a simple local web server.  They are often very difficult to install on your computer, since they require a number of dependencies such as PHP and MySQL.  Microsoft has worked to make this easier, releasing the Windows Web Platform Installer.  This lets you install many popular web apps and free tools in Windows with only a few clicks. Here we’re going to look at how to install WordPress and the free Visual Web Developer 2010 Express to edit web code with the Web Platform Installer.  But, if you’d rather install a different web app or tool, feel free to choose those as the installations are generally similar. Getting Started Head over to Microsoft’s Web development site and download the Web Platform Installer (link below).  This will download very quick, as it is just a small loader.  When you run this loader, it will download the Web Platform Installer files.  The Web Platform Installer works on XP, Vista, and Windows 7, as well as the related versions of Windows Server. After a couple moments, the Web Platform Installer will open and load information about the latest web offerings.    Now you can choose what you want to install.  You can quickly select the recommended products for several categories such as Web Server, Database, and more. Alternately, click Customize under the category and select exactly what you want to install.  Note that items already installed on your computer will be grayed out. We wanted to install Visual Web Developer 2010 Express, so select Customize under Tools, and select Visual Web Developer 2010 Express. Or, for more preset choices, select Options on the bottom of the window. You can choose to add Multimedia, Developer, and Enterprise tools to the lists, or add a new preset list from a feed. Choose Specific Web apps to Install We wanted to install WordPress, so instead of choosing a preset, select the Web Applications tab on the left.  Now you can choose from a variety of apps based on category, or you can view them all together in an A to Z, Most Popular, or Highest Rating list. Click the checkbox beside the app you want to install to select it, or click the “i” for more information. Here’s the More Information pane for WordPress.  If you’re ready to install it, click the checkbox. Now you can go back and add more web apps or tools to the install list if you like.  The Web Platform Installer will automatically find and select prerequisite apps such as MySQL, so you won’t need to worry about finding them. Once you’ve selected everything you want to install, click the Install button on the bottom of the window. The Web Platform Installer will now show you everything that’s selected, including components that it automatically selected.  Notice we only chose to install WordPress and Visual Web Developer 2010 Express, but it also has selected MySQL and PHP automatically.  Click I Accept to proceed. Enter an administrator password for MySQL before the setup begins. Now the Web Platform Installer will take over, automatically downloading, installing, and configuring all of your web apps.  It will also activate optional Windows components that may be needed on your computer.  This may take several minutes, depending on the components you selected and your internet speed.   Setting up Your Test Site Once the installation is finished, you’ll be asked to enter some information about your site.  You can simply accept the defaults or enter your own choices, and then click Continue. Now you’ll need to enter some information for your web apps.  When installing WordPress, you’ll need to choose a database and enter administrative usernames and passwords.  You may also be asked to enter extra information for additional security, but for a local-only test site this isn’t necessary.  Click Continue when you’re finished. You’ll need to wait a few more moments as it complete the setup of your web apps.  The good thing is, once it’s finished, they’ll be ready to go with only minimal configuration. And you’re finished!  The installer will let you know everything it installed, and if there were any problems.  In our test, Visual Web Developer 2010 Express failed to install successfully.  Often the problems may be with the download, so click Finish and then reselect the apps that didn’t install and run the installer again. Now you’re ready to run WordPress from your PC.  Click the Launch WordPress link or enter http://localhost:80/wordpress in your browser to get started. You’ll only have a little more setup to do on WordPress to get it running.  Once you’ve opened your WordPress page in your browser, enter a name for your blog and your email address, and click Install WordPress.   After a few seconds, you should see a Success! page with your username and a temporary password.  Copy the password, and then click Log In. Enter admin as the Username and paste the random generated password, and click Log In. WordPress will remind you to change the default password.  Click the Yes, Take me to my profile page link to do this. Enter something easier for you to remember, and click Update Profile. Now you’re ready to enjoy your new WordPress install on Windows.  You can add plugins and themes, and everything else you’d do with a normal WordPress site.  Here’s the dashboard running from localhost. And here’s the default blog running. Setting up Visual Web Developer 2010 Express As mentioned before, Visual Web Developer 2010 Express didn’t install correctly on our first try, but the second time it installed seamlessly.  Once it’s installed, launch it from your start menu as normal.  It may take a few minutes to load on the first run as it is finishing up setup. You may notice that the splash screen displayed while the program is loading says For Evaluation Purposes Only.  This is because you still need to register the program. You have 30 days to register the program, but let’s go ahead and do it to get this step out of the way.  Click Help in the menu bar, and select Register Product. Click Obtain a registration key online in the popup window. You’ll need to sign in with your Windows Live ID, and then fill out a quick form. When you’re done, copy the registration key displayed and paste it into the registration dialog in Visual Web Developer.   Now you’ve got a registered, free web development program with full standards compliance and IntelliSense to help you work smarter and faster.  And it works great with your local web apps, so you can create, tweak, and then deploy, all from your desktop with this simple installer! Install More Apps You can always run the Web Platform Installer again in the future and add more apps if you’d like.  The install adds a link to the Installer in the Start menu; just run it and repeat the steps above with your new selections. Also, from the installer, you can cleanup the setup files downloaded during the installation if you want.  Click the Options link in the bottom of the window, and then scroll down and select Delete installer cache folder. Uninstalling the apps is not as easy, unfortunately.  If you wish to uninstall the Web Platform Installer and everything you installed with it, you’ll need to uninstall each item individually.  One easy way to see what was all installed together is to sort the entries in Uninstall Programs by date.  In our case, we also installed some other applications on the same day, but it’s easier to see what was installed together. Or if you are not a fan of using Programs and Features to uninstall them, try out a program like Revo Uninstaller Pro. Conclusion Whether you’re a full-time web developer or just enjoy testing out the latest web apps, the Web Platform Installer makes it quick and easy to get your computer loaded up with the latest bits.  In fact, it’s easier to install these tools with all their dependencies than it is to install many standard boxed programs. If you’d like to take your web server anywhere you go and not have it confined to your desktop, then check out our article on how to Turn Your Flashdrive into a Portable Webserver. Link Download the Microsoft Web Platform Installer Similar Articles Productive Geek Tips Linux QuickTip: Downloading and Un-tarring in One StepQuick Tip: Set a Future Date for a Post in WordPressHow-To Geek SoftwareAdd Social Bookmarking (Digg This!) Links to your Wordpress BlogHow-To Geek Software: WordPress Comment Moderation Notifier 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 Windows Media Player Glass Icons (icons we like) 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

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  • September 2012 Release of the Ajax Control Toolkit

    - by Stephen.Walther
    I’m excited to announce the September 2012 release of the Ajax Control Toolkit! This is the first release of the Ajax Control Toolkit which supports the .NET 4.5 framework. We also continue to support ASP.NET 3.5 and ASP.NET 4.0. With this release, we’ve made several important bug fixes. The Superexpert team focused on fixing the highest voted issues associated with the CascadingDropDown control. I’ve created a list of these bug fixes later in this blog post. You can download the latest release of the Ajax Control Toolkit by visiting the following page at CodePlex: http://AjaxControlToolkit.CodePlex.com Alternatively, you can install the latest version of the Ajax Control Toolkit using NuGet by firing off the following command from the Package Manager Console: Install-Package AjaxControlToolkit Using the Ajax Control Toolkit with ASP.NET 4.5 Let me walk through the steps for using the Ajax Control Toolkit with ASP.NET 4.5. First, I’ll create a new ASP.NET 4.5 website with Visual Studio 2012. I’ll create the new website with the ASP.NET Web Forms Application template: When you create a new ASP.NET 4.5 site with the ASP.NET Web Forms Application template, you get a starter website. If you run the site, then you get a page with default content: Let me show you how you can add the Ajax Control Toolkit Calendar control to the homepage of this starter site. The first step is to use NuGet to install the Ajax Control Toolkit. Right-click the References folder in the Solution Explorer window and select the menu option Manage NuGet Packages. In the Manage NuGet Packages dialog, use the search box to search for the Ajax Control Toolkit (enter “AjaxControlToolkit”). After you find it, click the Install button to add the Ajax Control Toolkit to your project. That’s all you have to do to install the Ajax Control Toolkit! Now we are ready to start using the Ajax Control Toolkit controls. Open the default.aspx page so we can modify the contents of the page. Erase everything contained in the Content control with the ID of BodyContent. After erasing the content, declare the following two controls: <asp:TextBox ID="vacationDate" runat="server" /> <ajaxToolkit:CalendarExtender TargetControlID="vacationDate" runat="server" /> The first control is a standard ASP.NET TextBox control and the second control is an Ajax Control Toolkit Calendar control. You should get intellisense as you type out the Ajax Control Toolkit Calendar control. If you don’t, then close and re-open the Default.aspx page. Now, let’s run our app. Hit the F5 button or select Debug, Start Debugging from the Visual Studio menu. You will get the error message “MsAjaxBundle is not a valid script name”. Don’t despair! We need to update the Master Page so it uses the ToolkitScriptManager instead of the default ScriptManager. Open the Site.Master file and find where the ScriptManager is declared. The ScriptManager should look like this: <asp:ScriptManager runat="server"> <Scripts> <%--Framework Scripts--%> <asp:ScriptReference Name="MsAjaxBundle" /> <asp:ScriptReference Name="jquery" /> <asp:ScriptReference Name="jquery.ui.combined" /> <asp:ScriptReference Name="WebForms.js" Assembly="System.Web" Path="~/Scripts/WebForms/WebForms.js" /> <asp:ScriptReference Name="WebUIValidation.js" Assembly="System.Web" Path="~/Scripts/WebForms/WebUIValidation.js" /> <asp:ScriptReference Name="MenuStandards.js" Assembly="System.Web" Path="~/Scripts/WebForms/MenuStandards.js" /> <asp:ScriptReference Name="GridView.js" Assembly="System.Web" Path="~/Scripts/WebForms/GridView.js" /> <asp:ScriptReference Name="DetailsView.js" Assembly="System.Web" Path="~/Scripts/WebForms/DetailsView.js" /> <asp:ScriptReference Name="TreeView.js" Assembly="System.Web" Path="~/Scripts/WebForms/TreeView.js" /> <asp:ScriptReference Name="WebParts.js" Assembly="System.Web" Path="~/Scripts/WebForms/WebParts.js" /> <asp:ScriptReference Name="Focus.js" Assembly="System.Web" Path="~/Scripts/WebForms/Focus.js" /> <asp:ScriptReference Name="WebFormsBundle" /> <%--Site Scripts--%> </Scripts> </asp:ScriptManager> We need to make three changes to the ScriptManager: 1) We need to replace the asp:ScriptManager with the ajaxToolkit:ToolkitScriptManager 2) We need to remove the MsAjaxBundle bundle from the ScriptReferences 3) We need to remove the Assembly=”System.Web” attributes from the ScriptReferences After you make these three changes, the ToolkitScriptManager should looks like this: <ajaxToolkit:ToolkitScriptManager runat="server"> <Scripts> <%--Framework Scripts--%> <asp:ScriptReference Name="jquery" /> <asp:ScriptReference Name="jquery.ui.combined" /> <asp:ScriptReference Name="WebForms.js" Path="~/Scripts/WebForms/WebForms.js" /> <asp:ScriptReference Name="WebUIValidation.js" Path="~/Scripts/WebForms/WebUIValidation.js" /> <asp:ScriptReference Name="MenuStandards.js" Path="~/Scripts/WebForms/MenuStandards.js" /> <asp:ScriptReference Name="GridView.js" Path="~/Scripts/WebForms/GridView.js" /> <asp:ScriptReference Name="DetailsView.js" Path="~/Scripts/WebForms/DetailsView.js" /> <asp:ScriptReference Name="TreeView.js" Path="~/Scripts/WebForms/TreeView.js" /> <asp:ScriptReference Name="WebParts.js" Path="~/Scripts/WebForms/WebParts.js" /> <asp:ScriptReference Name="Focus.js" Path="~/Scripts/WebForms/Focus.js" /> <asp:ScriptReference Name="WebFormsBundle" /> <%--Site Scripts--%> </Scripts> </ajaxToolkit:ToolkitScriptManager> After we make these changes, the app should run successfully. You’ll get a page which contains a text field. When you click inside the text field, a popup calendar is displayed. Ajax Control Toolkit and jQuery You might have noticed that the ScriptManager includes a reference to jQuery by default. We did not remove that reference when we converted the ScriptManager to a ToolkitScriptManager. You can use the Ajax Control Toolkit and jQuery side-by-side. Here’s how you can modify the Default.aspx page so that it contains two popup calendars. The first popup calendar is created with the Ajax Control Toolkit and the second popup calendar is created with jQuery: <asp:TextBox ID="vacationDate" runat="server" /> <ajaxToolkit:CalendarExtender TargetControlID="vacationDate" runat="server" /> <input id="birthDate" /> <script> $("#birthDate").datepicker(); </script> Before you can start using jQuery UI plugins, you need to complete one more step. You need to add the jQuery UI themes bundle to the HEAD of the Site.Master page like this: <head runat="server"> <meta charset="utf-8" /> <title><%: Page.Title %> - My ASP.NET Application</title> <asp:PlaceHolder runat="server"> <%: Scripts.Render("~/bundles/modernizr") %> </asp:PlaceHolder> <webopt:BundleReference runat="server" Path="~/Content/css" /> <webopt:BundleReference runat="server" Path="~/Content/themes/base/css" /> <link href="~/favicon.ico" rel="shortcut icon" type="image/x-icon" /> <meta name="viewport" content="width=device-width" /> <asp:ContentPlaceHolder runat="server" ID="HeadContent" /> </head> The markup above includes a reference to the jQuery UI themes bundle: <webopt:BundleReference runat="server" Path="~/Content/themes/base/css" /> Now that we have made these changes, we can use the Ajax Control Toolkit and jQuery at the same time. When you run your app, you get two popup calendars. When you click in the first text field, the Ajax Control Toolkit calendar appears. When you click in the second text field, the jQuery UI popup calendar appears: Bug Fixes in this Release We made several important bug fixes with this release of the Ajax Control Toolkit and integrated several Pull Requests contributed by the community. Our primary focus during this sprint was fixing issues with the CascadingDropDown control. We fixed the following issues associated with the CascadingDropDown: · 9490 – Don’t disable dropdowns in CascadingDropDown · 14223 – CascadingDropDown Reset or Setting SelectedValue from WebMethod · 12189 – CascadingDropDown not obeying disabled state of DropDownList · 22942 – CascadingDropDown infinite loop (with solution) · 8671 – CascadingDropdown options is null or undefined · 14407 – CascadingDropDown: populated client event happens too often · 17148 – CascadingDropDown – Add “UseHttpGet” property · 10221 – No NotNull check in CascadingDropDown · 12228 – Provide property for case-insensitive DefaultValue lookup in CascadingDropdown We also fixed the following two issues which are not directly related to the CascadingDropDown control: · 27108 – CalendarExtender: Bug when selecting December shifts to January. · 27041 – Input controls with HTML5 types do not post back in Firefox, Chrome, Safari Finally, we integrated several Pull Requests submitted by the community (Thank you community!): · Added French localized resources for the AjaxFileUpload · Resolved an issue which prevented the AjaxFileUpload control from working with pages that require query string variables. · Extended the AjaxFileUploadEventArgs class to include the current file index in the queue and the total number of files in the queue. · Fixed an issue with TabContainer and TabPanel which caused the OnActiveTabChanged event to fire too often. Summary I’m happy to see the Ajax Control Toolkit move forward into the brave new world of ASP.NET 4.5! In this latest release, we focused on ensuring that the Ajax Control Toolkit works smoothly with ASP.NET 4.5 applications. We also fixed the highest voted bugs associated with the CascadingDropDown control and integrated several Pull Request submitted by the community. Once again, I want to thank the Superexpert team for their hard work on this release!

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  • Building the Ultimate SharePoint 2010 Development Environment

    - by Manesh Karunakaran
    It’s been more than a month since SharePoint 2010 RTMed. And a lot of people have downloaded and set up their very own SharePoint 2010 development rigs. And quite a few people have written blogs about setting up good development environments, there is even an MSDN article on it. Two of the blogs worth noting are from MVPs Sahil Malik and Wictor Wilén. Make sure that you check these out as well. Part of the bad side-effects of being a geek is the need to do the technical stuff the best way possible (pragmatic or otherwise), but the problem with this is that what is considered “best” is relative. Precisely the reason why you are reading this post now. Most of the posts that I read are out dated/need updations or are using the wrong OS’es or virtualization solutions (again, opinions vary) or using them the wrong way. Here’s a developer’s view of Building the Ultimate SharePoint 2010 Development Rig. If you are a sales guy, it’s time to close this window. Confusion 1: Which Host Operating System and Virtualization Solution to use? This point has been beaten to death in numerous blog posts in the past, if you have time to invest, read this excellent post by our very own SharePoint Joel on this subject. But if you are planning to build the Ultimate Development Rig, then Windows Server 2008 R2 with Hyper-V is the option that you should be looking at. I have been using this as my primary OS for about 6-7 months now, and I haven’t had any Driver issue or Application compatibility issue. In my experience all the Windows 7 drivers work fine with WIN2008 R2 also. You can enable Aero for eye candy (and the Windows 7 look and feel) and except for a few things like the Hibernation support (which a can be enabled if you really want it), Windows Server 2008 R2, is the best Workstation OS that I have used till date. But frankly the answer to this question of which OS to use depends primarily on one question - Are you willing to change your primary OS? If the answer to that is ‘Yes’, then Windows 2008 R2 with Hyper-V is the best option, if not look at vmWare or VirtualBox, both are equally good. Those who are familiar with a Virtual PC background might prefer Sun VirtualBox. Besides, these provide support for running 64 bit guest machines on 32 bit hosts if the underlying hardware is truly 64 bit. See my earlier post on this. Since we are going to make the ultimate rig, we will use Windows Server 2008 R2 with Hyper-V, for reasons mentioned above. Confusion 2: Should I use a multi-(virtual) server set up? A lot of people use multiple servers for their development environments - like Wictor Wilén is suggesting - one server hosting the Active directory, one hosting SharePoint Server and another one for SQL Server. True, this mimics the production environment the best possible way, but as somebody who has fallen for this set up earlier, I can tell you that you don’t really get anything by doing this. Microsoft has done well to ensure that if you can do it on one machine, you can do it in a farm environment as well. Besides, when you run multiple Server class machine instances in parallel, there are a lot of unwanted processor cycles wasted for no good use. In my personal experience, as somebody who needs to switch between MOSS 2007/SharePoint 2010 environments from time to time, the best possible solution is to Make the host Windows Server 2008 R2 machine your Domain Controller (AD Server) Make all your Virtual Guest OS’es join this domain. Have each Individual Guest OS Image have it’s own local SQL Server instance. The advantages are that you can reuse the users and groups in each of the Guest operating systems, you can manage the users in one place, AD is light weight and doesn't take too much resources on your host machine and also having separate SQL instances for each of the Development images gives you maximum flexibility in terms of configuration, for example your SharePoint rigs can have simpler DB configurations, compared to your MS BI blast pits. Confusion 3: Which Operating System should I use to run SharePoint 2010 Now that’s a no brainer. Use Windows 2008 R2 as your Guest OS. When you are building the ultimate rig, why compromise? If you are planning to run Windows Server 2008 as your Guest OS, there are a few patches that you need to install at different times during the installation, for that follow the steps mentioned here Okay now that we have made our choices, let’s get to the interesting part of building the rig, Step 1: Prepare the host machine – Install Windows Server 2008 R2 Install Windows Server 2008 R2 on your best Desktop/Laptop. If you have read this far, I am quite sure that you are somebody who can install an OS on your own, so go ahead and do that. Make sure that you run the compatibility wizard before you go ahead and nuke your current OS. There are plenty of blogs telling you how to make a good Windows 2008 R2 Workstation that feels and behaves like a Windows 7 machine, follow one and once you are done, head to Step 2. Step 2: Configure the host machine as a Domain Controller Before we begin this, let me tell you, this step is completely optional, you don’t really need to do this, you can simply use the local users on the Guest machines instead, but if this is a much cleaner approach to manage users and groups if you run multiple guest operating systems.  This post neatly explains how to configure your Windows Server 2008 R2 host machine as a Domain Controller. Follow those simple steps and you are good to go. If you are not able to get it to work, try this. Step 3: Prepare the guest machine – Install Windows Server 2008 R2 Open Hyper-V Manager Choose to Create a new Guest Operating system Allocate at least 2 GB of Memory to the Guest OS Choose the Windows 2008 R2 Installation Media Start the Virtual Machine to commence installation. Once the Installation is done, Activate the OS. Step 4: Make the Guest operating systems Join the Domain This step is quite simple, just follow these steps below, Fire up Hyper-V Manager, open your Guest OS Click on Start, and Right click on ‘Computer’ and choose ‘Properties’ On the window that pops-up, click on ‘Change Settings’ On the ‘System Properties’ Window that comes up, Click on the ‘Change’ button Now a window named ‘Computer Name/Domain Changes’ opens up, In the text box titled Domain, type in the Domain name from Step 2. Click Ok and windows will show you the welcome to domain message and ask you to restart the machine, click OK to restart. If the addition to domain fails, that means that you have not set up networking in Hyper-V for the Guest OS to communicate with the Host. To enable it, follow the steps I had mentioned in this post earlier. Step 5: Install SQL Server 2008 R2 on the Guest Machine SQL Server 2008 R2 gets installed with out hassle on Windows Server 2008 R2. SQL Server 2008 needs SP2 to work properly on WIN2008 R2. Also SQL Server 2008 R2 allows you to directly add PowerPivot support to SharePoint. Choose to install in SharePoint Integrated Mode in Reporting Server Configuration. Step 6: Install KB971831 and SharePoint 2010 Pre-requisites Now install the WCF Hotfix for Microsoft Windows (KB971831) from this location, and SharePoint 2010 Pre-requisites from the SP2010 Installation media. Step 7: Install and Configure SharePoint 2010 Install SharePoint 2010 from the installation media, after the installation is complete, you are prompted to start the SharePoint Products and Technologies Configuration Wizard. If you are using a local instance of Microsoft SQL Server 2008, install the Microsoft SQL Server 2008 KB 970315 x64 before starting the wizard. If your development environment uses a remote instance of Microsoft SQL Server 2008 or if it has a pre-existing installation of Microsoft SQL Server 2008 on which KB 970315 x64 has already been applied, this step is not necessary. With the wizard open, do the following: Install SQL Server 2008 KB 970315 x64. After the Microsoft SQL Server 2008 KB 970315 x64 installation is finished, complete the wizard. Alternatively, you can choose not to run the wizard by clearing the SharePoint Products and Technologies Configuration Wizard check box and closing the completed installation dialog box. Install SQL Server 2008 KB 970315 x64, and then manually start the SharePoint Products and Technologies Configuration Wizard by opening a Command Prompt window and executing the following command: C:\Program Files\Common Files\Microsoft Shared Debug\Web Server Extensions\14\BIN\psconfigui.exe The SharePoint Products and Technologies Configuration Wizard may fail if you are using a computer that is joined to a domain but that is not connected to a domain controller. Step 8: Install Visual Studio 2010 and SharePoint 2010 SDK Install Visual Studio 2010 Download and Install the Microsoft SharePoint 2010 SDK Step 9: Install PowerPivot for SharePoint and Configure Reporting Services Pop-In the SQLServer 2008 R2 installation media once again and install PowerPivot for SharePoint. This will get added as another instance named POWERPIVOT. Configure Reporting Services by following the steps mentioned here, if you need to get down to the details on how the integration between SharePoint 2010 and SQL Server 2008 R2 works, see Working Together: SQL Server 2008 R2 Reporting Services Integration in SharePoint 2010 an excellent article by Alan Le Marquand Step 10: Download and Install Sample Databases for Microsoft SQL Server 2008R2 SharePoint 2010 comes with a lot of cool stuff like PerformancePoint Services and BCS, if you need to try these out, you need to have data in your databases. So if you want to save yourself the trouble of creating sample data for your PerformancePoint and BCS experiments, download and install Sample Databases for Microsoft SQL Server 2008R2 from CodePlex. And you are done! Fire up your Visual Studio 2010 and Start Coding away!!

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  • Windows Azure Evolution &ndash; Preview Developer Portal

    - by Shaun
    With the MEET Windows Azure event on 7th June, there are many new features and updates in windows azure platform. In the coming several posts I will try to cover some of them. And in the first post here I would like to just have a quick walkthrough of the new preview developer portal.   History of the Developer Portal If you have been working with windows azure since 2009 or 2010, you should remember the first version of the developer portal. It was built in HTML with very limited features. I have the impression when I was using is old one. The layout is not that attractive and you have very limited features. On November, 2010 alone with the SDK 1.3 release, the developer portal was getting a big jump. In order to give more usability and features this it turned to be built on Silverlight. Hence it runs like a desktop application with many windows, lists, commands and context menus. From 2010 till now many features were involved into this portal, such as the remote desktop, co-admin, virtual connect, VM role, etc.. And the portal itself became more and more complicated. But it brought some problems by using the Silverlight. The first one is the browser capability. As you know in most mobile and tablet device the browser doesn’t allow the rich content plugin, such as Flash and Silverlight. This means people cannot open and configure their azure services from their iPad, iPhone and Windows Phone, etc., even though what they need may just be restart a hosted service, or view the status of their databases. Another problem is the performance. Silverlight provides rich experience to the users, but also needs more bandwidth. So in this upgrade the preview developer portal will be back to use HTML, with JavaScript, as a mobile friendly, cross browser, interactively web site.   Preview Portal vs. Silverlight Portal Before I started to talk about the new preview portal I’d better highlight that, this preview portal is a PREVIEW version, which means even though you can do almost all features that already in the old one, as long as some cool new features I will mention in the coming several posts, there are something still under developed and migrated. So sometimes you need to switch back to the old one. For example, in preview portal there is no co-admin manage function, no remote desktop function and the SQL database manage function will take you back to the old SQL Azure Manage Portal. But as Microsoft said these missing features will be moved in the preview portal in the couple of next few months. Since the public URL of the developer portal, https://windows.azure.com/, had been changed to point to this preview one, you need to click to preview button on top of the page and click the “Take me to the previous portal” link.   Overview There are four parts in the preview portal. On the top is the header which shows the account you are currently logging in. If you click on the header it will show the top menu of windows azure, where you can navigate to the windows azure home page, the price information page, community and account, etc.. The navigation bar is on the left hand side, with the categories listed below. ALL ITEMS All items in your windows azure account, includes the web sites, services, databases, etc.. WEB SITES The web sites in your windows azure account. It will only show the web sites you have. The linked resources will be shown if you drill down into a web site. VIRTUAL MACHINES The virtual machines that you had been deployed to azure. CLOUD SERVICES All windows azure hosted services in your account. SQL DATABASES All SQL databases (SQL Azure) in your account. STORAGE All windows azure storage services in your account. NETWORKS The virtual network (Windows Azure Connect) you had been created. The available items will be listed in the main part of the page based on which category your currently selected. If there’s no item it will show the link to you to quick create. At the bottom of the page there will be the command and information bar. Based on what is selected and what is performed by the user, it will show the related information and commands. For example, in the image below when I was creating a new web site, the information bar told me that my web site is being provisioned; and there are two commands in the command bar. And once it ready the command bar will show some commands that I can do to my new web site. The “Web Sites” is a new feature introduced alone with this upgrade. It gives us an easier and quicker way to establish a website from the scratch or from some existing library. I will introduce it more details in the coming next post. Also in the command bar you can create a service by clicking the NEW button. It will slide the creation panel up to you.   Where’s My Hosted Services The Windows Azure Hosted Services had been renamed to the Cloud Services. Create a new service would be very easy. Just click the NEW button at the bottom of the page, and select the CLOUD SERVICE and QIUICK CREATE. This will create a blank hosted service without deployment and certificate. It just needs you to specify the service URL and the affinity/region. Then the service will be shown in the list. If you clicked the item all information will be shown in the main part. Since there’s no package deployed to this service so currently we cannot see any information about it. But we can upload the package by using the command at the bottom. And as you can see, we could manage the configuration, instances, certificates and we can scale up and down (change the VM size), in and out (increase and decrease the instance count) to our service. Assuming I had created an ASP.NET MVC 3 web role project in Visual Studio and completed the package. Then I can click the UPLOAD button in this page to deploy my package. In the popping up window I just specify my deployment name, package file and configure file. Also I can check the box below so that it will NOT warn me if only one instance of this deployment. Once we clicked the OK button our package will be uploaded and provisioned by the platform. After a while we can see the service was ready from the information bar. We can have the basic information about this service and deployment if we to the dashboard page. For example the usage overview diagram, status, URL, public IP address, etc.. In the configure page we can view and change the CSCFG content such as the monitor setting, connection strings, OS family. In scale page we can increase and decrease the count of the instances. And in the instances page we can view all instances status. And, if your services is using some SQL databases and storages they will be shown as the linked resources under the linked resources page. And you can manage the certificates of this service as well under the certificates page.   How About My Storage Services The storage service can be managed by clicking into the STORAGES link in the navigation bar. And we can create a new storage service from the NEW button. After specify the storage name and region it will be previsioned by the platform. If you want to copy or manage the storage key you can just click the Manage Keys button at the bottom, which is very easy. What I want to highlight here is that, you can monitor your storage service by enabling the monitor configuration. Click the storage item in the list and navigate to the configure page. As you can see in the page you can enable the monitoring for blob, table and queue. And you can also enable the logging when any requests come to the storage. But as the tooltip shown in the page, enabling the monitoring and logging will increase the usage of the storage, which means increase the bill of them. So make sure you enable them properly.   And My SQL Databases (SQL Azure) The last thing I want to quick introduce is the SQL databases, which was formally named SQL Azure. You can create a new SQL Database Server and a new database by clicking the ADD button under the SQL Database navigation item. In the popping up windows just specify the database name, the edition, size, collation and the server. You can select an existing SQL Database Server if you have, or cerate a new one. If you selected to create a new server, there will be another step you need to do, which is specify the server login, password and the region. Once it ready you can mange your databases as well as the servers in the portal. In a particular server you can update the firewall settings in its Configure page. So, What Else There are some other area on the preview portal I didn’t cover, such as the virtual machines, virtual network and web sites. Regarding the virtual machines and web sites I will talk about them in the future separated post. Regarding the virtual network, it the Windows Azure Connect we are familiar with. But as I mention in the beginning of this post, the preview portal is still under developed. Some features are not available here. For example, you cannot manage the co-admin of your subscriptions, you cannot open the remote desktop on your hosted services, and you cannot navigate to the Windows Azure Service Bus, Access Control and Caching, which formally named Windows Azure AppFabric directly. In these cases you need to navigate back to the old portal. So in the coming several months we might need to use both these two sites.   Summary In this post I quick introduced the new windows azure developer portal. Since it had been rearranged and renamed I demonstrated some features that existing in the old portal, such as how to create and deploy a hosted service, how to provision a storage service and SQL database. All features in the old portal had been, is being and will be migrated into this new portal, but some of them were in a different category and page we need to figure out.   Hope this helps, Shaun All documents and related graphics, codes are provided "AS IS" without warranty of any kind. Copyright © Shaun Ziyan Xu. This work is licensed under the Creative Commons License.

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  • ODEE Green Field (Windows) Part 5 - Deployment and Validation

    - by AndyL-Oracle
    And here we are, almost finished with our installation of Oracle Documaker Enterprise Edition ("ODEE") in a Windows green field environment. Let's recap what we've done so far: In part 1, I went over the basic process that I intended to show with installing an ODEE on a green field server. I walked you through the basic installation of Oracle 11g database In part 2, I covered the installation of WebLogic application server. In part 3, I showed you how to install SOA Suite for WebLogic. In part 4, we did the first part of the installation of ODEE itself. What remains after all of that, is the deployment of the ODEE components onto the database and application server - so let's get to it! DATABASE First, we'll deploy the schemas to the database. The schemas are created during the ODEE installation according to the responses provided during the install process. To deploy the schemas, you'll need to login to the database server in your green field environment. Open a command line and CD into ODEE_HOME\documaker\database\oracle11g.Run SQLPLUS as SYSDBA and execute dmkr_admin.sql:  sqlplus / as sysdba @dmkr_admin.sql Execute dmkr_asline.sql, dmkr_admin_correspondence_example.sql.  If you require additional languages, run the appropriate SQL scripts (e.g. dmkr_asline_es.sql for Spanish). APPLICATION SERVER Next, we'll deploy the WebLogic domain and it's components - Documaker web services, Documaker Interactive, Documaker dashboard, and more. To deploy the components, you'll need to login to the application server in your green field environment. 1. Open Windows Explorer and navigate to ODEE_HOME\documaker\j2ee\weblogic\oracle11g\scripts.2. Using a text editor such as Notepad++, modify weblogic_installation_properties and set location of MIDDLEWARE_HOME and ODEE HOME. If you have used the defaults you’ll probably need to change the E: to C: and that’s it. Save the changes.3. Continuing in the same directory, use your text editor to modify set_middleware_env.cmd and set the drive and path to MIDDLEWARE_HOME. If you have used the defaults you’ll probably need to just change E: to C: and that’s it. Save the changes.4. In the same directory, execute wls_create_domain.cmd by double-clicking it. This should run to completion. If it does not, review any errors and correct them, and rerun the script.5. In the same directory, execute wls_add_correspondence.cmd by double-clicking it - again this should run to completion. 6. Next, we'll start the AdminServer - this is the main WebLogic domain server. To start it, use Windows Explorer and navigate to MIDDLEWARE_HOME\user_projects\domains\idocumaker_domain. Double-click startWebLogic.cmd and the server startup will begin. Once you see output that indicates that the server status changed to RUNNING you may proceed.  a. Note: if you saw database connection errors, you probably didn’t make sure your database name and connection type match. You can change this manually in the WebLogic Console. Open a browser and navigate to http://localhost:7001/console (replace localhost with the name of your application server host if you aren't opening the browser on the server), and login with the the weblogic credential you provided in the ODEE installation process. b. Once you're logged in, open Services?Data Sources. Select dmkr_admin and click Connection Pool.  c. The end of the URL should match the connection type you chose. If you chose ServiceName, the URL should be: jdbc:oracle:thin:@//<hostname>:1521/<serviceName> and if you chose SID, the URL should be: jdbc:oracle:thin:@//<hostname>:1521/<SIDname> d. An example serviceName is a fully qualified DNS-style name, e.g. "idmaker.us.oracle.com". (It does not need to actually resolve in DNS). An example SID is just a name, e.g. IDMAKER. e. Save the change and repeat for the data source dmkr_asline.  f. You will also need to make the same changes in the ODEE_HOME/documaker/docfactory/config/context/.bindings file - open the file in a text editor, locate the URL lines and make the appropriate change, then save the file.  7. Back in the ODEE_HOME\documaker\j2ee\weblogic\oracle11g\scripts directory, execute create_users_groups.cmd. 8. In the same directory, execute create_users_groups_correspondence_example.cmd. 9. Open a browser and navigate to http://localhost:7001/jpsquery. Replace localhost with the name of your application server host if you aren't running the browser on the application server. If you changed the default port for the AdminServer from 7001, use the port you changed it to. You should see output like this: 10. Start the WebLogic managed servers by opening a command prompt and navigating to MIDDLEWARE_HOME/user_projects/domains/idocumaker_domain/bin/. When you start the servers listed below, you will be prompted to enter the WebLogic credentials to start the server. You can prevent this by providing the credential in the startManagedwebLogic.cmd file for the WLS_USER and WLS_PASS values. Note that the credential will be stored in cleartext. To start the server, type in the command shown. a. Start the JMS Server: ./startManagedWebLogic.cmd jms_server b. Start Dashboard/Documaker Administrator: ./startManagedWebLogic.cmd dmkr_server c. Start Documaker Interactive for Correspondence: ./startManagedWebLogic.cmd idm_server SOA Composites  If you're planning on testing out the approval process components of BPEL that can be used with Documaker Interactive, then use the following steps to deploy the SOA composites. If you're not going to use BPEL, you can skip to the next section.1. Stop the servers listed in the previous section (Step 10) in the reverse order that they were started.2. Run the Domain configuration command: navigate to and execute MIDDLEWARE_HOME/wlserver_10.3/common/bin/config.cmd.3. Select Extend and click next. 4. Select the iDocumaker Domain and click Next. 5. Select the Oracle SOA Suite – 11.1.1.0 (this may automatically select other components which is OK). Click Next. 6. View the Configure JDBC resources screen. You should not make any changes. Click Next. 7. Check both connections and click Test Connections. After successful test, click Next. If the tests fail, something is broken. Go back to configure JDBC resources and check your service name/SID. 8. Check all schemas. Set a password (will be the same for all schemas). Enter the database information (service name, host name, port). Click Next. 9. Connections should test successfully. If not, go back and fix any errors. Click Next. 10. Click Next to pass through Optional Configuration. 11. Click Extend. 12. Click Done. 13. Open a terminal window and navigate to/execute: ODEE_HOME/documaker/j2ee/weblogic/oracle11g/bpel/antbuild.cmd14. Start the WebLogic Servers – AdminServer, jms_server, dmkr_server, idm_server. If you forgot how to do this, see the previous section Step 10. Note: if you previously changed the startManagedWebLogic.cmd script for WLS_USER and WLS_PASS you will need to make those changes again. 15. Start the WebLogic server soa_server1: MIDDLEWARE_HOME/user_projects/domains/idocumaker_domain/bin/startManagedWebLogic.cmd soa_server116. Open a browser to http://localhost:7001/console and login. 17. Navigate to Services?Data Sources and select DMKR_ASLINE. 18. Click the Targets tab. Check soa_server1, then click Save. Repeat for the DMKR_ADMIN data source. 19. Open a command prompt and navigate to ODEE_HOME/j2ee/weblogic/oracle11g/scripts, then execute deploy_soa.cmd. That's it! (As if that wasn't enough?) DOCUMAKER Deploy the sample MRL resources by navigating to/executing ODEE_HOME/documaker/mstrres/dmres/deploysamplemrl.bat. You should see approximately 500 resources deployed into the database. Start the Factory Services. Start?Run?services.msc. Locate the service named "ODDF xxxx" and right-click, select Start. Note that each Assembly Line has a separate Factory setup, including its own Factory service and Docupresentment service. The services are named for the assembly line and the machine on which they are installed (because you could have multiple machines servicing a single assembly line, so this allows for easy scripting to control all the services if you choose to do so. Repeat for the Docupresentment service. Note that each Assembly Line has a separate Docupresentment. Using Windows Explorer, navigate to ODEE_HOME/documaker/mstrres/dmres/input and select one of the XML files, and copy it into ODEE_HOME/documaker/hotdirectory. Note: if you chose a different hot directory during installation, copy the file there instead. Momentarily you should see the XML file disappear! Open browser and navigate to http://localhost:10001/DocumakerDashboard (previous versions 12.0-12.2 use http://localhost:10001/dashboard) and verify that job processed successfully. Note that some transactions may fail if you do not have a properly configured email server, and this is ok. You can set up a simple SMTP server (just search the internet for "SMTP developer" and you'll get several to choose from.  So... that's it? Where are we at this point? You now have a completely functional ODEE installation, from soup to nuts as they say. You can further expand your installation by doing some of the following activities: clustering WebLogic services configuring WebLogic for redundancy configuring Oracle 11g for RAC adding additional Factory servers for redundancy/processing capacity setting up a real MRL (instead of the sample resources) testing Documaker Web Services for job submission and more!  I certainly hope you've enjoyed this and find it useful. If you find yourself running into trouble, visit the Oracle Community for Documaker - there is plenty of activity there and you can ask questions. For more concentrated assistance, you can engage an Oracle consultant who is a subject matter expert to assist you. Feel free to email me [andy (dot) little (at) oracle (dot) com] and I can connect you with the appropriate resource to get started. Best of luck! -Andy 

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  • The fastest way to resize images from ASP.NET. And it’s (more) supported-ish.

    - by Bertrand Le Roy
    I’ve shown before how to resize images using GDI, which is fairly common but is explicitly unsupported because we know of very real problems that this can cause. Still, many sites still use that method because those problems are fairly rare, and because most people assume it’s the only way to get the job done. Plus, it works in medium trust. More recently, I’ve shown how you can use WPF APIs to do the same thing and get JPEG thumbnails, only 2.5 times faster than GDI (even now that GDI really ultimately uses WIC to read and write images). The boost in performance is great, but it comes at a cost, that you may or may not care about: it won’t work in medium trust. It’s also just as unsupported as the GDI option. What I want to show today is how to use the Windows Imaging Components from ASP.NET APIs directly, without going through WPF. The approach has the great advantage that it’s been tested and proven to scale very well. The WIC team tells me you should be able to call support and get answers if you hit problems. Caveats exist though. First, this is using interop, so until a signed wrapper sits in the GAC, it will require full trust. Second, the APIs have a very strong smell of native code and are definitely not .NET-friendly. And finally, the most serious problem is that older versions of Windows don’t offer MTA support for image decoding. MTA support is only available on Windows 7, Vista and Windows Server 2008. But on 2003 and XP, you’ll only get STA support. that means that the thread safety that we so badly need for server applications is not guaranteed on those operating systems. To make it work, you’d have to spin specialized threads yourself and manage the lifetime of your objects, which is outside the scope of this article. We’ll assume that we’re fine with al this and that we’re running on 7 or 2008 under full trust. Be warned that the code that follows is not simple or very readable. This is definitely not the easiest way to resize an image in .NET. Wrapping native APIs such as WIC in a managed wrapper is never easy, but fortunately we won’t have to: the WIC team already did it for us and released the results under MS-PL. The InteropServices folder, which contains the wrappers we need, is in the WicCop project but I’ve also included it in the sample that you can download from the link at the end of the article. In order to produce a thumbnail, we first have to obtain a decoding frame object that WIC can use. Like with WPF, that object will contain the command to decode a frame from the source image but won’t do the actual decoding until necessary. Getting the frame is done by reading the image bytes through a special WIC stream that you can obtain from a factory object that we’re going to reuse for lots of other tasks: var photo = File.ReadAllBytes(photoPath); var factory = (IWICComponentFactory)new WICImagingFactory(); var inputStream = factory.CreateStream(); inputStream.InitializeFromMemory(photo, (uint)photo.Length); var decoder = factory.CreateDecoderFromStream( inputStream, null, WICDecodeOptions.WICDecodeMetadataCacheOnLoad); var frame = decoder.GetFrame(0); We can read the dimensions of the frame using the following (somewhat ugly) code: uint width, height; frame.GetSize(out width, out height); This enables us to compute the dimensions of the thumbnail, as I’ve shown in previous articles. We now need to prepare the output stream for the thumbnail. WIC requires a special kind of stream, IStream (not implemented by System.IO.Stream) and doesn’t directlyunderstand .NET streams. It does provide a number of implementations but not exactly what we need here. We need to output to memory because we’ll want to persist the same bytes to the response stream and to a local file for caching. The memory-bound version of IStream requires a fixed-length buffer but we won’t know the length of the buffer before we resize. To solve that problem, I’ve built a derived class from MemoryStream that also implements IStream. The implementation is not very complicated, it just delegates the IStream methods to the base class, but it involves some native pointer manipulation. Once we have a stream, we need to build the encoder for the output format, which could be anything that WIC supports. For web thumbnails, our only reasonable options are PNG and JPEG. I explored PNG because it’s a lossless format, and because WIC does support PNG compression. That compression is not very efficient though and JPEG offers good quality with much smaller file sizes. On the web, it matters. I found the best PNG compression option (adaptive) to give files that are about twice as big as 100%-quality JPEG (an absurd setting), 4.5 times bigger than 95%-quality JPEG and 7 times larger than 85%-quality JPEG, which is more than acceptable quality. As a consequence, we’ll use JPEG. The JPEG encoder can be prepared as follows: var encoder = factory.CreateEncoder( Consts.GUID_ContainerFormatJpeg, null); encoder.Initialize(outputStream, WICBitmapEncoderCacheOption.WICBitmapEncoderNoCache); The next operation is to create the output frame: IWICBitmapFrameEncode outputFrame; var arg = new IPropertyBag2[1]; encoder.CreateNewFrame(out outputFrame, arg); Notice that we are passing in a property bag. This is where we’re going to specify our only parameter for encoding, the JPEG quality setting: var propBag = arg[0]; var propertyBagOption = new PROPBAG2[1]; propertyBagOption[0].pstrName = "ImageQuality"; propBag.Write(1, propertyBagOption, new object[] { 0.85F }); outputFrame.Initialize(propBag); We can then set the resolution for the thumbnail to be 96, something we weren’t able to do with WPF and had to hack around: outputFrame.SetResolution(96, 96); Next, we set the size of the output frame and create a scaler from the input frame and the computed dimensions of the target thumbnail: outputFrame.SetSize(thumbWidth, thumbHeight); var scaler = factory.CreateBitmapScaler(); scaler.Initialize(frame, thumbWidth, thumbHeight, WICBitmapInterpolationMode.WICBitmapInterpolationModeFant); The scaler is using the Fant method, which I think is the best looking one even if it seems a little softer than cubic (zoomed here to better show the defects): Cubic Fant Linear Nearest neighbor We can write the source image to the output frame through the scaler: outputFrame.WriteSource(scaler, new WICRect { X = 0, Y = 0, Width = (int)thumbWidth, Height = (int)thumbHeight }); And finally we commit the pipeline that we built and get the byte array for the thumbnail out of our memory stream: outputFrame.Commit(); encoder.Commit(); var outputArray = outputStream.ToArray(); outputStream.Close(); That byte array can then be sent to the output stream and to the cache file. Once we’ve gone through this exercise, it’s only natural to wonder whether it was worth the trouble. I ran this method, as well as GDI and WPF resizing over thirty twelve megapixel images for JPEG qualities between 70% and 100% and measured the file size and time to resize. Here are the results: Size of resized images   Time to resize thirty 12 megapixel images Not much to see on the size graph: sizes from WPF and WIC are equivalent, which is hardly surprising as WPF calls into WIC. There is just an anomaly for 75% for WPF that I noted in my previous article and that disappears when using WIC directly. But overall, using WPF or WIC over GDI represents a slight win in file size. The time to resize is more interesting. WPF and WIC get similar times although WIC seems to always be a little faster. Not surprising considering WPF is using WIC. The margin of error on this results is probably fairly close to the time difference. As we already knew, the time to resize does not depend on the quality level, only the size does. This means that the only decision you have to make here is size versus visual quality. This third approach to server-side image resizing on ASP.NET seems to converge on the fastest possible one. We have marginally better performance than WPF, but with some additional peace of mind that this approach is sanctioned for server-side usage by the Windows Imaging team. It still doesn’t work in medium trust. That is a problem and shows the way for future server-friendly managed wrappers around WIC. The sample code for this article can be downloaded from: http://weblogs.asp.net/blogs/bleroy/Samples/WicResize.zip The benchmark code can be found here (you’ll need to add your own images to the Images directory and then add those to the project, with content and copy if newer in the properties of the files in the solution explorer): http://weblogs.asp.net/blogs/bleroy/Samples/WicWpfGdiImageResizeBenchmark.zip WIC tools can be downloaded from: http://code.msdn.microsoft.com/wictools To conclude, here are some of the resized thumbnails at 85% fant:

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  • Metro: Creating an IndexedDbDataSource for WinJS

    - by Stephen.Walther
    The goal of this blog entry is to describe how you can create custom data sources which you can use with the controls in the WinJS library. In particular, I explain how you can create an IndexedDbDataSource which you can use to store and retrieve data from an IndexedDB database. If you want to skip ahead, and ignore all of the fascinating content in-between, I’ve included the complete code for the IndexedDbDataSource at the very bottom of this blog entry. What is IndexedDB? IndexedDB is a database in the browser. You can use the IndexedDB API with all modern browsers including Firefox, Chrome, and Internet Explorer 10. And, of course, you can use IndexedDB with Metro style apps written with JavaScript. If you need to persist data in a Metro style app written with JavaScript then IndexedDB is a good option. Each Metro app can only interact with its own IndexedDB databases. And, IndexedDB provides you with transactions, indices, and cursors – the elements of any modern database. An IndexedDB database might be different than the type of database that you normally use. An IndexedDB database is an object-oriented database and not a relational database. Instead of storing data in tables, you store data in object stores. You store JavaScript objects in an IndexedDB object store. You create new IndexedDB object stores by handling the upgradeneeded event when you attempt to open a connection to an IndexedDB database. For example, here’s how you would both open a connection to an existing database named TasksDB and create the TasksDB database when it does not already exist: var reqOpen = window.indexedDB.open(“TasksDB”, 2); reqOpen.onupgradeneeded = function (evt) { var newDB = evt.target.result; newDB.createObjectStore("tasks", { keyPath: "id", autoIncrement: true }); }; reqOpen.onsuccess = function () { var db = reqOpen.result; // Do something with db }; When you call window.indexedDB.open(), and the database does not already exist, then the upgradeneeded event is raised. In the code above, the upgradeneeded handler creates a new object store named tasks. The new object store has an auto-increment column named id which acts as the primary key column. If the database already exists with the right version, and you call window.indexedDB.open(), then the success event is raised. At that point, you have an open connection to the existing database and you can start doing something with the database. You use asynchronous methods to interact with an IndexedDB database. For example, the following code illustrates how you would add a new object to the tasks object store: var transaction = db.transaction(“tasks”, “readwrite”); var reqAdd = transaction.objectStore(“tasks”).add({ name: “Feed the dog” }); reqAdd.onsuccess = function() { // Tasks added successfully }; The code above creates a new database transaction, adds a new task to the tasks object store, and handles the success event. If the new task gets added successfully then the success event is raised. Creating a WinJS IndexedDbDataSource The most powerful control in the WinJS library is the ListView control. This is the control that you use to display a collection of items. If you want to display data with a ListView control, you need to bind the control to a data source. The WinJS library includes two objects which you can use as a data source: the List object and the StorageDataSource object. The List object enables you to represent a JavaScript array as a data source and the StorageDataSource enables you to represent the file system as a data source. If you want to bind an IndexedDB database to a ListView then you have a choice. You can either dump the items from the IndexedDB database into a List object or you can create a custom data source. I explored the first approach in a previous blog entry. In this blog entry, I explain how you can create a custom IndexedDB data source. Implementing the IListDataSource Interface You create a custom data source by implementing the IListDataSource interface. This interface contains the contract for the methods which the ListView needs to interact with a data source. The easiest way to implement the IListDataSource interface is to derive a new object from the base VirtualizedDataSource object. The VirtualizedDataSource object requires a data adapter which implements the IListDataAdapter interface. Yes, because of the number of objects involved, this is a little confusing. Your code ends up looking something like this: var IndexedDbDataSource = WinJS.Class.derive( WinJS.UI.VirtualizedDataSource, function (dbName, dbVersion, objectStoreName, upgrade, error) { this._adapter = new IndexedDbDataAdapter(dbName, dbVersion, objectStoreName, upgrade, error); this._baseDataSourceConstructor(this._adapter); }, { nuke: function () { this._adapter.nuke(); }, remove: function (key) { this._adapter.removeInternal(key); } } ); The code above is used to create a new class named IndexedDbDataSource which derives from the base VirtualizedDataSource class. In the constructor for the new class, the base class _baseDataSourceConstructor() method is called. A data adapter is passed to the _baseDataSourceConstructor() method. The code above creates a new method exposed by the IndexedDbDataSource named nuke(). The nuke() method deletes all of the objects from an object store. The code above also overrides a method named remove(). Our derived remove() method accepts any type of key and removes the matching item from the object store. Almost all of the work of creating a custom data source goes into building the data adapter class. The data adapter class implements the IListDataAdapter interface which contains the following methods: · change() · getCount() · insertAfter() · insertAtEnd() · insertAtStart() · insertBefore() · itemsFromDescription() · itemsFromEnd() · itemsFromIndex() · itemsFromKey() · itemsFromStart() · itemSignature() · moveAfter() · moveBefore() · moveToEnd() · moveToStart() · remove() · setNotificationHandler() · compareByIdentity Fortunately, you are not required to implement all of these methods. You only need to implement the methods that you actually need. In the case of the IndexedDbDataSource, I implemented the getCount(), itemsFromIndex(), insertAtEnd(), and remove() methods. If you are creating a read-only data source then you really only need to implement the getCount() and itemsFromIndex() methods. Implementing the getCount() Method The getCount() method returns the total number of items from the data source. So, if you are storing 10,000 items in an object store then this method would return the value 10,000. Here’s how I implemented the getCount() method: getCount: function () { var that = this; return new WinJS.Promise(function (complete, error) { that._getObjectStore().then(function (store) { var reqCount = store.count(); reqCount.onerror = that._error; reqCount.onsuccess = function (evt) { complete(evt.target.result); }; }); }); } The first thing that you should notice is that the getCount() method returns a WinJS promise. This is a requirement. The getCount() method is asynchronous which is a good thing because all of the IndexedDB methods (at least the methods implemented in current browsers) are also asynchronous. The code above retrieves an object store and then uses the IndexedDB count() method to get a count of the items in the object store. The value is returned from the promise by calling complete(). Implementing the itemsFromIndex method When a ListView displays its items, it calls the itemsFromIndex() method. By default, it calls this method multiple times to get different ranges of items. Three parameters are passed to the itemsFromIndex() method: the requestIndex, countBefore, and countAfter parameters. The requestIndex indicates the index of the item from the database to show. The countBefore and countAfter parameters represent hints. These are integer values which represent the number of items before and after the requestIndex to retrieve. Again, these are only hints and you can return as many items before and after the request index as you please. Here’s how I implemented the itemsFromIndex method: itemsFromIndex: function (requestIndex, countBefore, countAfter) { var that = this; return new WinJS.Promise(function (complete, error) { that.getCount().then(function (count) { if (requestIndex >= count) { return WinJS.Promise.wrapError(new WinJS.ErrorFromName(WinJS.UI.FetchError.doesNotExist)); } var startIndex = Math.max(0, requestIndex - countBefore); var endIndex = Math.min(count, requestIndex + countAfter + 1); that._getObjectStore().then(function (store) { var index = 0; var items = []; var req = store.openCursor(); req.onerror = that._error; req.onsuccess = function (evt) { var cursor = evt.target.result; if (index < startIndex) { index = startIndex; cursor.advance(startIndex); return; } if (cursor && index < endIndex) { index++; items.push({ key: cursor.value[store.keyPath].toString(), data: cursor.value }); cursor.continue(); return; } results = { items: items, offset: requestIndex - startIndex, totalCount: count }; complete(results); }; }); }); }); } In the code above, a cursor is used to iterate through the objects in an object store. You fetch the next item in the cursor by calling either the cursor.continue() or cursor.advance() method. The continue() method moves forward by one object and the advance() method moves forward a specified number of objects. Each time you call continue() or advance(), the success event is raised again. If the cursor is null then you know that you have reached the end of the cursor and you can return the results. Some things to be careful about here. First, the return value from the itemsFromIndex() method must implement the IFetchResult interface. In particular, you must return an object which has an items, offset, and totalCount property. Second, each item in the items array must implement the IListItem interface. Each item should have a key and a data property. Implementing the insertAtEnd() Method When creating the IndexedDbDataSource, I wanted to go beyond creating a simple read-only data source and support inserting and deleting objects. If you want to support adding new items with your data source then you need to implement the insertAtEnd() method. Here’s how I implemented the insertAtEnd() method for the IndexedDbDataSource: insertAtEnd:function(unused, data) { var that = this; return new WinJS.Promise(function (complete, error) { that._getObjectStore("readwrite").done(function(store) { var reqAdd = store.add(data); reqAdd.onerror = that._error; reqAdd.onsuccess = function (evt) { var reqGet = store.get(evt.target.result); reqGet.onerror = that._error; reqGet.onsuccess = function (evt) { var newItem = { key:evt.target.result[store.keyPath].toString(), data:evt.target.result } complete(newItem); }; }; }); }); } When implementing the insertAtEnd() method, you need to be careful to return an object which implements the IItem interface. In particular, you should return an object that has a key and a data property. The key must be a string and it uniquely represents the new item added to the data source. The value of the data property represents the new item itself. Implementing the remove() Method Finally, you use the remove() method to remove an item from the data source. You call the remove() method with the key of the item which you want to remove. Implementing the remove() method in the case of the IndexedDbDataSource was a little tricky. The problem is that an IndexedDB object store uses an integer key and the VirtualizedDataSource requires a string key. For that reason, I needed to override the remove() method in the derived IndexedDbDataSource class like this: var IndexedDbDataSource = WinJS.Class.derive( WinJS.UI.VirtualizedDataSource, function (dbName, dbVersion, objectStoreName, upgrade, error) { this._adapter = new IndexedDbDataAdapter(dbName, dbVersion, objectStoreName, upgrade, error); this._baseDataSourceConstructor(this._adapter); }, { nuke: function () { this._adapter.nuke(); }, remove: function (key) { this._adapter.removeInternal(key); } } ); When you call remove(), you end up calling a method of the IndexedDbDataAdapter named removeInternal() . Here’s what the removeInternal() method looks like: setNotificationHandler: function (notificationHandler) { this._notificationHandler = notificationHandler; }, removeInternal: function(key) { var that = this; return new WinJS.Promise(function (complete, error) { that._getObjectStore("readwrite").done(function (store) { var reqDelete = store.delete (key); reqDelete.onerror = that._error; reqDelete.onsuccess = function (evt) { that._notificationHandler.removed(key.toString()); complete(); }; }); }); } The removeInternal() method calls the IndexedDB delete() method to delete an item from the object store. If the item is deleted successfully then the _notificationHandler.remove() method is called. Because we are not implementing the standard IListDataAdapter remove() method, we need to notify the data source (and the ListView control bound to the data source) that an item has been removed. The way that you notify the data source is by calling the _notificationHandler.remove() method. Notice that we get the _notificationHandler in the code above by implementing another method in the IListDataAdapter interface: the setNotificationHandler() method. You can raise the following types of notifications using the _notificationHandler: · beginNotifications() · changed() · endNotifications() · inserted() · invalidateAll() · moved() · removed() · reload() These methods are all part of the IListDataNotificationHandler interface in the WinJS library. Implementing the nuke() Method I wanted to implement a method which would remove all of the items from an object store. Therefore, I created a method named nuke() which calls the IndexedDB clear() method: nuke: function () { var that = this; return new WinJS.Promise(function (complete, error) { that._getObjectStore("readwrite").done(function (store) { var reqClear = store.clear(); reqClear.onerror = that._error; reqClear.onsuccess = function (evt) { that._notificationHandler.reload(); complete(); }; }); }); } Notice that the nuke() method calls the _notificationHandler.reload() method to notify the ListView to reload all of the items from its data source. Because we are implementing a custom method here, we need to use the _notificationHandler to send an update. Using the IndexedDbDataSource To illustrate how you can use the IndexedDbDataSource, I created a simple task list app. You can add new tasks, delete existing tasks, and nuke all of the tasks. You delete an item by selecting an item (swipe or right-click) and clicking the Delete button. Here’s the HTML page which contains the ListView, the form for adding new tasks, and the buttons for deleting and nuking tasks: <!DOCTYPE html> <html> <head> <meta charset="utf-8" /> <title>DataSources</title> <!-- WinJS references --> <link href="//Microsoft.WinJS.1.0.RC/css/ui-dark.css" rel="stylesheet" /> <script src="//Microsoft.WinJS.1.0.RC/js/base.js"></script> <script src="//Microsoft.WinJS.1.0.RC/js/ui.js"></script> <!-- DataSources references --> <link href="indexedDb.css" rel="stylesheet" /> <script type="text/javascript" src="indexedDbDataSource.js"></script> <script src="indexedDb.js"></script> </head> <body> <div id="tmplTask" data-win-control="WinJS.Binding.Template"> <div class="taskItem"> Id: <span data-win-bind="innerText:id"></span> <br /><br /> Name: <span data-win-bind="innerText:name"></span> </div> </div> <div id="lvTasks" data-win-control="WinJS.UI.ListView" data-win-options="{ itemTemplate: select('#tmplTask'), selectionMode: 'single' }"></div> <form id="frmAdd"> <fieldset> <legend>Add Task</legend> <label>New Task</label> <input id="inputTaskName" required /> <button>Add</button> </fieldset> </form> <button id="btnNuke">Nuke</button> <button id="btnDelete">Delete</button> </body> </html> And here is the JavaScript code for the TaskList app: /// <reference path="//Microsoft.WinJS.1.0.RC/js/base.js" /> /// <reference path="//Microsoft.WinJS.1.0.RC/js/ui.js" /> function init() { WinJS.UI.processAll().done(function () { var lvTasks = document.getElementById("lvTasks").winControl; // Bind the ListView to its data source var tasksDataSource = new DataSources.IndexedDbDataSource("TasksDB", 1, "tasks", upgrade); lvTasks.itemDataSource = tasksDataSource; // Wire-up Add, Delete, Nuke buttons document.getElementById("frmAdd").addEventListener("submit", function (evt) { evt.preventDefault(); tasksDataSource.beginEdits(); tasksDataSource.insertAtEnd(null, { name: document.getElementById("inputTaskName").value }).done(function (newItem) { tasksDataSource.endEdits(); document.getElementById("frmAdd").reset(); lvTasks.ensureVisible(newItem.index); }); }); document.getElementById("btnDelete").addEventListener("click", function () { if (lvTasks.selection.count() == 1) { lvTasks.selection.getItems().done(function (items) { tasksDataSource.remove(items[0].data.id); }); } }); document.getElementById("btnNuke").addEventListener("click", function () { tasksDataSource.nuke(); }); // This method is called to initialize the IndexedDb database function upgrade(evt) { var newDB = evt.target.result; newDB.createObjectStore("tasks", { keyPath: "id", autoIncrement: true }); } }); } document.addEventListener("DOMContentLoaded", init); The IndexedDbDataSource is created and bound to the ListView control with the following two lines of code: var tasksDataSource = new DataSources.IndexedDbDataSource("TasksDB", 1, "tasks", upgrade); lvTasks.itemDataSource = tasksDataSource; The IndexedDbDataSource is created with four parameters: the name of the database to create, the version of the database to create, the name of the object store to create, and a function which contains code to initialize the new database. The upgrade function creates a new object store named tasks with an auto-increment property named id: function upgrade(evt) { var newDB = evt.target.result; newDB.createObjectStore("tasks", { keyPath: "id", autoIncrement: true }); } The Complete Code for the IndexedDbDataSource Here’s the complete code for the IndexedDbDataSource: (function () { /************************************************ * The IndexedDBDataAdapter enables you to work * with a HTML5 IndexedDB database. *************************************************/ var IndexedDbDataAdapter = WinJS.Class.define( function (dbName, dbVersion, objectStoreName, upgrade, error) { this._dbName = dbName; // database name this._dbVersion = dbVersion; // database version this._objectStoreName = objectStoreName; // object store name this._upgrade = upgrade; // database upgrade script this._error = error || function (evt) { console.log(evt.message); }; }, { /******************************************* * IListDataAdapter Interface Methods ********************************************/ getCount: function () { var that = this; return new WinJS.Promise(function (complete, error) { that._getObjectStore().then(function (store) { var reqCount = store.count(); reqCount.onerror = that._error; reqCount.onsuccess = function (evt) { complete(evt.target.result); }; }); }); }, itemsFromIndex: function (requestIndex, countBefore, countAfter) { var that = this; return new WinJS.Promise(function (complete, error) { that.getCount().then(function (count) { if (requestIndex >= count) { return WinJS.Promise.wrapError(new WinJS.ErrorFromName(WinJS.UI.FetchError.doesNotExist)); } var startIndex = Math.max(0, requestIndex - countBefore); var endIndex = Math.min(count, requestIndex + countAfter + 1); that._getObjectStore().then(function (store) { var index = 0; var items = []; var req = store.openCursor(); req.onerror = that._error; req.onsuccess = function (evt) { var cursor = evt.target.result; if (index < startIndex) { index = startIndex; cursor.advance(startIndex); return; } if (cursor && index < endIndex) { index++; items.push({ key: cursor.value[store.keyPath].toString(), data: cursor.value }); cursor.continue(); return; } results = { items: items, offset: requestIndex - startIndex, totalCount: count }; complete(results); }; }); }); }); }, insertAtEnd:function(unused, data) { var that = this; return new WinJS.Promise(function (complete, error) { that._getObjectStore("readwrite").done(function(store) { var reqAdd = store.add(data); reqAdd.onerror = that._error; reqAdd.onsuccess = function (evt) { var reqGet = store.get(evt.target.result); reqGet.onerror = that._error; reqGet.onsuccess = function (evt) { var newItem = { key:evt.target.result[store.keyPath].toString(), data:evt.target.result } complete(newItem); }; }; }); }); }, setNotificationHandler: function (notificationHandler) { this._notificationHandler = notificationHandler; }, /***************************************** * IndexedDbDataSource Method ******************************************/ removeInternal: function(key) { var that = this; return new WinJS.Promise(function (complete, error) { that._getObjectStore("readwrite").done(function (store) { var reqDelete = store.delete (key); reqDelete.onerror = that._error; reqDelete.onsuccess = function (evt) { that._notificationHandler.removed(key.toString()); complete(); }; }); }); }, nuke: function () { var that = this; return new WinJS.Promise(function (complete, error) { that._getObjectStore("readwrite").done(function (store) { var reqClear = store.clear(); reqClear.onerror = that._error; reqClear.onsuccess = function (evt) { that._notificationHandler.reload(); complete(); }; }); }); }, /******************************************* * Private Methods ********************************************/ _ensureDbOpen: function () { var that = this; // Try to get cached Db if (that._cachedDb) { return WinJS.Promise.wrap(that._cachedDb); } // Otherwise, open the database return new WinJS.Promise(function (complete, error, progress) { var reqOpen = window.indexedDB.open(that._dbName, that._dbVersion); reqOpen.onerror = function (evt) { error(); }; reqOpen.onupgradeneeded = function (evt) { that._upgrade(evt); that._notificationHandler.invalidateAll(); }; reqOpen.onsuccess = function () { that._cachedDb = reqOpen.result; complete(that._cachedDb); }; }); }, _getObjectStore: function (type) { type = type || "readonly"; var that = this; return new WinJS.Promise(function (complete, error) { that._ensureDbOpen().then(function (db) { var transaction = db.transaction(that._objectStoreName, type); complete(transaction.objectStore(that._objectStoreName)); }); }); }, _get: function (key) { return new WinJS.Promise(function (complete, error) { that._getObjectStore().done(function (store) { var reqGet = store.get(key); reqGet.onerror = that._error; reqGet.onsuccess = function (item) { complete(item); }; }); }); } } ); var IndexedDbDataSource = WinJS.Class.derive( WinJS.UI.VirtualizedDataSource, function (dbName, dbVersion, objectStoreName, upgrade, error) { this._adapter = new IndexedDbDataAdapter(dbName, dbVersion, objectStoreName, upgrade, error); this._baseDataSourceConstructor(this._adapter); }, { nuke: function () { this._adapter.nuke(); }, remove: function (key) { this._adapter.removeInternal(key); } } ); WinJS.Namespace.define("DataSources", { IndexedDbDataSource: IndexedDbDataSource }); })(); Summary In this blog post, I provided an overview of how you can create a new data source which you can use with the WinJS library. I described how you can create an IndexedDbDataSource which you can use to bind a ListView control to an IndexedDB database. While describing how you can create a custom data source, I explained how you can implement the IListDataAdapter interface. You also learned how to raise notifications — such as a removed or invalidateAll notification — by taking advantage of the methods of the IListDataNotificationHandler interface.

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  • A WPF Image Button

    - by psheriff
    Instead of a normal button with words, sometimes you want a button that is just graphical. Yes, you can put an Image control in the Content of a normal Button control, but you still have the button outline, and trying to change the style can be rather difficult. Instead I like creating a user control that simulates a button, but just accepts an image. Figure 1 shows an example of three of these custom user controls to represent minimize, maximize and close buttons for a borderless window. Notice the highlighted image button has a gray rectangle around it. You will learn how to highlight using the VisualStateManager in this blog post.Figure 1: Creating a custom user control for things like image buttons gives you complete control over the look and feel.I would suggest you read my previous blog post on creating a custom Button user control as that is a good primer for what I am going to expand upon in this blog post. You can find this blog post at http://weblogs.asp.net/psheriff/archive/2012/08/10/create-your-own-wpf-button-user-controls.aspx.The User ControlThe XAML for this image button user control contains just a few controls, plus a Visual State Manager. The basic outline of the user control is shown below:<Border Grid.Row="0"        Name="borMain"        Style="{StaticResource pdsaButtonImageBorderStyle}"        MouseEnter="borMain_MouseEnter"        MouseLeave="borMain_MouseLeave"        MouseLeftButtonDown="borMain_MouseLeftButtonDown">  <VisualStateManager.VisualStateGroups>  ... MORE XAML HERE ...  </VisualStateManager.VisualStateGroups>  <Image Style="{StaticResource pdsaButtonImageImageStyle}"         Visibility="{Binding Path=Visibility}"         Source="{Binding Path=ImageUri}"         ToolTip="{Binding Path=ToolTip}" /></Border>There is a Border control named borMain and a single Image control in this user control. That is all that is needed to display the buttons shown in Figure 1. The definition for this user control is in a DLL named PDSA.WPF. The Style definitions for both the Border and the Image controls are contained in a resource dictionary names PDSAButtonStyles.xaml. Using a resource dictionary allows you to create a few different resource dictionaries, each with a different theme for the buttons.The Visual State ManagerTo display the highlight around the button as your mouse moves over the control, you will need to add a Visual State Manager group. Two different states are needed; MouseEnter and MouseLeave. In the MouseEnter you create a ColorAnimation to modify the BorderBrush color of the Border control. You specify the color to animate as “DarkGray”. You set the duration to less than a second. The TargetName of this storyboard is the name of the Border control “borMain” and since we are specifying a single color, you need to set the TargetProperty to “BorderBrush.Color”. You do not need any storyboard for the MouseLeave state. Leaving this VisualState empty tells the Visual State Manager to put everything back the way it was before the MouseEnter event.<VisualStateManager.VisualStateGroups>  <VisualStateGroup Name="MouseStates">    <VisualState Name="MouseEnter">      <Storyboard>        <ColorAnimation             To="DarkGray"            Duration="0:0:00.1"            Storyboard.TargetName="borMain"            Storyboard.TargetProperty="BorderBrush.Color" />      </Storyboard>    </VisualState>    <VisualState Name="MouseLeave" />  </VisualStateGroup></VisualStateManager.VisualStateGroups>Writing the Mouse EventsTo trigger the Visual State Manager to run its storyboard in response to the specified event, you need to respond to the MouseEnter event on the Border control. In the code behind for this event call the GoToElementState() method of the VisualStateManager class exposed by the user control. To this method you will pass in the target element (“borMain”) and the state (“MouseEnter”). The VisualStateManager will then run the storyboard contained within the defined state in the XAML.private void borMain_MouseEnter(object sender,  MouseEventArgs e){  VisualStateManager.GoToElementState(borMain,    "MouseEnter", true);}You also need to respond to the MouseLeave event. In this event you call the VisualStateManager as well, but specify “MouseLeave” as the state to go to.private void borMain_MouseLeave(object sender, MouseEventArgs e){  VisualStateManager.GoToElementState(borMain,     "MouseLeave", true);}The Resource DictionaryBelow is the definition of the PDSAButtonStyles.xaml resource dictionary file contained in the PDSA.WPF DLL. This dictionary can be used as the default look and feel for any image button control you add to a window. <ResourceDictionary  ... >  <!-- ************************* -->  <!-- ** Image Button Styles ** -->  <!-- ************************* -->  <!-- Image/Text Button Border -->  <Style TargetType="Border"         x:Key="pdsaButtonImageBorderStyle">    <Setter Property="Margin"            Value="4" />    <Setter Property="Padding"            Value="2" />    <Setter Property="BorderBrush"            Value="Transparent" />    <Setter Property="BorderThickness"            Value="1" />    <Setter Property="VerticalAlignment"            Value="Top" />    <Setter Property="HorizontalAlignment"            Value="Left" />    <Setter Property="Background"            Value="Transparent" />  </Style>  <!-- Image Button -->  <Style TargetType="Image"         x:Key="pdsaButtonImageImageStyle">    <Setter Property="Width"            Value="40" />    <Setter Property="Margin"            Value="6" />    <Setter Property="VerticalAlignment"            Value="Top" />    <Setter Property="HorizontalAlignment"            Value="Left" />  </Style></ResourceDictionary>Using the Button ControlOnce you make a reference to the PDSA.WPF DLL from your WPF application you will see the “PDSAucButtonImage” control appear in your Toolbox. Drag and drop the button onto a Window or User Control in your application. I have not referenced the PDSAButtonStyles.xaml file within the control itself so you do need to add a reference to this resource dictionary somewhere in your application such as in the App.xaml.<Application.Resources>  <ResourceDictionary>    <ResourceDictionary.MergedDictionaries>      <ResourceDictionary         Source="/PDSA.WPF;component/PDSAButtonStyles.xaml" />    </ResourceDictionary.MergedDictionaries>  </ResourceDictionary></Application.Resources>This will give your buttons a default look and feel unless you override that dictionary on a specific Window or User Control or on an individual button. After you have given a global style to your application and you drag your image button onto a window, the following will appear in your XAML window.<my:PDSAucButtonImage ... />There will be some other attributes set on the above XAML, but you simply need to set the x:Name, the ToolTip and ImageUri properties. You will also want to respond to the Click event procedure in order to associate an action with clicking on this button. In the sample code you download for this blog post you will find the declaration of the Minimize button to be the following:<my:PDSAucButtonImage       x:Name="btnMinimize"       Click="btnMinimize_Click"       ToolTip="Minimize Application"       ImageUri="/PDSA.WPF;component/Images/Minus.png" />The ImageUri property is a dependency property in the PDSAucButtonImage user control. The x:Name and the ToolTip we get for free. You have to create the Click event procedure yourself. This is also created in the PDSAucButtonImage user control as follows:private void borMain_MouseLeftButtonDown(object sender,  MouseButtonEventArgs e){  RaiseClick(e);}public delegate void ClickEventHandler(object sender,  RoutedEventArgs e);public event ClickEventHandler Click;protected void RaiseClick(RoutedEventArgs e){  if (null != Click)    Click(this, e);}Since a Border control does not have a Click event you will create one by using the MouseLeftButtonDown on the border to fire an event you create called “Click”.SummaryCreating your own image button control can be done in a variety of ways. In this blog post I showed you how to create a custom user control and simulate a button using a Border and Image control. With just a little bit of code to respond to the MouseLeftButtonDown event on the border you can raise your own Click event. Dependency properties, such as ImageUri, allow you to set attributes on your custom user control. Feel free to expand on this button by adding additional dependency properties, change the resource dictionary, and even the animation to make this button look and act like you want.NOTE: You can download the sample code for this article by visiting my website at http://www.pdsa.com/downloads. Select “Tips & Tricks”, then select “A WPF Image  Button” from the drop down list.

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  • Caveats with the runAllManagedModulesForAllRequests in IIS 7/8

    - by Rick Strahl
    One of the nice enhancements in IIS 7 (and now 8) is the ability to be able to intercept non-managed - ie. non ASP.NET served - requests from within ASP.NET managed modules. This opened up a ton of new functionality that could be applied across non-managed content using .NET code. I thought I had a pretty good handle on how IIS 7's Integrated mode pipeline works, but when I put together some samples last tonight I realized that the way that managed and unmanaged requests fire into the pipeline is downright confusing especially when it comes to the runAllManagedModulesForAllRequests attribute. There are a number of settings that can affect whether a managed module receives non-ASP.NET content requests such as static files or requests from other frameworks like PHP or ASP classic, and this is topic of this blog post. Native and Managed Modules The integrated mode IIS pipeline for IIS 7 and later - as the name suggests - allows for integration of ASP.NET pipeline events in the IIS request pipeline. Natively IIS runs unmanaged code and there are a host of native mode modules that handle the core behavior of IIS. If you set up a new IIS site or application without managed code support only the native modules are supported and fired without any interaction between native and managed code. If you use the Integrated pipeline with managed code enabled however things get a little more confusing as there both native modules and .NET managed modules can fire against the same IIS request. If you open up the IIS Modules dialog you see both managed and unmanaged modules. Unmanaged modules point at physical files on disk, while unmanaged modules point at .NET types and files referenced from the GAC or the current project's BIN folder. Both native and managed modules can co-exist and execute side by side and on the same request. When running in IIS 7 the IIS pipeline actually instantiates a the ASP.NET  runtime (via the System.Web.PipelineRuntime class) which unlike the core HttpRuntime classes in ASP.NET receives notification callbacks when IIS integrated mode events fire. The IIS pipeline is smart enough to detect whether managed handlers are attached and if they're none these notifications don't fire, improving performance. The good news about all of this for .NET devs is that ASP.NET style modules can be used for just about every kind of IIS request. All you need to do is create a new Web Application and enable ASP.NET on it, and then attach managed handlers. Handlers can look at ASP.NET content (ie. ASPX pages, MVC, WebAPI etc. requests) as well as non-ASP.NET content including static content like HTML files, images, javascript and css resources etc. It's very cool that this capability has been surfaced. However, with that functionality comes a lot of responsibility. Because every request passes through the ASP.NET pipeline if managed modules (or handlers) are attached there are possible performance implications that come with it. Running through the ASP.NET pipeline does add some overhead. ASP.NET and Your Own Modules When you create a new ASP.NET project typically the Visual Studio templates create the modules section like this: <system.webServer> <validation validateIntegratedModeConfiguration="false" /> <modules runAllManagedModulesForAllRequests="true" > </modules> </system.webServer> Specifically the interesting thing about this is the runAllManagedModulesForAllRequest="true" flag, which seems to indicate that it controls whether any registered modules always run, even when the value is set to false. Realistically though this flag does not control whether managed code is fired for all requests or not. Rather it is an override for the preCondition flag on a particular handler. With the flag set to the default true setting, you can assume that pretty much every IIS request you receive ends up firing through your ASP.NET module pipeline and every module you have configured is accessed even by non-managed requests like static files. In other words, your module will have to handle all requests. Now so far so obvious. What's not quite so obvious is what happens when you set the runAllManagedModulesForAllRequest="false". You probably would expect that immediately the non-ASP.NET requests no longer get funnelled through the ASP.NET Module pipeline. But that's not what actually happens. For example, if I create a module like this:<add name="SharewareModule" type="HowAspNetWorks.SharewareMessageModule" /> by default it will fire against ALL requests regardless of the runAllManagedModulesForAllRequests flag. Even if the value runAllManagedModulesForAllRequests="false", the module is fired. Not quite expected. So what is the runAllManagedModulesForAllRequests really good for? It's essentially an override for managedHandler preCondition. If I declare my handler in web.config like this:<add name="SharewareModule" type="HowAspNetWorks.SharewareMessageModule" preCondition="managedHandler" /> and the runAllManagedModulesForAllRequests="false" my module only fires against managed requests. If I switch the flag to true, now my module ends up handling all IIS requests that are passed through from IIS. The moral of the story here is that if you intend to only look at ASP.NET content, you should always set the preCondition="managedHandler" attribute to ensure that only managed requests are fired on this module. But even if you do this, realize that runAllManagedModulesForAllRequests="true" can override this setting. runAllManagedModulesForAllRequests and Http Application Events Another place the runAllManagedModulesForAllRequest attribute affects is the Global Http Application object (typically in global.asax) and the Application_XXXX events that you can hook up there. So while the events there are dynamically hooked up to the application class, they basically behave as if they were set with the preCodition="managedHandler" configuration switch. The end result is that if you have runAllManagedModulesForAllRequests="true" you'll see every Http request passed through the Application_XXXX events, and you only see ASP.NET requests with the flag set to "false". What's all that mean? Configuring an application to handle requests for both ASP.NET and other content requests can be tricky especially if you need to mix modules that might require both. Couple of things are important to remember. If your module doesn't need to look at every request, by all means set a preCondition="managedHandler" on it. This will at least allow it to respond to the runAllManagedModulesForAllRequests="false" flag and then only process ASP.NET requests. Look really carefully to see whether you actually need runAllManagedModulesForAllRequests="true" in your applications as set by the default new project templates in Visual Studio. Part of the reason, this is the default because it was required for the initial versions of IIS 7 and ASP.NET 2 in order to handle MVC extensionless URLs. However, if you are running IIS 7 or later and .NET 4.0 you can use the ExtensionlessUrlHandler instead to allow you MVC functionality without requiring runAllManagedModulesForAllRequests="true": <handlers> <remove name="ExtensionlessUrlHandler-Integrated-4.0" /> <add name="ExtensionlessUrlHandler-Integrated-4.0" path="*." verb="GET,HEAD,POST,DEBUG,PUT,DELETE,PATCH,OPTIONS" type="System.Web.Handlers.TransferRequestHandler" preCondition="integratedMode,runtimeVersionv4.0" /> </handlers> Oddly this is the default for Visual Studio 2012 MVC template apps, so I'm not sure why the default template still adds runAllManagedModulesForAllRequests="true" is - it should be enabled only if there's a specific need to access non ASP.NET requests. As a side note, it's interesting that when you access a static HTML resource, you can actually write into the Response object and get the output to show, which is trippy. I haven't looked closely to see how this works - whether ASP.NET just fires directly into the native output stream or whether the static requests are re-routed directly through the ASP.NET pipeline once a managed code module is detected. This doesn't work for all non ASP.NET resources - for example, I can't do the same with ASP classic requests, but it makes for an interesting demo when injecting HTML content into a static HTML page :-) Note that on the original Windows Server 2008 and Vista (IIS 7.0) you might need a HotFix in order for ExtensionLessUrlHandler to work properly for MVC projects. On my live server I needed it (about 6 months ago), but others have observed that the latest service updates have integrated this functionality and the hotfix is not required. On IIS 7.5 and later I've not needed any patches for things to just work. Plan for non-ASP.NET Requests It's important to remember that if you write a .NET Module to run on IIS 7, there's no way for you to prevent non-ASP.NET requests from hitting your module. So make sure you plan to support requests to extensionless URLs, to static resources like files. Luckily ASP.NET creates a full Request and full Response object for you for non ASP.NET content. So even for static files and even for ASP classic for example, you can look at Request.FilePath or Request.ContentType (in post handler pipeline events) to determine what content you are dealing with. As always with Module design make sure you check for the conditions in your code that make the module applicable and if a filter fails immediately exit - minimize the code that runs if your module doesn't need to process the request.© Rick Strahl, West Wind Technologies, 2005-2012Posted in IIS7   ASP.NET   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

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  • Performance Enhancement in Full-Text Search Query

    - by Calvin Sun
    Ever since its first release, we are continuing consolidating and developing InnoDB Full-Text Search feature. There is one recent improvement that worth blogging about. It is an effort with MySQL Optimizer team that simplifies some common queries’ Query Plans and dramatically shorted the query time. I will describe the issue, our solution and the end result by some performance numbers to demonstrate our efforts in continuing enhancement the Full-Text Search capability. The Issue: As we had discussed in previous Blogs, InnoDB implements Full-Text index as reversed auxiliary tables. The query once parsed will be reinterpreted into several queries into related auxiliary tables and then results are merged and consolidated to come up with the final result. So at the end of the query, we’ll have all matching records on hand, sorted by their ranking or by their Doc IDs. Unfortunately, MySQL’s optimizer and query processing had been initially designed for MyISAM Full-Text index, and sometimes did not fully utilize the complete result package from InnoDB. Here are a couple examples: Case 1: Query result ordered by Rank with only top N results: mysql> SELECT FTS_DOC_ID, MATCH (title, body) AGAINST ('database') AS SCORE FROM articles ORDER BY score DESC LIMIT 1; In this query, user tries to retrieve a single record with highest ranking. It should have a quick answer once we have all the matching documents on hand, especially if there are ranked. However, before this change, MySQL would almost retrieve rankings for almost every row in the table, sort them and them come with the top rank result. This whole retrieve and sort is quite unnecessary given the InnoDB already have the answer. In a real life case, user could have millions of rows, so in the old scheme, it would retrieve millions of rows' ranking and sort them, even if our FTS already found there are two 3 matched rows. Apparently, the million ranking retrieve is done in vain. In above case, it should just ask for 3 matched rows' ranking, all other rows' ranking are 0. If it want the top ranking, then it can just get the first record from our already sorted result. Case 2: Select Count(*) on matching records: mysql> SELECT COUNT(*) FROM articles WHERE MATCH (title,body) AGAINST ('database' IN NATURAL LANGUAGE MODE); In this case, InnoDB search can find matching rows quickly and will have all matching rows. However, before our change, in the old scheme, every row in the table was requested by MySQL one by one, just to check whether its ranking is larger than 0, and later comes up a count. In fact, there is no need for MySQL to fetch all rows, instead InnoDB already had all the matching records. The only thing need is to call an InnoDB API to retrieve the count The difference can be huge. Following query output shows how big the difference can be: mysql> select count(*) from searchindex_inno where match(si_title, si_text) against ('people')  +----------+ | count(*) | +----------+ | 666877 | +----------+ 1 row in set (16 min 17.37 sec) So the query took almost 16 minutes. Let’s see how long the InnoDB can come up the result. In InnoDB, you can obtain extra diagnostic printout by turning on “innodb_ft_enable_diag_print”, this will print out extra query info: Error log: keynr=2, 'people' NL search Total docs: 10954826 Total words: 0 UNION: Searching: 'people' Processing time: 2 secs: row(s) 666877: error: 10 ft_init() ft_init_ext() keynr=2, 'people' NL search Total docs: 10954826 Total words: 0 UNION: Searching: 'people' Processing time: 3 secs: row(s) 666877: error: 10 Output shows it only took InnoDB only 3 seconds to get the result, while the whole query took 16 minutes to finish. So large amount of time has been wasted on the un-needed row fetching. The Solution: The solution is obvious. MySQL can skip some of its steps, optimize its plan and obtain useful information directly from InnoDB. Some of savings from doing this include: 1) Avoid redundant sorting. Since InnoDB already sorted the result according to ranking. MySQL Query Processing layer does not need to sort to get top matching results. 2) Avoid row by row fetching to get the matching count. InnoDB provides all the matching records. All those not in the result list should all have ranking of 0, and no need to be retrieved. And InnoDB has a count of total matching records on hand. No need to recount. 3) Covered index scan. InnoDB results always contains the matching records' Document ID and their ranking. So if only the Document ID and ranking is needed, there is no need to go to user table to fetch the record itself. 4) Narrow the search result early, reduce the user table access. If the user wants to get top N matching records, we do not need to fetch all matching records from user table. We should be able to first select TOP N matching DOC IDs, and then only fetch corresponding records with these Doc IDs. Performance Results and comparison with MyISAM The result by this change is very obvious. I includes six testing result performed by Alexander Rubin just to demonstrate how fast the InnoDB query now becomes when comparing MyISAM Full-Text Search. These tests are base on the English Wikipedia data of 5.4 Million rows and approximately 16G table. The test was performed on a machine with 1 CPU Dual Core, SSD drive, 8G of RAM and InnoDB_buffer_pool is set to 8 GB. Table 1: SELECT with LIMIT CLAUSE mysql> SELECT si_title, match(si_title, si_text) against('family') as rel FROM si WHERE match(si_title, si_text) against('family') ORDER BY rel desc LIMIT 10; InnoDB MyISAM Times Faster Time for the query 1.63 sec 3 min 26.31 sec 127 You can see for this particular query (retrieve top 10 records), InnoDB Full-Text Search is now approximately 127 times faster than MyISAM. Table 2: SELECT COUNT QUERY mysql>select count(*) from si where match(si_title, si_text) against('family‘); +----------+ | count(*) | +----------+ | 293955 | +----------+ InnoDB MyISAM Times Faster Time for the query 1.35 sec 28 min 59.59 sec 1289 In this particular case, where there are 293k matching results, InnoDB took only 1.35 second to get all of them, while take MyISAM almost half an hour, that is about 1289 times faster!. Table 3: SELECT ID with ORDER BY and LIMIT CLAUSE for selected terms mysql> SELECT <ID>, match(si_title, si_text) against(<TERM>) as rel FROM si_<TB> WHERE match(si_title, si_text) against (<TERM>) ORDER BY rel desc LIMIT 10; Term InnoDB (time to execute) MyISAM(time to execute) Times Faster family 0.5 sec 5.05 sec 10.1 family film 0.95 sec 25.39 sec 26.7 Pizza restaurant orange county California 0.93 sec 32.03 sec 34.4 President united states of America 2.5 sec 36.98 sec 14.8 Table 4: SELECT title and text with ORDER BY and LIMIT CLAUSE for selected terms mysql> SELECT <ID>, si_title, si_text, ... as rel FROM si_<TB> WHERE match(si_title, si_text) against (<TERM>) ORDER BY rel desc LIMIT 10; Term InnoDB (time to execute) MyISAM(time to execute) Times Faster family 0.61 sec 41.65 sec 68.3 family film 1.15 sec 47.17 sec 41.0 Pizza restaurant orange county california 1.03 sec 48.2 sec 46.8 President united states of america 2.49 sec 44.61 sec 17.9 Table 5: SELECT ID with ORDER BY and LIMIT CLAUSE for selected terms mysql> SELECT <ID>, match(si_title, si_text) against(<TERM>) as rel  FROM si_<TB> WHERE match(si_title, si_text) against (<TERM>) ORDER BY rel desc LIMIT 10; Term InnoDB (time to execute) MyISAM(time to execute) Times Faster family 0.5 sec 5.05 sec 10.1 family film 0.95 sec 25.39 sec 26.7 Pizza restaurant orange county califormia 0.93 sec 32.03 sec 34.4 President united states of america 2.5 sec 36.98 sec 14.8 Table 6: SELECT COUNT(*) mysql> SELECT count(*) FROM si_<TB> WHERE match(si_title, si_text) against (<TERM>) LIMIT 10; Term InnoDB (time to execute) MyISAM(time to execute) Times Faster family 0.47 sec 82 sec 174.5 family film 0.83 sec 131 sec 157.8 Pizza restaurant orange county califormia 0.74 sec 106 sec 143.2 President united states of america 1.96 sec 220 sec 112.2  Again, table 3 to table 6 all showing InnoDB consistently outperform MyISAM in these queries by a large margin. It becomes obvious the InnoDB has great advantage over MyISAM in handling large data search. Summary: These results demonstrate the great performance we could achieve by making MySQL optimizer and InnoDB Full-Text Search more tightly coupled. I think there are still many cases that InnoDB’s result info have not been fully taken advantage of, which means we still have great room to improve. And we will continuously explore the area, and get more dramatic results for InnoDB full-text searches. Jimmy Yang, September 29, 2012

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  • Solving Big Problems with Oracle R Enterprise, Part II

    - by dbayard
    Part II – Solving Big Problems with Oracle R Enterprise In the first post in this series (see https://blogs.oracle.com/R/entry/solving_big_problems_with_oracle), we showed how you can use R to perform historical rate of return calculations against investment data sourced from a spreadsheet.  We demonstrated the calculations against sample data for a small set of accounts.  While this worked fine, in the real-world the problem is much bigger because the amount of data is much bigger.  So much bigger that our approach in the previous post won’t scale to meet the real-world needs. From our previous post, here are the challenges we need to conquer: The actual data that needs to be used lives in a database, not in a spreadsheet The actual data is much, much bigger- too big to fit into the normal R memory space and too big to want to move across the network The overall process needs to run fast- much faster than a single processor The actual data needs to be kept secured- another reason to not want to move it from the database and across the network And the process of calculating the IRR needs to be integrated together with other database ETL activities, so that IRR’s can be calculated as part of the data warehouse refresh processes In this post, we will show how we moved from sample data environment to working with full-scale data.  This post is based on actual work we did for a financial services customer during a recent proof-of-concept. Getting started with the Database At this point, we have some sample data and our IRR function.  We were at a similar point in our customer proof-of-concept exercise- we had sample data but we did not have the full customer data yet.  So our database was empty.  But, this was easily rectified by leveraging the transparency features of Oracle R Enterprise (see https://blogs.oracle.com/R/entry/analyzing_big_data_using_the).  The following code shows how we took our sample data SimpleMWRRData and easily turned it into a new Oracle database table called IRR_DATA via ore.create().  The code also shows how we can access the database table IRR_DATA as if it was a normal R data.frame named IRR_DATA. If we go to sql*plus, we can also check out our new IRR_DATA table: At this point, we now have our sample data loaded in the database as a normal Oracle table called IRR_DATA.  So, we now proceeded to test our R function working with database data. As our first test, we retrieved the data from a single account from the IRR_DATA table, pull it into local R memory, then call our IRR function.  This worked.  No SQL coding required! Going from Crawling to Walking Now that we have shown using our R code with database-resident data for a single account, we wanted to experiment with doing this for multiple accounts.  In other words, we wanted to implement the split-apply-combine technique we discussed in our first post in this series.  Fortunately, Oracle R Enterprise provides a very scalable way to do this with a function called ore.groupApply().  You can read more about ore.groupApply() here: https://blogs.oracle.com/R/entry/analyzing_big_data_using_the1 Here is an example of how we ask ORE to take our IRR_DATA table in the database, split it by the ACCOUNT column, apply a function that calls our SimpleMWRR() calculation, and then combine the results. (If you are following along at home, be sure to have installed our myIRR package on your database server via  “R CMD INSTALL myIRR”). The interesting thing about ore.groupApply is that the calculation is not actually performed in my desktop R environment from which I am running.  What actually happens is that ore.groupApply uses the Oracle database to perform the work.  And the Oracle database is what actually splits the IRR_DATA table by ACCOUNT.  Then the Oracle database takes the data for each account and sends it to an embedded R engine running on the database server to apply our R function.  Then the Oracle database combines all the individual results from the calls to the R function. This is significant because now the embedded R engine only needs to deal with the data for a single account at a time.  Regardless of whether we have 20 accounts or 1 million accounts or more, the R engine that performs the calculation does not care.  Given that normal R has a finite amount of memory to hold data, the ore.groupApply approach overcomes the R memory scalability problem since we only need to fit the data from a single account in R memory (not all of the data for all of the accounts). Additionally, the IRR_DATA does not need to be sent from the database to my desktop R program.  Even though I am invoking ore.groupApply from my desktop R program, because the actual SimpleMWRR calculation is run by the embedded R engine on the database server, the IRR_DATA does not need to leave the database server- this is both a performance benefit because network transmission of large amounts of data take time and a security benefit because it is harder to protect private data once you start shipping around your intranet. Another benefit, which we will discuss in a few paragraphs, is the ability to leverage Oracle database parallelism to run these calculations for dozens of accounts at once. From Walking to Running ore.groupApply is rather nice, but it still has the drawback that I run this from a desktop R instance.  This is not ideal for integrating into typical operational processes like nightly data warehouse refreshes or monthly statement generation.  But, this is not an issue for ORE.  Oracle R Enterprise lets us run this from the database using regular SQL, which is easily integrated into standard operations.  That is extremely exciting and the way we actually did these calculations in the customer proof. As part of Oracle R Enterprise, it provides a SQL equivalent to ore.groupApply which it refers to as “rqGroupEval”.  To use rqGroupEval via SQL, there is a bit of simple setup needed.  Basically, the Oracle Database needs to know the structure of the input table and the grouping column, which we are able to define using the database’s pipeline table function mechanisms. Here is the setup script: At this point, our initial setup of rqGroupEval is done for the IRR_DATA table.  The next step is to define our R function to the database.  We do that via a call to ORE’s rqScriptCreate. Now we can test it.  The SQL you use to run rqGroupEval uses the Oracle database pipeline table function syntax.  The first argument to irr_dataGroupEval is a cursor defining our input.  You can add additional where clauses and subqueries to this cursor as appropriate.  The second argument is any additional inputs to the R function.  The third argument is the text of a dummy select statement.  The dummy select statement is used by the database to identify the columns and datatypes to expect the R function to return.  The fourth argument is the column of the input table to split/group by.  The final argument is the name of the R function as you defined it when you called rqScriptCreate(). The Real-World Results In our real customer proof-of-concept, we had more sophisticated calculation requirements than shown in this simplified blog example.  For instance, we had to perform the rate of return calculations for 5 separate time periods, so the R code was enhanced to do so.  In addition, some accounts needed a time-weighted rate of return to be calculated, so we extended our approach and added an R function to do that.  And finally, there were also a few more real-world data irregularities that we needed to account for, so we added logic to our R functions to deal with those exceptions.  For the full-scale customer test, we loaded the customer data onto a Half-Rack Exadata X2-2 Database Machine.  As our half-rack had 48 physical cores (and 96 threads if you consider hyperthreading), we wanted to take advantage of that CPU horsepower to speed up our calculations.  To do so with ORE, it is as simple as leveraging the Oracle Database Parallel Query features.  Let’s look at the SQL used in the customer proof: Notice that we use a parallel hint on the cursor that is the input to our rqGroupEval function.  That is all we need to do to enable Oracle to use parallel R engines. Here are a few screenshots of what this SQL looked like in the Real-Time SQL Monitor when we ran this during the proof of concept (hint: you might need to right-click on these images to be able to view the images full-screen to see the entire image): From the above, you can notice a few things (numbers 1 thru 5 below correspond with highlighted numbers on the images above.  You may need to right click on the above images and view the images full-screen to see the entire image): The SQL completed in 110 seconds (1.8minutes) We calculated rate of returns for 5 time periods for each of 911k accounts (the number of actual rows returned by the IRRSTAGEGROUPEVAL operation) We accessed 103m rows of detailed cash flow/market value data (the number of actual rows returned by the IRR_STAGE2 operation) We ran with 72 degrees of parallelism spread across 4 database servers Most of our 110seconds was spent in the “External Procedure call” event On average, we performed 8,200 executions of our R function per second (110s/911k accounts) On average, each execution was passed 110 rows of data (103m detail rows/911k accounts) On average, we did 41,000 single time period rate of return calculations per second (each of the 8,200 executions of our R function did rate of return calculations for 5 time periods) On average, we processed over 900,000 rows of database data in R per second (103m detail rows/110s) R + Oracle R Enterprise: Best of R + Best of Oracle Database This blog post series started by describing a real customer problem: how to perform a lot of calculations on a lot of data in a short period of time.  While standard R proved to be a very good fit for writing the necessary calculations, the challenge of working with a lot of data in a short period of time remained. This blog post series showed how Oracle R Enterprise enables R to be used in conjunction with the Oracle Database to overcome the data volume and performance issues (as well as simplifying the operations and security issues).  It also showed that we could calculate 5 time periods of rate of returns for almost a million individual accounts in less than 2 minutes. In a future post, we will take the same R function and show how Oracle R Connector for Hadoop can be used in the Hadoop world.  In that next post, instead of having our data in an Oracle database, our data will live in Hadoop and we will how to use the Oracle R Connector for Hadoop and other Oracle Big Data Connectors to move data between Hadoop, R, and the Oracle Database easily.

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  • Stumbling Through: Visual Studio 2010 (Part IV)

    So finally we get to the fun part the fruits of all of our middle-tier/back end labors of generating classes to interface with an XML data source that the previous posts were about can now be presented quickly and easily to an end user.  I think.  Well see.  Well be using a WPF window to display all of our various MFL information that weve collected in the two XML files, and well provide a means of adding, updating and deleting each of these entities using as little code as possible.  Additionally, I would like to dig into the performance of this solution as well as the flexibility of it if were were to modify the underlying XML schema.  So first things first, lets create a WPF project and include our xml data in a data folder within.  On the main window, well drag out the following controls: A combo box to contain all of the teams A list box to show the players of the selected team, along with add/delete player buttons A text box tied to the selected players name, with a save button to save any changes made to the player name A combo box of all the available positions, tied to the currently selected players position A data grid tied to the statistics of the currently selected player, with add/delete statistic buttons This monstrosity of a form and its associated project will look like this (dont forget to reference the DataFoundation project from the Presentation project): To get to the visual data binding, as we learned in a previous post, you have to first make sure the project containing your bindable classes is compiled.  Do so, and then open the Data Sources pane to add a reference to the Teams and Positions classes in the DataFoundation project: Why only Team and Position?  Well, we will get to Players from Teams, and Statistics from Players so no need to make an interface for them as well see in a second.  As for Positions, well need a way to bind the dropdown to ALL positions they dont appear underneath any of the other classes so we need to reference it directly.  After adding these guys, expand every node in your Data Sources pane and see how the Team node allows you to drill into Players and then Statistics.  This is why there was no need to bring in a reference to those classes for the UI we are designing: Now for the seriously hard work of binding all of our controls to the correct data sources.  Drag the following items from the Data Sources pane to the specified control on the window design canvas: Team.Name > Teams combo box Team.Players.Name > Players list box Team.Players.Name > Player name text box Team.Players.Statistics > Statistics data grid Position.Name > Positions combo box That is it!  Really?  Well, no, not really there is one caveat here in that the Positions combo box is not bound the selected players position.  To do so, we will apply a binding to the position combo boxs SelectedValue to point to the current players PositionId value: That should do the trick now, all we need to worry about is loading the actual data.  Sadly, it appears as if we will need to drop to code in order to invoke our IO methods to load all teams and positions.  At least Visual Studio kindly created the stubs for us to do so, ultimately the code should look like this: Note the weirdness with the InitializeDataFiles call that is my current means of telling an IO where to load the data for each of the entities.  I havent thought of a more intuitive way than that yet, but do note that all data is loaded from Teams.xml besides for positions, which is loaded from Lookups.xml.   I think that may be all we need to do to at least load all of the data, lets run it and see: Yay!  All of our glorious data is being displayed!  Er, wait, whats up with the position dropdown?  Why is it red?  Lets select the RB and see if everything updates: Crap, the position didnt update to reflect the selected player, but everything else did.  Where did we go wrong in binding the position to the selected player?  Thinking about it a bit and comparing it to how traditional data binding works, I realize that we never set the value member (or some similar property) to tell the control to join the Id of the source (positions) to the position Id of the player.  I dont see a similar property to that on the combo box control, but I do see a property named SelectedValuePath that might be it, so I set it to Id and run the app again: Hey, all right!  No red box around the positions combo box.  Unfortunately, selecting the RB does not update the dropdown to point to Runningback.  Hmmm.  Now what could it be?  Maybe the problem is that we are loading teams before we are loading positions, so when it binds position Id, all of the positions arent loaded yet.  I went to the code behind and switched things so position loads first and no dice.  Same result when I run.  Why?  WHY?  Ok, ok, calm down, take a deep breath.  Get something with caffeine or sugar (preferably both) and think rationally. Ok, gigantic chocolate chip cookie and a mountain dew chaser have never let me down in the past, so dont fail me now!  Ah ha!  of course!  I didnt even have to finish the mountain dew and I think Ive got it:  Data Context.  By default, when setting on the selected value binding for the dropdown, the data context was list_team.  I dont even know what the heck list_team is, we want it to be bound to our team players view source resource instead, like this: Running it now and selecting the various players: Done and done.  Everything read and bound, thank you caffeine and sugar!  Oh, and thank you Visual Studio 2010.  Lets wire up some of those buttons now There has got to be a better way to do this, but it works for now.  What the add player button does is add a new player object to the currently selected team.  Unfortunately, I couldnt get the new object to automatically show up in the players list (something about not using an observable collection gotta look into this) so I just save the change immediately and reload the screen.  Terrible, but it works: Lets go after something easier:  The save button.  By default, as we type in new text for the players name, it is showing up in the list box as updated.  Cool!  Why couldnt my add new player logic do that?  Anyway, the save button should be as simple as invoking MFL.IO.Save for the selected player, like this: MFL.IO.Save((MFL.Player)lbTeamPlayers.SelectedItem, true); Surprisingly, that worked on the first try.  Lets see if we get as lucky with the Delete player button: MFL.IO.Delete((MFL.Player)lbTeamPlayers.SelectedItem); Refresh(); Note the use of the Refresh method again I cant seem to figure out why updates to the underlying data source are immediately reflected, but adds and deletes are not.  That is a problem for another day, and again my hunch is that I should be binding to something more complex than IEnumerable (like observable collection). Now that an example of the basic CRUD methods are wired up, I want to quickly investigate the performance of this beast.  Im going to make a special button to add 30 teams, each with 50 players and 10 seasons worth of stats.  If my math is right, that will end up with 15000 rows of data, a pretty hefty amount for an XML file.  The save of all this new data took a little over a minute, but that is acceptable because we wouldnt typically be saving batches of 15k records, and the resulting XML file size is a little over a megabyte.  Not huge, but big enough to see some read performance numbers or so I thought.  It reads this file and renders the first team in under a second.  That is unbelievable, but we are lazy loading and the file really wasnt that big.  I will increase it to 50 teams with 100 players and 20 seasons each - 100,000 rows.  It took a year and a half to save all of that data, and resulted in an 8 megabyte file.  Seriously, if you are loading XML files this large, get a freaking database!  Despite this, it STILL takes under a second to load and render the first team, which is interesting mostly because I thought that it was loading that entire 8 MB XML file behind the scenes.  I have to say that I am quite impressed with the performance of the LINQ to XML approach, particularly since I took no efforts to optimize any of this code and was fairly new to the concept from the start.  There might be some merit to this little project after all Look out SQL Server and Oracle, use XML files instead!  Next up, I am going to completely pull the rug out from under the UI and change a number of entities in our model.  How well will the code be regenerated?  How much effort will be required to tie things back together in the UI?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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  • IIS7 web farm - local or shared content?

    - by rbeier
    We're setting up an IIS7 web farm with two servers. Should each server have its own local copy of the content, or should they pull content directly from a UNC share? What are the pros and cons of each approach? We currently have a single live server WEB1, with content stored locally on a separate partition. A job periodically syncs WEB1 to a standby server WEB2, using robocopy for content and msdeploy for config. If WEB1 goes down, Nagios notifies us, and we manually run a script to move the IP addresses to WEB2's network interface. Both servers are actually VMs running on separate VMWare ESX 4 hosts. The servers are domain-joined. We have around 50-60 live sites on WEB1 - mostly ASP.NET, with a few that are just static HTML. Most are low-traffic "microsites". A few have moderate traffic, but none are massive. We'd like to change this so both WEB1 and WEB2 are actively serving content. This is mainly for reliability - if WEB1 goes down, we don't want to have to manually intervene to fail things over. Spreading the load is also nice, but the load is not high enough right now for us to need this. We're planning to configure our firewall to balance traffic across the two servers. It will detect when a server goes down and will send all the traffic to the remaining live server. We're planning to use sticky sessions for now... eventually we may move to SQL Server session state and stateless load balancing. But we need a way for the servers to share content. We were originally planning to move all the content to a UNC share. Our storage provider says they can set up a highly available SMB share for us. So if we go the UNC route, the storage shouldn't be a single point of failure. But we're wondering about the downsides to this approach: We'll need to change the physical paths for each site and virtual directory. There are also some projects that have absolute paths in their web.config files - we'll have to update those as well. We'll need to create a domain user for the web servers to access the share, and grant that user appropriate permissions. I haven't looked into this yet - I'm not sure if the application pool identity needs to be changed to this user, or if there's another way to tell IIS to use this account when connecting to the share. Sites will no longer be able to access their content if there's ever an Active Directory problem. In general, it just seems a lot more complicated, with more moving parts that could break. Our storage provider would create a volume for us on their redundant SAN. If I understand correctly, this SAN volume would be mounted on a VM running in their redundant VMWare environment; this VM would then expose the SMB share to our web servers. On the other hand, a benefit of the shared content approach is that we'd only need to deploy code to one place, and there would never be a temporary inconsistency between multiple copies of the content. This thread is pretty interesting, though some of these people are working at a much larger scale. I've just been discussing content so far, but we also need to think about configuration. I don't know if we can just use DFS replication for the applicationHost.config and other files, or if it's best to use the shared configuration feature with the config on a UNC share. What do you think? Thanks for your help, Richard

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  • Exchange 2010 OWA - a few questions about using multiple mailboxes

    - by Alexey Smolik
    We have an Exchange 2010 SP2 deployment and we need that our users could access multiple mailboxes in OWA. The problem is that a user (eg John Smith) needs to access not just somebody else's (eg Tom Anderson) mailboxes, but his OWN mailboxes, e.g. in different domains: [email protected], [email protected], [email protected], etc. Of course it is preferable for the user to work with all of his mailboxes from a single window. Such mailboxes can be added as multiple Exchange accounts in Outlook, that works almost fine. But in OWA, there are problems: 1) In the left pane - as I've learned - we can open only Inbox folders from other mailboxes. No way to view all folders like in Outlook? 2) With Send-As permissions set, when trying to send a message from another address, that message is saved in the Sent Items folder of the mailbox that is opened in OWA, and not in the mailbox the message is sent from. The same thing with the trash can. Is there a way to fix that? Also, this problem exists in desktop Outlook when mailboxes are added automatically via the Auto Mapping feature, so that we need to turn it off and add the accounts manually. Is there a simpler workaround? 3) Okay, suppose we only open Inbox folders in the left pane. The problem is that the mailbox names shown there are formed from Display Name attributes. But those names are all identical! All the mailboxes are owned by John Smith, so they should be all named John Smith - so that letter recepient sees "John Smith" in the "from" field, no matter what mailbox it is sent from. Also, the user knows what's his name - no need to tell him. He wants to know what mailbox he works with. So we need a way to either: a) customize OWA to show mailbox email address instead of user Display Name, or b) make Exchange use another attribute to put in the "from" field when sending letters 4) Okay, we can switch between mailboxes using "Open Other Mailbox" in the upper-right corner menu. But: a) To select a mailbox we need to enter its name (or first letters). It there a way to show a list of links to mailboxes the user has full access to? Eg in the page header... b) If we start entering the first letters, we see a popup list with possible mailboxes to be opened. But there are all mailboxes (apparently from GAL), not only mailboxes the user has permission to open! How to filter that popup list? c) The same problem as in (3) with mailbox naming. We can see the opened mailbox email address ONLY in the page URL, which is insufficient for many users. In the left pane we see "John Smith" which is useless. 5) Each mailbox is tied with a separate user in AD. If one has several mailboxes, we need to have additional dummy AD accounts, create additional OUs to store them, etc. That's not very nice, is there any standartized, optimal way to build such a structure? We would really appreciate any answers or additional info for any of these questions. Thank you in advance.

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  • What Apache/PHP configurations do you know and how good are they?

    - by FractalizeR
    Hello. I wanted to ask you about PHP/Apache configuration methods you know, their pros and cons. I will start myself: ---------------- PHP as Apache module---------------- Pros: good speed since you don't need to start exe every time especially in mpm-worker mode. You can also use various PHP accelerators in this mode like APC or eAccelerator. Cons: if you are running apache in mpm-worker mode, you may face stability issues because every glitch in any php script will lead to unstability to the whole thread pool of that apache process. Also in this mode all scripts are executed on behalf of apache user. This is bad for security. mpm-worker configuration requires PHP compiled in thread-safe mode. At least CentOS and RedHat default repositories doesn't have thread-safe PHP version so on these OSes you need to compile at least PHP yourself (there is a way to activate worker mpm on Apache). The use of thread-safe PHP binaries is considered experimental and unstable. Plus, many PHP extensions does not support thread-safe mode or were not well-tested in thread-safe mode. ---------------- PHP as CGI ---------------- This seems to be the slowest default configuration which seems to be a "con" itself ;) ---------------- PHP as CGI via mod_suphp ---------------- Pros: suphp allows you to execute php scipts on behalf of the script file owner. This way you can securely separate different sites on the same machine. Also, suphp allows to use different php.ini files per virtual host. Cons: PHP in CGI mode means less performance. In this mode you can't use php accelerators like APC because each time new process is spawned to handle script rendering the cache of previous process useless. BTW, do you know the way to apply some accelerator in this config? I heard something about using shm for php bytecode cache. Also, you cannot configure PHP via .htaccess files in this mode. You will need to install PECL htscanner for this if you need to set various per-script options via .htaccess (php_value / php_flag directives) ---------------- PHP as CGI via suexec ---------------- This configuration looks the same as with suphp, but I heard, that it's slower and less safe. Almost same pros and cons apply. ---------------- PHP as FastCGI ---------------- Pros: FastCGI standard allows single php process to handle several scripts before php process is killed. This way you gain performance since no need to spin up new php process for each script. You can also use PHP accelerators in this configuration (see cons section for comment). Also, FCGI almost like suphp also allows php processes to be executed on behalf of some user. mod_fcgid seems to have the most complete fcgi support and flexibility for apache. Cons: The use of php accelerator in fastcgi mode will lead to high memory consumption because each PHP process will have his own bytecode cache (unless there is some accelerator that can use shared memory for bytecode cache. Is there such?). FastCGI is also a little bit complex to configure. You need to create various configuration files and make some configuration modifications. It seems, that fastcgi is the most stable, secure, fast and flexible PHP configuration, however, a bit difficult to be configured. But, may be, I missed something? Comments are welcome!

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  • simple and reliable centralized logging inside Amazon VPC

    - by Nakedible
    I need to set up centralized logging for a set of servers (10-20) in an Amazon VPC. The logging should be as to not lose any log messages in case any single server goes offline - or in the case that an entire availability zone goes offline. It should also tolerate packet loss and other normal network conditions without losing or duplicating messages. It should store the messages durably, at the minimum on two different EBS volumes in two availability zones, but S3 is a good place as well. It should also be realtime so that the messages arrive within seconds of their generation to two different availability zones. I also need to sync logfiles not generated via syslog, so a syslog-only centralized logging solution would not fulfill all the needs, although I guess that limitation could be worked around. I have already reviewed a few solutions, and I will list them here: Flume to Flume to S3: I could set up two logservers as Flume hosts which would store log messages either locally or in S3, and configure all the servers with Flume to send all messages to both servers, using the end-to-end reliability options. That way the loss of a single server shouldn't cause lost messages and all messages would arrive in two availability zones in realtime. However, there would need to be some way to join the logs of the two servers, deduplicating all the messages delivered to both. This could be done by adding a unique id on the sending side to each message and then write some manual deduplication runs on the logfiles. I haven't found an easy solution to the duplication problem. Logstash to Logstash to ElasticSearch: I could install Logstash on the servers and have them deliver to a central server via AMQP, with the durability options turned on. However, for this to work I would need to use some of the clustering capable AMQP implementations, or fan out the deliver just as in the Flume case. AMQP seems to be a yet another moving part with several implementations and no real guidance on what works best this sort of setup. And I'm not entirely convinced that I could get actual end-to-end durability from logstash to elasticsearch, assuming crashing servers in between. The fan-out solutions run in to the deduplication problem again. The best solution that would seem to handle all the cases, would be Beetle, which seems to provide high availability and deduplication via a redis store. However, I haven't seen any guidance on how to set this up with Logstash and Redis is one more moving part again for something that shouldn't be terribly difficult. Logstash to ElasticSearch: I could run Logstash on all the servers, have all the filtering and processing rules in the servers themselves and just have them log directly to a removet ElasticSearch server. I think this should bring me reliable logging and I can use the ElasticSearch clustering features to share the database transparently. However, I am not sure if the setup actually survives Logstash restarts and intermittent network problems without duplicating messages in a failover case or similar. But this approach sounds pretty promising. rsync: I could just rsync all the relevant log files to two different servers. The reliability aspect should be perfect here, as the files should be identical to the source files after a sync is done. However, doing an rsync several times per second doesn't sound fun. Also, I need the logs to be untamperable after they have been sent, so the rsyncs would need to be in append-only mode. And log rotations mess things up unless I'm careful. rsyslog with RELP: I could set up rsyslog to send messages to two remote hosts via RELP and have a local queue to store the messages. There is the deduplication problem again, and RELP itself might also duplicate some messages. However, this would only handle the things that log via syslog. None of these solutions seem terribly good, and they have many unknowns still, so I am asking for more information here from people who have set up centralized reliable logging as to what are the best tools to achieve that goal.

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  • How to troubleshoot problem with OpenVpn Appliance Server not able to connect

    - by Peter
    1) I have a Windows Server 2008 Standard SP2 2) I am running Hyper-V and have the OpenSvn Appliance Server virtual running 3) I have configured it as it said, only issue was that the legacy network adapter does not have a setting the instructions mention "Enable spoofing of MAC Addresses". My understand is that before R2, this was on by default. 4) Server is running, web interfaces look good 5) I am trying to connect from a Vista 64 box and cannot 5a) If I set to UPD I am stuck at Authorizing and client log looks like: 10/11/09 15:00:42: INFO: OvpnConfig: connect... 10/11/09 15:00:42: INFO: Gui listen socket at 34567 10/11/09 15:00:42: INFO: sending start command to instantiator... 10/11/09 15:00:42: INFO: start 34567 ?C:\Users\Peter\AppData\Roaming\OpenVPNTech\config?02369512D0C82A04B88093022DA0226202218022A902264022AE022B? 10/11/09 15:00:42: INFO: Got line from MI->>INFO:OpenVPN Management Interface Version 1 -- type 'help' for more info 10/11/09 15:00:42: INFO: Got line from MI->>HOLD:Waiting for hold release 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: real-time state notification set to ON 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: bytecount interval changed 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: hold flag set to OFF 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: hold release succeeded 10/11/09 15:00:43: INFO: Got line from MI->>PASSWORD:Need 'Auth' username/password 10/11/09 15:00:43: INFO: Processing PASSWORD. 10/11/09 15:00:43: INFO: OvpnClient: setting need auth to true. 10/11/09 15:00:43: INFO: OvpnConfig: Setting need auth to true. 10/11/09 15:00:43: INFO: Got auth request from active_config from 0 10/11/09 15:00:47: INFO: Sending Credentials.... 10/11/09 15:00:47: INFO: Sending 25 bytes for username. 10/11/09 15:00:47: INFO: Sent 25 bytes for username. 10/11/09 15:00:47: INFO: Sending 30 bytes for password. 10/11/09 15:00:47: INFO: Sent 30 bytes for password. 10/11/09 15:00:48: INFO: Got line from MI->SUCCESS: 'Auth' username entered, but not yet verified 10/11/09 15:00:48: INFO: Got line from MI->SUCCESS: 'Auth' password entered, but not yet verified 10/11/09 15:00:48: INFO: Got line from MI->>STATE:1255287647,WAIT,,, 10/11/09 15:00:48: INFO: Got line from MI->>BYTECOUNT:0,42 10/11/09 15:00:48: INFO: Got line from MI->>BYTECOUNT:54,42 10/11/09 15:00:48: INFO: Got line from MI->>STATE:1255287648,AUTH,,, 10/11/09 15:00:50: INFO: Got line from MI->>BYTECOUNT:2560,2868 10/11/09 15:00:52: INFO: Got line from MI->>BYTECOUNT:2560,3378 5b) I setup server for tcp and try to connect, I get a loop of authorizing and reconnecting. Log looks like: 10/11/09 15:00:42: INFO: Got line from MI->>HOLD:Waiting for hold release 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: real-time state notification set to ON 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: bytecount interval changed 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: hold flag set to OFF 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: hold release succeeded 10/11/09 15:00:43: INFO: Got line from MI->>PASSWORD:Need 'Auth' username/password 10/11/09 15:00:43: INFO: Processing PASSWORD. 10/11/09 15:00:43: INFO: OvpnClient: setting need auth to true. 10/11/09 15:00:43: INFO: OvpnConfig: Setting need auth to true. 10/11/09 15:00:43: INFO: Got auth request from active_config from 0 10/11/09 15:00:47: INFO: Sending Credentials.... 10/11/09 15:00:47: INFO: Sending 25 bytes for username. 10/11/09 15:00:47: INFO: Sent 25 bytes for username. 10/11/09 15:00:47: INFO: Sending 30 bytes for password. 10/11/09 15:00:47: INFO: Sent 30 bytes for password. 10/11/09 15:00:48: INFO: Got line from MI->SUCCESS: 'Auth' username entered, but not yet verified 10/11/09 15:00:48: INFO: Got line from MI->SUCCESS: 'Auth' password entered, but not yet verified 10/11/09 15:00:48: INFO: Got line from MI->>STATE:1255287647,WAIT,,, 10/11/09 15:00:48: INFO: Got line from MI->>BYTECOUNT:0,42 10/11/09 15:00:48: INFO: Got line from MI->>BYTECOUNT:54,42 10/11/09 15:00:48: INFO: Got line from MI->>STATE:1255287648,AUTH,,, 10/11/09 15:00:50: INFO: Got line from MI->>BYTECOUNT:2560,2868 10/11/09 15:00:52: INFO: Got line from MI->>BYTECOUNT:2560,3378 10/11/09 15:00:54: INFO: Got line from MI->>BYTECOUNT:2560,3888 ... Is there anyway to turn on robust logging on the server to understand what is happening? Any ideas on how to hunt this down?

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  • How to troubleshoot problem with OpenVPN Appliance Server not able to connect

    - by Peter
    1) I have a Windows Server 2008 Standard SP2 2) I am running Hyper-V and have the OpenVPN Appliance Server virtual running 3) I have configured it as it said, only issue was that the legacy network adapter does not have a setting the instructions mention "Enable spoofing of MAC Addresses". My understand is that before R2, this was on by default. 4) Server is running, web interfaces look good 5) I am trying to connect from a Vista 64 box and cannot 5a) If I set to UPD I am stuck at Authorizing and client log looks like: 10/11/09 15:00:42: INFO: OvpnConfig: connect... 10/11/09 15:00:42: INFO: Gui listen socket at 34567 10/11/09 15:00:42: INFO: sending start command to instantiator... 10/11/09 15:00:42: INFO: start 34567 ?C:\Users\Peter\AppData\Roaming\OpenVPNTech\config?02369512D0C82A04B88093022DA0226202218022A902264022AE022B? 10/11/09 15:00:42: INFO: Got line from MI->>INFO:OpenVPN Management Interface Version 1 -- type 'help' for more info 10/11/09 15:00:42: INFO: Got line from MI->>HOLD:Waiting for hold release 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: real-time state notification set to ON 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: bytecount interval changed 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: hold flag set to OFF 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: hold release succeeded 10/11/09 15:00:43: INFO: Got line from MI->>PASSWORD:Need 'Auth' username/password 10/11/09 15:00:43: INFO: Processing PASSWORD. 10/11/09 15:00:43: INFO: OvpnClient: setting need auth to true. 10/11/09 15:00:43: INFO: OvpnConfig: Setting need auth to true. 10/11/09 15:00:43: INFO: Got auth request from active_config from 0 10/11/09 15:00:47: INFO: Sending Credentials.... 10/11/09 15:00:47: INFO: Sending 25 bytes for username. 10/11/09 15:00:47: INFO: Sent 25 bytes for username. 10/11/09 15:00:47: INFO: Sending 30 bytes for password. 10/11/09 15:00:47: INFO: Sent 30 bytes for password. 10/11/09 15:00:48: INFO: Got line from MI->SUCCESS: 'Auth' username entered, but not yet verified 10/11/09 15:00:48: INFO: Got line from MI->SUCCESS: 'Auth' password entered, but not yet verified 10/11/09 15:00:48: INFO: Got line from MI->>STATE:1255287647,WAIT,,, 10/11/09 15:00:48: INFO: Got line from MI->>BYTECOUNT:0,42 10/11/09 15:00:48: INFO: Got line from MI->>BYTECOUNT:54,42 10/11/09 15:00:48: INFO: Got line from MI->>STATE:1255287648,AUTH,,, 10/11/09 15:00:50: INFO: Got line from MI->>BYTECOUNT:2560,2868 10/11/09 15:00:52: INFO: Got line from MI->>BYTECOUNT:2560,3378 5b) I setup server for tcp and try to connect, I get a loop of authorizing and reconnecting. Log looks like: 10/11/09 15:00:42: INFO: Got line from MI->>HOLD:Waiting for hold release 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: real-time state notification set to ON 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: bytecount interval changed 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: hold flag set to OFF 10/11/09 15:00:43: INFO: Got line from MI->SUCCESS: hold release succeeded 10/11/09 15:00:43: INFO: Got line from MI->>PASSWORD:Need 'Auth' username/password 10/11/09 15:00:43: INFO: Processing PASSWORD. 10/11/09 15:00:43: INFO: OvpnClient: setting need auth to true. 10/11/09 15:00:43: INFO: OvpnConfig: Setting need auth to true. 10/11/09 15:00:43: INFO: Got auth request from active_config from 0 10/11/09 15:00:47: INFO: Sending Credentials.... 10/11/09 15:00:47: INFO: Sending 25 bytes for username. 10/11/09 15:00:47: INFO: Sent 25 bytes for username. 10/11/09 15:00:47: INFO: Sending 30 bytes for password. 10/11/09 15:00:47: INFO: Sent 30 bytes for password. 10/11/09 15:00:48: INFO: Got line from MI->SUCCESS: 'Auth' username entered, but not yet verified 10/11/09 15:00:48: INFO: Got line from MI->SUCCESS: 'Auth' password entered, but not yet verified 10/11/09 15:00:48: INFO: Got line from MI->>STATE:1255287647,WAIT,,, 10/11/09 15:00:48: INFO: Got line from MI->>BYTECOUNT:0,42 10/11/09 15:00:48: INFO: Got line from MI->>BYTECOUNT:54,42 10/11/09 15:00:48: INFO: Got line from MI->>STATE:1255287648,AUTH,,, 10/11/09 15:00:50: INFO: Got line from MI->>BYTECOUNT:2560,2868 10/11/09 15:00:52: INFO: Got line from MI->>BYTECOUNT:2560,3378 10/11/09 15:00:54: INFO: Got line from MI->>BYTECOUNT:2560,3888 ... Is there anyway to turn on robust logging on the server to understand what is happening? Any ideas on how to hunt this down?

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  • Integrating JavaScript Unit Tests with Visual Studio

    - by Stephen Walther
    Modern ASP.NET web applications take full advantage of client-side JavaScript to provide better interactivity and responsiveness. If you are building an ASP.NET application in the right way, you quickly end up with lots and lots of JavaScript code. When writing server code, you should be writing unit tests. One big advantage of unit tests is that they provide you with a safety net that enable you to safely modify your existing code – for example, fix bugs, add new features, and make performance enhancements -- without breaking your existing code. Every time you modify your code, you can execute your unit tests to verify that you have not broken anything. For the same reason that you should write unit tests for your server code, you should write unit tests for your client code. JavaScript is just as susceptible to bugs as C#. There is no shortage of unit testing frameworks for JavaScript. Each of the major JavaScript libraries has its own unit testing framework. For example, jQuery has QUnit, Prototype has UnitTestJS, YUI has YUI Test, and Dojo has Dojo Objective Harness (DOH). The challenge is integrating a JavaScript unit testing framework with Visual Studio. Visual Studio and Visual Studio ALM provide fantastic support for server-side unit tests. You can easily view the results of running your unit tests in the Visual Studio Test Results window. You can set up a check-in policy which requires that all unit tests pass before your source code can be committed to the source code repository. In addition, you can set up Team Build to execute your unit tests automatically. Unfortunately, Visual Studio does not provide “out-of-the-box” support for JavaScript unit tests. MS Test, the unit testing framework included in Visual Studio, does not support JavaScript unit tests. As soon as you leave the server world, you are left on your own. The goal of this blog entry is to describe one approach to integrating JavaScript unit tests with MS Test so that you can execute your JavaScript unit tests side-by-side with your C# unit tests. The goal is to enable you to execute JavaScript unit tests in exactly the same way as server-side unit tests. You can download the source code described by this project by scrolling to the end of this blog entry. Rejected Approach: Browser Launchers One popular approach to executing JavaScript unit tests is to use a browser as a test-driver. When you use a browser as a test-driver, you open up a browser window to execute and view the results of executing your JavaScript unit tests. For example, QUnit – the unit testing framework for jQuery – takes this approach. The following HTML page illustrates how you can use QUnit to create a unit test for a function named addNumbers(). <!DOCTYPE HTML PUBLIC "-//W3C//DTD HTML 4.01 Transitional//EN" "http://www.w3.org/TR/html4/loose.dtd"> <html> <head> <title>Using QUnit</title> <link rel="stylesheet" href="http://github.com/jquery/qunit/raw/master/qunit/qunit.css" type="text/css" /> </head> <body> <h1 id="qunit-header">QUnit example</h1> <h2 id="qunit-banner"></h2> <div id="qunit-testrunner-toolbar"></div> <h2 id="qunit-userAgent"></h2> <ol id="qunit-tests"></ol> <div id="qunit-fixture">test markup, will be hidden</div> <script type="text/javascript" src="http://code.jquery.com/jquery-latest.js"></script> <script type="text/javascript" src="http://github.com/jquery/qunit/raw/master/qunit/qunit.js"></script> <script type="text/javascript"> // The function to test function addNumbers(a, b) { return a+b; } // The unit test test("Test of addNumbers", function () { equals(4, addNumbers(1,3), "1+3 should be 4"); }); </script> </body> </html> This test verifies that calling addNumbers(1,3) returns the expected value 4. When you open this page in a browser, you can see that this test does, in fact, pass. The idea is that you can quickly refresh this QUnit HTML JavaScript test driver page in your browser whenever you modify your JavaScript code. In other words, you can keep a browser window open and keep refreshing it over and over while you are developing your application. That way, you can know very quickly whenever you have broken your JavaScript code. While easy to setup, there are several big disadvantages to this approach to executing JavaScript unit tests: You must view your JavaScript unit test results in a different location than your server unit test results. The JavaScript unit test results appear in the browser and the server unit test results appear in the Visual Studio Test Results window. Because all of your unit test results don’t appear in a single location, you are more likely to introduce bugs into your code without noticing it. Because your unit tests are not integrated with Visual Studio – in particular, MS Test -- you cannot easily include your JavaScript unit tests when setting up check-in policies or when performing automated builds with Team Build. A more sophisticated approach to using a browser as a test-driver is to automate the web browser. Instead of launching the browser and loading the test code yourself, you use a framework to automate this process. There are several different testing frameworks that support this approach: · Selenium – Selenium is a very powerful framework for automating browser tests. You can create your tests by recording a Firefox session or by writing the test driver code in server code such as C#. You can learn more about Selenium at http://seleniumhq.org/. LTAF – The ASP.NET team uses the Lightweight Test Automation Framework to test JavaScript code in the ASP.NET framework. You can learn more about LTAF by visiting the project home at CodePlex: http://aspnet.codeplex.com/releases/view/35501 jsTestDriver – This framework uses Java to automate the browser. jsTestDriver creates a server which can be used to automate multiple browsers simultaneously. This project is located at http://code.google.com/p/js-test-driver/ TestSwam – This framework, created by John Resig, uses PHP to automate the browser. Like jsTestDriver, the framework creates a test server. You can open multiple browsers that are automated by the test server. Learn more about TestSwarm by visiting the following address: https://github.com/jeresig/testswarm/wiki Yeti – This is the framework introduced by Yahoo for automating browser tests. Yeti uses server-side JavaScript and depends on Node.js. Learn more about Yeti at http://www.yuiblog.com/blog/2010/08/25/introducing-yeti-the-yui-easy-testing-interface/ All of these frameworks are great for integration tests – however, they are not the best frameworks to use for unit tests. In one way or another, all of these frameworks depend on executing tests within the context of a “living and breathing” browser. If you create an ASP.NET Unit Test then Visual Studio will launch a web server before executing the unit test. Why is launching a web server so bad? It is not the worst thing in the world. However, it does introduce dependencies that prevent your code from being tested in isolation. One of the defining features of a unit test -- versus an integration test – is that a unit test tests code in isolation. Another problem with launching a web server when performing unit tests is that launching a web server can be slow. If you cannot execute your unit tests quickly, you are less likely to execute your unit tests each and every time you make a code change. You are much more likely to fall into the pit of failure. Launching a browser when performing a JavaScript unit test has all of the same disadvantages as launching a web server when performing an ASP.NET unit test. Instead of testing a unit of JavaScript code in isolation, you are testing JavaScript code within the context of a particular browser. Using the frameworks listed above for integration tests makes perfect sense. However, I want to consider a different approach for creating unit tests for JavaScript code. Using Server-Side JavaScript for JavaScript Unit Tests A completely different approach to executing JavaScript unit tests is to perform the tests outside of any browser. If you really want to test JavaScript then you should test JavaScript and leave the browser out of the testing process. There are several ways that you can execute JavaScript on the server outside the context of any browser: Rhino – Rhino is an implementation of JavaScript written in Java. The Rhino project is maintained by the Mozilla project. Learn more about Rhino at http://www.mozilla.org/rhino/ V8 – V8 is the open-source Google JavaScript engine written in C++. This is the JavaScript engine used by the Chrome web browser. You can download V8 and embed it in your project by visiting http://code.google.com/p/v8/ JScript – JScript is the JavaScript Script Engine used by Internet Explorer (up to but not including Internet Explorer 9), Windows Script Host, and Active Server Pages. Internet Explorer is still the most popular web browser. Therefore, I decided to focus on using the JScript Script Engine to execute JavaScript unit tests. Using the Microsoft Script Control There are two basic ways that you can pass JavaScript to the JScript Script Engine and execute the code: use the Microsoft Windows Script Interfaces or use the Microsoft Script Control. The difficult and proper way to execute JavaScript using the JScript Script Engine is to use the Microsoft Windows Script Interfaces. You can learn more about the Script Interfaces by visiting http://msdn.microsoft.com/en-us/library/t9d4xf28(VS.85).aspx The main disadvantage of using the Script Interfaces is that they are difficult to use from .NET. There is a great series of articles on using the Script Interfaces from C# located at http://www.drdobbs.com/184406028. I picked the easier alternative and used the Microsoft Script Control. The Microsoft Script Control is an ActiveX control that provides a higher level abstraction over the Window Script Interfaces. You can download the Microsoft Script Control from here: http://www.microsoft.com/downloads/en/details.aspx?FamilyID=d7e31492-2595-49e6-8c02-1426fec693ac After you download the Microsoft Script Control, you need to add a reference to it to your project. Select the Visual Studio menu option Project, Add Reference to open the Add Reference dialog. Select the COM tab and add the Microsoft Script Control 1.0. Using the Script Control is easy. You call the Script Control AddCode() method to add JavaScript code to the Script Engine. Next, you call the Script Control Run() method to run a particular JavaScript function. The reference documentation for the Microsoft Script Control is located at the MSDN website: http://msdn.microsoft.com/en-us/library/aa227633%28v=vs.60%29.aspx Creating the JavaScript Code to Test To keep things simple, let’s imagine that you want to test the following JavaScript function named addNumbers() which simply adds two numbers together: MvcApplication1\Scripts\Math.js function addNumbers(a, b) { return 5; } Notice that the addNumbers() method always returns the value 5. Right-now, it will not pass a good unit test. Create this file and save it in your project with the name Math.js in your MVC project’s Scripts folder (Save the file in your actual MVC application and not your MVC test application). Creating the JavaScript Test Helper Class To make it easier to use the Microsoft Script Control in unit tests, we can create a helper class. This class contains two methods: LoadFile() – Loads a JavaScript file. Use this method to load the JavaScript file being tested or the JavaScript file containing the unit tests. ExecuteTest() – Executes the JavaScript code. Use this method to execute a JavaScript unit test. Here’s the code for the JavaScriptTestHelper class: JavaScriptTestHelper.cs   using System; using System.IO; using Microsoft.VisualStudio.TestTools.UnitTesting; using MSScriptControl; namespace MvcApplication1.Tests { public class JavaScriptTestHelper : IDisposable { private ScriptControl _sc; private TestContext _context; /// <summary> /// You need to use this helper with Unit Tests and not /// Basic Unit Tests because you need a Test Context /// </summary> /// <param name="testContext">Unit Test Test Context</param> public JavaScriptTestHelper(TestContext testContext) { if (testContext == null) { throw new ArgumentNullException("TestContext"); } _context = testContext; _sc = new ScriptControl(); _sc.Language = "JScript"; _sc.AllowUI = false; } /// <summary> /// Load the contents of a JavaScript file into the /// Script Engine. /// </summary> /// <param name="path">Path to JavaScript file</param> public void LoadFile(string path) { var fileContents = File.ReadAllText(path); _sc.AddCode(fileContents); } /// <summary> /// Pass the path of the test that you want to execute. /// </summary> /// <param name="testMethodName">JavaScript function name</param> public void ExecuteTest(string testMethodName) { dynamic result = null; try { result = _sc.Run(testMethodName, new object[] { }); } catch { var error = ((IScriptControl)_sc).Error; if (error != null) { var description = error.Description; var line = error.Line; var column = error.Column; var text = error.Text; var source = error.Source; if (_context != null) { var details = String.Format("{0} \r\nLine: {1} Column: {2}", source, line, column); _context.WriteLine(details); } } throw new AssertFailedException(error.Description); } } public void Dispose() { _sc = null; } } }     Notice that the JavaScriptTestHelper class requires a Test Context to be instantiated. For this reason, you can use the JavaScriptTestHelper only with a Visual Studio Unit Test and not a Basic Unit Test (These are two different types of Visual Studio project items). Add the JavaScriptTestHelper file to your MVC test application (for example, MvcApplication1.Tests). Creating the JavaScript Unit Test Next, we need to create the JavaScript unit test function that we will use to test the addNumbers() function. Create a folder in your MVC test project named JavaScriptTests and add the following JavaScript file to this folder: MvcApplication1.Tests\JavaScriptTests\MathTest.js /// <reference path="JavaScriptUnitTestFramework.js"/> function testAddNumbers() { // Act var result = addNumbers(1, 3); // Assert assert.areEqual(4, result, "addNumbers did not return right value!"); }   The testAddNumbers() function takes advantage of another JavaScript library named JavaScriptUnitTestFramework.js. This library contains all of the code necessary to make assertions. Add the following JavaScriptnitTestFramework.js to the same folder as the MathTest.js file: MvcApplication1.Tests\JavaScriptTests\JavaScriptUnitTestFramework.js var assert = { areEqual: function (expected, actual, message) { if (expected !== actual) { throw new Error("Expected value " + expected + " is not equal to " + actual + ". " + message); } } }; There is only one type of assertion supported by this file: the areEqual() assertion. Most likely, you would want to add additional types of assertions to this file to make it easier to write your JavaScript unit tests. Deploying the JavaScript Test Files This step is non-intuitive. When you use Visual Studio to run unit tests, Visual Studio creates a new folder and executes a copy of the files in your project. After you run your unit tests, your Visual Studio Solution will contain a new folder named TestResults that includes a subfolder for each test run. You need to configure Visual Studio to deploy your JavaScript files to the test run folder or Visual Studio won’t be able to find your JavaScript files when you execute your unit tests. You will get an error that looks something like this when you attempt to execute your unit tests: You can configure Visual Studio to deploy your JavaScript files by adding a Test Settings file to your Visual Studio Solution. It is important to understand that you need to add this file to your Visual Studio Solution and not a particular Visual Studio project. Right-click your Solution in the Solution Explorer window and select the menu option Add, New Item. Select the Test Settings item and click the Add button. After you create a Test Settings file for your solution, you can indicate that you want a particular folder to be deployed whenever you perform a test run. Select the menu option Test, Edit Test Settings to edit your test configuration file. Select the Deployment tab and select your MVC test project’s JavaScriptTest folder to deploy. Click the Apply button and the Close button to save the changes and close the dialog. Creating the Visual Studio Unit Test The very last step is to create the Visual Studio unit test (the MS Test unit test). Add a new unit test to your MVC test project by selecting the menu option Add New Item and selecting the Unit Test project item (Do not select the Basic Unit Test project item): The difference between a Basic Unit Test and a Unit Test is that a Unit Test includes a Test Context. We need this Test Context to use the JavaScriptTestHelper class that we created earlier. Enter the following test method for the new unit test: [TestMethod] public void TestAddNumbers() { var jsHelper = new JavaScriptTestHelper(this.TestContext); // Load JavaScript files jsHelper.LoadFile("JavaScriptUnitTestFramework.js"); jsHelper.LoadFile(@"..\..\..\MvcApplication1\Scripts\Math.js"); jsHelper.LoadFile("MathTest.js"); // Execute JavaScript Test jsHelper.ExecuteTest("testAddNumbers"); } This code uses the JavaScriptTestHelper to load three files: JavaScripUnitTestFramework.js – Contains the assert functions. Math.js – Contains the addNumbers() function from your MVC application which is being tested. MathTest.js – Contains the JavaScript unit test function. Next, the test method calls the JavaScriptTestHelper ExecuteTest() method to execute the testAddNumbers() JavaScript function. Running the Visual Studio JavaScript Unit Test After you complete all of the steps described above, you can execute the JavaScript unit test just like any other unit test. You can use the keyboard combination CTRL-R, CTRL-A to run all of the tests in the current Visual Studio Solution. Alternatively, you can use the buttons in the Visual Studio toolbar to run the tests: (Unfortunately, the Run All Impacted Tests button won’t work correctly because Visual Studio won’t detect that your JavaScript code has changed. Therefore, you should use either the Run Tests in Current Context or Run All Tests in Solution options instead.) The results of running the JavaScript tests appear side-by-side with the results of running the server tests in the Test Results window. For example, if you Run All Tests in Solution then you will get the following results: Notice that the TestAddNumbers() JavaScript test has failed. That is good because our addNumbers() function is hard-coded to always return the value 5. If you double-click the failing JavaScript test, you can view additional details such as the JavaScript error message and the line number of the JavaScript code that failed: Summary The goal of this blog entry was to explain an approach to creating JavaScript unit tests that can be easily integrated with Visual Studio and Visual Studio ALM. I described how you can use the Microsoft Script Control to execute JavaScript on the server. By taking advantage of the Microsoft Script Control, we were able to execute our JavaScript unit tests side-by-side with all of our other unit tests and view the results in the standard Visual Studio Test Results window. You can download the code discussed in this blog entry from here: http://StephenWalther.com/downloads/Blog/JavaScriptUnitTesting/JavaScriptUnitTests.zip Before running this code, you need to first install the Microsoft Script Control which you can download from here: http://www.microsoft.com/downloads/en/details.aspx?FamilyID=d7e31492-2595-49e6-8c02-1426fec693ac

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  • How To Activate Your Free Office 2007 to 2010 Tech Guarantee Upgrade

    - by Matthew Guay
    Have you purchased Office 2007 since March 5th, 2010?  If so, here’s how you can activate and download your free upgrade to Office 2010! Microsoft Office 2010 has just been released, and today you can purchase upgrades from most retail stores or directly from Microsoft via download.  But if you’ve purchased a new copy of Office 2007 or a new computer that came with Office 2007 since March 5th, 2010, then you’re entitled to an absolutely free upgrade to Office 2010.  You’ll need enter information about your Office 2007 and then download the upgrade, so we’ll step you through the process. Getting Started First, if you’ve recently purchased Office 2007 but haven’t installed it, you’ll need to go ahead and install it before you can get your free Office 2010 upgrade.  Install it as normal.   Once Office 2007 is installed, run any of the Office programs.  You’ll be prompted to activate Office.  Make sure you’re connected to the internet, and then click Next to activate. Get your Free Upgrade to Office 2010 Now you’re ready to download your upgrade to Office 2010.  Head to the Office Tech Guarantee site (link below), and click Upgrade now. You’ll need to enter some information about your Office 2007.  Check that you purchased your copy of Office 2007 after March 5th, select your computer manufacturer, and check that you agree to the terms. Now you’re going to need the Product ID number from Office 2007.  To find this, open Word or any other Office 2007 application.  Click the Office Orb, and select Options on the bottom. Select the Resources button on the left, and then click About. Near the bottom of this dialog, you’ll see your Product ID.  This should be a number like: 12345-123-1234567-12345   Go back to the Office Tech Guarantee signup page in your browser, and enter this Product ID.  Select the language of your edition of Office 2007, enter the verification code, and then click Submit. It may take a few moments to validate your Product ID. When it is finished, you’ll be taken to an order page that shows the edition of Office 2010 you’re eligible to receive.  The upgrade download is free, but if you’d like to purchase a backup DVD of Office 2010, you can add it to your order for $13.99.  Otherwise, simply click Continue to accept. Do note that the edition of Office 2010 you receive may be different that the edition of Office 2007 you purchased, as the number of editions has been streamlined in the Office 2010 release.  Here’s a chart you can check to see what edition you’ll receive.  Note that you’ll still be allowed to install Office on the same number of computers; for example, Office 2007 Home and Student allows you to install it on up to 3 computers in the same house, and your Office 2010 upgrade will allow the same. Office 2007 Edition Office 2010 Upgrade You’ll Receive Office 2007 Home and Student Office Home and Student 2010 Office Basic 2007Office Standard 2007 Office Home and Business 2010 Office Small Business 2007Office Professional 2007Office Ultimate 2007 Office Professional 2010 Office Professional 2007 AcademicOffice Ultimate 2007 Academic Office Professional Academic 2010 Sign in with your Windows Live ID, or create a new one if you don’t already have one. Enter your name, select your country, and click Create My Account.  Note that Office will send Office 2010 tips to your email address; if you don’t wish to receive them, you can unsubscribe from the emails later.   Finally, you’re ready to download Office 2010!  Click the Download Now link to start downloading Office 2010.  Your Product Key will appear directly above the Download link, so you can copy it and then paste it in the installer when your download is finished.  You will additionally receive an email with the download links and product key, so if your download fails you can always restart it from that link. If your edition of Office 2007 included the Office Business Contact Manager, you will be able to download it from the second Download link.  And, of course, even if you didn’t order a backup DVD, you can always burn the installers to a DVD for a backup.   Install Office 2010 Once you’re finished downloading Office 2010, run the installer to get it installed on your computer.  Enter your Product Key from the Tech Guarantee website as above, and click Continue. Accept the license agreement, and then click Upgrade to upgrade to the latest version of Office.   The installer will remove all of your Office 2007 applications, and then install their 2010 counterparts.  If you wish to keep some of your Office 2007 applications instead, click Customize and then select to either keep all previous versions or simply keep specific applications. By default, Office 2010 will try to activate online automatically.  If it doesn’t activate during the install, you’ll need to activate it when you first run any of the Office 2010 apps.   Conclusion The Tech Guarantee makes it easy to get the latest version of Office if you recently purchased Office 2007.  The Tech Guarantee program is open through the end of September, so make sure to grab your upgrade during this time.  Actually, if you find a great deal on Office 2007 from a major retailer between now and then, you could also take advantage of this program to get Office 2010 cheaper. And if you need help getting started with Office 2010, check out our articles that can help you get situated in your new version of Office! Link Activate and Download Your free Office 2010 Tech Guarantee Upgrade Similar Articles Productive Geek Tips Remove Office 2010 Beta and Reinstall Office 2007Upgrade Office 2003 to 2010 on XP or Run them Side by SideCenter Pictures and Other Objects in Office 2007 & 2010Change the Default Color Scheme in Office 2010Show Two Time Zones in Your Outlook 2007 Calendar 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 Windows Media Player Plus! – Cool WMP Enhancer Get Your Team’s World Cup Schedule In Google Calendar Backup Drivers With Driver Magician TubeSort: YouTube Playlist Organizer XPS file format & XPS Viewer Explained Microsoft Office Web Apps Guide

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