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  • Why is my concurrency capacity so low for my web app on a LAMP EC2 instance?

    - by AMF
    I come from a web developer background and have been humming along building my PHP app, using the CakePHP framework. The problem arose when I began the ab (Apache Bench) testing on the Amazon EC2 instance in which the app resides. I'm getting pretty horrendous average page load times, even though I'm running a c1.medium instance (2 cores, 2GB RAM), and I think I'm doing everything right. I would run: ab -n 200 -c 20 http://localhost/heavy-but-view-cached-page.php Here are the results: Concurrency Level: 20 Time taken for tests: 48.197 seconds Complete requests: 200 Failed requests: 0 Write errors: 0 Total transferred: 392111200 bytes HTML transferred: 392047600 bytes Requests per second: 4.15 [#/sec] (mean) Time per request: 4819.723 [ms] (mean) Time per request: 240.986 [ms] (mean, across all concurrent requests) Transfer rate: 7944.88 [Kbytes/sec] received While the ab test is running, I run VMStat, which shows that Swap stays at 0, CPU is constantly at 80-100% (although I'm not sure I can trust this on a VM), RAM utilization ramps up to about 1.6G (leaving 400M free). Load goes up to about 8 and site slows to a crawl. Here's what I think I'm doing right on the code side: In Chrome browser uncached pages typically load in 800-1000ms, and cached pages load in 300-500ms. Not stunning, but not terrible either. Thanks to view caching, there might be at most one DB query per page-load to write session data. So we can rule out a DB bottleneck. I have APC on. I am using Memcached to serve the view cache and other site caches. xhprof code profiler shows that cached pages take up 10MB-40MB in memory and 100ms - 1000ms in wall time. Pages that would be the worst offenders would look something like this in xhprof: Total Incl. Wall Time (microsec): 330,143 microsecs Total Incl. CPU (microsecs): 320,019 microsecs Total Incl. MemUse (bytes): 36,786,192 bytes Total Incl. PeakMemUse (bytes): 46,667,008 bytes Number of Function Calls: 5,195 My Apache config: KeepAlive On MaxKeepAliveRequests 100 KeepAliveTimeout 3 <IfModule mpm_prefork_module> StartServers 5 MinSpareServers 5 MaxSpareServers 10 MaxClients 120 MaxRequestsPerChild 1000 </IfModule> Is there something wrong with the server? Some gotcha with the EC2? Or is it my code? Some obvious setting I should look into? Too many DNS lookups? What am I missing? I really want to get to 1,000 concurrency capacity, but at this rate, it ain't gonna happen.

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  • function to org-sort by three (3) criteria: due date / priority / title

    - by lawlist
    Is anyone aware of an org-sort function / modification that can refile / organize a group of TODO so that it sorts them by three (3) criteria: first sort by due date, second sort by priority, and third sort by by title of the task? EDIT: If anyone can please help me to modify this so that undated TODO are sorted last, that would be greatly appreciated -- at the present time, undated TODO are not being sorted: ;; multiple sort (defun org-sort-multi (&rest sort-types) "Multiple sorts on a certain level of an outline tree, or plain list items. SORT-TYPES is a list where each entry is either a character or a cons pair (BOOL . CHAR), where BOOL is whether or not to sort case-sensitively, and CHAR is one of the characters defined in `org-sort-entries-or-items'. Entries are applied in back to front order. Example: To sort first by TODO status, then by priority, then by date, then alphabetically (case-sensitive) use the following call: (org-sort-multi '(?d ?p ?t (t . ?a)))" (interactive) (dolist (x (nreverse sort-types)) (when (char-valid-p x) (setq x (cons nil x))) (condition-case nil (org-sort-entries (car x) (cdr x)) (error nil)))) ;; sort current level (defun lawlist-sort (&rest sort-types) "Sort the current org level. SORT-TYPES is a list where each entry is either a character or a cons pair (BOOL . CHAR), where BOOL is whether or not to sort case-sensitively, and CHAR is one of the characters defined in `org-sort-entries-or-items'. Entries are applied in back to front order. Defaults to \"?o ?p\" which is sorted by TODO status, then by priority" (interactive) (when (equal mode-name "Org") (let ((sort-types (or sort-types (if (or (org-entry-get nil "TODO") (org-entry-get nil "PRIORITY")) '(?d ?t ?p) ;; date, time, priority '((nil . ?a)))))) (save-excursion (outline-up-heading 1) (let ((start (point)) end) (while (and (not (bobp)) (not (eobp)) (<= (point) start)) (condition-case nil (outline-forward-same-level 1) (error (outline-up-heading 1)))) (unless (> (point) start) (goto-char (point-max))) (setq end (point)) (goto-char start) (apply 'org-sort-multi sort-types) (goto-char end) (when (eobp) (forward-line -1)) (when (looking-at "^\\s-*$") ;; (delete-line) ) (goto-char start) ;; (dotimes (x ) (org-cycle)) )))))

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  • mdadm raid1 fails to resync

    - by JuanD
    Hello, I'm trying to solve this problem I'm having with an mdadm raid1. I have an ubuntu 9.04 server running on a software 2-drive raid1 with mdadm. Yesterday, one of the drives failed, and so I replaced it with a brand new drive of the same size. I removed the faulty drive, copied the partition from the remaining good drive to the new drive and then added it to the raid. It re-synced and the system worked fine, until the drive that hadn't failed, was also labeled failed. Now I had the raid running solely on the new drive. So I purchased another drive and repeated the procedure above. So now I had 2 brand new drives and the raid was syncing. However, after a few minutes I checked /proc/mdstat and the raid was no longer syncing. mdadm --detail /dev/md1 shows: (sdb is the first new drive, and sdc is the second new drive) root@dola:/home/jjaramillo# mdadm --detail /dev/md1 /dev/md1: Version : 00.90 Creation Time : Sat Dec 20 00:42:05 2008 Raid Level : raid1 Array Size : 974711680 (929.56 GiB 998.10 GB) Used Dev Size : 974711680 (929.56 GiB 998.10 GB) Raid Devices : 2 Total Devices : 2 Preferred Minor : 1 Persistence : Superblock is persistent Update Time : Wed Jun 2 10:09:35 2010 State : clean, degraded Active Devices : 1 Working Devices : 2 Failed Devices : 0 Spare Devices : 1 UUID : bba497c6:5029ba0b:bfa4f887:c0dc8f3d Events : 0.5395594 Number Major Minor RaidDevice State 2 8 35 0 spare rebuilding /dev/sdc3 1 8 19 1 active sync /dev/sdb3 I've tried removing and re-adding the drive a few times, but the same thing happens. The raid fails to resync. I've looked at /var/log/messages, and found the following: Jun 2 07:57:36 dola kernel: [35708.917337] sd 5:0:0:0: [sdb] Unhandled sense code Jun 2 07:57:36 dola kernel: [35708.917339] sd 5:0:0:0: [sdb] Result: hostbyte=DID_OK driverbyte=DRIVER_SENSE Jun 2 07:57:36 dola kernel: [35708.917342] sd 5:0:0:0: [sdb] Sense Key : Medium Error [current] [descriptor] Jun 2 07:57:36 dola kernel: [35708.917346] Descriptor sense data with sense descriptors (in hex): Jun 2 07:57:36 dola kernel: [35708.917348] 72 03 11 04 00 00 00 0c 00 0a 80 00 00 00 00 00 Jun 2 07:57:36 dola kernel: [35708.917357] 00 43 9e 47 Jun 2 07:57:36 dola kernel: [35708.917360] sd 5:0:0:0: [sdb] Add. Sense: Unrecovered read error - auto reallocate failed So it looks like there's some kind of error on sdb (the first new drive). My question is, what would be the best approach to get the raid up and running again? I've thought about dd'ing the /dev/md1 to a blank hard drive, then re-doing the raid from scratch and loading the data back, but there could be an easier solution.. Any help would be appreciated.

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  • Lenovo Windows 8 EFI restore from image

    - by anderhil
    First time here. I have bought e530 with windows 8 and the first hour of work with it i have a problem. I have ssd with windows 7 which i want to use with my new e530. I have made a sysprep of win 7 and installed ssd to the e530. The HDD which was inside e530 i want to use as second hdd instead of my DVD Drive. I connected this HDD through usb-to-sata adapter to copy some files from ssd to the hdd. Unfortunately it didn't see the file system on the HDD (but first time i have booted to it and first boot into Windows 8) I've made some mistakes and i corrupt the filesystem on the hdd. I tried bunch of tricks to recover the GPT, but it didn't work. I have managed to recover the Lenovo_Recovery partition to my ssd using recovery tools. And now I'm stuck, with this new things to me - EFI, GPT, etc i don't how this stuff works, and i have been trying to understand this for hours - but nothing seems to work. I want to restore the Windows 8 to the hdd, so it is there alive. What i have done so far: Formated the HDD I took the PBRALL file from the Lenovo_Recovery " convert gpt create partition Primary size=1000 ID=DE94BBA4-06D1-4D40-A16A-BFD50179D6AC gpt attributes=0x8000000000000001 assign letter=W format quick LABEL=WINRE_DRV create partition efi size=260 assign letter=s format quick fs=fat32 LABEL=SYSTEM_DRV create par msr size=128 create partition primary noerr assign letter=t format quick LABEL=Windows8_OS shrink desired=12197 create partition Primary ID=DE94BBA4-06D1-4D40-A16A-BFD50179D6AC gpt attributes=0x8000000000000001 assign letter=q format quick LABEL=Lenovo_Recovery " it recreated the partitions copied contents of SDRIVE.zip to SYSTEM_DRV partition copied contents of WDRIVE.zip to WINRE_DRV partition Copied restored Lenovo_Recovery back to Lenovo_Recovery partition So now I have 3 system partitions: SYSTE_DRV BOOT boot.sdi EFI BOOT bootx64.efi LenovoBT.efi Lenovo ... Microsoft ... WINRE_DRV\Recovery\WindowsRE\winre.wim Lenovo_Recovery (whic contains install.wim and bunch of other things) So i put back the HDD inside the laptop and tried to boot - but nothing works. It just doesn't boot to anything - no errors - nothing at all. When I choose this HDD manually for boot - just black screen blinks and that's all - it returns back to the devices boot menu. SYSTEM_DRV is EFI partition, so I don't understand why it doesn't boot, it has files needed inside. Can anybody tell me what should be done to make it boot to recovery console or smth like that? How to restore the Windows 8 from the Lenovo_Recovery install.wim image? As I understand I have all the files where they should be, but why it doesn't work? How to troubleshoot such things? Also, if somebody has good link where EFI booting process is explained in details that would be great. Cause i still don't understand how it knows what partition to boot?

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  • Windows 7 immediately disconnects a USB drive

    - by Daniel Saner
    I am having a problem with Windows 7 x64 consistently disconnecting one specific USB mass storage drive immediately after it is connected. The drive in question is a Cowon C2 digital music player which works in standard mass storage controller mode (i.e. no device-specific drivers needed/available). When I connect the player, Windows plays the "USB connect" sound and the device appears (under its correct name) in the device manager, but it never appears as a drive. The player itself displays "USB Connected" for a split-second before reporting that it has been disconnected again. Since the player, by design, reboots after it has been disconnected, Windows plays the "USB disconnect" sound before restarting the whole cycle once the player has powered back on. I am connecting the player through an Intel X79 Chipset motherboard (Gigabyte GA-X79-UD3) to Windows 7 Pro 64-bit. The player used to work fine the first few times I connected it, showing up as an external drive; it only recently stopped working. It is not a problem with the player, since it works fine when connected to another computer, even such running the exact same operating system. It is also not a problem with the USB controller, since the issue is the same on both the Intel USB 2.0 and the Fresco Logic FL1009 USB 3.0 controller ports. I have also not had the problem with any other drive so far. Among the things I have tried so far: Disabling USB legacy mode in BIOS Disabling energy-saving power down for all USB controllers in Windows' device manager Removing and reinstalling Windows' USB mass storage driver Removing and reinstalling Intel and Fresco Logic USB controller driver Restoring the player to factory defaults None of these made a difference. Again, the player used to work fine on the exact same system just days ago; I didn't install any new hardware or drivers on it since then. I would be very grateful for any hints on what else to try. Edit: Here is another new hint; I found out that when I connect the drive before booting Windows, it is available in Windows Explorer as it should, and does not automatically disconnect. If I remove and reconnect it though, the infinite connect/disconnect-loop starts anew.

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  • How to Create Auto Playlists in Windows Media Player 12

    - by DigitalGeekery
    Are you getting tired of the same old playlists in Windows Media Player? Today we’ll show you how to create dynamic auto playlists based on criteria you choose in WMP 12 in Windows 7. Auto Playlists In Library view, click on Create playlist dropdown arrow and select Create auto playlist. On the New Auto Playlist window type in a name for the playlist in the text box. Now we need to choose our criteria by which to filter your playlist. Select Click here to add criteria. For our example, we will create a playlist of songs that were added to the library in the last week from the Alternative genre. So, we will first select Date Added from the dropdown list. Many criteria will have addition options to configure. In the example below you will see that we have a few options to fine tune.   We will filter all the songs added to the library in the last 7 days. We will select Is After from the first dropdown list. Then select Last 7 Days from the second dropdown list. You can add multiple criteria to further filter your playlist. If you can’t find the criteria you are looking for, select “More” at the bottom of the dropdown list.   This will pull up a filter window with all the criteria. Select a filter and then click OK when finished.   From the Genre dropdown, we will select Alternative. If you’d like to add Pictures, Videos, or TV Shows to your auto playlists you can do so by selecting them from the dropdown list under And also include. You will then be able to select criteria for your pictures, videos, or TV shows from the dropdown list.   Finally, you can also add restrictions to your music such as the number of items, duration, or total size. We will limit the duration of our playlist to one hour by selecting Limit Total Duration To… Then type in 1 hour…Click OK.   Our library is automatically filtered and a playlist is created based on the criteria we selected. When additional songs are added to the Windows Media Player library, any of new songs that fit the criteria will automatically be added to the New Songs playlist. You can also save a copy of an auto playlist as a regular playlist. Switch to Playlists view by clicking Playlists from either the top menu or the navigation bar. Select the Play tab and then click Clear list to remove any tracks from the list pane.   Right-click on the playlist you want to save, select Add to, and then Play list. The songs from your auto playlist will appear as an Unsaved list on the list pane. Click Save list. Type in a name for your playlist. Your auto playlist will continue to change as you add or remove items from your Media Player library that meet the criteria you established. The new saved playlist we just created will stay as it is currently. Editing a Auto playlist is easy. Right-click on the playlist and select Edit. Now you are ready to enjoy your playlist. Conclusion Auto playlists are great way to keep your playlists fresh in Windows Media Player 12. Users can get creative and experiment with the wide variety of criteria to customize their listening experience. If you are new to playlists in Windows Media Player, you may want to check our our previous post on how to create custom playlists in Windows Media Player 12. Are you looking to get better sound from WMP 12? Take a look at how to improve playback using enhancements in Windows Media Player 12. Similar Articles Productive Geek Tips Create Custom Playlists in Windows Media Player 12Fixing When Windows Media Player Library Won’t Let You Add FilesInstall and Use the VLC Media Player on Ubuntu LinuxMake Windows Media Player Automatically Open in Mini Player ModeMake VLC Player Look like Windows Media Player 10 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 VMware Workstation 7 Acronis Online Backup DVDFab 6 Revo Uninstaller Pro Error Goblin Explains Windows Error Codes Twelve must-have Google Chrome plugins Cool Looking Skins for Windows Media Player 12 Move the Mouse Pointer With Your Face Movement Using eViacam Boot Windows Faster With Boot Performance Diagnostics Create Ringtones For Your Android Phone With RingDroid

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  • How To Configure Remote Desktop To Hyper-V Guest Virtual Machines

    - by Brian Jackett
    Configuring Remote Desktop (RDP) from a host Hyper-V machine to a guest virtual machine can be tricky, so this post is dedicated to the issues and resolution steps I went through to allow RDP.  Cutting to the point, below are the things to look for followed by some explanation about my scenario if you care to read.  This is not an exhaustive list of what is required, just the items that were causing problems for my particular scenario. Requirements Allow Remote Desktop Connections in guest OS. The network adapter type must allow communication with host machine (e.g. use an “Internal” virtual adapter.) If running Server 2008 R2 on guest, network discovery mode must be turned on. If running Server 2008 R2 on guest, the services supporting network discovery mode must be running: - DNS Client - Function Discovery Resource Publication - SSDP Discovery - UPnP Device Host My Environment     A quick word about my environment.  I am running Windows Server 2008 R2 with Hyper V on my laptop and numerous guest VMs running Windows Server 2003 R2 or Windows Server 2008 R2.  I run a domain controller VM and then 1 or 2 SharePoint servers depending on my work needs.  I’ve found this setup to work well except when it comes to the display window for my VMs. The Issue     Ever since I began running Hyper-V I haven’t been able to RDP to my guest VMs which means the resolution for my connection windows ha been limited to what the native Hyper-V connections allow.  During personal use I can put the resolution up to 1152 x 864, but during presentations I am usually limited to a measly 800 x 600.  That is until today when I decided to fully investigate why I couldn’t connect via RDP.     First a thank you to John Ross (@johnrossjr), Christina Wheeler (@cwheeler76) and Clayton Cobb (@warrtalon) for various suggestions while I was researching tonight.  As it turns out I had not 1, not 2, but 3 items preventing me from using RDP.  Let’s dig into the requirements above. Allow RDP Connection     This item I had previously taken care of, but it bears repeating because by default Windows Server 2008 R2 does not allow RDP connections.  Change the setting from “Don’t allow…” to whichever “Allow connections…” setting suits your needs.  I chose the less secure option as this is just my dev laptop. Network Adapter Type     When I originally configured my VMs I configured each to use 2 network adapters: one using the physical ethernet adapter for internet use and a virtual private adapter for communication between the VMs.  The connection for the ethernet adapter is an "”External” adapter and thus doesn’t connect between the host and guest.  The virtual private adapter allowed communication ONLY between the VMs and not to my host.  There is a third option “Internal” which allows communication between VMs as well as to the host.  After finding out this distinction I promptly created an Internal network adapter and assigned that to my VMs. Turn On Network Discovery     Seems like a pretty common sense thing, but in order to allow remote desktop connections the target computer must able to be found by the source computer (explained here.)  One of the settings that controls if a computer can be found on the network is aptly named Network Discovery.  By default Windows Server 2008 R2 turns Network Discovery off for security purposes.  To enable it open up the Network and Sharing Center.  Click “Change Advanced Sharing Settings” on the left.  On the following screen select “Turn on network discovery” for the currently used profile and click Save Settings.  You may notice though that your selection to turn on network discovery doesn’t save.  If this is the case then you most likely don’t have the supporting services running (as was my case.) Network Discovery Supporting Services     There are a total of 4 services (listed again below) that need to be running before you can turn on network discovery (explained here.)  The below images highlight these services.  In my guest VM I found that I had DNS Client already running while the other 3 were disabled.  I set them all to enabled and started the ones that were stopped.  After this change I returned to the Sharing settings screen and found that Network Discovery was turned on.  I’m not sure whether this was picking up my attempt to turn it on previously or if starting those services turned it on.  Either way the end result was a success. - DNS Client - Function Discovery Resource Publication - SSDP Discovery - UPnP Device Host Before and After Results     The first image is the smaller square shaped viewing window used by the Hyper-V native connection.  The second is the full-screen RDP connection in all its widescreen glory. Conclusion     Over the past few months I’ve found Hyper-V to be very useful for virtualizing my development environments, but I’ve also had a steep learning curve to get various items configured just right.  Allowing RDP connections to guest VMs was one area that I hadn’t been able to get right for the longest time.  Now that I resolved these issues I hope that others can avoid the pitfalls that I ran into.  If you know of any other items I left off feel free to let me know.        -Frog Out   Links Turning on Network Discovery http://sqlblog.com/blogs/john_paul_cook/archive/2009/08/15/remote-desktop-connection-on-windows-server-2008-r2.aspx Services required for Network Discovery http://social.technet.microsoft.com/Forums/en-US/winservergen/thread/2e1fea01-3f2b-4c46-a631-a8db34ed4f84

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  • MVC Portable Areas Enhancement &ndash; Embedded Resource Controller

    - by Steve Michelotti
    MvcContrib contains a feature called Portable Areas which I’ve recently blogged about. In short, portable areas provide a way to distribute MVC binary components as simple .NET assemblies where the aspx/ascx files are actually compiled into the assembly as embedded resources. This is an extremely cool feature but once you start building robust portable areas, you’ll also want to be able to access other external files like css and javascript.  After my recent post suggesting portable areas be expanded to include other embedded resources, Eric Hexter asked me if I’d like to contribute the code to MvcContrib (which of course I did!). Embedded resources are stored in a case-sensitive way in .NET assemblies and the existing embedded view engine inside MvcContrib already took this into account. Obviously, we’d want the same case sensitivity handling to be taken into account for any embedded resource so my job consisted of 1) adding the Embedded Resource Controller, and 2) a little refactor to extract the logic that deals with embedded resources so that the embedded view engine and the embedded resource controller could both leverage it and, therefore, keep the code DRY. The embedded resource controller targets these scenarios: External image files that are referenced in an <img> tag External files referenced like css or JavaScript files Image files referenced inside css files Embedded Resources Walkthrough This post will describe a walkthrough of using the embedded resource controller in your portable areas to include the scenarios outlined above. I will build a trivial “Quick Links” widget to illustrate the concepts. The portable area registration is the starting point for all portable areas. The MvcContrib.PortableAreas.EmbeddedResourceController is optional functionality – you must opt-in if you want to use it.  To do this, you simply “register” it by providing a route in your area registration that uses it like this: 1: context.MapRoute("ResourceRoute", "quicklinks/resource/{resourceName}", 2: new { controller = "EmbeddedResource", action = "Index" }, 3: new string[] { "MvcContrib.PortableAreas" }); First, notice that I can specify any route I want (e.g., “quicklinks/resources/…”).  Second, notice that I need to include the “MvcContrib.PortableAreas” namespace as the fourth parameter so that the framework is able to find the EmbeddedResourceController at runtime. The handling of embedded views and embedded resources have now been merged.  Therefore, the call to: 1: RegisterTheViewsInTheEmmeddedViewEngine(GetType()); has now been removed (breaking change).  It has been replaced with: 1: RegisterAreaEmbeddedResources(); Other than that, the portable area registration remains unchanged. The solution structure for the static files in my portable area looks like this: I’ve got a css file in a folder called “Content” as well as a couple of image files in a folder called “images”. To reference these in my aspx/ascx code, all of have to do is this: 1: <link href="<%= Url.Resource("Content.QuickLinks.css") %>" rel="stylesheet" type="text/css" /> 2: <img src="<%= Url.Resource("images.globe.png") %>" /> This results in the following HTML mark up: 1: <link href="/quicklinks/resource/Content.QuickLinks.css" rel="stylesheet" type="text/css" /> 2: <img src="/quicklinks/resource/images.globe.png" /> The Url.Resource() method is now included in MvcContrib as well. Make sure you import the “MvcContrib” namespace in your views. Next, I have to following html to render the quick links: 1: <ul class="links"> 2: <li><a href="http://www.google.com">Google</a></li> 3: <li><a href="http://www.bing.com">Bing</a></li> 4: <li><a href="http://www.yahoo.com">Yahoo</a></li> 5: </ul> Notice the <ul> tag has a class called “links”. This is defined inside my QuickLinks.css file and looks like this: 1: ul.links li 2: { 3: background: url(/quicklinks/resource/images.navigation.png) left 4px no-repeat; 4: padding-left: 20px; 5: margin-bottom: 4px; 6: } On line 3 we’re able to refer to the url for the background property. As a final note, although we already have complete control over the location of the embedded resources inside the assembly, what if we also want control over the physical URL routes as well. This point was raised by John Nelson in this post. This has been taken into account as well. For example, suppose you want your physical url to look like this: 1: <img src="/quicklinks/images/globe.png" /> instead of the same corresponding URL shown above (i.e., “/quicklinks/resources/images.globe.png”). You can do this easily by specifying another route for it which includes a “resourcePath” parameter that is pre-pended. Here is the complete code for the area registration with the custom route for the images shown on lines 9-11: 1: public class QuickLinksRegistration : PortableAreaRegistration 2: { 3: public override void RegisterArea(System.Web.Mvc.AreaRegistrationContext context, IApplicationBus bus) 4: { 5: context.MapRoute("ResourceRoute", "quicklinks/resource/{resourceName}", 6: new { controller = "EmbeddedResource", action = "Index" }, 7: new string[] { "MvcContrib.PortableAreas" }); 8:   9: context.MapRoute("ResourceImageRoute", "quicklinks/images/{resourceName}", 10: new { controller = "EmbeddedResource", action = "Index", resourcePath = "images" }, 11: new string[] { "MvcContrib.PortableAreas" }); 12:   13: context.MapRoute("quicklink", "quicklinks/{controller}/{action}", 14: new {controller = "links", action = "index"}); 15:   16: this.RegisterAreaEmbeddedResources(); 17: } 18:   19: public override string AreaName 20: { 21: get 22: { 23: return "QuickLinks"; 24: } 25: } 26: } The Quick Links portable area results in the following requests (including custom route formats): The complete code for this post is now included in the Portable Areas sample solution in the latest MvcContrib source code. You can get the latest code now.  Portable Areas open up exciting new possibilities for MVC development!

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  • MEF CompositionInitializer for WPF

    - by Reed
    The Managed Extensibility Framework is an amazingly useful addition to the .NET Framework.  I was very excited to see System.ComponentModel.Composition added to the core framework.  Personally, I feel that MEF is one tool I’ve always been missing in my .NET development. Unfortunately, one perfect scenario for MEF tends to fall short of it’s full potential is in Windows Presentation Foundation development.  In particular, there are many times when the XAML parser constructs objects in WPF development, which makes composition of those parts difficult.  The current release of MEF (Preview Release 9) addresses this for Silverlight developers via System.ComponentModel.Composition.CompositionInitializer.  However, there is no equivalent class for WPF developers. The CompositionInitializer class provides the means for an object to compose itself.  This is very useful with WPF and Silverlight development, since it allows a View, such as a UserControl, to be generated via the standard XAML parser, and still automatically pull in the appropriate ViewModel in an extensible manner.  Glenn Block has demonstrated the usage for Silverlight in detail, but the same issues apply in WPF. As an example, let’s take a look at a very simple case.  Take the following XAML for a Window: <Window x:Class="WpfApplication1.MainView" xmlns="http://schemas.microsoft.com/winfx/2006/xaml/presentation" xmlns:x="http://schemas.microsoft.com/winfx/2006/xaml" Title="MainWindow" Height="220" Width="300"> <Grid> <TextBlock Text="{Binding TheText}" /> </Grid> </Window> This does nothing but create a Window, add a simple TextBlock control, and use it to display the value of our “TheText” property in our DataContext class.  Since this is our main window, WPF will automatically construct and display this Window, so we need to handle constructing the DataContext and setting it ourselves. We could do this in code or in XAML, but in order to do it directly, we would need to hard code the ViewModel type directly into our XAML code, or we would need to construct the ViewModel class and set it in the code behind.  Both have disadvantages, and the disadvantages grow if we’re using MEF to compose our ViewModel. Ideally, we’d like to be able to have MEF construct our ViewModel for us.  This way, it can provide any construction requirements for our ViewModel via [ImportingConstructor], and it can handle fully composing the imported properties on our ViewModel.  CompositionInitializer allows this to occur. We use CompositionInitializer within our View’s constructor, and use it for self-composition of our View.  Using CompositionInitializer, we can modify our code behind to: public partial class MainView : Window { public MainView() { InitializeComponent(); CompositionInitializer.SatisfyImports(this); } [Import("MainViewModel")] public object ViewModel { get { return this.DataContext; } set { this.DataContext = value; } } } We then can add an Export on our ViewModel class like so: [Export("MainViewModel")] public class MainViewModel { public string TheText { get { return "Hello World!"; } } } MEF will automatically compose our application, decoupling our ViewModel injection to the DataContext of our View until runtime.  When we run this, we’ll see: There are many other approaches for using MEF to wire up the extensible parts within your application, of course.  However, any time an object is going to be constructed by code outside of your control, CompositionInitializer allows us to continue to use MEF to satisfy the import requirements of that object. In order to use this from WPF, I’ve ported the code from MEF Preview 9 and Glenn Block’s (now obsolete) PartInitializer port to Windows Presentation Foundation.  There are some subtle changes from the Silverlight port, mainly to handle running in a desktop application context.  The default behavior of my port is to construct an AggregateCatalog containing a DirectoryCatalog set to the location of the entry assembly of the application.  In addition, if an “Extensions” folder exists under the entry assembly’s directory, a second DirectoryCatalog for that folder will be included.  This behavior can be overridden by specifying a CompositionContainer or one or more ComposablePartCatalogs to the System.ComponentModel.Composition.Hosting.CompositionHost static class prior to the first use of CompositionInitializer. Please download CompositionInitializer and CompositionHost for VS 2010 RC, and contact me with any feedback. Composition.Initialization.Desktop.zip Edit on 3/29: Glenn Block has since updated his version of CompositionInitializer (and ExportFactory<T>!), and made it available here: http://cid-f8b2fd72406fb218.skydrive.live.com/self.aspx/blog/Composition.Initialization.Desktop.zip This is a .NET 3.5 solution, and should soon be pushed to CodePlex, and made available on the main MEF site.

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  • C# 5 Async, Part 1: Simplifying Asynchrony – That for which we await

    - by Reed
    Today’s announcement at PDC of the future directions C# is taking excite me greatly.  The new Visual Studio Async CTP is amazing.  Asynchronous code – code which frustrates and demoralizes even the most advanced of developers, is taking a huge leap forward in terms of usability.  This is handled by building on the Task functionality in .NET 4, as well as the addition of two new keywords being added to the C# language: async and await. This core of the new asynchronous functionality is built upon three key features.  First is the Task functionality in .NET 4, and based on Task and Task<TResult>.  While Task was intended to be the primary means of asynchronous programming with .NET 4, the .NET Framework was still based mainly on the Asynchronous Pattern and the Event-based Asynchronous Pattern. The .NET Framework added functionality and guidance for wrapping existing APIs into a Task based API, but the framework itself didn’t really adopt Task or Task<TResult> in any meaningful way.  The CTP shows that, going forward, this is changing. One of the three key new features coming in C# is actually a .NET Framework feature.  Nearly every asynchronous API in the .NET Framework has been wrapped into a new, Task-based method calls.  In the CTP, this is done via as external assembly (AsyncCtpLibrary.dll) which uses Extension Methods to wrap the existing APIs.  However, going forward, this will be handled directly within the Framework.  This will have a unifying effect throughout the .NET Framework.  This is the first building block of the new features for asynchronous programming: Going forward, all asynchronous operations will work via a method that returns Task or Task<TResult> The second key feature is the new async contextual keyword being added to the language.  The async keyword is used to declare an asynchronous function, which is a method that either returns void, a Task, or a Task<T>. Inside the asynchronous function, there must be at least one await expression.  This is a new C# keyword (await) that is used to automatically take a series of statements and break it up to potentially use discontinuous evaluation.  This is done by using await on any expression that evaluates to a Task or Task<T>. For example, suppose we want to download a webpage as a string.  There is a new method added to WebClient: Task<string> WebClient.DownloadStringTaskAsync(Uri).  Since this returns a Task<string> we can use it within an asynchronous function.  Suppose, for example, that we wanted to do something similar to my asynchronous Task example – download a web page asynchronously and check to see if it supports XHTML 1.0, then report this into a TextBox.  This could be done like so: private async void button1_Click(object sender, RoutedEventArgs e) { string url = "http://reedcopsey.com"; string content = await new WebClient().DownloadStringTaskAsync(url); this.textBox1.Text = string.Format("Page {0} supports XHTML 1.0: {1}", url, content.Contains("XHTML 1.0")); } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Let’s walk through what’s happening here, step by step.  By adding the async contextual keyword to the method definition, we are able to use the await keyword on our WebClient.DownloadStringTaskAsync method call. When the user clicks this button, the new method (Task<string> WebClient.DownloadStringTaskAsync(string)) is called, which returns a Task<string>.  By adding the await keyword, the runtime will call this method that returns Task<string>, and execution will return to the caller at this point.  This means that our UI is not blocked while the webpage is downloaded.  Instead, the UI thread will “await” at this point, and let the WebClient do it’s thing asynchronously. When the WebClient finishes downloading the string, the user interface’s synchronization context will automatically be used to “pick up” where it left off, and the Task<string> returned from DownloadStringTaskAsync is automatically unwrapped and set into the content variable.  At this point, we can use that and set our text box content. There are a couple of key points here: Asynchronous functions are declared with the async keyword, and contain one or more await expressions In addition to the obvious benefits of shorter, simpler code – there are some subtle but tremendous benefits in this approach.  When the execution of this asynchronous function continues after the first await statement, the initial synchronization context is used to continue the execution of this function.  That means that we don’t have to explicitly marshal the call that sets textbox1.Text back to the UI thread – it’s handled automatically by the language and framework!  Exception handling around asynchronous method calls also just works. I’d recommend every C# developer take a look at the documentation on the new Asynchronous Programming for C# and Visual Basic page, download the Visual Studio Async CTP, and try it out.

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  • SSAS: Using fake dimension and scopes for dynamic ranges

    - by DigiMortal
    In one of my BI projects I needed to find count of objects in income range. Usual solution with range dimension was useless because range where object belongs changes in time. These ranges depend on calculation that is done over incomes measure so I had really no option to use some classic solution. Thanks to SSAS forums I got my problem solved and here is the solution. The problem – how to create dynamic ranges? I have two dimensions in SSAS cube: one for invoices related to objects rent and the other for objects. There is measure that sums invoice totals and two calculations. One of these calculations performs some computations based on object income and some other object attributes. Second calculation uses first one to define income ranges where object belongs. What I need is query that returns me how much objects there are in each group. I cannot use dimension for range because on one date object may belong to one range and two days later to another income range. By example, if object is not rented out for two days it makes no money and it’s income stays the same as before. If object is rented out after two days it makes some income and this income may move it to another income range. Solution – fake dimension and scopes Thanks to Gerhard Brueckl from pmOne I got everything work fine after some struggling with BI Studio. The original discussion he pointed out can be found from SSAS official forums thread Create a banding dimension that groups by a calculated measure. Solution was pretty simple by nature – we have to define fake dimension for our range and use scopes to assign values for object count measure. Object count measure is primitive – it just counts objects and that’s it. We will use it to find out how many objects belong to one or another range. We also need table for fake ranges and we have to fill it with ranges used in ranges calculation. After creating the table and filling it with ranges we can add fake range dimension to our cube. Let’s see now how to solve the problem step-by-step. Solving the problem Suppose you have ranges calculation defined like this: CASE WHEN [Measures].[ComplexCalc] < 0 THEN 'Below 0'WHEN [Measures].[ComplexCalc] >=0 AND  [Measures].[ComplexCalc] <=50 THEN '0 - 50'...END Let’s create now new table to our analysis database and name it as FakeIncomeRange. Here is the definition for table: CREATE TABLE [FakeIncomeRange] (     [range_id] [int] IDENTITY(1,1) NOT NULL,     [range_name] [nvarchar](50) NOT NULL,     CONSTRAINT [pk_fake_income_range] PRIMARY KEY CLUSTERED      (         [range_id] ASC     ) ) Don’t forget to fill this table with range labels you are using in ranges calculation. To use ranges from table we have to add this table to our data source view and create new dimension. We cannot bind this table to other tables but we have to leave it like it is. Our dimension has two attributes: ID and Name. The next thing to create is calculation that returns objects count. This calculation is also fake because we override it’s values for all ranges later. Objects count measure can be defined as calculation like this: COUNT([Object].[Object].[Object].members) Now comes the most crucial part of our solution – defining the scopes. Based on data used in this posting we have to define scope for each of our ranges. Here is the example for first range. SCOPE([FakeIncomeRange].[Name].&[Below 0], [Measures].[ObjectCount])     This=COUNT(            FILTER(                [Object].[Object].[Object].members,                 [Measures].[ComplexCalc] < 0          )     ) END SCOPE To get these scopes defined in cube we need MDX script blocks for each line given here. Take a look at the screenshot to get better idea what I mean. This example is given from SQL Server books online to avoid conflicts with NDA. :) From previous example the lines (MDX scripts) are: Line starting with SCOPE Block for This = Line with END SCOPE And now it is time to deploy and process our cube. Although you may see examples where there are semicolons in the end of statements you don’t need them. Visual Studio BI tools generate separate command from each script block so you don’t need to worry about it.

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  • Parallelism in .NET – Part 12, More on Task Decomposition

    - by Reed
    Many tasks can be decomposed using a Data Decomposition approach, but often, this is not appropriate.  Frequently, decomposing the problem into distinctive tasks that must be performed is a more natural abstraction. However, as I mentioned in Part 1, Task Decomposition tends to be a bit more difficult than data decomposition, and can require a bit more effort.  Before we being parallelizing our algorithm based on the tasks being performed, we need to decompose our problem, and take special care of certain considerations such as ordering and grouping of tasks. Up to this point in this series, I’ve focused on parallelization techniques which are most appropriate when a problem space can be decomposed by data.  Using PLINQ and the Parallel class, I’ve shown how problem spaces where there is a collection of data, and each element needs to be processed, can potentially be parallelized. However, there are many other routines where this is not appropriate.  Often, instead of working on a collection of data, there is a single piece of data which must be processed using an algorithm or series of algorithms.  Here, there is no collection of data, but there may still be opportunities for parallelism. As I mentioned before, in cases like this, the approach is to look at your overall routine, and decompose your problem space based on tasks.  The idea here is to look for discrete “tasks,” individual pieces of work which can be conceptually thought of as a single operation. Let’s revisit the example I used in Part 1, an application startup path.  Say we want our program, at startup, to do a bunch of individual actions, or “tasks”.  The following is our list of duties we must perform right at startup: Display a splash screen Request a license from our license manager Check for an update to the software from our web server If an update is available, download it Setup our menu structure based on our current license Open and display our main, welcome Window Hide the splash screen The first step in Task Decomposition is breaking up the problem space into discrete tasks. This, naturally, can be abstracted as seven discrete tasks.  In the serial version of our program, if we were to diagram this, the general process would appear as: These tasks, obviously, provide some opportunities for parallelism.  Before we can parallelize this routine, we need to analyze these tasks, and find any dependencies between tasks.  In this case, our dependencies include: The splash screen must be displayed first, and as quickly as possible. We can’t download an update before we see whether one exists. Our menu structure depends on our license, so we must check for the license before setting up the menus. Since our welcome screen will notify the user of an update, we can’t show it until we’ve downloaded the update. Since our welcome screen includes menus that are customized based off the licensing, we can’t display it until we’ve received a license. We can’t hide the splash until our welcome screen is displayed. By listing our dependencies, we start to see the natural ordering that must occur for the tasks to be processed correctly. The second step in Task Decomposition is determining the dependencies between tasks, and ordering tasks based on their dependencies. Looking at these tasks, and looking at all the dependencies, we quickly see that even a simple decomposition such as this one can get quite complicated.  In order to simplify the problem of defining the dependencies, it’s often a useful practice to group our tasks into larger, discrete tasks.  The goal when grouping tasks is that you want to make each task “group” have as few dependencies as possible to other tasks or groups, and then work out the dependencies within that group.  Typically, this works best when any external dependency is based on the “last” task within the group when it’s ordered, although that is not a firm requirement.  This process is often called Grouping Tasks.  In our case, we can easily group together tasks, effectively turning this into four discrete task groups: 1. Show our splash screen – This needs to be left as its own task.  First, multiple things depend on this task, mainly because we want this to start before any other action, and start as quickly as possible. 2. Check for Update and Download the Update if it Exists - These two tasks logically group together.  We know we only download an update if the update exists, so that naturally follows.  This task has one dependency as an input, and other tasks only rely on the final task within this group. 3. Request a License, and then Setup the Menus – Here, we can group these two tasks together.  Although we mentioned that our welcome screen depends on the license returned, it also depends on setting up the menu, which is the final task here.  Setting up our menus cannot happen until after our license is requested.  By grouping these together, we further reduce our problem space. 4. Display welcome and hide splash - Finally, we can display our welcome window and hide our splash screen.  This task group depends on all three previous task groups – it cannot happen until all three of the previous groups have completed. By grouping the tasks together, we reduce our problem space, and can naturally see a pattern for how this process can be parallelized.  The diagram below shows one approach: The orange boxes show each task group, with each task represented within.  We can, now, effectively take these tasks, and run a large portion of this process in parallel, including the portions which may be the most time consuming.  We’ve now created two parallel paths which our process execution can follow, hopefully speeding up the application startup time dramatically. The main point to remember here is that, when decomposing your problem space by tasks, you need to: Define each discrete action as an individual Task Discover dependencies between your tasks Group tasks based on their dependencies Order the tasks and groups of tasks

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  • UCM 11g is 4 days old!

    - by kyle.hatlestad
    Ok...so I missed posting a blog entry when UCM 11g and the entire ECM suite released on Tuesday. Hopefully you've already seen the announcements on any number of the Oracle ECM blogs out there such as ECM Alerts, Fusion ECM, bex huff, or C4. So I won't bore you with the same talking points like 179 million check-ins per day or 124 web site page hits per second. Instead, I thought I'd show some screenshots of the new features in UCM and URM 11g. WebLogic Server and Enterprise Manager So probably the biggest change in 11g is UCM and URM now run on top of the WebLogic Server application server. This is a huge step as ECM is now on a standard platform with the rest of Oracle Fusion Middleware which makes installation, configuration, and integration consistent among all the products. From a feature perspective, it's also beneficial because it's now integrated with Oracle Enterprise Manager. Enterprise Manager provides a lot of provisioning control over servers as well as performance monitoring and access to logs and debugging information. Desktop Integration Suite Desktop Integration Suite got a complete overhaul for 11g. It exposes a lot more features within Windows Explorer such as saved searches, workflow queue, and checked-out items. It also now support metadata pop-up screens to let users fill in additional metadata when they drag-n-drop files in! And the integration within Office applications has changed significantly by introducing a dedicated UCM menu to do open, save, compare, etc. Site Studio for External Applications In UCM Site Studio 10gR4, a major architectural shift was introduced which brought several new objects such as elements, region definitions, region templates, and placeholder definitions. This truly separated the content from the display and from the definition. It also allowed separation of the content from needing to be rendered on a complete Site Studio page. Well, the new Site Studio for External Applications takes advantage of that architecture and introduces pre-built tags and plug-ins to JDeveloper to allow to go from simply adding a content area to your web application page to building an entire web site, just like you would have done in Site Studio Designer. In addition to these changes, enhancements to the core Site Studio have been added as well. One of the big ones is called Designer Mode which allows power-users to bypass the standard rules defined by the placeholder definition or template and perform any number of additional actions. This reduces the need to go back to Site Studio Designer or JDeveloper to make more advanced changes to the site. Dashboards As part of the updated records management functionality in both UCM and URM, users can now set a dashboard view on their home page to surface common functions in a single view. It has pre-built "portlets" users can choose from to display and organize they way they want. Behind the scenes, these dashboards are stored as Content Folios. So the dashboards themselves are content items that can be revisioned and shared between users. And new dashboard portlets can be easily added (like the User Profile one in the screenshots) by getting a copy of an existing one, modifying the display, and then checking it in as a new one to select from. URM Interface Enhancements URM includes several new UI and usability enhancements in 11g. There is a new view for physical records, a place to configure "favorite" items to quickly get to, and new placement of the records management menu. BI Publisher Reports Records management in UCM and URM now offer reports generated through embedded BI Publisher. Templates are controlled by rich text files checked directly into the repository, so they can be easily modified. Other Features A new Inbound Refinery conversion option is available that does native Microsoft Office HTML conversion. If your IBR is on Windows and you have the native applications loaded, the IBR can use them to produce HTML. A new GUI template editor for Dynamic Converter is available. It's written in Java so is available through all the supported browsers and platforms. The original ActiveX based editor is also still available. The Component Manager interface has changed to help provide an easier and more descriptive way to enable core components that are installed along with UCM. All of the supported components are immediately available to turn on and do not have to be installed separately as in previous versions. My Downloads is located in the My Content Server menu and provides for easy download of client installs including Desktop Integration Suite and Site Studio Designer. Well, hopefully that gives you a taste for some of the new things in 11g. We're all pretty excited here at Oracle about all the new changes and enhancements. Over the next few months I hope to highlight some of these features more in-depth, so keep your eye out for those posts.

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  • SQL SERVER – Database Dynamic Caching by Automatic SQL Server Performance Acceleration

    - by pinaldave
    My second look at SafePeak’s new version (2.1) revealed to me few additional interesting features. For those of you who hadn’t read my previous reviews SafePeak and not familiar with it, here is a quick brief: SafePeak is in business of accelerating performance of SQL Server applications, as well as their scalability, without making code changes to the applications or to the databases. SafePeak performs database dynamic caching, by caching in memory result sets of queries and stored procedures while keeping all those cache correct and up to date. Cached queries are retrieved from the SafePeak RAM in microsecond speed and not send to the SQL Server. The application gets much faster results (100-500 micro seconds), the load on the SQL Server is reduced (less CPU and IO) and the application or the infrastructure gets better scalability. SafePeak solution is hosted either within your cloud servers, hosted servers or your enterprise servers, as part of the application architecture. Connection of the application is done via change of connection strings or adding reroute line in the c:\windows\system32\drivers\etc\hosts file on all application servers. For those who would like to learn more on SafePeak architecture and how it works, I suggest to read this vendor’s webpage: SafePeak Architecture. More interesting new features in SafePeak 2.1 In my previous review of SafePeak new I covered the first 4 things I noticed in the new SafePeak (check out my article “SQLAuthority News – SafePeak Releases a Major Update: SafePeak version 2.1 for SQL Server Performance Acceleration”): Cache setup and fine-tuning – a critical part for getting good caching results Database templates Choosing which database to cache Monitoring and analysis options by SafePeak Since then I had a chance to play with SafePeak some more and here is what I found. 5. Analysis of SQL Performance (present and history): In SafePeak v.2.1 the tools for understanding of performance became more comprehensive. Every 15 minutes SafePeak creates and updates various performance statistics. Each query (or a procedure execute) that arrives to SafePeak gets a SQL pattern, and after it is used again there are statistics for such pattern. An important part of this product is that it understands the dependencies of every pattern (list of tables, views, user defined functions and procs). From this understanding SafePeak creates important analysis information on performance of every object: response time from the database, response time from SafePeak cache, average response time, percent of traffic and break down of behavior. One of the interesting things this behavior column shows is how often the object is actually pdated. The break down analysis allows knowing the above information for: queries and procedures, tables, views, databases and even instances level. The data is show now on all arriving queries, both read queries (that can be cached), but also any types of updates like DMLs, DDLs, DCLs, and even session settings queries. The stats are being updated every 15 minutes and SafePeak dashboard allows going back in time and investigating what happened within any time frame. 6. Logon trigger, for making sure nothing corrupts SafePeak cache data If you have an application with many parts, many servers many possible locations that can actually update the database, or the SQL Server is accessible to many DBAs or software engineers, each can access some database directly and do some changes without going thru SafePeak – this can create a potential corruption of the data stored in SafePeak cache. To make sure SafePeak cache is correct it needs to get all updates to arrive to SafePeak, and if a DBA will access the database directly and do some changes, for example, then SafePeak will simply not know about it and will not clean SafePeak cache. In the new version, SafePeak brought a new feature called “Logon Trigger” to solve the above challenge. By special click of a button SafePeak can deploy a special server logon trigger (with a CLR object) on your SQL Server that actually monitors all connections and informs SafePeak on any connection that is coming not from SafePeak. In SafePeak dashboard there is an interface that allows to control which logins can be ignored based on login names and IPs, while the rest will invoke cache cleanup of SafePeak and actually locks SafePeak cache until this connection will not be closed. Important to note, that this does not interrupt any logins, only informs SafePeak on such connection. On the Dashboard screen in SafePeak you will be able to see those connections and then decide what to do with them. Configuration of this feature in SafePeak dashboard can be done here: Settings -> SQL instances management -> click on instance -> Logon Trigger tab. Other features: 7. User management ability to grant permissions to someone without changing its configuration and only use SafePeak as performance analysis tool. 8. Better reports for analysis of performance using 15 minute resolution charts. 9. Caching of client cursors 10. Support for IPv6 Summary SafePeak is a great SQL Server performance acceleration solution for users who want immediate results for sites with performance, scalability and peak spikes challenges. Especially if your apps are packaged or 3rd party, since no code changes are done. SafePeak can significantly increase response times, by reducing network roundtrip to the database, decreasing CPU resource usage, eliminating I/O and storage access. SafePeak team provides a free fully functional trial www.safepeak.com/download and actually provides a one-on-one assistance during such trial. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: About Me, Pinal Dave, PostADay, SQL, SQL Authority, SQL Performance, SQL Query, SQL Server, SQL Tips and Tricks, SQL Utility, T SQL, Technology

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  • MVC 2 Editor Template for Radio Buttons

    - by Steve Michelotti
    A while back I blogged about how to create an HTML Helper to produce a radio button list.  In that post, my HTML helper was “wrapping” the FluentHtml library from MvcContrib to produce the following html output (given an IEnumerable list containing the items “Foo” and “Bar”): 1: <div> 2: <input id="Name_Foo" name="Name" type="radio" value="Foo" /><label for="Name_Foo" id="Name_Foo_Label">Foo</label> 3: <input id="Name_Bar" name="Name" type="radio" value="Bar" /><label for="Name_Bar" id="Name_Bar_Label">Bar</label> 4: </div> With the release of MVC 2, we now have editor templates we can use that rely on metadata to allow us to customize our views appropriately.  For example, for the radio buttons above, we want the “id” attribute to be differentiated and unique and we want the “name” attribute to be the same across radio buttons so the buttons will be grouped together and so model binding will work appropriately. We also want the “for” attribute in the <label> element being set to correctly point to the id of the corresponding radio button.  The default behavior of the RadioButtonFor() method that comes OOTB with MVC produces the same value for the “id” and “name” attributes so this isn’t exactly what I want out the the box if I’m trying to produce the HTML mark up above. If we use an EditorTemplate, the first gotcha that we run into is that, by default, the templates just work on your view model’s property. But in this case, we *also* was the list of items to populate all the radio buttons. It turns out that the EditorFor() methods do give you a way to pass in additional data. There is an overload of the EditorFor() method where the last parameter allows you to pass an anonymous object for “extra” data that you can use in your view – it gets put on the view data dictionary: 1: <%: Html.EditorFor(m => m.Name, "RadioButtonList", new { selectList = new SelectList(new[] { "Foo", "Bar" }) })%> Now we can create a file called RadioButtonList.ascx that looks like this: 1: <%@ Control Language="C#" Inherits="System.Web.Mvc.ViewUserControl" %> 2: <% 3: var list = this.ViewData["selectList"] as SelectList; 4: %> 5: <div> 6: <% foreach (var item in list) { 7: var radioId = ViewData.TemplateInfo.GetFullHtmlFieldId(item.Value); 8: var checkedAttr = item.Selected ? "checked=\"checked\"" : string.Empty; 9: %> 10: <input type="radio" id="<%: radioId %>" name="<%: ViewData.TemplateInfo.HtmlFieldPrefix %>" value="<%: item.Value %>" <%: checkedAttr %>/> 11: <label for="<%: radioId %>"><%: item.Text %></label> 12: <% } %> 13: </div> There are several things to note about the code above. First, you can see in line #3, it’s getting the SelectList out of the view data dictionary. Then on line #7 it uses the GetFullHtmlFieldId() method from the TemplateInfo class to ensure we get unique IDs. We pass the Value to this method so that it will produce IDs like “Name_Foo” and “Name_Bar” rather than just “Name” which is our property name. However, for the “name” attribute (on line #10) we can just use the normal HtmlFieldPrefix property so that we ensure all radio buttons have the same name which corresponds to the view model’s property name. We also get to leverage the fact the a SelectListItem has a Boolean Selected property so we can set the checkedAttr variable on line #8 and use it on line #10. Finally, it’s trivial to set the correct “for” attribute for the <label> on line #11 since we already produced that value. Because the TemplateInfo class provides all the metadata for our view, we’re able to produce this view that is widely re-usable across our application. In fact, we can create a couple HTML helpers to better encapsulate this call and make it more user friendly: 1: public static MvcHtmlString RadioButtonList<TModel, TProperty>(this HtmlHelper<TModel> htmlHelper, Expression<Func<TModel, TProperty>> expression, params string[] items) 2: { 3: return htmlHelper.RadioButtonList(expression, new SelectList(items)); 4: } 5:   6: public static MvcHtmlString RadioButtonList<TModel, TProperty>(this HtmlHelper<TModel> htmlHelper, Expression<Func<TModel, TProperty>> expression, IEnumerable<SelectListItem> items) 7: { 8: var func = expression.Compile(); 9: var result = func(htmlHelper.ViewData.Model); 10: var list = new SelectList(items, "Value", "Text", result); 11: return htmlHelper.EditorFor(expression, "RadioButtonList", new { selectList = list }); 12: } This allows us to simply the call like this: 1: <%: Html.RadioButtonList(m => m.Name, "Foo", "Bar" ) %> In that example, the values for the radio button are hard-coded and being passed in directly. But if you had a view model that contained a property for the collection of items you could call the second overload like this: 1: <%: Html.RadioButtonList(m => m.Name, Model.FooBarList ) %> The Editor templates introduced in MVC 2 definitely allow for much more flexible views/editors than previously available. By knowing about the features you have available to you with the TemplateInfo class, you can take these concepts and customize your editors with extreme flexibility and re-usability.

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  • [Windows 8] Update TextBox’s binding on TextChanged

    - by Benjamin Roux
    Since UpdateSourceTrigger is not available in WinRT we cannot update the text’s binding of a TextBox at will (or at least not easily) especially when using MVVM (I surely don’t want to write behind-code to do that in each of my apps !). Since this kind of demand is frequent (for example to disable of button if the TextBox is empty) I decided to create some attached properties to to simulate this missing behavior. namespace Indeed.Controls { public static class TextBoxEx { public static string GetRealTimeText(TextBox obj) { return (string)obj.GetValue(RealTimeTextProperty); } public static void SetRealTimeText(TextBox obj, string value) { obj.SetValue(RealTimeTextProperty, value); } public static readonly DependencyProperty RealTimeTextProperty = DependencyProperty.RegisterAttached("RealTimeText", typeof(string), typeof(TextBoxEx), null); public static bool GetIsAutoUpdate(TextBox obj) { return (bool)obj.GetValue(IsAutoUpdateProperty); } public static void SetIsAutoUpdate(TextBox obj, bool value) { obj.SetValue(IsAutoUpdateProperty, value); } public static readonly DependencyProperty IsAutoUpdateProperty = DependencyProperty.RegisterAttached("IsAutoUpdate", typeof(bool), typeof(TextBoxEx), new PropertyMetadata(false, OnIsAutoUpdateChanged)); private static void OnIsAutoUpdateChanged(DependencyObject sender, DependencyPropertyChangedEventArgs e) { var value = (bool)e.NewValue; var textbox = (TextBox)sender; if (value) { Observable.FromEventPattern<TextChangedEventHandler, TextChangedEventArgs>( o => textbox.TextChanged += o, o => textbox.TextChanged -= o) .Do(_ => textbox.SetValue(TextBoxEx.RealTimeTextProperty, textbox.Text)) .Subscribe(); } } } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } The code is composed of two attached properties. The first one “RealTimeText” reflects the text in real time (updated after each TextChanged event). The second one is only used to enable the functionality. To subscribe to the TextChanged event I used Reactive Extensions (Rx-Metro package in Nuget). If you’re not familiar with this framework just replace the code with a simple: textbox.TextChanged += textbox.SetValue(TextBoxEx.RealTimeTextProperty, textbox.Text); .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } To use these attached properties, it’s fairly simple <TextBox Text="{Binding Path=MyProperty, Mode=TwoWay}" ic:TextBoxEx.IsAutoUpdate="True" ic:TextBoxEx.RealTimeText="{Binding Path=MyProperty, Mode=TwoWay}" /> .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Just make sure to create a binding (in TwoWay) for both Text and RealTimeText. Hope this helps !

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  • Linux-Containers — Part 1: Overview

    - by Lenz Grimmer
    "Containers" by Jean-Pierre Martineau (CC BY-NC-SA 2.0). Linux Containers (LXC) provide a means to isolate individual services or applications as well as of a complete Linux operating system from other services running on the same host. To accomplish this, each container gets its own directory structure, network devices, IP addresses and process table. The processes running in other containers or the host system are not visible from inside a container. Additionally, Linux Containers allow for fine granular control of resources like RAM, CPU or disk I/O. Generally speaking, Linux Containers use a completely different approach than "classicial" virtualization technologies like KVM or Xen (on which Oracle VM Server for x86 is based on). An application running inside a container will be executed directly on the operating system kernel of the host system, shielded from all other running processes in a sandbox-like environment. This allows a very direct and fair distribution of CPU and I/O-resources. Linux containers can offer the best possible performance and several possibilities for managing and sharing the resources available. Similar to Containers (or Zones) on Oracle Solaris or FreeBSD jails, the same kernel version runs on the host as well as in the containers; it is not possible to run different Linux kernel versions or other operating systems like Microsoft Windows or Oracle Solaris for x86 inside a container. However, it is possible to run different Linux distribution versions (e.g. Fedora Linux in a container on top of an Oracle Linux host), provided it supports the version of the Linux kernel that runs on the host. This approach has one caveat, though - if any of the containers causes a kernel crash, it will bring down all other containers (and the host system) as well. For example, Oracle's Unbreakable Enterprise Kernel Release 2 (2.6.39) is supported for both Oracle Linux 5 and 6. This makes it possible to run Oracle Linux 5 and 6 container instances on top of an Oracle Linux 6 system. Since Linux Containers are fully implemented on the OS level (the Linux kernel), they can be easily combined with other virtualization technologies. It's certainly possible to set up Linux containers within a virtualized Linux instance that runs inside Oracle VM Server for Oracle VM Virtualbox. Some use cases for Linux Containers include: Consolidation of multiple separate Linux systems on one server: instances of Linux systems that are not performance-critical or only see sporadic use (e.g. a fax or print server or intranet services) do not necessarily need a dedicated server for their operations. These can easily be consolidated to run inside containers on a single server, to preserve energy and rack space. Running multiple instances of an application in parallel, e.g. for different users or customers. Each user receives his "own" application instance, with a defined level of service/performance. This prevents that one user's application could hog the entire system and ensures, that each user only has access to his own data set. It also helps to save main memory — if multiple instances of a same process are running, the Linux kernel can share memory pages that are identical and unchanged across all application instances. This also applies to shared libraries that applications may use, they are generally held in memory once and mapped to multiple processes. Quickly creating sandbox environments for development and testing purposes: containers that have been created and configured once can be archived as templates and can be duplicated (cloned) instantly on demand. After finishing the activity, the clone can safely be discarded. This allows to provide repeatable software builds and test environments, because the system will always be reset to its initial state for each run. Linux Containers also boot significantly faster than "classic" virtual machines, which can save a lot of time when running frequent build or test runs on applications. Safe execution of an individual application: if an application running inside a container has been compromised because of a security vulnerability, the host system and other containers remain unaffected. The potential damage can be minimized, analyzed and resolved directly from the host system. Note: Linux Containers on Oracle Linux 6 with the Unbreakable Enterprise Kernel Release 2 (2.6.39) are still marked as Technology Preview - their use is only recommended for testing and evaluation purposes. The Open-Source project "Linux Containers" (LXC) is driving the development of the technology behind this, which is based on the "Control Groups" (CGroups) and "Name Spaces" functionality of the Linux kernel. Oracle is actively involved in the Linux Containers development and contributes patches to the upstream LXC code base. Control Groups provide means to manage and monitor the allocation of resources for individual processes or process groups. Among other things, you can restrict the maximum amount of memory, CPU cycles as well as the disk and network throughput (in MB/s or IOP/s) that are available for an application. Name Spaces help to isolate process groups from each other, e.g. the visibility of other running processes or the exclusive access to a network device. It's also possible to restrict a process group's access and visibility of the entire file system hierarchy (similar to a classic "chroot" environment). CGroups and Name Spaces provide the foundation on which Linux containers are based on, but they can actually be used independently as well. A more detailed description of how Linux Containers can be created and managed on Oracle Linux will be explained in the second part of this article. Additional links related to Linux Containers: OTN Article: The Role of Oracle Solaris Zones and Linux Containers in a Virtualization Strategy Linux Containers on Wikipedia - Lenz Grimmer Follow me on: Personal Blog | Facebook | Twitter | Linux Blog |

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  • Change the User Interface Language in Ubuntu

    - by Matthew Guay
    Would you like to use your Ubuntu computer in another language?  Here’s how you can easily change your interface language in Ubuntu. Ubuntu’s default install only includes a couple languages, but it makes it easy to find and add a new interface language to your computer.  To get started, open the System menu, select Administration, and then click Language Support. Ubuntu may ask if you want to update or add components to your current default language when you first open the dialog.  Click Install to go ahead and install the additional components, or you can click Remind Me Later to wait as these will be installed automatically when you add a new language. Now we’re ready to find and add an interface language to Ubuntu.  Click Install / Remove Languages to add the language you want. Find the language you want in the list, and click the check box to install it.  Ubuntu will show you all the components it will install for the language; this often includes spellchecking files for OpenOffice as well.  Once you’ve made your selection, click Apply Changes to install your new language.  Make sure you’re connected to the internet, as Ubuntu will have to download the additional components you’ve selected. Enter your system password when prompted, and then Ubuntu will download the needed languages files and install them.   Back in the main Language & Text dialog, we’re now ready to set our new language as default.  Find your new language in the list, and then click and drag it to the top of the list. Notice that Thai is the first language listed, and English is the second.  This will make Thai the default language for menus and windows in this account.  The tooltip reminds us that this setting does not effect system settings like currency or date formats. To change these, select the Text Tab and pick your new language from the drop-down menu.  You can preview the changes in the bottom Example box. The changes we just made will only affect this user account; the login screen and startup will not be affected.  If you wish to change the language in the startup and login screens also, click Apply System-Wide in both dialogs.  Other user accounts will still retain their original language settings; if you wish to change them, you must do it from those accounts. Once you have your new language settings all set, you’ll need to log out of your account and log back in to see your new interface language.  When you re-login, Ubuntu may ask you if you want to update your user folders’ names to your new language.  For example, here Ubuntu is asking if we want to change our folders to their Thai equivalents.  If you wish to do so, click Update or its equivalents in your language. Now your interface will be almost completely translated into your new language.  As you can see here, applications with generic names are translated to Thai but ones with specific names like Shutter keep their original name. Even the help dialogs are translated, which makes it easy for users around to world to get started with Ubuntu.  Once again, you may notice some things that are still in English, but almost everything is translated. Adding a new interface language doesn’t add the new language to your keyboard, so you’ll still need to set that up.  Check out our article on adding languages to your keyboard to get this setup. If you wish to revert to your original language or switch to another new language, simply repeat the above steps, this time dragging your original or new language to the top instead of the one you chose previously. Conclusion Ubuntu has a large number of supported interface languages to make it user-friendly to people around the globe.  And since you can set the language for each user account, it’s easy for multi-lingual individuals to share the same computer. Or, if you’re using Windows, check out our article on how you can Change the User Interface Language in Vista or Windows 7, too! Similar Articles Productive Geek Tips Restart the Ubuntu Gnome User Interface QuicklyChange the User Interface Language in Vista or Windows 7Create a Samba User on UbuntuInstall Samba Server on UbuntuSee Which Groups Your Linux User Belongs To 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 VMware Workstation 7 Acronis Online Backup DVDFab 6 Revo Uninstaller Pro FetchMp3 Can Download Videos & Convert Them to Mp3 Use Flixtime To Create Video Slideshows Creating a Password Reset Disk in Windows Bypass Waiting Time On Customer Service Calls With Lucyphone MELTUP – "The Beginning Of US Currency Crisis And Hyperinflation" Enable or Disable the Task Manager Using TaskMgrED

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  • SQL SERVER – Spatial Database Queries – What About BLOB – T-SQL Tuesday #006

    - by pinaldave
    Michael Coles is one of the most interesting book authors I have ever met. He has a flair of writing complex stuff in a simple language. There are a very few people like that.  I really enjoyed reading his recent book, Expert SQL Server 2008 Encryption. I strongly suggest taking a look at it. This blog is written in response to T-SQL Tuesday #006: “What About BLOB? by Michael Coles. Spatial Database is my favorite subject. Since I did my TechEd India 2010 presentation, I have enjoyed this subject a lot. Before I continue this blog post, there are a few other blog posts, so I suggest you read them.  To help build the environment run the queries, I am going to present them in this single blog post. SQL SERVER – What is Spatial Database? – Developing with SQL Server Spatial and Deep Dive into Spatial Indexing This blog post explains the basics of Spatial Database and also provides a good introduction to Indexing concept. SQL SERVER – World Shapefile Download and Upload to Database – Spatial Database This blog post will enable you with how to load the shape file into database. SQL SERVER – Spatial Database Definition and Research Documents This blog post links to the white paper about Spatial Database written by Microsoft experts. SQL SERVER – Introduction to Spatial Coordinate Systems: Flat Maps for a Round Planet This blog post links to the white paper explaining coordinate system, as written by Microsoft experts. After reading the above listed blog posts, I am very confident that you are ready to run the following script. Once you create a database using the World Shapefile, as mentioned in the second link above,you can display the image of India just like the following. Please note that this is not an accurate political map. The boundary of this map has many errors and it is just a representation. You can run the following query to generate the map of India from the database spatial which you have created after following the instructions here. USE Spatial GO -- India Map SELECT [CountryName] ,[BorderAsGeometry] ,[Border] FROM [Spatial].[dbo].[Countries] WHERE Countryname = 'India' GO Now, let us find the longitude and latitude of the two major IT cities of India, Hyderabad and Bangalore. I find their values as the following: the values of longitude-latitude for Bangalore is 77.5833300000 13.0000000000; for Hyderabad, longitude-latitude is 78.4675900000 17.4531200000. Now, let us try to put these values on the India Map and see their location. -- Bangalore DECLARE @GeoLocation GEOGRAPHY SET @GeoLocation = GEOGRAPHY::STPointFromText('POINT(77.5833300000 13.0000000000)',4326).STBuffer(20000); -- Hyderabad DECLARE @GeoLocation1 GEOGRAPHY SET @GeoLocation1 = GEOGRAPHY::STPointFromText('POINT(78.4675900000 17.4531200000)',4326).STBuffer(20000); -- Bangalore and Hyderabad on Map of India SELECT name, [GeoLocation] FROM [IndiaGeoNames] I WHERE I.[GeoLocation].STDistance(@GeoLocation) <= 0 UNION ALL SELECT name, [GeoLocation] FROM [IndiaGeoNames] I WHERE I.[GeoLocation].STDistance(@GeoLocation1) <= 0 UNION ALL SELECT '',[Border] FROM [Spatial].[dbo].[Countries] WHERE Countryname = 'India' GO Now let us quickly draw a straight line between them. DECLARE @GeoLocation GEOGRAPHY SET @GeoLocation = GEOGRAPHY::STPointFromText('POINT(78.4675900000 17.4531200000)',4326).STBuffer(10000); DECLARE @GeoLocation1 GEOGRAPHY SET @GeoLocation1 = GEOGRAPHY::STPointFromText('POINT(77.5833300000 13.0000000000)',4326).STBuffer(10000); DECLARE @GeoLocation2 GEOGRAPHY SET @GeoLocation2 = GEOGRAPHY::STGeomFromText('LINESTRING(78.4675900000 17.4531200000, 77.5833300000 13.0000000000)',4326) SELECT name, [GeoLocation] FROM [IndiaGeoNames] I WHERE I.[GeoLocation].STDistance(@GeoLocation) <= 0 UNION ALL SELECT name, [GeoLocation] FROM [IndiaGeoNames] I1 WHERE I1.[GeoLocation].STDistance(@GeoLocation1) <= 0 UNION ALL SELECT '' name, @GeoLocation2 UNION ALL SELECT '',[Border] FROM [Spatial].[dbo].[Countries] WHERE Countryname = 'India' GO Let us use the distance function of the spatial database and find the straight line distance between this two cities. -- Distance Between Hyderabad and Bangalore DECLARE @GeoLocation GEOGRAPHY SET @GeoLocation = GEOGRAPHY::STPointFromText('POINT(78.4675900000 17.4531200000)',4326) DECLARE @GeoLocation1 GEOGRAPHY SET @GeoLocation1 = GEOGRAPHY::STPointFromText('POINT(77.5833300000 13.0000000000)',4326) SELECT @GeoLocation.STDistance(@GeoLocation1)/1000 'KM'; GO The result of above query is as displayed in following image. As per SQL Server, the distance between these two cities is 501 KM, but according to what I know, the distance between those two cities is around 562 KM by road. However, please note that roads are not straight and they have lots of turns, whereas this is a straight-line distance. What would be more accurate is the distance between these two cities by air travel. When we look at the air travel distance between Bangalore and Hyderabad, the total distance covered is 495 KM, which is very close to what SQL Server has estimated, which is 501 KM. Bravo! SQL Server has accurately provided the distance between two of the cities. SQL Server Spatial Database can be very useful simply because it is very easy to use, as demonstrated above. I appreciate your comments, so let me know what your thoughts and opinions about this are. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Pinal Dave, SQL, SQL Authority, SQL Query, SQL Scripts, SQL Server, SQL Tips and Tricks, T SQL, Technology Tagged: Spatial Database

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  • Kaiden and the Arachnoid Cyst

    - by Martin Hinshelwood
    Some of you may remember when my son Kaiden was born I posted pictures of him and his sister. Kaiden is now 15 months old and is progressing perfectly in every area except that and we had been worried that he was not walking yet. We were only really concerned as his sister was walking at 8 months. Figure: Kai as his usual self   Jadie and I were concerned over that and that he had a rather large head (noggin) so we talked to various GP’s and our health visitor who immediately dismissed our concerns every time. That was until about two months ago when we happened to get a GP whose daughter had Hyper Mobility and she recognised the symptoms immediately. We were referred to the Southbank clinic who were lovely and the paediatrician confirmed that he had Hyper Mobility after testing all of his faculties. This just means that his joints are overly mobile and would need a little physiotherapy to help him out. At the end the paediatrician remarked offhand that he has a rather large head and wanted to measure it. Sure enough he was a good margin above the highest percentile mark for his height and weight. The paediatrician showed the measurements to a paediatric consultant who, as a precautionary measure, referred us for an MRI at Yorkhill Children's hospital. Now, Yorkhill has always been fantastic to us, and this was no exception. You know we have NEVER had a correct diagnosis for the kids (with the exception of the above) from a GP and indeed twice have been proscribed incorrect medication that made the kids sicker! We now always go strait to Yorkhill to save them having to fix GP mistakes as well. Monday 24th May, 7pm The scan went fantastically, with Kaiden sleeping in the MRI machine for all but 5 minutes at the end where he waited patiently for it to finish. We were not expecting anything to be wrong as this was just a precautionary scan to make sure that nothing in his head was affecting his gross motor skills. After the scan we were told to expect a call towards the end of the week… Tuesday 25th May, 12pm The very next day we got a call from Southbank who said that they has found an Arachnoid Cyst and could we come in the next day to see a Consultant and that Kai would need an operation. Wednesday 26th May, 12:30pm We went into the Southbank clinic and spoke to the paediatric consultant who assured us that it was operable but that it was taking up considerable space in Kai’s head. Cerebrospinal fluid is building up as a cyst is blocking the channels it uses to drain. Thankfully they told us that prospects were good and that Kai would expect to make a full recovery before showing us the MRI pictures. Figure: Normal brain MRI cross section. This normal scan shows the spaces in the middle of the brain that contain and produce the Cerebrospinal fluid. Figure: Normal Cerebrospinal Flow This fluid is needed by the brain but is drained in the middle down the spinal column. Figure: Kai’s cyst blocking the four channels. I do not think that I need to explain the difference between the healthy picture and Kai’s picture. However you can see in this first picture the faint outline of the cyst in the middle that is blocking the four channels from draining. After seeing the scans a Neurosurgeon has decided that he is not acute, but needs an operation to unblock the flow. Figure: OMFG! You can see in the second picture the effect of the build up of fluid. If I was not horrified by the first picture I was seriously horrified by this one. What next? Kai is not presenting the symptoms of vomiting or listlessness that would show an immediate problem and as such we will get an appointment to see the Paediatric Neurosurgeon at the Southern General hospital in about 4 weeks. This timescale is based on the Neurosurgeon seeing the scans. After that Kai will need an operation to release the pressure and either remove the cyst completely or put in a permanent shunt (tube from brain to stomach) to bypass the blockage. We have updated his notes for the referral with additional recent information on top of the scan that the consultant things will help improve the timescales, but that is just a guess.   All we can do now is wait and see, and be watchful for tell tail signs of listlessness, eye problems and vomiting that would signify a worsening of his condition.   Technorati Tags: Personal

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  • Connecting to DB2 from SSIS

    - by Christopher House
    The project I'm currently working on involves moving various pieces of data from a legacy DB2 environment to some SQL Server and flat file locations.  Most of the data flows are real time, so they were a natural fit for the client's MQSeries on their iSeries servers and BizTalk to handle the messaging.  Some of the data flows, however, are daily batch type transmissions.  For the daily batch transmissions, it was decided that we'd use SSIS to pull the data direct from DB2 to either a SQL Server or flat file.  I'm not at all an SSIS guy, I've done a bit here and there, but mainly for situations were we needed to move data from a dev environment to QA, mostly informal stuff like that.  And, as much as I'm not an SSIS guy, I'm even less a DB2/iSeries guy.  Prior to this engagement, my knowledge of DB2 was limited to the fact that it's an IBM product and that it was probably a DBMS flatform (that's what the DB in DB2 means, right?).   One of my first goals when I came onto this project was to develop of POC SSIS package to pull some data from DB2 and dump it to a flat file.  It sounded like a pretty straight forward task.  As always, the devil is in the details.  Configuring the DB2 connection manager took a bit of trial and error.  As such, I thought I'd post my experiences here in hopes that they might save someone the efforts I went through.  That being said, please keep in mind, as I pointed out, I'm not at all a DB2 guy, so my terminology and explanations may not be 100% spot on. Before you get started, you need to figure out how you're going to connect to DB2.  From the research I did, it looks like there are a few options.  IBM has both an OLE DB and .Net data provider which can be found here.  I installed their client access tools and tried to use both the .Net and OLE DB providers but I received an error message from both when attempting to connect to the iSeries that indicated I needed a license for a product called DB2 Connect.  I inquired with one of my client's iSeries resources about a license for this product and it appears they didn't have one, so that meant the IBM drivers were out.  The other option that I found quite a bit of discussion around was Microsoft's OLE DB Provider for DB2.  This driver is part of the feature pack for SQL Server 2008 Enterprise Edition and can be downloaded here. As it turns out, I already had Microsoft's driver installed on my dev VM, which stuck me as odd since I hadn't installed it.  I discovered that the driver is installed with the BizTalk adapter pack for host systems, which was also installed on my VM.  However, it looks like the version used by the adapter pack is newer than the version provided in the SQL Server feature pack.   Once you get the driver installed, create a connection manager in your package just like you normally would and select the Microsoft OLE DB Provider for DB2 from the list of available drivers. After you select the driver, you'll need to enter in your host name, login credentials and initial catalog. A couple of things to note here.  First, the Initial catalog needs to be the same as your host name.  Not sure why that is, but trust me, it just does.  Second, for credentials, in my environment, we're using what the client's iSeries people refer to as "profiles".  I guess this is similar to SQL auth in the SQL Server world.  In other words, they've given me a username and password for connecting to DB, so I've entered it here. Next, click the Data Links button.  On the Data Links screen, enter your package collection on the first tab. Package collection is one of those DB2 concepts I'm still trying to figure out.  From the little bit I've read, packages are used to control SQL compilation and each DB2 connection needs one.  The package collection, I believe, controls where your package is created.  One of the iSeries folks I've been working with told me that I should always use QGPL for my package collection, as QGPL is "general purpose" and doesn't require any additional authority. Next click the ellipsis next to the Network drop-down.  Here you'll want to enter your host name again. Again, not sure why you need to do this, but trust me, my connection wouldn't work until I entered my hostname here. Finally, go to the Advanced tab, select your DBMS platform and check Process binary as character. My environment is DB2 on the iSeries and iSeries is the replacement for AS/400, so I selected DB2/AS400 for my platform.  Process binary as character was necessary to handle some of the DB2 data types.  I had a few columns that showed all their data as "System.Byte[]".  Checking Process binary as character resolved this. At this point, you should be good to go.  You can go back to the Connection tab on the Data Links dialog to perform a couple of tests to validate your configuration.  The Test Connection button is obvious, this just verifies you can connect to the host using the configuration data you've entered.  The Packages button will attempt to connect to the host and create the packages required to execute queries. This isn't meant to be a comprehensive look SSIS and DB2, these are just some of the notes I've come up with since I've started working with DB2 and SSIS.  I'm sure as I continue developing my packages, I'll find more quirks and will post them here.

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  • ASP.NET MVC 3: Razor’s @: and <text> syntax

    - by ScottGu
    This is another in a series of posts I’m doing that cover some of the new ASP.NET MVC 3 features: New @model keyword in Razor (Oct 19th) Layouts with Razor (Oct 22nd) Server-Side Comments with Razor (Nov 12th) Razor’s @: and <text> syntax (today) In today’s post I’m going to discuss two useful syntactical features of the new Razor view-engine – the @: and <text> syntax support. Fluid Coding with Razor ASP.NET MVC 3 ships with a new view-engine option called “Razor” (in addition to the existing .aspx view engine).  You can learn more about Razor, why we are introducing it, and the syntax it supports from my Introducing Razor blog post.  Razor minimizes the number of characters and keystrokes required when writing a view template, and enables a fast, fluid coding workflow. Unlike most template syntaxes, you do not need to interrupt your coding to explicitly denote the start and end of server blocks within your HTML. The Razor parser is smart enough to infer this from your code. This enables a compact and expressive syntax which is clean, fast and fun to type. For example, the Razor snippet below can be used to iterate a list of products: When run, it generates output like:   One of the techniques that Razor uses to implicitly identify when a code block ends is to look for tag/element content to denote the beginning of a content region.  For example, in the code snippet above Razor automatically treated the inner <li></li> block within our foreach loop as an HTML content block because it saw the opening <li> tag sequence and knew that it couldn’t be valid C#.  This particular technique – using tags to identify content blocks within code – is one of the key ingredients that makes Razor so clean and productive with scenarios involving HTML creation. Using @: to explicitly indicate the start of content Not all content container blocks start with a tag element tag, though, and there are scenarios where the Razor parser can’t implicitly detect a content block. Razor addresses this by enabling you to explicitly indicate the beginning of a line of content by using the @: character sequence within a code block.  The @: sequence indicates that the line of content that follows should be treated as a content block: As a more practical example, the below snippet demonstrates how we could output a “(Out of Stock!)” message next to our product name if the product is out of stock: Because I am not wrapping the (Out of Stock!) message in an HTML tag element, Razor can’t implicitly determine that the content within the @if block is the start of a content block.  We are using the @: character sequence to explicitly indicate that this line within our code block should be treated as content. Using Code Nuggets within @: content blocks In addition to outputting static content, you can also have code nuggets embedded within a content block that is initiated using a @: character sequence.  For example, we have two @: sequences in the code snippet below: Notice how within the second @: sequence we are emitting the number of units left within the content block (e.g. - “(Only 3 left!”). We are doing this by embedding a @p.UnitsInStock code nugget within the line of content. Multiple Lines of Content Razor makes it easy to have multiple lines of content wrapped in an HTML element.  For example, below the inner content of our @if container is wrapped in an HTML <p> element – which will cause Razor to treat it as content: For scenarios where the multiple lines of content are not wrapped by an outer HTML element, you can use multiple @: sequences: Alternatively, Razor also allows you to use a <text> element to explicitly identify content: The <text> tag is an element that is treated specially by Razor. It causes Razor to interpret the inner contents of the <text> block as content, and to not render the containing <text> tag element (meaning only the inner contents of the <text> element will be rendered – the tag itself will not).  This makes it convenient when you want to render multi-line content blocks that are not wrapped by an HTML element.  The <text> element can also optionally be used to denote single-lines of content, if you prefer it to the more concise @: sequence: The above code will render the same output as the @: version we looked at earlier.  Razor will automatically omit the <text> wrapping element from the output and just render the content within it.  Summary Razor enables a clean and concise templating syntax that enables a very fluid coding workflow.  Razor’s smart detection of <tag> elements to identify the beginning of content regions is one of the reasons that the Razor approach works so well with HTML generation scenarios, and it enables you to avoid having to explicitly mark the beginning/ending of content regions in about 95% of if/else and foreach scenarios. Razor’s @: and <text> syntax can then be used for scenarios where you want to avoid using an HTML element within a code container block, and need to more explicitly denote a content region. Hope this helps, Scott P.S. In addition to blogging, I am also now using Twitter for quick updates and to share links. Follow me at: twitter.com/scottgu

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  • Chester Devs Presentation and source code &ndash; &lsquo;Event Store - an introduction to a DSD for event sourcing and notifications&rsquo;

    - by Liam Westley
    Originally posted on: http://geekswithblogs.net/twickers/archive/2013/11/11/chester-devs-presentation-and-source-code-ndash-lsquoevent-store.aspxThank you everyone at Chester Devs Thanks to Fran Hoey and all the people from Chester Devs. It was a hard drive up and back but the enthusiasm of the audience, with some great questions does make it worthwhile. Presentation and source code My presentation, source code, Event Store runners and text files containing the various command line parameters used for curl is now available on GitHub; https://github.com/westleyl/ChesterDevs-EventStore. Don’t worry if you don’t have a GitHub account, you don’t need one, you can just click on the Download Zip button on the right hand menu to download all the files as a single ZIP file.  If all you want is the PowerPoint presentation, go to https://github.com/westleyl/ChesterDevs-EventStore/blob/master/Powerpoint/Huddle-EventStore.pptx, and click on the View Raw button. Downloading and installing Event Store and Tools Download Event Store http://download.geteventstore.com – I unzipped these files into C:\EventStore\v2.0.1 Download Curl from http://curl.haxx.se/download.html – I downloaded Win64 Generic (with SSL) and unzipped these files into C:\curl version 7.31.0 Running the tools I used in my presentation Demonstration 1 (running Event Store) You can use one of my Event Store runner command files to run the single node version of Event Store, using default ports of 2213 for HTTP and 1113  for TCP, and with a wildcard HTTP pattern.  Both take a single command line parameter to specify the location of the data and log files.  The runners assume the single node executable is located in C:\EventStore\v2.0.1, and will placed data files and logs beneath C:\EventStore\Data, i.e. RunEventStore.cmd TestData1 This will create data files in C:\EventStore\Data\TestData1\Data and log files in C:\EventStore\Data\TestData1\logs. If, when running Event Store you may see the following message, [03288,15,06:23:00.622] Failed to start http server Access is denied You will either need to run Event Store in an administrator console window, or you can use the netsh command to create a firewall permission to allow HTTP listening (this will need to be run, once, in an administrator console window), netsh http add urlacl url=http://*:2213/ user=liam You can always delete this later by running the delete; netsh http delete urlacl url=http://*:2213/ If you want to confirm that everything is running OK, open the management console in a browser by navigating to http://127.0.0.1:2213. If at any point you are asked for a user name and password use the default of ‘admin’/‘changeit’. Demonstration 2 (reading and adding data, curl) In my second demonstration I used curl directly from the console to read streams, write events and then read back those events. On GitHub I have included is a set of curl commands, CurlCommandLine.txt, and a sample data file, SampleData.json, to load an event into a DDDNorth3 stream. As there is not much data in the Event Store at this point I used the $stats-127.0.0.1:2113 which is a stream containing performance statistics for Event Store and is updated every 30 seconds (default). Demonstration 3 (projections) On GitHub I have included a sample projection, Projection-ByRoom.txt, which will create streams based on the room on which a session was held on the DDDNorth3 agenda. Browse to the management console, http://127.0.0.1:2213.  Click on Projections, New Projection, give it a name, Sessions-ByRoom, and copy in the JavaScript in the Projection-ByRoom.txt file.  Select Continuous, tick Emit Enabled and then click on Post. It should run immediately. You may by challenged for the administration login for the management console, if so use the default user name and password; 'admin'/'changeit'. Demonstration 4 (C# client) The final demonstration was the Visual Studio 2012 project using the Event Store client – referenced directly as C:\EventStore\v2.0.1\EventStore.ClientAPI.dll, although you can switch this to the latest Event Store client NuGet package. The source code provides a console app for viewing projections with the projection manager (HTTP connection), as well as containing a full set of data for the entire DDDNorth3 agenda.  It also deals with the strategy for reading newest events backwards to older events and ignoring older events that have been superseded. Resources Event Store home page: http://www.geteventstore.com/ Event Store source code on GitHub: https://github.com/eventstore/eventstore Event Store documentation on GitHub: https://github.com/eventstore/eventstore/wiki (includes index to @RobAshton’s blog series on Event Store at https://github.com/eventstore/eventstore/wiki#rob-ashton---projections-series) Event Store forum in Google Groups: https://groups.google.com/forum/?fromgroups#!forum/event-store TopShelf Windows service wrapper is available on github: https://gist.github.com/trbngr/5083266

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  • Install SharePoint 2013 on a two server farm

    - by sreejukg
    When SharePoint 2010 was released, I published an article on how to install SharePoint on a two server farm. You can find that article from the below link. http://weblogs.asp.net/sreejukg/archive/2010/09/28/install-sharepoint-2010-in-a-farm-environment.aspx Now it is the time for SharePoint 2013. SharePoint 2013 brings lots of improvements to the topologies, but still supports two-server architecture. Be noted that “two-server architecture” is meant for small implementations with limited service applications. Refer the below link to understand more about the SharePoint architecture http://technet.microsoft.com/en-us/sharepoint/fp123594.aspx A two tier farm consists of a database server and a web/application server as follows. In this article I am going to explain how to install SharePoint in a two server farm. I prepared 2 servers, both of them joined to a domain(SP2013Domain), and in one server I installed SQL Server 2012 (Server name: SP2013_DB). Now I am going to install SharePoint 2013 in the second server (Server Name: SP2013). The following domain accounts are created for the installation.   User Account Purpose Server roles required SQLService - SQL Server service account - This account is used as the service account for SQL Server. - domain user account / local account spSetup - You will be running SharePoint setup and SharePoint products and configuration wizard using this account. -domain user account - Member of the Administrators group on each server on which Setup is run(In our case SP2013) - SQL Server login on the computer running SQL Server - Member of the Server admin SQL Server security role spDataaccess - Configure and manage server farm. This - Application pool identity for central admin website - Microsoft SharePoint Foundation Workflow Timer Service Domain user account (Other permissions will be set to this account automatically)   The above are the minimum list of accounts needed for SharePoint 2013 installation. Now you need additional accounts for services, application pool identities for web applications etc. Refer the service accounts requirements for SharePoint from the below link. http://technet.microsoft.com/en-us/library/cc263445.aspx In order to install SharePoint 2013 login to the server using setup account(spsetup). Now run the setup from the installation media. First you need to install the pre-requisites. During the installation process, the server may restart several times. The installation wizard will guide you through the installation. In the next step, you need to agree on the terms and conditions as usual. Once you click next, the installation will start immediately. The installation wizard will let you know the progress of the installation. During the installation you may receive notifications to restart the server, you need to just click the finish button so that the system will be restarted. Once all the pre-requisites are installed, you will get the success message as below. Click finish to close the dialog. Now from the media, run the setup again and this time you choose install SharePoint server. In the next screen, you need to enter the product key, and then click continue. Now you need to agree on the terms and conditions for SharePoint 2013, and click continue. Choose the file location as per your policies and click on the install now button. You will see the installation progress. Once completed, you will see the installation completed dialog. Make sure you select the run products and configuration wizard option and click close. From the start screen, click next to start the configuration wizard. You will receive warning telling you some of the services will be stopped during the installation. Select “create new server farm” radio button and click next. In the next step, you need to enter the configuration database settings. Enter the database server details and then specify the database access account. You need to specify the farm account(spdataaccess). The wizard will grant additional privileges to the account as needed. In the next step you need to specify the passphrase, you need to note this as you need this passphrase if you add additional server to the farm. In the next step, you need to enter the central administration website port and security settings. You can choose a port or just keep it as suggested by the wizard. Click next, you will see the summary of what you have been selected. Verify the selected settings and if you want to change any, just click back and change them, or click continue to start the configuration. The configuration may take some time, you can view the progress, in case of any error, you will get the log file, you need to fix any error and again start the configuration wizard. Once the configuration successful, you will see the success message. Just click finish. Now you can browse the central administration website. It is good to check the health analyzer to review whether there are any errors/warnings. No warnings/errors indicate a good installation. Two-Server architecture is the least configuration for production environments. For small firms with less number of employees can implement SharePoint 2013 using this topology and as the workload increases, they can add more servers to the farm without reconstructing everything.

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  • It’s nice to be important, but it’s more important to be nice

    - by BuckWoody
    I’ve been a little “preachy” lately, telling you that you should let people finish their sentences, and always check a problem out before you tell a user that their issue is “impossible”. Well, I’ll round that out with one more tip today. Keep in mind that all of these things are actions I’ve been guilty of, hopefully in the past. I’m kind of a “work in progress”. And yes, I know these tips are coming from someone who picks on people in presentations, but that is of course done in fun, and (hopefully) with the audience’s knowledge.   (No, this isn’t aimed at any one person or event in particular – I just see it happen a lot)   I’ve seen, unfortunately over and over, someone in authority react badly to someone who is incorrect, or at least perceived to be incorrect. This might manifest itself in a comment, post, question or whatever, but the point is that I’ve seen really intelligent people literally attack someone they view as getting something wrong. Don’t misunderstand me; if someone posts that you should always drop a production database in the middle of the day I think you should certainly speak up and mention that this might be a bad idea!  No, I’m talking about generalizations or even incorrect statements done in good faith. Let me explain with an example.   Suppose someone makes the statement: “If you don’t have enough space on your system, you can just use a DBCC command to shrink the database”. Let’s take two responses to this statement.   Response One: “That’s insane. Everyone knows that shrinking a database is a stupid idea, you’re just going to fragment your indexes all over the place.” Response Two: “That’s an interesting take – in my experience and from what I’ve read here (someurl.com) I think this might not be a universal best practice.”   Of course, both responses let the person making the statement and those reading it know that you don’t agree, and that it’s probably wrong. But the person you responded to and the general audience hearing you (or reading your response) might form two different opinions of you.   The first response says to me “this person really needs to be right, and takes arguments personally. They aren’t thinking of the other person at all, or the folks reading or hearing the exchange. They turned an incorrect technical statement into a personal attack. They haven’t left the other party any room to ‘save face’, and they have potentially turned what could be a positive learning experience for everyone into a negative. Also, they sound more than just a little arrogant.”   The second response says to me “this person has left room for everyone to save face, has presented evidence to the contrary and is thinking about moving the ball forward and getting it right rather than attacking someone for getting it wrong.” It’s the idea of questioning a statement rather than attacking a person.   Perhaps you have a different take. Maybe you think the “direct” approach is best – and maybe that’s worked for you. Something to consider is what you’ve really accomplished while using that first method. Sure, the info you provide is correct, and perhaps someone out there won’t shrink a database because of your response – but perhaps you’ve turned a lot more people off, and now they won’t listen to your other valuable information. You’ll be an expert, but another one of the nameless, arrogant jerks in technology. And I don’t think anyone likes to be thought of that way.   OK, I’ll get down off of the high-horse now. And I’ll keep the title of this entry (said to me by my grandmother when I was a little kid) in mind when I dismount. Share this post: email it! | bookmark it! | digg it! | reddit! | kick it! | live it!

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