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  • Scaling a video processing application on EC2?

    - by Stpn
    I am approaching the need to scale a video-processign application that runs on EC2. So far the setup is one machine: Backbonejs frontend Rails 3.2 Postgresql Resque + S3 for storage The flow of the app is as follows: 1) Request from frontend. Upload a video. 2) Storing video 3) Quering external APIs. 4) Processing / encoding videos. 5) Post to frontend. I can separate the backend and frontend without any problems, but when it comes to distributing the backend between several servers I am a bit puzzled. I can probably come up with a temporary solution (like just duplicating apps making several instances), but since I don't really have expertise in backend system administration, there can be some fundamental mistakes.. Also I would rather have something that is scalable. I wonder if anyone can give some feedback on the following plan: A) Frontend machine. Just frontend, talks to backend via REST Api of sorts. B) Backend server (BS), main database. Gets request from 1), posts to 2) saves uploads to 3) C) S3 storage. D) Server for quering APIs. Basically just a Resque workers, that post info back to 2) E) Server for video encoding. Processes videos uploaded on 3) and uploads them back. So I will have: A)frontend \ \ B)MAIN_APP/DB ----- C)S3 Storage (Files) / \ / / \ / D)ExternalAPI_queries E)Video_Processing (redundant DB) (redundant DB) All this will supposedly talk to each other via HTTP requests. My reason for this is that Video Processing part is really the most resource-intensive and I would just run barebones application that accepts requests and starts processing them. Questions: 1) In this setup I will have the main database at B) and all other servers will communicate with it via HTTP requests (and store duplicates of databases also I guess..for safety reasons). Is it the right approach or should I have 1 database that everyone connects to (how then?) 2) Is it a good idea to separate API queries from Video Processing part? Logically they are very close (processing is determined by the result of API queries), but resource-wise Video Processing is waaay more intensive. 3) what should I use to distribute calls between backend apps based on load?

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  • BizTalk 2009 - Custom Functoid Wizard

    - by StuartBrierley
    When creating BizTalk maps you may find that there are times when you need perform tasks that the standard functoids do not cover.  At other times you may find yourself reapeating a pattern of standard functoids over and over again, adding visual complexity to an otherwise simple process.  In these cases you may find it preferable to create your own custom functoids.  In the past I have created a number of custom functoids from scratch, but recently I decided to try out the Custom Functoid Wizard for BizTalk 2009. After downloading and installing the wizard you should start Visual Studio and select to create a new BizTalk Server Functoid Project. Following the splash screen you will be presented with the General Properties screen, where you can set the classname, namespace, assembly name and strong name key file. The next screen is the first set of properties for the functoid.  First of all is the fuctoid ID; this must be a value above 6000. You should also then set the name, tooltip and description of the functoid.  The name will appear in the visual studio toolbox and the tooltip on hover over in the toolbox.  The descrition will be shown when you configure the functoid inputs when using it in a map; as such it should provide a decent level of information to allow the functoid to be used. Next you must set the category, exception mesage, icon and implementation language.  The category will affect the positioning of the functoid within the toolbox and also some of the behaviours of the functoid. We must then define the parameters and connections for our new functoid.  Here you can define the names and types of your input parameters along with the minimum and maximum number of input connections.  You will also need to define the types of connections accepted and the output type of the functoid. Finally you can click finish and your custom functoid project will be created. The results of this process can be seen in the solution explorer, where you will see that a project, functoid class file and a resource file have been created for you. If you open the class file you will see that the following code has been created for you: The "base" function sets all the properties that you previsouly detailed in the custom functoid wizard.  public TestFunctoids():base()  {    int functoidID;    // This has to be a number greater than 6000    functoidID = System.Convert.ToInt32(resmgr.GetString("FunctoidId"));    this.ID = functoidID;    // Set Resource strings, bitmaps    SetupResourceAssembly(ResourceName, Assembly.GetExecutingAssembly());    SetName("FunctoidName");                     SetTooltip("FunctoidToolTip");    SetDescription("FunctoidDescription");    SetBitmap("FunctoidBitmap");    // Minimum and maximum parameters that the functoid accepts    this.SetMinParams(2);    this.SetMaxParams(2);    /// Function name that needs to be called when this Functoid is invoked.    /// Put this in GAC.    SetExternalFunctionName(GetType().Assembly.FullName,     "MyCompany.BizTalk.Functoids.TestFuntoids.TestFunctoids", "Execute");    // Category for this functoid.    this.Category = FunctoidCategory.String;    // Input and output Connection type    this.OutputConnectionType = ConnectionType.AllExceptRecord;    AddInputConnectionType(ConnectionType.AllExceptRecord);   } The "Execute" function provides a skeleton function that contains the code to be executed by your new functoid.  The inputs and outputs should match those you defined in the Custom Functoid Wizard.   public System.Int32 Execute(System.Int32 Cool)   {    ResourceManager resmgr = new ResourceManager(ResourceName, Assembly.GetExecutingAssembly());    try    {     // TODO: Implement Functoid Logic    }    catch (Exception e)    {     throw new Exception(resmgr.GetString("FunctoidException"), e);    }   } Opening the resource file you will see some of the various string values that you defined in the Custom Functoid Wizard - Name, Tooltip, Description and Exception. You can also select to look at the image resources.  This will display the embedded icon image for the functoid.  To change this right click the icon and select "Import from File". Once you have completed the skeleton code you can then look at trying out your functoid. To do this you will need to build the project, copy the compiled DLL to C:\Program Files\Microsoft BizTalk Server 2009\Developer Tools\Mapper Extensions and then refresh the toolbox in visual studio.

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  • Linux software Raid 10 no superblock

    - by Shoshomiga
    I have a software raid 10 with 6 x 2tb hard drives (raid 1 for /boot), ubuntu 10.04 is the os. I had a raid controller failure that put 2 drives out of sync, crashed the system and initially the os didnt boot up and went into initramfs instead, saying that drives were busy but I eventually managed to bring the raid up by stopping and assembling the drives. The os booted up and said that there were filesystem errors, I chose to ignore because it would remount the fs in read-only mode if there was a problem. Everything seemed to be working fine and the 2 drives started to rebuild, I was sure that it was a sata controller failure because I had dma errors in my log files. The os crashed soon after that with ext errors. Now its not bringing up the raid, it says that there is no superblock on /dev/sda2. I tried to reassemble manually with all the device names but it still would not bring up the raid 10 complaining about the missing superblock on sda2, and sda1 was also dropped from the raid 1. When I did examine on the raid10 it says that 1 of the initially failed drives is a spare, the other is spare rebuilding and sda2 is removed. It seems that sda decided to fail right when the system was vulnerable to it because when I boot up a live cd it spews out sda unrecoverable read failures. I have been trying to fix this all week but I'm not sure where to go with this now, I ordered more hard drives because I didn't have a complete backup, but its too late for that now and the only thing I could do is mirror all the hard drives onto the new ones (I'm not sure whether sda was mirrored without errors). On the internet I read that you can recover from this by recreating the array with the same options as when it was made, however because sda is failing I cant use it and I don't want to risk using its mirror instead, so I'm waiting to get another hard drive. I'm also not sure whether to include the out of sync drives or if I can actually use those instead to recover the array. Sorry if this is a mess to read but I've been trying to fix this all day and its late at night now, any thoughts on this would be greatly appreciated. I also did a memtest and changed the motherboard in addition to everything else. EDIT: This is my partition layout Disk /dev/sdb: 2000.4 GB, 2000398934016 bytes 255 heads, 63 sectors/track, 243201 cylinders, total 3907029168 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 4096 bytes I/O size (minimum/optimal): 4096 bytes / 4096 bytes Disk identifier: 0x0009c34a Device Boot Start End Blocks Id System /dev/sdb1 * 2048 511999 254976 83 Linux /dev/sdb2 512000 3904980991 1952234496 83 Linux /dev/sdb3 3904980992 3907028991 1024000 82 Linux swap / Solaris

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  • Thinkpad speaker turns mute - Linux Codec issue?

    - by Curlew
    At some point a few days ago the speakers on my Lenovo Thinkpad T410 (Model number: 2537A11) suddenly stopped working randomly. This error happens every time I watch a video or listen to a music file. The sound just abruptly stops. At the moment, I can't produce a single sound no matter what I do. I am using Debian GNU/Linux on this laptop and there doesn't appear to be anything else wrong (the fan is working, no abnormal heat (staying around ~40°C), no other obvious errors or problems). Here is the output of a nice program someone pointed me to: martin@martin:~/Downloads$ sudo python run.py --monitor Using temporary directory: /dev/shm/hda-analyzer You may remove this directory when finished or if you like to download the most recent copy of hda-analyzer tool. Downloading file hda_analyzer.py Downloading file hda_guilib.py Downloading file hda_codec.py Downloading file hda_proc.py Downloading file hda_graph.py Downloading file hda_mixer.py Downloaded all files, executing hda_analyzer.py Watching 1 cards ====================================== Sound is working normally and then it stops and the following lines appear: Diff for codec 0/0 (0x14f15069): --- +++ @@ -164,17 +164,17 @@ Power: setting=D0, actual=D0 Node 0x1f [Pin Complex] wcaps 0x400501: Stereo Pincap 0x00000010: OUT Pin Default 0x901701f0: [Fixed] Speaker at Int N/A Conn = Analog, Color = Unknown DefAssociation = 0xf, Sequence = 0x0 Misc = NO_PRESENCE Pin-ctls: 0x40: OUT - Power: setting=D0, actual=D0 + Power: setting=D3, actual=D3 Connection: 2 0x10* 0x11 Node 0x20 [Pin Complex] wcaps 0x400781: Stereo Digital Pincap 0x00000010: OUT Pin Default 0x40f001f0: [N/A] Other at Ext N/A Conn = Unknown, Color = Unknown DefAssociation = 0xf, Sequence = 0x0 Misc = NO_PRESENCE And now there is also an error in the dmesg output hda-intel: IRQ timing workaround is activated for card #0. Suggest a bigger bdl_pos_adj. I changed the bdl_pos_adj to various numbers (-1, 0, 64, 1024) and either there is no change at all or dmesg reports that the adjustment is too big. I wonder if this bdl_pos_adj is the real reason for the error. Here is my hardware information provided by alsa-info.sh website. Okay, i did some serious testing and even installed Windows and now i officially conclude that this is a hard-ware related issue with my Laptop speakers. Reason: The error occurs in my installed Debian Linux, an Ubuntu Live distribution and Windows XP No error-message appears in all of the OS. The sound just keeps running and i can't hear a thing. I tested different setups, including OSS, ALSA and the pulseaudio server on top If i use my new usb-headphones i can hear sound all the time without any sudden silences. So obviously, although hard to believe, my laptop speakers are not okay (never heard of similar cases). I'll award the bounty to anyone who can point me to good tutorials or the procedure how to exchange my T410 speakers (i still have warranty. The laptop was bought in Germany, but now i am in Denmark). Or to someone who can explain me the output from hda-analyzer (big log above).

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  • Business Case for investing time developing Stubs and BizUnit Tests

    - by charlie.mott
    I was recently in a position where I had to justify why effort should be spent developing Stubbed Integration Tests for BizTalk solutions. These tests are usually developed using the BizUnit framework. I assumed that most seasoned BizTalk developers would consider this best practice. Even though Microsoft suggest use of BizUnit on MSDN, I've not found a single site listing the justifications for investing time writing stubs and BizUnit tests. Stubs Stubs should be developed to isolate your development team from external dependencies. This is described by Michael Stephenson here. Failing to do this can result in the following problems: In contract-first scenarios, the external system interface will have been defined.  But the interface may not have been setup or even developed yet for the BizTalk developers to work with. By the time you open the target location to see the data BizTalk has sent, it may have been swept away. If you are relying on the UI of the target system to see the data BizTalk has sent, what do you do if it fails to arrive? It may take time for the data to be processed or it may be scheduled to be processed later. Learning how to use the source\target systems and investigations into where things go wrong in these systems will slow down the BizTalk development effort. By the time the data is visible in a UI it may have undergone further transformations. In larger development teams working together, do you all use the same source and target instances. How do you know which data was created by whose tests? How do you know which event log error message are whose?  Another developer may have “cleaned up” your data. It is harder to write BizUnit tests that clean up the data\logs after each test run. What if your B2B partners' source or target system cannot support the sort of testing you want to do. They may not even have a development or test instance that you can work with. Their single test instance may be used by the SIT\UAT teams. There may be licencing costs of setting up an instances of the external system. The stubs I like to use are generic stubs that can accept\return any message type.  Usually I need to create one per protocol. They should be driven by BizUnit steps to: validates the data received; and select a response messages (or error response). Once built, they can be re-used for many integration tests and from project to project. I’m not saying that developers should never test against a real instance.  Every so often, you still need to connect to real developer or test instances of the source and target endpoints\services. The interface developers may ask you to send them some data to see if everything still works.  Or you might want some messages sent to BizTalk to get confidence that everything still works beyond BizTalk. Tests Automated “Stubbed Integration Tests” are usually built using the BizUnit framework. These facilitate testing of the entire integration process from source stub to target stub. It will ensure that all of the BizTalk components are configured together correctly to meet all the requirements. More fine grained unit testing of individual BizTalk components is still encouraged.  But BizUnit provides much the easiest way to test some components types (e.g. Orchestrations). Using BizUnit with the Behaviour Driven Development approach described by Mike Stephenson delivers the following benefits: source: http://biztalkbddsample.codeplex.com – Video 1. Requirements can be easily defined using Given/When/Then Requirements are close to the code so easier to manage as features and scenarios Requirements are defined in domain language The feature files can be used as part of the documentation The documentation is accurate to the build of code and can be published with a release The scenarios are effective to document the scenarios and are not over excessive The scenarios are maintained with the code There’s an abstraction between the intention and implementation of tests making them easier to understand The requirements drive the testing These same tests can also be used to drive load testing as described here. If you don't do this ... If you don't follow the above “Stubbed Integration Tests” approach, the developer will need to manually trigger the tests. This has the following risks: Developers are unlikely to check all the scenarios each time and all the expected conditions each time. After the developer leaves, these manual test steps may be lost. What test scenarios are there?  What test messages did they use for each scenario? There is no mechanism to prove adequate test coverage. A test team may attempt to automate integration test scenarios in a test environment through the triggering of tests from a source system UI. If this is a replacement for BizUnit tests, then this carries the following risks: It moves the tests downstream, so problems will be found later in the process. Testers may not check all the expected conditions within the BizTalk infrastructure such as: event logs, suspended messages, etc. These automated tests may also get in the way of manual tests run on these environments.

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  • How to Remove Extensions From, and Force the Trailing Slash at the End of URLs?

    - by Kronbernkzion
    Example of current file structure: example.com/foo.php example.com/bar.html example.com/directory/ example.com/directory/foo.php example.com/directory/bar.html example.com/cgi-bin/directory/foo.cgi I want to remove HTML, PHP and CGI extensions from, and then force the trailing slash at the end of URLs. So, it could look like this: example.com/foo/ example.com/bar/ example.com/directory/ example.com/directory/foo/ example.com/directory/bar/ example.com/cgi-bin/directory/foo/ I've searched for solution for 17 hours straight and visited more than a few hundred pages on various blogs and forums. I'm not joking. So I think I've done my research. Here is the code that sits in my .htaccess file right now: RewriteCond %{REQUEST_FILENAME} !-d RewriteCond %{REQUEST_FILENAME}\.html -f RewriteRule ^(([^/]+/)*[^./]+)/$ $1.html RewriteCond %{REQUEST_FILENAME} !-f RewriteCond %{REQUEST_FILENAME} !-d RewriteCond %{REQUEST_URI} !(\.[a-zA-Z0-9]|/)$ RewriteRule (.*)$ /$1/ [R=301,L] As you can see, this code only removes .html (and I'm not very happy with it because I think it could be done a lot simpler). I can remove the extension from PHP files when I rename them to .html through .htaccess, but that's not what I want. I want to remove it straight. This is the first thing I don't know how to do. The second thing is actually very annoying. My .htaccess file with code above, adds .html/ to every string entered after example.com/directory/foo/. So if I enter example.com/directory/foo/bar (obviously /bar doesn't exist since foo is a file), instead of just displaying message that page is not found, it converts it to example.com/directory/foo/bar.html/, then searches for a file for a few seconds and then displays the not found message. This, of course, is bad behavior. So, once again, I need the code in .htaccess to do the following things: Remove .html extension Remove .php extension Remove .cgi extension Force the trailing slash at the end of URLs Requests should behave correctly (no adding trailing slashes or extensions to strings if file or directory doesn't exist on server) Code should be as simple as possible I would very much appreciate any help. And to first person that gives me the solution, I'll send two $50 iTunes Store gift cards for US store. If this offend anyone, I am truly sorry and I apologize. Thanks in advance. And sorry for such a long post.

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  • How to setup GIT repo on server with need for working dir (non- bare)

    - by OrangeTux
    I want to have configurate a GIT repo for a website. Multiple users will have a clone of the repo on their local machine and on the end of each day they push their work to the server. I can setup a bare repo, but I want a working dir/non-bare repository. The idea is that the working dir of the repository will the root folder for the website. At the end of each day all changes will be visible directly. But I can't find a way to do this. Initializing the server repo with git init gives the following error when a client is trying to push some files: git push origin master [email protected]'s password: Counting objects: 3, done. Writing objects: 100% (3/3), 227 bytes, done. Total 3 (delta 0), reused 0 (delta 0) remote: error: refusing to update checked out branch: refs/heads/master remote: error: By default, updating the current branch in a non-bare repository remote: error: is denied, because it will make the index and work tree inconsistent remote: error: with what you pushed, and will require 'git reset --hard' to match remote: error: the work tree to HEAD. remote: error: remote: error: You can set 'receive.denyCurrentBranch' configuration variable to remote: error: 'ignore' or 'warn' in the remote repository to allow pushing into remote: error: its current branch; however, this is not recommended unless you remote: error: arranged to update its work tree to match what you pushed in some remote: error: other way. remote: error: remote: error: To squelch this message and still keep the default behaviour, set remote: error: 'receive.denyCurrentBranch' configuration variable to 'refuse'. To ssh://[email protected]/home/orangetux/www/ ! [remote rejected] master -> master (branch is currently checked out) error: failed to push some refs to 'ssh://[email protected]/home/orangetux/www/' So I'm wondering if this the right way to setup a GIT repo for a website? If so, how do I have to do this? If not, what is a better way to setup a GIT repo for the development of a website? EDIT you can't push to a non-bare repository Oke, clear. But whats the way to solve my problem? Create a bare repository on the server and have a clone of this repo on the same server in the htdocs folder? This looks a bit clumsy to me. To see the result of a commit I've to clone the repository each time.

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  • amazon ec2 ubuntu with gitlab and nginx - cant load?

    - by thebluefox
    Ok, so I've spooled up an Amazon EC2 server running Ubuntu, and then followed the instructions below to install GitLab; http://doc.gitlab.com/ce/install/installation.html The only step I've not been able to complete is running the following check on the status; sudo -u git -H bundle exec rake gitlab:check RAILS_ENV=production I get the following error; rake aborted! Errno::ENOMEM: Cannot allocate memory - whoami Which I presume is becuase my EC2 is just running a free tier setup, so isn't that well spec'd. Regardless, I've been trying to access this through my browser. I've set up the elastic IP and pointed my domain at it (for the purpose of this, lets say its git.mydom.co.uk). Doing a whois on this domain shows me its pointing to the right place. For some reason though, I get the "Oops, Chrome could not connect to git.mydom.co.uk". Now - for a period of time I was getting the Nginx holding page (telling me I still needed to perform configuration). This though disappeared after removing the default file from /etc/nginx/sites-enabled/ (after reading this could be issue on a troubleshooting page). Since then, I've had nothing, even when I symlinked the file back in from /sites-available. I've tried changing the owner of the git.mydom.co.uk file sat inside /sites-enabled and /sites-available to www-data, as suggested here, but I could only change the permission of the file in /sites-available, and not the symlinked one in /sites-enabled. The content of this file is as follows; upstream gitlab { server unix:/home/git/gitlab/tmp/sockets/gitlab.socket; } server { listen *:80 default_server; # e.g., listen 192.168.1.1:80; In most cases *:80 is a good idea server_name git.mydom.co.uk; # e.g., server_name source.example.com; server_tokens off; # don't show the version number, a security best practice root /home/git/gitlab/public; # Increase this if you want to upload large attachments # Or if you want to accept large git objects over http client_max_body_size 20m; # individual nginx logs for this gitlab vhost access_log /var/log/nginx/gitlab_access.log; error_log /var/log/nginx/gitlab_error.log; location / { # serve static files from defined root folder;. # @gitlab is a named location for the upstream fallback, see below try_files $uri $uri/index.html $uri.html @gitlab; } All the paths mentioned in here look ok...I'm about at the end of my knowledge now!

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  • NFS4 / ZFS: revert ACL to clean/inherited state

    - by Keiichi
    My problem is identical to this Windows question, but pertains NFS4 (Linux) and the underlying ZFS (OpenIndiana) we are using. We have this ZFS shared via NFS4 and CIFS for Linux and Windows users respectively. It would be nice for both user groups to benefit from ACLs, but the one missing puzzle piece goes thusly: Each user has a home, where he sets a top-level, inherited ACL. He can later on refine permissions for the contained files/folders iteratively. Over time, sometimes permissions need to be generalized again to avoid increasing pollution of ACL entries. You can tweak the ACL of every single file if need be to obtain the wanted permissions, but that defeats the purpose of inherited ACLs. So, how can an ACL be completely cleared like in the question linked above? I have found nothing about what a blank, inherited ACL should look like. This usecase simply does not seem to exist. In fact, the solaris chmod manpage clearly states A- Removes all ACEs for current ACL on file and replaces current ACL with new ACL that represents only the current mode of the file. I.e. we get three new ACL entries filled with stuff representing the permission bits, which is rather useless for cleaning up. If I try to manually remove every ACE, on the last one I get chmod A0- <file> chmod: ERROR: Can't remove all ACL entries from a file Which by the way makes me think: and why not? In fact, I really want the whole file-specific ACL gone. The same holds for linux, which enumerates ACEs starting with 1(!), and verbalizes its woes less diligently nfs4_setacl -x 1 <file> Failed setxattr operation: Unknown error 524 So, what is the idea behind ACLs under Solaris/NFS? Can they never be cleaned up? Why does the recursion option for the ACL setting commands pollute all children instead of setting a single ACL and making the children inherit? Is this really the intention of the designers? I can clean up the ACLs using a windows client perfectly well, but am I supposed to tell the linux users they have to switch OS just to consolidate permissions?

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  • Common reasons for the &lsquo;Sys is undefined&rsquo; error in ASP.NET Ajax applications

      In this blog I will try to summarize the most common reasons for getting the famous 'Sys is undefined' error when running an Ajax enabled web site or application (there are almost one milion results on Google for that phrase). Where does it come from? In every Ajax web pages source you will see a code like this: <script type="text/javascript"> //<![CDATA[ Sys.WebForms.PageRequestManager._initialize('ScriptManager1', document.getElementById('form1')); Sys.WebForms.PageRequestManager.getInstance()._updateControls([], [], [], 90); //]]> </script>   This is the initialization script of the ScriptManager. So, if for some reason the Sys namespace is not available when the code executes you get the Sys is undefined error. Here are the most common reasons and solutions for that problem:   1. The error occurs when you have added a control from RadControls for ASP.NET AJAX, but your application is not configured to use ASP.NET AJAX. For example, in VS 2005 you created a new Blank Site instead of a new Ajax-Enabled Web Site and the Sys is undefined message pops up. To fix it you need to follow the steps described at Configuring ASP.NET Ajax article (check the topic called Adding ASP.NET AJAX Configuration Elements to an Existing Web Site) or simply create the Ajax-Enabled Web Site. You can also check my other blog post on the matter: Visual Studio 2008: Where is the new ASP.NET Ajax-Enabled Web Site template?   2. Authentication - as the website denies access to all pages to unauthorized users, access to the Telerik.Web.UI.WebResource.axd handler is unauthorized (this is the default handler of RadScriptManager). This causes the handler to serve the content of the login page instead of the combined scripts, hence the error. To solve it - add a <location> section to the application configuration file to allow access to Telerik.Web.UI.WebResource.axd to all users, like: <configuration> ... <location path="Telerik.Web.UI.WebResource.axd"> <system.web> <authorization> <allow users="*"/> </authorization> </system.web> </location> ... </configuration>   Note that the access to the standard ScriptResource.axd and WebResource.axd is automatically allowed for all users (authenticated and unauthenticated), so if you use the ScriptManager instead of RadScriptManager - you will not face this problem. The authentication problem does not manifest when you disable script combining or use the CDN. Adding the above configuration section will make it work with RadScriptManagers combined script.   3. The IE6 browser fails to load the compressed script. The problem does not appear in any other browser. There is a well known bug in the older versions of IE6 which lose the first 2,048 bytes of data that are sent back from a Web server that uses HTTP compression. Latest versions of RadScriptManager does not compress the output at all if the client is IE6, but in the previous versions you need to manually disable the output compression to prevent the error. So, if you get the Sys is undefined error in IE6 - update to the latest version of RadControls or simply disable the output compression.   4. Requests to the *.axd files returns Error Code 404 - Not Found. This could  be fixed easily: Check in the IIS management console that the .axd extension (the default HTTP handler extension) is allowed:     Also check if the Verify if file exists checkbox is unchecked (click on the Edit button appearing in the previous screenshot to check). More information can be found in our troubleshooting article and from the ASP.NET QA team blog post   5. The virtual directory in IIS is not marked as Web Application. Converting it to Web Application should fix the problem.   6. Check for the <xhtmlConformance mode="Legacy"/> option in your web.config and remove it. It would be rather rare to become a victim of this exact case, but still have it in mind. Scott Guthrie describes it in more details   In the above points I mentioned several times the terms web resources, javascript output, compressed script. If you want to find out more about these please see the Web Resources Demystified series of my friend and colleague Atanas Korchev   I hope that one of the above solutions will help you get rid of the Sys is undefined error.   Did you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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  • How do I (robustly) remotely execute tasks on Windows workstations in a domain?

    - by Zac B
    I'm not even sure if "robustly" is a word. Anyway. Context: We have a few hundred Windows 7 workstations on a LAN. We use AD/GPO management pretty heavily, but there are a lot of periodic and/or manual maintenance tasks we need to do that can't be done via GPO/scheduled task. For example, say I want to execute program X (which runs silently, in the background, and doesn't bother the user) on workstation Y, or say I want to execute task A on a workstation group B either on a schedule or on demand. Kicking the users off of their computers to do this (i.e. using RDP) is a no-no, and doesn't work on groups anyway. Question: What's the best way to do this that is robust enough that, after setup, I could give it to beginner support people (read: people who are phobic of the command line, and get confused with GUI interfaces more complicated than Firefox)? I'm a competent programmer, and, if there is a robust set of tools or framework out there for this type of task, I'd consider hacking something together myself if it didn't take too long. If there's some combination of tools or techniques that others use to make remote-workstation-administration doable by beginners, I have yet to find it. For those who care about the "why": I'm midlevel IT, and was told to implement a remote management solution that allows arbitrary/scheduled remote execution, with confirmation that programs actually ran remotely, and the ability to view what they returned. "Why?" I asked, "Can't I just use PsExec and the task scheduler on a dispatcher machine?" "No," I was told, "'Joe' the second-week tech is going to be in charge of this one, and he needs something simple with a GUI." What I've tried: I've played with making a bunch of one-clickable "transfer files to remote computer and run them with PsExec" batch/VB scrips, but those tend to break down and don't easily support running on customizable groups. I've played a little bit with the Windows version of Puppet, but it doesn't support arbitrary-time remote execution (it's ability to group computers into a tree/node structure is really nice though). I've used an older version of Altiris, and, while it does a lot of what I want, it's interface is awful, it's slow, crashes a lot, and is probably too expensive for management. SwiftWater's DMS solution does some of what I want, but it's very underdeveloped, closed-source (not a deal breaker but not ideal), and I get the impression that support and reliability are lacking.

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  • When upgrading from Vista to Windows 7 on a DELL laptop, how do I know which drivers to reinstall an

    - by msorens
    According to Dell's upgrade page for Vista to Windows 7, after using the upgrade assistant the final step is to install drivers. They refer to this page for the order of driver installation, listing 9 items. From there I go to the Dell Drivers and Downloads page, enter my system tag, and get a list of the downloads available for my specific box. That page, by the way, has a link to driver install instructions that lists 10 rather than 9 items. Going to Drivers Help in the side panel and clicking on "In what order should drivers be installed?" shows yet a third list, this one containing 13 items. Not surprisingly, the order of these 3 lists of drivers are not quite the same for the common items! Furthermore, of the 26 files Dell's site recommends for my machine, there are several not shown on any of the 3 lists! I can make determinations for some of these: 6 of them are "applications" so I know which of those I want and that they could probably be safely installed after all drivers. BIOS: I would think this should be unaffected by an OS upgrade so could be skipped. Two tools in the diagnostics category: could probably be done after all drivers. That leaves just a CD/DVD driver and a webcam driver unaccounted for. So my two related questions are these: How critical is the driver installation order and which one do I follow? (Keep in mind this is for an upgrade, not a fresh install.) Where in the order do I insert the CD/DVD and the webcam drivers (if needed) ? Dell's driver download page provides (in theory) the list of all downloads relevant to my specific machine, via the service tag. But do I actually need to reinstall all of them? some? none? How does one determine this? They do label each with Recommended or Optional, so do I need to reinstall all the recommended ones? (Part of the reason for my perplexed frown is that I wonder why I would need to reinstall a CD/DVD driver since I would already be using the drive to install the OS!)

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  • Software Raid 10 corrupted superblock after dual disk failure, how do I recover it?

    - by Shoshomiga
    I have a software raid 10 with 6 x 2tb hard drives (raid 1 for /boot), ubuntu 10.04 is the os. I had a raid controller failure that put 2 drives out of sync, crashed the system and initially the os didnt boot up and went into initramfs instead, saying that drives were busy but I eventually managed to bring the raid up by stopping and assembling the drives. The os booted up and said that there were filesystem errors, I chose to ignore because it would remount the fs in read-only mode if there was a problem. Everything seemed to be working fine and the 2 drives started to rebuild, I was sure that it was a sata controller failure because I had dma errors in my log files. The os crashed soon after that with ext errors. Now its not bringing up the raid, it says that there is no superblock on /dev/sda2, even if I assemble manually with all the device names. I also did a memtest and changed the motherboard in addition to everything else. EDIT: This is my partition layout Disk /dev/sdb: 2000.4 GB, 2000398934016 bytes 255 heads, 63 sectors/track, 243201 cylinders, total 3907029168 sectors Units = sectors of 1 * 512 = 512 bytes Sector size (logical/physical): 512 bytes / 4096 bytes I/O size (minimum/optimal): 4096 bytes / 4096 bytes Disk identifier: 0x0009c34a Device Boot Start End Blocks Id System /dev/sdb1 * 2048 511999 254976 83 Linux /dev/sdb2 512000 3904980991 1952234496 83 Linux /dev/sdb3 3904980992 3907028991 1024000 82 Linux swap / Solaris All 6 disks have the same layout, partition #1 is for raid 1 /boot, partition #2 is for raid 10 far plan, partition #3 is swap, but sda did not have swap enabled EDIT2: This is the output of mdadm --detail /dev/md1 Layout : near=1, far=2 Chunk Size : 64k UUID : a0feff55:2018f8ff:e368bf24:bd0fce41 Events : 0.3112126 Number Major Minor RaidDevice State 0 8 34 0 spare rebuilding /dev/sdc2 1 0 0 1 removed 2 8 18 2 active sync /dev/sdb2 3 8 50 3 active sync /dev/sdd2 4 0 0 4 removed 5 8 82 5 active sync /dev/sdf2 6 8 66 - spare /dev/sde2 EDIT3: I ran ddrescue and it has copied everything from sda except a single 4096 byte sector that I suspect is the raid superblock EDIT4: Here is some more info too long to fit here lshw: http://pastebin.com/2eKrh7nF mdadm --detail /dev/sd[abcdef]1 (raid1): http://pastebin.com/cgMQWerS mdadm --detail /dev/sd[abcdef]2 (raid10): http://pastebin.com/V5dtcGPF dumpe2fs of /dev/sda2 (from the ddrescue cloned drive): http://pastebin.com/sp0GYcJG I tried to recreate md1 based on this info with the command mdadm --create /dev/md1 -v --assume-clean --level=10 --raid-devices=6 --chunk=64K --layout=f2 /dev/sda2 missing /dev/sdc2 /dev/sdd2 missing /dev/sdf2 But I can't mount it, I also tried to recreate it based on my initial mdadm --detail /dev/md1 but it still doesn't mount It also warns me that /dev/sda2 is an ext2fs file system but I guess its because of ddrescue

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  • How can the Private Bytes of a process be significantly less than its effect on the system commit charge?

    - by bacar
    On a 64-bit Windows Server 2003, I can see using taskmgr or process explorer that the total commit charge is around 3.5GB, yet when I sum the Private Bytes consumed by each process (by running pslist -m and adding all values under the Priv column) the total comes in at 1.6GB. I know which process seems to be causing this (sqlservr.exe) as when I kill the process, the commit charge drops dramatically. However the process in question is consuming only ~220MB of Private Bytes yet killing the process drops the commit charge by ~1.6GB. How is this possible? How can the commit charge be so significantly greater than Private Bytes, which should represent the amount of committed memory? If some other factor contributes to the commit charge, what is that factor and how can I view its impact in process explorer? Note: I claim that I understand the difference between reserved and committed memory already: my investigations above relate specifically to Private Bytes which includes only committed memory and excludes reserved memory. the Virtual Size of the process in this case is over 4GB, but this should be irrelevant - Virtual Size in procexp represents reserved, not committed memory, and should not contribute to the commit charge. I'm particularly interested in generalised answers to this question: I'm assuming that if sqlservr.exe can behave in this way, that any process potentially could. Further Investigations I notice that pointing Sysinternals VMMap at this process reports a committed "Private Data" of 1.6GB despite Procexp's reported a Private Bytes of 220MB. This is particularly strange given that the documentation for this field in the "Windows® Sysinternals Administrator's Reference" states that: Private Data memory is memory that is allocated by VirtualAlloc and that is not further handled by the Heap Manager or the .NET runtime, or assigned to the Stack category... VMMap’s definition of “Private Data” is more granular than that of Process Explorer’s “private bytes.” Procexp’s “private bytes” includes all private committed memory belonging to the process. i.e. that VMMap's committed "Private Data" should be smaller than procexp's "Private Bytes". Also, after reading the 'Process committed memory' section of Mark Russinovich's excellent Pushing the Limits of Windows: Virtual Memory, he highlights two cases which won't show up in Private Bytes: File mapping views with copy-on-write semantics (however, according to VMMap there is no significant space allocated to Mapped Files). pagefile-backed virtual memory (however, I tried testlimit with the -l flag as suggested, and no significant memory is consumed by pagefile-backed sections)

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  • Creating a really public Windows network share

    - by Timur Aydin
    I want to create a shared folder under Windows (actually, Windows XP, Vista, and Win 7) which can be mounted from a linux system without prompting for a username/password. But before attempting this, I first wanted to establish that this works between two Windows 7 machines. So, on machine A (The server that will hold the public share), I created a folder and set its permissions such that Everyone has read/write access. Then I visited Control Panel - Network and Sharing Center - Advanced Sharing Settings and then selected "Turn off password protected sharing". Then, on machine B (The client that wants to access the public share with no username/password prompt), I tried to "map network driver" and I was immediately prompted by a password prompt. Some search on google suggested changing "Acconts: Limit local account use of blank passwords to console logon only" to "Disabled". Tried that, no luck, still getting username/password prompt. If I enter the username/password, I am not prompted for it again and can use the share as long as the session is active. But still, I really need to access the share without any username/password transaction whatsoever and this is not just a convenience related thing. Here is the actual reason: The device that will access this windows network share is an embedded system running uclinux. It will mount this share locally and then play media files. Its only user interface is a javascript based web page. So, if there is going to be any username/password transaction, I would have to ask the user to enter them over the web page, which will be ridiculously insecure and completely exposed to packet sniffing. After hours of doing experiments, I have found one way to make this happen, but I am not really very fond of it... I first create a new user (shareuser) and give it a password (sharepass). Then I open Group Policy Editor and set "Deny log on locally" to "A\shareuser". Then, I create a folder on A and share it so that shareuser has Read access to it. This way, shareuser cannot login to A, but can access the shared folder. And, if someone discovers the shareuser/sharepass through network sniffing, they can just access the shared folder, but can't logon to A. The same thing can be achieved by enabling the Guest user and then going to Group Policy Editor and deleting the "Guest" from the "Deny access to this computer from the network" setting. Again, Guest can mount the public share, but logging in to A as Guest won't be possible, because Guest is already not allowed to log in by default. So my question would be, how can I create a network share that is truly public, so that it can be mounted from a linux machine without requiring a password? Sorry for the long question, but I wanted to explain the reason for really needing this...

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  • Localhost has just stopped working (using xampp)

    - by Joe Taylor
    I installed Xampp to use for local development of a Drupal site. Its been working fine out of the box until now. The main Xampp localhost welcome menu loads, however my subdirectory (localhost/drupal) doesn't. It just spins in the browser for ages and nothing happens. Just a blank screen. I've tried the edit people suggest in the hosts file but that hasn't work and I'm getting no errors so not sure what to do. Anyone have any ideas what might be wrong? PS I'm running Windows 7 edit: Log files: Fatal error: Allowed memory size of 134217728 bytes exhausted (tried to allocate 123731968 bytes) in C:\xampp\apps\drupal\htdocs\sites\all\themes\directory\node--job.tpl.php on line 41 Fatal error: Allowed memory size of 134217728 bytes exhausted (tried to allocate 123731968 bytes) in C:\xampp\apps\drupal\htdocs\sites\all\themes\directory\node--job.tpl.php on line 41 [Tue Nov 05 20:52:07.242454 2013] [ssl:warn] [pid 8432:tid 260] AH01909: RSA certificate configured for www.example.com:443 does NOT include an ID which matches the server name [Tue Nov 05 20:52:07.331459 2013] [core:warn] [pid 8432:tid 260] AH00098: pid file C:/xampp/apache/logs/httpd.pid overwritten -- Unclean shutdown of previous Apache run? [Tue Nov 05 20:52:07.820487 2013] [ssl:warn] [pid 8432:tid 260] AH01909: RSA certificate configured for www.example.com:443 does NOT include an ID which matches the server name [Tue Nov 05 20:52:07.898492 2013] [mpm_winnt:notice] [pid 8432:tid 260] AH00455: Apache/2.4.4 (Win32) OpenSSL/0.9.8y PHP/5.4.16 configured -- resuming normal operations [Tue Nov 05 20:52:07.898492 2013] [mpm_winnt:notice] [pid 8432:tid 260] AH00456: Server built: Feb 23 2013 13:07:34 [Tue Nov 05 20:52:07.898492 2013] [core:notice] [pid 8432:tid 260] AH00094: Command line: 'c:\xampp\apache\bin\httpd.exe -d C:/xampp/apache' [Tue Nov 05 20:52:07.905492 2013] [mpm_winnt:notice] [pid 8432:tid 260] AH00418: Parent: Created child process 7588 [Tue Nov 05 20:52:08.882548 2013] [ssl:warn] [pid 7588:tid 272] AH01909: RSA certificate configured for www.example.com:443 does NOT include an ID which matches the server name [Tue Nov 05 20:52:09.467582 2013] [ssl:warn] [pid 7588:tid 272] AH01909: RSA certificate configured for www.example.com:443 does NOT include an ID which matches the server name [Tue Nov 05 20:52:09.534585 2013] [mpm_winnt:notice] [pid 7588:tid 272] AH00354: Child: Starting 150 worker threads. Fatal error: Allowed memory size of 134217728 bytes exhausted (tried to allocate 123731968 bytes) in C:\xampp\apps\drupal\htdocs\sites\all\themes\directory\node--job.tpl.php on line 41 Fatal error: Allowed memory size of 134217728 bytes exhausted (tried to allocate 123731968 bytes) in C:\xampp\apps\drupal\htdocs\sites\all\themes\directory\node--job.tpl.php on line 41

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  • System failure - need diagnostic recommendation

    - by Ladislav Mrnka
    I have big problem with my computer. Configuration is: Intel i7 + 6x2GB OCZ DDR3 Motheboard: Asus P6T Deluxe V2, HDD controller configured to AHCI Main drive: OCZ Vertex 2 (SSD) - contains all installed programs and system Second drive: Samsung SpinPoint - contains User profiles, ProgramData, virtual machines and databases Third drive: Samsung SpinPoint - data drive + backups OS: Windows 7 Ultimate x64 I have never had any problem with this computer until now. During weekend my computer completely crashed without any reason. Each time I tried to boot to Windows I got BSOD with message BAD_SYSTEM_CONFIG_INFO and automatic restart (I didn't install any new SW or HW). But after restart main OCZ drive was disconnected (not detected by BIOS). When I turned off and on computer, the drive was again connected. It also happend every single time I tried to repair installation somehow. It ended with some error and after restart drive was disconnected. The only thing which worked was format + fresh install. After installing almost everything I wanted to install Visual Studio 2010 Ultimate (complete installation without SQL Server Express). During installation of VS itself I always get BSOD - it is too fast so I'm not able to read description. After restart it searches for all disk drives for really long time and sometimes it changes boot drive so the system is not able to start - Bootmgr not found. After reconfiguring BIOS the system starts. There is no event describing the failure in Event viewer. Installing VS 2010 is absolutely necessary for me. I need help with diagnostic. I need to find where is the problem - I expect that the problem is in OCZ drive or in HDD controller on motherboard but I don't know how to find it. All components still have valid warranty. Can you recommend me some approach or tools to find the problem? Edit: I'm still looking for source of the problem. New information is that Windows are not able to perform check disk (Chkdsk) on the SSD system drive. After restarting it always starts checking drive and in part where files are checked it fails with BSOD - BAD_SYSTEM_CONFIG_INFO. After next restart and skipping check disk tests the system runs.

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  • Add Machine Key to machine.config in Load Balancing environment to multiple versions of .net framework

    - by davidb
    I have two web servers behind a F5 load balancer. Each web server has identical applications to the other. There was no issue until the config of the load balancer changed from source address persistence to least connections. Now in some applications I receieve this error Server Error in '/' Application. Validation of viewstate MAC failed. If this application is hosted by a Web Farm or cluster, ensure that configuration specifies the same validationKey and validation algorithm. AutoGenerate cannot be used in a cluster. Description: An unhandled exception occurred during the execution of the current web request. Please review the stack trace for more information about the error and where it originated in the code. Exception Details: System.Web.HttpException: Validation of viewstate MAC failed. If this application is hosted by a Web Farm or cluster, ensure that configuration specifies the same validationKey and validation algorithm. AutoGenerate cannot be used in a cluster. Source Error: The source code that generated this unhandled exception can only be shown when compiled in debug mode. To enable this, please follow one of the below steps, then request the URL: Add a "Debug=true" directive at the top of the file that generated the error. Example: or: 2) Add the following section to the configuration file of your application: Note that this second technique will cause all files within a given application to be compiled in debug mode. The first technique will cause only that particular file to be compiled in debug mode. Important: Running applications in debug mode does incur a memory/performance overhead. You should make sure that an application has debugging disabled before deploying into production scenario. How do I add a machine key to the machine.config file? Do I do it at server level in IIS or at website/application level for each site? Does the validation and decryption keys have to be the same across both web servers or are they different? Should they be different for each machine.config version of .net? I cannot find any documentation of this scenario.

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  • Migrating R Scripts from Development to Production

    - by Mark Hornick
    Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 “How do I move my R scripts stored in one database instance to another? I have my development/test system and want to migrate to production.” Users of Oracle R Enterprise Embedded R Execution will often store their R scripts in the R Script Repository in Oracle Database, especially when using the ORE SQL API. From previous blog posts, you may recall that Embedded R Execution enables running R scripts managed by Oracle Database using both R and SQL interfaces. In ORE 1.3.1., the SQL API requires scripts to be stored in the database and referenced by name in SQL queries. The SQL API enables seamless integration with database-based applications and ease of production deployment. Loading R scripts in the repository Before talking about migration, we’ll first introduce how users store R scripts in Oracle Database. Users can add R scripts to the repository in R using the function ore.scriptCreate, or SQL using the function sys.rqScriptCreate. For the sample R script     id <- 1:10     plot(1:100,rnorm(100),pch=21,bg="red",cex =2)     data.frame(id=id, val=id / 100) users wrap this in a function and store it in the R Script Repository with a name. In R, this looks like ore.scriptCreate("RandomRedDots", function () { line-height: 115%; font-family: "Courier New";">     id <- 1:10     plot(1:100,rnorm(100),pch=21,bg="red",cex =2)     data.frame(id=id, val=id / 100)) }) In SQL, this looks like begin sys.rqScriptCreate('RandomRedDots',  'function(){     id <- 1:10     plot(1:100,rnorm(100),pch=21,bg="red",cex =2)     data.frame(id=id, val=id / 100)   }'); end; / The R function ore.scriptDrop and SQL function sys.rqScriptDrop can be used to drop these scripts as well. Note that the system will give an error if the script name already exists. Accessing R scripts once they’ve been loaded If you’re not using a source code control system, it is possible that your R scripts can be misplaced or files modified, making what is stored in Oracle Database to only or best copy of your R code. If you’ve loaded your R scripts to the database, it is straightforward to access these scripts from the database table SYS.RQ_SCRIPTS. For example, select * from sys.rq_scripts where name='myScriptName'; From R, scripts in the repository can be loaded into the R client engine using a function similar to the following: ore.scriptLoad <- function(name) { query <- paste("select script from sys.rq_scripts where name='",name,"'",sep="") str.f <- OREbase:::.ore.dbGetQuery(query) assign(name,eval(parse(text = str.f)),pos=1) } ore.scriptLoad("myFunctionName") This function is also useful if you want to load an existing R script from the repository into another R script in the repository – think modular coding style. Just include this function in the body of the other function and load the named script. Migrating R scripts from one database instance to another To move a set of functions from one system to another, the following script loads the functions from one R script repository into the client R engine, then connects to the target database and creates the scripts there with the same names. scriptNames <- OREbase:::.ore.dbGetQuery("select name from sys.rq_scripts where name not like 'RQG$%' and name not like 'RQ$%'")$NAME for(s in scriptNames) { cat(s,"\n") ore.scriptLoad(s) } ore.disconnect() ore.connect("rquser","orcl","localhost","rquser") for(s in scriptNames) { cat(s,"\n") ore.scriptDrop(s) ore.scriptCreate(s,get(s)) } Best Practice When naming R scripts, keep in mind that the name can be up to 128 characters. As such, consider organizing scripts in a directory structure manner. For example, if an organization has multiple groups or applications sharing the same database and there are multiple components, use “/” to facilitate the function organization: line-height: 115%;">ore.scriptCreate("/org1/app1/component1/myFuntion1", myFunction1) ore.scriptCreate("/org1/app1/component1/myFuntion2", myFunction2) ore.scriptCreate("/org1/app2/component2/myFuntion2", myFunction2) ore.scriptCreate("/org2/app2/component1/myFuntion3", myFunction3) ore.scriptCreate("/org3/app2/component1/myFuntion4", myFunction4) Users can then query for all functions using the path prefix when looking up functions. /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;}

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  • Error in eclipse on run android project

    - by Larz
    I am trying to get a simple hello world android project working in eclipse using an android emulator. I have been using the examples on developer.android.com. I actually did have a hello world app working. I then modified it's xml files to have a text input field and a button as in the second example shows on that site. This failed to run on the emulator. I then went back and tried to create another simple hello world project, but it fails to run. The console says "Waiting for HOME ('android.process.acore') to be launched, but nothing happens or sometimes a messenger in the emulator says "unfortunately Android Wear has stopped". Below is a sample error filter on the log file. I find trying to debug this is something new to me and I am not sure the best way to go about it. I am just trying to learn some basic android developer skills. 05-30 16:19:07.336: E/SELinux(469): SELinux: Loaded file_contexts from /file_contexts, 05-30 16:19:07.336: E/SELinux(469): digest= 05-30 16:19:07.376: E/SELinux(469): b0 05-30 16:19:07.376: E/SELinux(469): 4b 05-30 16:19:07.756: E/SELinux(469): 03 05-30 16:19:07.756: E/SELinux(469): 4a 05-30 16:19:07.826: E/SELinux(469): 73 05-30 16:19:07.886: E/SELinux(469): ab 05-30 16:19:07.886: E/SELinux(469): 6d 05-30 16:19:07.896: E/SELinux(469): 46 05-30 16:19:07.896: E/SELinux(469): b4 05-30 16:19:07.896: E/SELinux(469): a5 05-30 16:19:07.896: E/SELinux(469): 73 05-30 16:19:07.896: E/SELinux(469): 8a 05-30 16:19:07.896: E/SELinux(469): ee 05-30 16:19:07.896: E/SELinux(469): ac 05-30 16:19:07.906: E/SELinux(469): 68 05-30 16:19:07.906: E/SELinux(469): ff 05-30 16:19:07.906: E/SELinux(469): 04 05-30 16:19:07.906: E/SELinux(469): dc 05-30 16:19:07.906: E/SELinux(469): b8 05-30 16:19:07.906: E/SELinux(469): a2 05-30 16:19:11.806: E/SensorManager(511): sensor or listener is null 05-30 16:19:16.196: E/BluetoothAdapter(378): Bluetooth binder is null 05-30 16:19:16.206: E/BluetoothAdapter(378): Bluetooth binder is null 05-30 16:19:17.186: E/WVMExtractor(54): Failed to open libwvm.so: dlopen failed: library "libwvm.so" not found 05-30 16:19:17.776: E/AudioCache(54): Error 1, -2147483648 occurred 05-30 16:19:17.796: E/SoundPool(378): Unable to load sample: (null) 05-30 16:19:18.536: E/AudioCache(54): Error 1, -2147483648 occurred 05-30 16:19:18.546: E/SoundPool(378): Unable to load sample: (null)

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  • Tip #19 Module Private Visibility in OSGi

    - by ByronNevins
    I hate public and protected methods and classes.  It requires so much work to change them in a huge project like GlassFish.  Not to mention that you may well have to support those APIs forever.  They are highly overused in GlassFish.  In fact I'd bet that > 95% of classes are marked as public for no good reason.  It's just (bad) habit is my guess. private and default visibility (I call it package-private) is easier to maintain.  It is much much easier to change such classes and methods around.  If you have ANY public method or public class in GlassFish you'll need to grep through a tremendous amount of source code to find all callers.  But even that won't be theoretically reliable.  What if a caller is using reflection to access public methods?  You may never find such usages. If you have package private methods, it's easy.  Simply grep through all the code in that one package.  As long as that package compiles ok you're all set.  There can' be any compile errors anywhere else.  It's a waste of time to even look around or build the "outside" world.  So you may be thinking: "Aha!  I'll just make my module have one giant package with all the java files.  Then I can use the default visibility and maintenance will be much easier.  But there's a problem.  You are wasting a very nice feature of java -- organizing code into separate packages.  It also makes the code much more encapsulated.  Unfortunately to share code between the packages you have no choice but to declare public visibility. What happens in practice is that a module ends up having tons of public classes and methods that are used exclusively inside the module.  Which finally brings me to the point of this blog:  If Only There Was A Module-Private Visibility Available Well, surprise!  There is such a mechanism.  If your project is running under OSGi that is.  Like GlassFish does!  With this mechanism you can easily add another level of visibility by telling OSGi exactly which public you want to be exposed outside of the module.  You get the best of both worlds: Better encapsulation of your code so that maintenance is easier and productivity is increased. Usage of public visibility inside the module so that you can encapsulate intra-module better with packages. How I do this in GlassFish: Carefully plan out at least one package that will contain "true" publics.  This is the package that will be exported by OSGi.  I recommend just one package. Here is how to tell OSGi to use it in GlassFish -- edit osgi.bundle like so:-exportcontents:     org.glassfish.mymodule.truepublics;  version=${project.osgi.version} Now all publics declared in any other packages will be visible module-wide but not outside the module. There is one caveat: Accessing "module-private" items outside of the module is controlled at run-time, not compile-time.  The compiler has no clue that a public in a dependent module isn't really public.  it will happily compile it.  At runtime you will definitely see fireworks.  The good news is that you don't have to wait for the code path that tries to use the "module-private" items to fire.  OSGi will complain loudly when that module gets loaded.  OSGi will refuse to load it.  You will see an error like this: remote failure: Error while loading FOO: Exception while adding the new configuration : Error occurred during deployment: Exception while loading the app : org.osgi.framework.BundleException: Unresolved constraint in bundle com.oracle.glassfish.miscreant.code [115]: Unable to resolve 115.0: missing requirement [115.0] osgi.wiring.package; (osgi.wiring.package=org.glassfish.mymodule.unexported). Please see server.log for more details. That is if you accidentally change code in module B to use a public that is really a "module-private" in module A, then you will see the error immediately when you try to test whatever you were changing in module B.

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  • SSH Connection Refused - Debug using Recovery Console

    - by olrehm
    Hey everyone, I have found a ton of questions answered about debugging why one cannot connect via SSH, but they all seem to require that you can still access the system - or say that without that nothing can be done. In my case, I cannot access the system directly, but I do have access to the filesystem using a recovery console. So this is the situation: My provider made some kernel update today and in the process also rebooted my server. For some reason, I cannot connect via SSH anymore, but instead get a ssh: connect to host mydomain.de port 22: Connection refused I do not know whether sshd is just not running, or whether something (e.g. iptables) blocks my ssh connection attempts. I looked at the logfiles, none of the files in /var/log contain any mentioning on ssh, and /var/log/auth.log is empty. Before the kernel update, I could log in just fine and used certificates so that I would not need a password everytime I connect from my local machine. What I tried so far: I looked in /etc/rc*.d/ for a link to the /etc/init.d/ssh script and found none. So I am expecting that sshd is not started properly on boot. Since I cannot run any programs in my system, I cannot use update-rc to change this. I tried to make a link manually using ln -s /etc/init.d/ssh /etc/rc6.d/K09sshd and restarted the server - this did not fix the problem. I do not know wether it is at all possible to do it like this and whether it is correct to create it in rc6.d and whether the K09 is correct. I just copied that from apache. I also tried to change my /etc/iptables.rules file to allow everything: # Generated by iptables-save v1.4.0 on Thu Dec 10 18:05:32 2009 *mangle :PREROUTING ACCEPT [7468813:1758703692] :INPUT ACCEPT [7468810:1758703548] :FORWARD ACCEPT [3:144] :OUTPUT ACCEPT [7935930:3682829426] :POSTROUTING ACCEPT [7935933:3682829570] COMMIT # Completed on Thu Dec 10 18:05:32 2009 # Generated by iptables-save v1.4.0 on Thu Dec 10 18:05:32 2009 *filter :INPUT ACCEPT [7339662:1665166559] :FORWARD ACCEPT [3:144] :OUTPUT ACCEPT [7935930:3682829426] -A INPUT -i lo -j ACCEPT -A INPUT -p tcp -m tcp --dport 25 -j ACCEPT -A INPUT -p tcp -m tcp --dport 993 -j ACCEPT -A INPUT -p tcp -m tcp --dport 22 -j ACCEPT -A INPUT -p tcp -m tcp --dport 143 -j ACCEPT -A INPUT -m conntrack --ctstate RELATED,ESTABLISHED -j ACCEPT -A INPUT -p tcp -m tcp --dport 80 -j ACCEPT -A INPUT -p tcp --dport 8080 -s localhost -j ACCEPT -A INPUT -m limit --limit 5/min -j LOG --log-prefix "iptables denied: " --log-level 7 -A INPUT -j ACCEPT -A FORWARD -j ACCEPT -A OUTPUT -j ACCEPT COMMIT # Completed on Thu Dec 10 18:05:32 2009 # Generated by iptables-save v1.4.0 on Thu Dec 10 18:05:32 2009 *nat :PREROUTING ACCEPT [101662:5379853] :POSTROUTING ACCEPT [393275:25394346] :OUTPUT ACCEPT [393273:25394250] COMMIT # Completed on Thu Dec 10 18:05:32 2009 I am not sure this is done correctly or has any effect at all. I also did not find any mentioning of iptables in any file in /var/log. So what else can I do? Thank you for your help.

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  • Apache returns 403 Forbidden for alternative port vhost

    - by Wesley
    I'm having an issue getting vhosts to work on Apache 2.2, Debian 6. I have two VirtualHosts, one on port 80 and one on port 8888. The port 80 one has been created automatically by DirectAdmin, the 8888 is a custom one. It's configuration is as follows. <VirtualHost *:8888 > DocumentRoot /home/user/public_html/development ServerName www.myserver.nl ServerAlias myserver.nl <Directory "/home/user/public_html/development"> Options +Indexes +FollowSymLinks +MultiViews AllowOverride All Order Allow,deny Allow from all </Directory> </VirtualHost> Of course I also have a NameVirtualHost *:8888 The port 80 DocumentRoot is /home/user/public_html/production, which is perfectly accessible and works like a charm. The port 8888 docroot of /home/user/public_html/development is 403 forbidden though. I have compared the permissions for both folders. They seem fine to me. drwxr-xr-x 2 root root 4096 Aug 17 16:14 development drwxr-xr-x 4 root root 4096 Aug 18 04:29 production Also, the index.php file which is supposed to display when accessing through port 8888, located in /development/: -rwxr-xr-x 1 root root 41 Aug 17 16:14 index.html I have looked at my error_log and found many of the following entries, only being added to the log file when accessing through port 8888. [Sat Aug 18 04:35:09 2012] [error] [client 27.32.156.232] Symbolic link not allowed or link target not accessible: /home/user/public_html /home/user/public_html is a symbolic link that refers to /home/user/domains/mydomain/public_html. The symbolic link has the following permissions: lrwxrwxrwx 1 admin admin 29 Aug 17 15:56 public_html -> ./domains/mydomain/public_html I'm at a loss. It seems that everything is readable or executable. I've set the Directory to FollowSymLinks in the httpd.conf file, but that doesn't seem to make a difference. If I change that directory tag to <Directory "/home/admin/public_html"> (so it has FollowSymLinks on that as well) it still does not work. Any help is greatly appreciated. If I need to post more information, let me know. I'm pretty much a beginner at this stuff. .. .. UPDATE: I ended up changing the configuration to directly go to the actual path of the files, avoiding the public_html symlink altogether. That worked. Thanks for the suggestions folks. DocumentRoot /home/user/domains/mydomain/public_html/development instead of DocumentRoot /home/user/public_html/development

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  • Email client wont connect to SMTP Authentication server

    - by Jason
    Im having trouble installing SMTH Auth for my ubuntu email server. I have followed ubuntu own guide for SMTH AUT (https://help.ubuntu.com/14.04/serverguide/postfix.html). But my email client thunderbird is giving this error " lost connection to SMTP-client 127.0.0.1." I cant add new users to thundbird either because of this connection problem. Do i have to alter any setting on my Thunderbird perhaps since ? I did try to make thunderbird use SSL for imap as well but that neither works. I restarted postfix and dovecot to find errors but both run just fine. Prior to SMTP auth changes thunderbird could connect just fine to my server and send mails. This is my main.cf file in postfix. It looks just like the one on ubuntu guide above. readme_directory = no # TLS parameters #smtpd_use_tls=yes smtpd_tls_session_cache_database = btree:${data_directory}/smtpd_scache smtp_tls_session_cache_database = btree:${data_directory}/smtp_scache myhostname = mail.mysite.com mydomain = mysite.com alias_maps = hash:/etc/aliases alias_database = hash:/etc/aliases myorigin = $mydomain mydestination = mysite.com #relayhost = smtp.192.168.10.1.com mynetworks = 127.0.0.0/8 [::ffff:127.0.0.0]/104 [::1]/128 192.168.10.0/24 mailbox_size_limit = 0 recipient_delimiter = + inet_interfaces = all home_mailbox = Maildir/ mailbox_command = #SMTP AUTH smtpd_sasl_type = dovecot smtpd_recipient_restrictions=permit_mynetworks, permit_sasl_authenticated,reject_unauth_destination smtpd_sasl_local_domain = smtpd_sasl_auth_enable = yes smtpd_sasl_security_options = noanonymous broken_sasl_auth_clients = yes smtpd_tls_auth_only = no smtp_tls_security_level = may smtpd_tls_security_level = may smtp_tls_note_starttls_offer = yes smtpd_tls_key_file = /etc/ssl/private/smtpd.key smtpd_tls_cert_file = /etc/ssl/certs/smtpd.crt smtpd_tls_CAfile = /etc/ssl/certs/cacert.pem smtpd_tls_loglevel = 1 smtpd_tls_received_header = yes This my dovecot configuration at 10-master.conf service imap-login { inet_listener imap { #port = 143 } inet_listener imaps { #port = 993 #ssl = yes } # Number of connections to handle before starting a new process. Typically # the only useful values are 0 (unlimited) or 1. 1 is more secure, but 0 # is faster. <doc/wiki/LoginProcess.txt> #service_count = 1 # Number of processes to always keep waiting for more connections. #process_min_avail = 0 # If you set service_count=0, you probably need to grow this. #vsz_limit = $default_vsz_limit } service pop3-login { inet_listener pop3 { #port = 110 } inet_listener pop3s { #port = 995 #ssl = yes } } service lmtp { unix_listener lmtp { #mode = 0666 } # Create inet listener only if you can't use the above UNIX socket #inet_listener lmtp { # Avoid making LMTP visible for the entire internet #address = #port = #} } service imap { # Most of the memory goes to mmap()ing files. You may need to increase this # limit if you have huge mailboxes. #vsz_limit = $default_vsz_limit # Max. number of IMAP processes (connections) #process_limit = 1024 } service pop3 { # Max. number of POP3 processes (connections) #process_limit = 1024 } service auth { unix_listener auth-userdb { #mode = 0600 #user = #group = } # Postfix smtp-auth unix_listener /var/spool/postfix/private/auth { mode = 0660 user = postfix } } service dict { # If dict proxy is used, mail processes should have access to its socket. # For example: mode=0660, group=vmail and global mail_access_groups=vmail unix_listener dict { #mode = 0600 #user = #group = } } I did add auth_mechanisms = plain login to 10-auth.conf as well.

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  • How to make creating viewmodels at runtime less painfull

    - by Mr Happy
    I apologize for the long question, it reads a bit as a rant, but I promise it's not! I've summarized my question(s) below In the MVC world, things are straightforward. The Model has state, the View shows the Model, and the Controller does stuff to/with the Model (basically), a controller has no state. To do stuff the Controller has some dependencies on web services, repository, the lot. When you instantiate a controller you care about supplying those dependencies, nothing else. When you execute an action (method on Controller), you use those dependencies to retrieve or update the Model or calling some other domain service. If there's any context, say like some user wants to see the details of a particular item, you pass the Id of that item as parameter to the Action. Nowhere in the Controller is there any reference to any state. So far so good. Enter MVVM. I love WPF, I love data binding. I love frameworks that make data binding to ViewModels even easier (using Caliburn Micro a.t.m.). I feel things are less straightforward in this world though. Let's do the exercise again: the Model has state, the View shows the ViewModel, and the ViewModel does stuff to/with the Model (basically), a ViewModel does have state! (to clarify; maybe it delegates all the properties to one or more Models, but that means it must have a reference to the model one way or another, which is state in itself) To do stuff the ViewModel has some dependencies on web services, repository, the lot. When you instantiate a ViewModel you care about supplying those dependencies, but also the state. And this, ladies and gentlemen, annoys me to no end. Whenever you need to instantiate a ProductDetailsViewModel from the ProductSearchViewModel (from which you called the ProductSearchWebService which in turn returned IEnumerable<ProductDTO>, everybody still with me?), you can do one of these things: call new ProductDetailsViewModel(productDTO, _shoppingCartWebService /* dependcy */);, this is bad, imagine 3 more dependencies, this means the ProductSearchViewModel needs to take on those dependencies as well. Also changing the constructor is painfull. call _myInjectedProductDetailsViewModelFactory.Create().Initialize(productDTO);, the factory is just a Func, they are easily generated by most IoC frameworks. I think this is bad because Init methods are a leaky abstraction. You also can't use the readonly keyword for fields that are set in the Init method. I'm sure there are a few more reasons. call _myInjectedProductDetailsViewModelAbstractFactory.Create(productDTO); So... this is the pattern (abstract factory) that is usually recommended for this type of problem. I though it was genious since it satisfies my craving for static typing, until I actually started using it. The amount of boilerplate code is I think too much (you know, apart from the ridiculous variable names I get use). For each ViewModel that needs runtime parameters you'll get two extra files (factory interface and implementation), and you need to type the non-runtime dependencies like 4 extra times. And each time the dependencies change, you get to change it in the factory as well. It feels like I don't even use an DI container anymore. (I think Castle Windsor has some kind of solution for this [with it's own drawbacks, correct me if I'm wrong]). do something with anonymous types or dictionary. I like my static typing. So, yeah. Mixing state and behavior in this way creates a problem which don't exist at all in MVC. And I feel like there currently isn't a really adequate solution for this problem. Now I'd like to observe some things: People actually use MVVM. So they either don't care about all of the above, or they have some brilliant other solution. I haven't found an indepth example of MVVM with WPF. For example, the NDDD-sample project immensely helped me understand some DDD concepts. I'd really like it if someone could point me in the direction of something similar for MVVM/WPF. Maybe I'm doing MVVM all wrong and I should turn my design upside down. Maybe I shouldn't have this problem at all. Well I know other people have asked the same question so I think I'm not the only one. To summarize Am I correct to conclude that having the ViewModel being an integration point for both state and behavior is the reason for some difficulties with the MVVM pattern as a whole? Is using the abstract factory pattern the only/best way to instantiate a ViewModel in a statically typed way? Is there something like an in depth reference implementation available? Is having a lot of ViewModels with both state/behavior a design smell?

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