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• Web Application Integration Steps in OAM 11gR2 (High Level)

  - by Venkata Srikanth

  Install OAM, Webtier (OHS) and WebGate as per the standard installation steps. Create a WebGate instance (i.e deploy WebGate) A WebGate instance must be created that will copy required bits of agent from WEBGATE_HOME to WebGate instance location that shares the same INSTANCE_HOME with OHS ./deployWebGateInstance.sh –w /Oracle/Middleware/Oracle_WT1/instances/instance1/config/ohs1 –oh /Oracle/Middleware/Oracle_OAMWebGate1 Note: Here –w flag indicates OHS instance folder and –oh indicates the WebGate Oracle home Configure WebGate In the webgate configuration the EditHttpdConf utility will copy OUI instantiated apache_webgate.template from WEBGATE_HOME to webgate instance location (renamed to webgate.conf), and update httpd.conf with one additional line to include webgate.conf. export LD_LIBRARY_PATH=$ LD_LIBRARY_PATH:/Oracle/Middleware/Oracle_WT1/lib Navigate to /Oracle/Middleware/Oracle_OAMWebGate1/webgate/ohs/tools/setup/InstallTools ./EditHttpdConf –w /Oracle/Middleware/Oracle_WT1/instances/instace1/config/OHS/ohs1 –oh /Oracle/Middleware/Oracle_OAMWebGate1 –o webgate.conf Register WebGate Use RREG tool to register the OAM 11G WebGate Navigate to /Oracle/Middleware/Oracle_IDM1/oam/server/rreg/input Edit OAM11Grequest.xml. Change the specific xml content to include the weblogic admin URL, agentBaseURL, host identifier etc.. Navigate to /Oracle/Middleware/Oracle_IDM1/oam/server/rreg/bin Set permissions to oamreg.sh à chmod 777 oamreg.sh Edit oamreg.sh and set OAM_REG_HOME=/Oracle/Middleware/Oracle_IDM1/oam/server/rreg ./oamreg.sh inband input/OAM11Grequest.xml Enter the WebLogic admin credentials when prompted. After performing the above steps, there will be two artifcats created under Oracle/Middleware/Oracle_IDM1/oam/server/rreg/output, namely ObAccessClient.xml (Stroing webgate config parameters) and cwallet.sso (storing the agent key). These files must be copied to WebGate instance config folder (/Oracle/Middleware/Oracle_WT1/instances/instance1/config/ohs1/webgate/config) Restart OHS Deploy the web application (myApp) in WebLogic application server Proxy Configuration in OHS The mod_wl_ohs module enables requests to be proxied from Oracle HTTP Server 11g to Oracle WebLogic Server. Navigate to /Oracle/Middleware/Oracle_WT1/instances/instance1/config/OHS/ohs1 Edit mod_wl_ohs.conf file to include the following: <IfModule weblogic_module> WebLogicHost <WEBLOGIC_HOST> WebLogicPort <WEBLOGIC_PORT> # Debug ON # WLLogFile /tmp/weblogic.log MatchExpression *.jsp </IfModule> <Location /myApp> SetHandler weblogic-handler # PathTrim /weblogic # ErrorPage http:/WEBLOGIC_HOME:WEBLOGIC_PORT/ </Location> Note: Here WEBLOGIC_HOST and WEBLOGIC_PORT are the WebLogic admin server host and port respectively Restart OHS. Now if we access the web application URL with OHS host and port (Ex: http://OHS_HOST:<OHS_PORT>/myApp) so that the requests will be proxied to WebLogic server. Create a new application domain Login to OAM Admin Console Navigate to Shared Componentsà Authentication Schemesà Create Authentication Scheme (Ex: LDAP Auth Scheme. Here the scheme is assoicated with LDAP Authentication Module) Navigate to Policy Configuration à Application Domain à Create Application Domain Enter the Application Domain Name and Click Apply. Navigate to Resources tab and add the resource urls (Web Application URLs that needs to be protected) Navigate to Authentication Policy tab à Create a new authentication ploicy by providing the Resource URLs (The sample Web Application URLs) and Authentication Scheme. Navigate to Authorization Policy tab à Create a new authorization policy à Enter authorization policy name and navigate to Resource Tab à Attach the Reource URL, Host Identifiers here. Navigate to Conditions tab à Add the conditions like whom to allow and whom to deny access. Navigate to Rules tab à Crate the Allow Rule and Deny Rule with the available conditions from the previous step so that the Authorization Policy may authorize the logins. Navigate to Resources tab and attach the Authentication and Authorization plocies created in the above steps. Test the Web Application Integration.

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• Best depth sorting method for a Top Down 2D game using a 3D physics engine

  - by Alic44

  I've spent many days googling this and still have issues with my game engine I'd like to ask about, which I haven't seen addressed before. I think the problem is that my game is an unusual combination of a completely 2D graphical approach using XNA's SpriteBatch, and a completely 3D engine (the amazing BEPU physics engine) with rotation mostly disabled. In essence, my question is similar to this one (the part about "faux 3D"), but the difference is that in my game, the player as well as every other creature is represented by 3D objects, and they can all jump, pick up other objects, and throw them around. What this means is that sorting by one value, such as a Z position (how far north/south a character is on the screen) won't work, because as soon as a smaller creature jumps on top of a larger creature, or a box, and walks backwards, the moment its z value is less than that other creature, it will appear to be behind the object it is actually standing on. I actually originally solved this problem by splitting every object in the game into physics boxes which MUST have a Y height equal to their Z depth. I then based the depth sorting value on the object's y position (how high it is off the ground) PLUS its z position (how far north or south it is on the screen). The problem with this approach is that it requires all moving objects in the game to be split graphically into chunks which match up with a physical box which has its y dimension equal to its z dimension. Which is stupid. So, I got inspired last night to rewrite with a fresh approach. My new method is a little more complex, but I think a little more sane: every object which needs to be sorted by depth in the game exposes the interface IDepthDrawable and is added to a list owned by the DepthDrawer object. IDepthDrawable contains: public interface IDepthDrawable { Rectangle Bounds { get; } //possibly change this to a class if struct copying of the xna Rectangle type becomes an issue DepthDrawShape DepthShape { get; } void Draw(SpriteBatch spriteBatch); } The Bounds Rectangle of each IDepthDrawable object represents the 2D Axis-Aligned Bounding Box it will take up when drawn to the screen. Anything that doesn't intersect the screen will be culled at this stage and the remaining on-screen IDepthDrawables will be Bounds tested for intersections with each other. This is where I get a little less sure of what I'm doing. Each group of collisions will be added to a list or other collection, and each list will sort itself based on its DepthShape property, which will have access to the object-to-be-drawn's physics information. For starting out, lets assume everything in the game is an axis aligned 3D Box shape. Boxes are pretty easy to sort. Something like: if (depthShape1.Back > depthShape2.Front) //if depthShape1 is in front of depthShape2. //depthShape1 goes on top. else if (depthShape1.Bottom > depthShape2.Top) //if depthShape1 is above depthShape2. //depthShape1 goes on top. //if neither of these are true, depthShape2 must be in front or above. So, by sorting draw order by several different factors from the physics engine, I believe I can get a really correct draw order. My question is, is this a good way of going about this, or is there some tried and true, tested way which is completely different and has somehow completely eluded me on the internets? And, if this does seem like a good way to remake my draw order sorting, what's the right sorting algorithm for reordering the Bounds Rectangle collision lists, and how do you deal with a Bounds Rectangle colliding with two different object which don't collide with eachother. I know these are solved problems, but I've only been programming for a year so any specific input here will be greatly appreciated. Thanks for reading this far, ye who made it -- sorry it was so long!

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• Restrict number of characters to be typed for af:autoSuggestBehavior

  - by Arunkumar Ramamoorthy

  When using AutoSuggestBehavior for a UI Component, the auto suggest list is displayed as soon as the user starts typing in the field. In this article, we will find how to restrict the autosuggest list to be displayed till the user types in couple of characters. This would be more useful in the low latency networks and also the autosuggest list is bigger. We could display a static message to let the user know that they need to type in more characters to get a list for picking a value from. Final output we would expect is like the below image Lets see how we can implement this. Assuming we have an input text for the users to enter the country name and an autosuggest behavior is added to it. <af:inputText label="Country" id="it1"> <af:autoSuggestBehavior /> </af:inputText> Also, assuming we have a VO (we'll name it as CountryView for this example), with a view criteria to filter out the VO based on the bind variable passed. Now, we would generate View Impl class from the java node (including bind variables) and then expose the setter method of the bind variable to client interface. In the View layer, we would create a tree binding for the VO and the method binding for the setter method of the bind variable exposed above, in the pagedef file As we've already added an input text and an autosuggestbehavior for the test, we would not need to build the suggested items for the autosuggest list.Let us add a method in the backing bean to return us List of select items to be bound to the autosuggest list. padding: 5px; background-color: #fbfbfb; min-height: 40px; width: 544px; height: 168px; overflow: auto;"> public List onSuggest(String searchTerm) { ArrayList<SelectItem> selectItems = new ArrayList<SelectItem>(); if(searchTerm.length()>1) { //get access to the binding context and binding container at runtime BindingContext bctx = BindingContext.getCurrent(); BindingContainer bindings = bctx.getCurrentBindingsEntry(); //set the bind variable value that is used to filter the View Object //query of the suggest list. The View Object instance has a View //Criteria assigned OperationBinding setVariable = (OperationBinding) bindings.get("setBind_CountryName"); setVariable.getParamsMap().put("value", searchTerm); setVariable.execute(); //the data in the suggest list is queried by a tree binding. JUCtrlHierBinding hierBinding = (JUCtrlHierBinding) bindings.get("CountryView1"); //re-query the list based on the new bind variable values hierBinding.executeQuery(); //The rangeSet, the list of queries entries, is of type //JUCtrlValueBndingRef. List<JUCtrlValueBindingRef> displayDataList = hierBinding.getRangeSet(); for (JUCtrlValueBindingRef displayData : displayDataList){ Row rw = displayData.getRow(); //populate the SelectItem list selectItems.add(new SelectItem( (String)rw.getAttribute("Name"), (String)rw.getAttribute("Name"))); } } else{ SelectItem a = new SelectItem("","Type in two or more characters..","",true); selectItems.add(a); } return selectItems; } So, what we are doing in the above method is, to check the length of the search term and if it is more than 1 (i.e 2 or more characters), the return the actual suggest list. Otherwise, create a read only select item new SelectItem("","Type in two or more characters..","",true); and add it to the list of suggested items to be displayed. The last parameter for the SelectItem (boolean) is to make it as readOnly, so that users would not be able to select this static message from the displayed list. Finally, bind this method to the input text's autosuggestbehavior's suggestedItems property. <af:inputText label="Country" id="it1"> <af:autoSuggestBehavior suggestedItems="#{AutoSuggestBean.onSuggest}"/> </af:inputText>

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• Create a Remote Git Repository from an Existing XCode Repository

  - by codeWithoutFear

  Introduction Distributed version control systems (VCS’s), like Git, provide a rich set of features for managing source code.  Many development tools, including XCode, provide built-in support for various VCS’s.  These tools provide simple configuration with limited customization to get you up and running quickly while still providing the safety net of basic version control. I hate losing (and re-doing) work.  I have OCD when it comes to saving and versioning source code.  Save early, save often, and commit to the VCS often.  I also hate merging code.  Smaller and more frequent commits enable me to minimize merge time and effort as well. The work flow I prefer even for personal exploratory projects is: Make small local changes to the codebase to create an incrementally improved (and working) system. Commit these changes to the local repository.  Local repositories are quick to access, function even while offline, and provides the confidence to continue making bold changes to the system.  After all, I can easily recover to a recent working state. Repeat 1 & 2 until the codebase contains “significant” functionality and I have connectivity to the remote repository. Push the accumulated changes to the remote repository.  The smaller the change set, the less likely extensive merging will be required.  Smaller is better, IMHO. The remote repository typically has a greater degree of fault tolerance and active management dedicated to it.  This can be as simple as a network share that is backed up nightly or as complex as dedicated hardware with specialized server-side processing and significant administrative monitoring. XCode’s out-of-the-box Git integration enables steps 1 and 2 above.  Time Machine backups of the local repository add an additional degree of fault tolerance, but do not support collaboration or take advantage of managed infrastructure such as on-premises or cloud-based storage. Creating a Remote Repository These are the steps I use to enable the full workflow identified above.  For simplicity the “remote” repository is created on the local file system.  This location could easily be on a mounted network volume. Create a Test Project My project is called HelloGit and is located at /Users/Don/Dev/HelloGit.  Be sure to commit all outstanding changes.  XCode always leaves a single changed file for me after the project is created and the initial commit is submitted. Clone the Local Repository We want to clone the XCode-created Git repository to the location where the remote repository will reside.  In this case it will be /Users/Don/Dev/RemoteHelloGit. Open the Terminal application. Clone the local repository to the remote repository location: git clone /Users/Don/Dev/HelloGit /Users/Don/Dev/RemoteHelloGit Convert the Remote Repository to a Bare Repository The remote repository only needs to contain the Git database.  It does not need a checked out branch or local files. Go to the remote repository folder: cd /Users/Don/Dev/RemoteHelloGit Indicate the repository is “bare”: git config --bool core.bare true Remove files, leaving the .git folder: rm -R * Remove the “origin” remote: git remote rm origin Configure the Local Repository The local repository should reference the remote repository.  The remote name “origin” is used by convention to indicate the originating repository.  This is set automatically when a repository is cloned.  We will use the “origin” name here to reflect that relationship. Go to the local repository folder: cd /Users/Don/Dev/HelloGit Add the remote: git remote add origin /Users/Don/Dev/RemoteHelloGit Test Connectivity Any changes made to the local Git repository can be pushed to the remote repository subject to the merging rules Git enforces. Create a new local file: date > date.txt /li> Add the new file to the local index: git add date.txt Commit the change to the local repository: git commit -m "New file: date.txt" Push the change to the remote repository: git push origin master Now you can save, commit, and push/pull to your OCD hearts’ content! Code without fear! --Don

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• JPedal Action for Converting PDF to JavaFX

  - by Geertjan

  The question of the day comes from Mark Stephens, from JPedal (JPedal is the leading 100% Java PDF library, providing a Java PDF viewer, PDF to image conversion, PDF printing or adding PDF search and PDF extraction features), in the form of a screenshot: The question is clear. By looking at the annotations above, you can see that Mark has an ActionListener that has been bound to the right-click popup menu on PDF files. Now he needs to get hold of the file to which the Action has been bound. How, oh  how, can one get hold of that file? Well, it's simple. Leave everything you see above exactly as it is but change the Java code section to this: public final class PDF2JavaFXContext implements ActionListener {     private final DataObject context;     public PDF2JavaFXContext(DataObject context) {         this.context = context;     }     public void actionPerformed(ActionEvent ev) {         FileObject fo = context.getPrimaryFile();         File theFile = FileUtil.toFile(fo);         //do something with your file...     } } The point is that the annotations at the top of the class bind the Action to either Actions.alwaysEnabled, which is a factory method for creating always-enabled Actions, or Actions.context, which is a factory method for creating context-sensitive Actions. How does the Action get bound to the factory method? The annotations are converted, when the module is compiled, into XML registration entries in the "generated-layer.xml", which you can find in your "build" folder, in the Files window, after building the module. In Mark's case, since the Action should be context-sensitive to PDF files, he needs to bind his PDF2JavaFXContext ActionListener (which should probably be named "PDF2JavaFXActionListener", since the class is an ActionListener) to Actions.context. All he needs to do that is pass in the object he wants to work with into the constructor of the ActionListener. Now, when the module is built, the annotation processor is going to take the annotations and convert them to XML registration entries, but the constructor will also be checked to see whether it is empty or not. In this case, the constructor isn't empty, hence the Action should be context-sensitive and so the ActionListener is bound to Actions.context. The Actions.context will do all the enablement work for Mark, so that he will not need to provide any code for enabling/disabling the Action. The Action will be enabled whenever a DataObject is selected. Since his Action is bound to Nodes in the Projects window that represent PDF files, the Action will always be enabled whenever Mark right-clicks on a PDF Node, since the Node exposes its own DataObject. Once Mark has access to the DataObject, he can get the underlying FileObject via getPrimaryFile and he can then convert the FileObject to a java.io.File via FileUtil.getConfigFile. Once he's got the java.io.File, he can do with it whatever he needs. Further reading: http://bits.netbeans.org/dev/javadoc/

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• questions about dual-boot install Ubuntu 10.04 and Windows 7 on same hard drive

  - by Tim

  I'd like to dual-boot install Ubuntu 10.04 on the same hard drive as Windows 7 which has already been installed. As to sources on the internet: I found a website iinet about dual-boot installation of Ubuntu 10.10 and Windows 7 on the same hard drive, which I think more specific than the one on Ubuntu Community without specific version of the OSes. Since I am installing Ubuntu 10.04 instead of 10.10, my question is whether their installers are same or almost same and if I can follow iinet for my dual-boot installation? Or are there better websites for information about dual-boot installtion of Ubuntu 10.04 and Windows 7? As to shrinking Windows partitions to make free space for Ubuntu partitions: iinet uses the partition software in Ubuntu's installer to shrink the Windows partition. But I saw in many website that the partition software in Ubuntu's installer cannot guarantee shrinking Windows 7 partitions successfully, so they recommended in general to shrink Windows partitions under Windows itself using its softwares. For example, in Ubuntu Community, it says: Some people think that the Windows partition must be resized only from within Windows Vista and Windows 7 using the shrink/resize option. ... If you use GParted Partition Editor in the Ubuntu Live CD be careful. So I was wondering which way to go in my situation? As to partition for bootloader files: In iinet, I don't see there is a partition created and dedicated to boot files (i.e. Grub files). However, I saw in many websites strongly suggesting using a boot partition for Grub files, especially for the purpose of separation and protection from installed OS files. I was wondering which way I should choose and why? As to installing bootloader Grub, in iinet, I see that to install Grub it only needs to specify the hard drive device for bootloader installation. However, in ubuntuguide(for more than 2 OSes and Ubuntu 9.04), some commands are needed to run in order to put Grub configuration files in MBR, and OS partition, for the chain-load process (where to find the files for the next stage). In Ubuntu Community, there are some related sentences which I don't quite understand how to do in practice: the only thing in your computer outside of Ubuntu that needs to be changed is a small code in the MBR (Master Boot Record) of the first hard disk. The MBR code is changed to point to the boot loader in Ubuntu. If you have a problem with changing the MBR code, you might prefer to just install the code for pointing to GRUB to the first sector of your Ubuntu partition instead. If you do that during the Ubuntu installation process, then Ubuntu won't boot until you configure some other boot manager to point to Ubuntu's boot sector. Windows Vista no longer utilizes boot.ini, ntdetect.com, and ntldr when booting. Instead, Vista stores all data for its new boot manager in a boot folder. Windows Vista ships with an command line utility called bcdedit.exe, which requires administrator credentials to use. You may want to read http://go.microsoft.com/fwlink/?LinkId=112156 about it. Using a command line utility always has its learning curve, so a more productive and better job can be done with a free utility called EasyBCD, developed and mastered in during the times of Vista Beta already. EasyBCD is user friendly and many Vista users highly recommend EasyBCD. In what is quoted above, I was wondering how exactly I should change the MBR code to point to the bootloader in Ubuntu? if I fail to change MBR code, are the other suggested boot managers being bcdedit.exe and EasyBCD in Windows? With the three sources above, which one shall I follow? Thanks and regards

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• Running a WebLogic Portal (WLP) 10.3.4 Domain as a Windows Service

  - by user647124

  To start a WLP server as a Windows service it is simplest to make your own script based on the provided standard script located at WL_HOME\server\bin\installSvc.cmd. The standard script works fine for a plain WLS domain, but lacks some classpath and options necessary for WLP.Start by making a copy of the installSvc.cmd script and naming it something specific to your domain.Next, just under SETLOCAL you will find where WL_HOME is defined. Here you will add the definitions you would normally add in a script that later calls installSvc.cmd (as per the standard documentation). set DOMAIN_NAME=gnma_test_domainset USERDOMAIN_HOME=D:\my_test_domainset SERVER_NAME=AdminServerset WLS_USER=weblogicset WLS_PW=gnmaAdmin01set PRODUCTION_MODE=trueset MEM_ARGS=-Xms512m –Xmx512mset MW_HOME=C:\Oracle\Middleware Note: I had heard of people using this approach who had issues with the length of the command line. This may be due to their use of the default domain path. In the example above, I use a shorter path.At this point, edit the DOMAIN_HOME\bin\startWebLogic.cmd and set it to echo both the classpath and the options. Then start the domain and capture the output of those echoes, then shut the domain back down. Now REM out the existing CLASSPATH definition, then use the outputs you captured earlier to set the CLASSPATH and JAVA_OPTIONS like this: REM set CLASSPATH=%WEBLOGIC_CLASSPATH%;%CLASSPATH%; C:\Oracle\Middleware\wlportal_10.3\portal\lib\security\wsrp-security-providers.jarset CLASSPATH=%MW_HOME%\patch_wls1034\profiles\default\sys_manifest_classpath\weblogic_patch.jar;%MW_HOME%\patch_wlp1034\profiles\default\sys_manifest_classpath\weblogic_patch.jar;%MW_HOME%\patch_oepe1111\profiles\default\sys_manifest_classpath\weblogic_patch.jar;%MW_HOME%\patch_ocm1033\profiles\default\sys_manifest_classpath\weblogic_patch.jar;%MW_HOME%\JROCKI~1.1-3\lib\tools.jar;%WL_HOME%\server\lib\weblogic_sp.jar;%WL_HOME%\server\lib\weblogic.jar;%MW_HOME%\modules\features\weblogic.server.modules_10.3.4.0.jar;%WL_HOME%\server\lib\webservices.jar;%MW_HOME%\modules\ORGAPA~1.1/lib/ant-all.jar;%MW_HOME%\modules\NETSFA~1.0_1/lib/ant-contrib.jar;%WL_HOME%\common\derby\lib\derbyclient.jar;%WL_HOME%\server\lib\xqrl.jar;%WL_HOME%\server\lib\xquery.jar;%WL_HOME%\server\lib\binxml.jarset JAVA_OPTIONS= -Xverify:none -ea -da:com.bea... -da:javelin... -da:weblogic... -ea:com.bea.wli... -ea:com.bea.broker... -ea:com.bea.sbconsole... -Dplatform.home=%WL_HOME% -Dwls.home=%WL_HOME%\server -Dweblogic.home=%WL_HOME%\server -Dweblogic.wsee.bind.suppressDeployErrorMessage=true -Dweblogic.wsee.skip.async.response=true -Dweblogic.management.discover=true -Dwlw.iterativeDev=true -Dwlw.testConsole=true -Dwlw.logErrorsToConsole=true -Dweblogic.ext.dirs=%MW_HOME%\patch_wls1034\profiles\default\sysext_manifest_classpath;%MW_HOME%\patch_wlp1034\profiles\default\sysext_manifest_classpath;%MW_HOME%\patch_oepe1111\profiles\default\sysext_manifest_classpath;%MW_HOME%\patch_ocm1033\profiles\default\sysext_manifest_classpath;%MW_HOME%\wlportal_10.3\p13n\lib\system;%MW_HOME%\wlportal_10.3\light-portal\lib\system;%MW_HOME%\wlportal_10.3\portal\lib\system;%MW_HOME%\wlportal_10.3\info-mgmt\lib\system;%MW_HOME%\wlportal_10.3\analytics\lib\system;%MW_HOME%\wlportal_10.3\apps\lib\system;%MW_HOME%\wlportal_10.3\info-mgmt\deprecated\lib\system;%MW_HOME%\wlportal_10.3\content-mgmt\lib\system -Dweblogic.alternateTypesDirectory=%MW_HOME%\wlportal_10.3\portal\lib\securityAnd that's it. Looks really simple, but it took me quite some time to gather all the necessary pieces in order to make it work. Hopefully you find this before you went through half as much research.The example here uses a domain with only the Admin server and no managed servers. For a variety of reasons I only want the Admin server to be run as a service. The standard documentation along with the example above should allow you to expand this to include managed servers should you feel the need.

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• RewriteRule not working for certain URLs

  - by keiki

  There are a few domains pointing towards the same server, and of course I need them all redirect to only one of them. Redirects work, but only for certain URLs. What works: http://www.domain.com, http://domain.com, domain.com/index.html, domain.com/index.php, , domain.com/nonExistentDirectory, and if I click in the menu the following URLs are also redirected correctly: domain.com/foo/bar, domain.com/foo/bar.html or .php or other extension. What doesn't work: domain.com/existentDirectory, domain.com/foo/bar (if I type the URL in the address bar). If anyone will have the time and skill and will to tell me where's the mistake, I'll be deeply grateful. Here's my .htaccess file: AddHandler x-httpd-php .html .htm <ifModule mod_gzip.c> mod_gzip_on Yes mod_gzip_dechunk Yes mod_gzip_item_include file \.(html?|txt|css|js|php|pl)$ mod_gzip_item_include handler ^cgi-script$ mod_gzip_item_include mime ^text/.* mod_gzip_item_include mime ^application/x-javascript.* mod_gzip_item_exclude mime ^image/.* mod_gzip_item_exclude rspheader ^Content-Encoding:.*gzip.* </ifModule> <ifModule mod_expires.c> ExpiresActive On ExpiresDefault "access plus 1 seconds" ExpiresByType text/html "access plus 1 seconds" ExpiresByType image/gif "access plus 2592000 seconds" ExpiresByType image/jpeg "access plus 2592000 seconds" ExpiresByType image/png "access plus 2592000 seconds" ExpiresByType text/css "access plus 2592000 seconds" ExpiresByType text/javascript "access plus 216000 seconds" ExpiresByType application/x-javascript "access plus 216000 seconds" </ifModule> <ifModule mod_headers.c> <filesMatch "\\.(ico|pdf|flv|jpg|jpeg|png|gif|swf)$"> Header set Cache-Control "max-age=2592000, public" </filesMatch> <filesMatch "\\.(css)$"> Header set Cache-Control "max-age=2592000, public" </filesMatch> <filesMatch "\\.(js)$"> Header set Cache-Control "max-age=216000, private" </filesMatch> <filesMatch "\\.(xml|txt)$"> Header set Cache-Control "max-age=216000, public, must-revalidate" </filesMatch> <filesMatch "\\.(html|htm|php)$"> Header set Cache-Control "max-age=1, private, must-revalidate" </filesMatch> </ifModule> <ifModule mod_headers.c> Header unset ETag </ifModule> FileETag None <ifModule mod_headers.c> Header unset Last-Modified </ifModule> # BEGIN WordPress <IfModule mod_rewrite.c> RewriteEngine On RewriteBase / RewriteCond %{REQUEST_FILENAME} !-f RewriteCond %{REQUEST_FILENAME} !-d RewriteRule . /index.php [L] </IfModule> # END WordPress RewriteCond %{HTTP_HOST} ^foo\.com$ [OR] RewriteCond %{HTTP_HOST} ^www\.foo\.com$ RewriteRule (.*) http://domain.com/$1 [R=301,L,QSA] RewriteCond %{HTTP_HOST} ^foo1\.com$ [OR] RewriteCond %{HTTP_HOST} ^www\.foo1\.com$ RewriteRule (.*) http://domain.com/$1 [R=301,L,QSA] RewriteCond %{HTTP_HOST} ^foo2\.com$ [OR] RewriteCond %{HTTP_HOST} ^www\.foo2\.com$ RewriteRule (.*) http://domain.com/$1 [R=301,L,QSA] RewriteCond %{HTTP_HOST} ^foo3\.com$ [OR] RewriteCond %{HTTP_HOST} ^www\.foo3\.com$ RewriteRule (.*) http://domain.com/$1 [R=301,L,QSA] RewriteCond %{HTTP_HOST} ^foo8\.com$ [OR] RewriteCond %{HTTP_HOST} ^www\.foo8\.com$ RewriteRule (.*) http://domain.com/$1 [R=301,L,QSA] Thinking that the above version was overkill, I've also tried to redirect all the requests for domains different than the main on to be redirected to it like this: RewriteCond %{HTTP_HOST} !^domain\.com$ [NC] RewriteRule ^(.*)$ http://domain.com [L,R=301] Is it also wrong? Because it doesn't work either! P.S. @Sodved I've tried that and it doesn't help (I comment here because I can't seem to be able to comment your answer.) Removing the following piece of code didn't solve the issue either, so the problem must be somewhere else: # BEGIN WordPress <IfModule mod_rewrite.c> RewriteEngine On RewriteBase / RewriteCond %{REQUEST_FILENAME} !-f RewriteCond %{REQUEST_FILENAME} !-d RewriteRule . /index.php [L] </IfModule> # END WordPress New details: using this tool for checking the redirects I got the following results for the URLs that are not redirected: Checked link: http://domain.com/aDirectory/ Type of link: direct link (note the trailing slash above) and: Checked link: http://domain.com/aDirectory Type of redirect: 301 Moved Permanently Redirected to: http://domain.com/aDirectory/ (no trailing slash here) I hope/suspect I'm getting closer to the cause of this behavior.

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• Roles / Profiles / Perspectives in NetBeans IDE 7.1

  - by Geertjan

  With a check out of main-silver from yesterday, I'm able to use the brand new "role" attribute in @TopComponent.Registration, as you can see below, in the bit in bold: @ConvertAsProperties(dtd = "-//org.role.demo.ui//Admin//EN", autostore = false) @TopComponent.Description(preferredID = "AdminTopComponent", //iconBase="SET/PATH/TO/ICON/HERE", persistenceType = TopComponent.PERSISTENCE_ALWAYS) @TopComponent.Registration(mode = "editor", openAtStartup = true, role="admin") public final class AdminTopComponent extends TopComponent { And here's a window for general users of the application, with the "role" attribute set to "user": @ConvertAsProperties(dtd = "-//org.role.demo.ui//User//EN", autostore = false) @TopComponent.Description(preferredID = "UserTopComponent", //iconBase="SET/PATH/TO/ICON/HERE", persistenceType = TopComponent.PERSISTENCE_ALWAYS) @TopComponent.Registration(mode = "explorer", openAtStartup = true, role="user") public final class UserTopComponent extends TopComponent { So, I have two windows. One is assigned to the "admin" role, the other to the "user" role. In the "ModuleInstall" class, I add a "WindowSystemListener" and set "user" as the application's role: public class Installer extends ModuleInstall implements WindowSystemListener { @Override public void restored() { WindowManager.getDefault().addWindowSystemListener(this); } @Override public void beforeLoad(WindowSystemEvent event) { WindowManager.getDefault().setRole("user"); WindowManager.getDefault().removeWindowSystemListener(this); } @Override public void afterLoad(WindowSystemEvent event) { } @Override public void beforeSave(WindowSystemEvent event) { } @Override public void afterSave(WindowSystemEvent event) { } } So, when the application starts, the "UserTopComponent" is shown, not the "AdminTopComponent". Next, I have two Actions, for switching between the two roles, as shown below: @ActionID(category = "Window", id = "org.role.demo.ui.SwitchToAdminAction") @ActionRegistration(displayName = "#CTL_SwitchToAdminAction") @ActionReferences({ @ActionReference(path = "Menu/Window", position = 250) }) @Messages("CTL_SwitchToAdminAction=Switch To Admin") public final class SwitchToAdminAction extends AbstractAction { @Override public void actionPerformed(ActionEvent e) { WindowManager.getDefault().setRole("admin"); } @Override public boolean isEnabled() { return !WindowManager.getDefault().getRole().equals("admin"); } } @ActionID(category = "Window", id = "org.role.demo.ui.SwitchToUserAction") @ActionRegistration(displayName = "#CTL_SwitchToUserAction") @ActionReferences({ @ActionReference(path = "Menu/Window", position = 250) }) @Messages("CTL_SwitchToUserAction=Switch To User") public final class SwitchToUserAction extends AbstractAction { @Override public void actionPerformed(ActionEvent e) { WindowManager.getDefault().setRole("user"); } @Override public boolean isEnabled() { return !WindowManager.getDefault().getRole().equals("user"); } } When I select one of the above actions, the role changes, and the other window is shown. I could, of course, add a Login dialog to the "SwitchToAdminAction", so that authentication is required in order to switch to the "admin" role. Now, let's say I am now in the "user" role. So, the "UserTopComponent" shown above is now opened. I decide to also open another window, the Properties window, as below... ...and, when I am in the "admin" role, when the "AdminTopComponent" is open, I decide to also open the Output window, as below... Now, when I switch from one role to the other, the additional window/s I opened will also be opened, together with the explicit members of the currently selected role. And, the main window position and size are also persisted across roles. When I look in the "build" folder of my project in development, I see two different Windows2Local folders, one per role, automatically created by the fact that there is something to be persisted for a particular role, e.g., when a switch to a different role is done: And, with that, we now clearly have roles/profiles/perspectives in NetBeans Platform applications from NetBeans Platform 7.1 onwards.

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• Scenarios for Throwing Exceptions

  - by Joe Mayo

  I recently came across a situation where someone had an opinion that differed from mine of when an exception should be thrown. This particular case was an issue opened on LINQ to Twitter for an Exception on EndSession.  The premise of the issue was that the poster didn’t feel an exception should be raised, regardless of authentication status.  As first, this sounded like a valid point.  However, I went back to review my code and decided not to make any changes. Here's my rationale: 1. The exception doesn’t occur if the user is authenticated when EndAccountSession is called. 2. The exception does occur if the user is not authenticated when EndAccountSession is called. 3. The exception represents the fact that EndAccountSession is not able to fulfill its intended purpose - to end the session.  If a session never existed, then it would not be possible to perform the requested action.  Therefore, an exception is appropriate. To help illustrate how to handle this situation, I've modified the following code in Program.cs in the LinqToTwitterDemo project to illustrate the situation: static void EndSession(ITwitterAuthorizer auth) { using (var twitterCtx = new TwitterContext(auth, "https://api.twitter.com/1/", "https://search.twitter.com/")) { try { //Log twitterCtx.Log = Console.Out; var status = twitterCtx.EndAccountSession(); Console.WriteLine("Request: {0}, Error: {1}" , status.Request , status.Error); } catch (TwitterQueryException tqe) { var webEx = tqe.InnerException as WebException; if (webEx != null) { var webResp = webEx.Response as HttpWebResponse; if (webResp != null && webResp.StatusCode == HttpStatusCode.Unauthorized) Console.WriteLine("Twitter didn't recognize you as having been logged in. Therefore, your request to end session is illogical.\n"); } var status = tqe.Response; Console.WriteLine("Request: {0}, Error: {1}" , status.Request , status.Error); } } } As expected, LINQ to Twitter wraps the exception in a TwitterQueryException as the InnerException.  The TwitterQueryException serves a very useful purpose through it's Response property.  Notice in the example above that the response has Request and Error proprieties.  These properties correspond to the information that Twitter returns as part of it's response payload.  This is often useful while debugging to help you understand why Twitter was unable to perform the  requested action.  Other times, it's cryptic, but that's another story.  At least you have some way of knowing in your code how to anticipate and handle these situations, along with having extra information to debug with. To sum things up, there are two points to make: when and why an exception should be raised and when to wrap and re-throw an exception in a custom exception type. I felt it was necessary to allow the exception to be raised because the called method was unable to perform the task it was designed for.  I also felt that it is inappropriate for a general library to do anything with exceptions because that could potentially hide a problem from the caller.  A related point is that it should be the exclusive decision of the application that uses the library on what to do with an exception.  Another aspect of this situation is that I wrapped the exception in a custom exception and re-threw.  This is a tough call because I don’t want to hide any stack trace information.  However, the need to make the exception more meaningful by including vital information returned from Twitter swayed me in the direction to design an interface that was as helpful as possible to library consumers.  As shown in the code above, you can dig into the exception and pull out a lot of good information, such as the fact that the underlying HTTP response was a 401 Unauthorized.  In all, trade-offs are seldom perfect for all cases, but combining the fact that the method was unable to perform its intended function, this is a library, and the extra information can be more helpful, it seemed to be the better design. @JoeMayo

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• "previousMode": Controling the Pin Action of a TopComponent

  - by Geertjan

  An excellent thing I learned today is that you, as a developer of a NetBeans module or NetBeans Platform application, can control the pin button. Up until today, whenever I had a TopComponent defined to appear in "rightSlidingSide" mode and then I clicked the "pin" button, as shown here... ...the TopComponent would then find itself pinned in the "explorer" mode. Would make more sense if it would be pinned in the "properties" mode, which is the docked mode closest to the "rightSlidingSide" mode. Not being able to control the "pin" button has been a recurring question (including in my own head) over several years. But the NetBeans Team's window system guru Stan Aubrecht informed me today that a "previousMode" attribute exists in the "tc-ref" file of the TopComponent. Since a few releases, that file is generated via the annotations in the TopComponent. However, "previousMode" is currently not one of the attributes exposed by the @TopComponent.Registration annotation. Therefore, what I did was this: Set "rightSlidingSide" in the "mode" attribute of the @TopComponent.Registration. Build the module. Find the "generated-layer.xml" (in the Files window) and move the layer registration of the TopComponent, including its action and menu item for opening the TopComponent, into my own manual layer within the module. Then remove all the TopComponent annotations from the TopComponent, though you can keep @ConvertAsProperties and @Messages. Then add the "previousMode" attribute, as highlighted below, into my own layer file, i.e., within the tags copied from the "generated-layer.xml": <folder name="Modes"> <folder name="rightSlidingSide"> <file name="ComparatorTopComponent.wstcref"> <![CDATA[<?xml version="1.0" encoding="UTF-8"?> <!DOCTYPE tc-ref PUBLIC "-//NetBeans//DTD Top Component in Mode Properties 2.0//EN" "http://www.netbeans.org/dtds/tc-ref2_0.dtd"> <tc-ref version="2.0"> <tc-id id="ComparatorTopComponent"/> <state opened="false"/> <previousMode name="properties" index="0" /> </tc-ref> ]]> </file> </folder> </folder> Now when you run the application and pin the specific TopComponent defined above, i.e., in the above case, named "ComparatorTopComponent", you will find it is pinned into the "properties" mode! That's pretty cool and if you agree, then you're a pretty cool NetBeans Platform developer, and I'd love to find out more about the application/s you're creating on the NetBeans Platform! Meanwhile, I'm going to create an issue for exposing the "previousMode" attribute in the @TopComponent.Registration annotation.

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• Syntax of passing lambda

  - by Astara

  Right now, I'm working on refactoring a program that calls its parts by polling to a more event-driven structure. I've created sched and task classes with the sced to become a base class of the current main loop. The tasks will be created for each meter so they can be called off of that instead of polling. Each of the events main calls are a type of meter that gather info and display it. When the program is coming up, all enabled meters get 'constructed' by a main-sub. In that sub, I want to store off the "this" pointer associated with the meter, as well as the common name for the "action routine. void MeterMaker::Meter_n_Task (Meter * newmeter,) { push(newmeter); // handle non-timed draw events Task t = new Task(now() + 0.5L); t.period={0,1U}; t.work_meter = newmeter; t.work = [&newmeter](){newmeter.checkevent();};<<--attempt at lambda t.flags = T_Repeat; t.enable_task(); _xos->sched_insert(t); } A sample call to it: Meter_n_Task(new CPUMeter(_xos, "CPU ")); 've made the scheduler a base class of the main routine (that handles the loop), and I've tried serveral variations to get the task class to be a base of the meter class, but keep running into roadblocks. It's alot like "whack-a-mole" -- pound in something to fix something one place, and then a new probl pops out elsewhere. Part of the problem, is that the sched.h file that is trying to hold the Task Q, includes the Task header file. The task file Wants to refer to the most "base", Meter class. The meter class pulls in the main class of the parent as it passes a copy of the parent to the children so they can access the draw routines in the parent. Two references in the task file are for the 'this' pointer of the meter and the meter's update sub (to be called via this). void *this_data= NULL; void (*this_func)() = NULL; Note -- I didn't really want to store these in the class, as I wanted to use a lamdba in that meter&task routine above to store a routine+context to be used to call the meter's action routine. Couldn't figure out the syntax. But am running into other syntax problems trying to store the pointers...such as g++: COMPILE lsched.cc In file included from meter.h:13:0, from ltask.h:17, from lsched.h:13, from lsched.cc:13: xosview.h:30:47: error: expected class-name before ‘{’ token class XOSView : public XWin, public Scheduler { Like above where it asks for a class, where the classname "Scheduler" is. !?!? Huh? That IS a class name. I keep going in circles with things that don't make sense... Ideally I'd get the lamba to work right in the Meter_n_Task routine at the top. I wanted to only store 1 pointer in the 'Task' class that was a pointer to my lambda that would have already captured the "this" value ... but couldn't get that syntax to work at all when I tried to start it into a var in the 'Task' class. This project, FWIW, is my teething project on the new C++... (of course it's simple!.. ;-))... I've made quite a bit of progress in other areas in the code, but this lambda syntax has me stumped...its at times like thse that I appreciate the ease of this type of operation in perl. Sigh. Not sure the best way to ask for help here, as this isn't a simple question. But thought I'd try!... ;-) Too bad I can't attach files to this Q.

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• Is there a theory for "transactional" sequences of failing and no-fail actions?

  - by Ross Bencina

  My question is about writing transaction-like functions that execute sequences of actions, some of which may fail. It is related to the general C++ principle "destructors can't throw," no-fail property, and maybe also with multi-phase transactions or exception safety. However, I'm thinking about it in language-neutral terms. My concern is with correctly designing error handling in C++ functions that must be reliable. I would like to know what the concepts below are called so that I can learn more about them. I'm sorry that I can't ask the question more directly. Since I don't know this area I have provided an example to explain my question. The question is at the end. Here goes: Consider a sequence of steps or actions executed sequentially, where actions belong to one of two classes: those that always succeed, and those that may fail. In the examples below: S stands for an action that always succeeds (called "no-fail" in some settings). F stands for an action that may fail (for example, it might fail to allocate memory or do I/O that could fail). Consider a sequences of actions (executed sequentially from left to right): S->S->S->S Since each action in the sequence above succeeds, the whole sequence succeeds. On the other hand, the following sequence may fail because the last action may fail: S->S->S->F So, claim: a sequence has the no-fail (S) property if and only if all of its actions are no-fail. Now, I'm interested in action sequences that form "atomic transactions", with "failure atomicity," i.e. where either the whole sequence completes successfully, or there is no effect. I.e. if some action fails, the earlier ones must be rolled back. This requires that any successfully executed actions prior to a failing action must always be able to be rolled back. Consider the sequence: S->S->S->F S<-S<-S In the example above, the first row is the forward path of the transaction, and the second row are inverse actions (executed from right to left) that can be used to roll back if the final top row actions fails. It seems to me that for a transaction to support failure atomicity, the following invariant must hold: Claim: To support failure atomicity (either completion or complete roll-back on failure) all actions preceding the latest failable (F) action on the forward path (marked * in the example below) must have no-fail (S) inverses. The following is an example of a sequence that supports failure atomicity: * S->F->F->F S<-S<-S Further, if we want the transaction to be able to attempt cancellation mid-way through, but still guarantee either full completion or full rollback then we need the following property: Claim: To support failure atomicity and cancellation mid-way through execution, in the face of errors in the inverse (cancellation) path, all actions following the earliest failable (F) inverse on the reverse path (marked *) must be no-fail (S). F->F->F->S->S S<-S<-F<-F * I believe that these two conditions guarantee that an abortable/cancelable transaction will never get "stuck". My questions are: What is the study and theory of these properties called? are my claims correct? and what else is there to know? UPDATE 1: Updated terminology: what I previously called "robustness" is called atomicity in the database literature. UPDATE 2: Added explicit reference to failure atomicity, which seems to be a thing.

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• Incorrect lighting results with deferred rendering

  - by Lasse

  I am trying to render a light-pass to a texture which I will later apply on the scene. But I seem to calculate the light position wrong. I am working on view-space. In the image above, I am outputting the attenuation of a point light which is currently covering the whole screen. The light is at 0,10,0 position, and I transform it to view-space first: Vector4 pos; Vector4 tmp = new Vector4 (light.Position, 1); // Transform light position for shader Vector4.Transform (ref tmp, ref Camera.ViewMatrix, out pos); shader.SendUniform ("LightViewPosition", ref pos); Now to me that does not look as it should. What I think it should look like is that the white area should be on the center of the scene. The camera is at the corner of the scene, and it seems as if the light would move along with the camera. Here's the fragment shader code: void main(){ // default black color vec3 color = vec3(0); // Pixel coordinates on screen without depth vec2 PixelCoordinates = gl_FragCoord.xy / ScreenSize; // Get pixel position using depth from texture vec4 depthtexel = texture( DepthTexture, PixelCoordinates ); float depthSample = unpack_depth(depthtexel); // Get pixel coordinates on camera-space by multiplying the // coordinate on screen-space by inverse projection matrix vec4 world = (ImP * RemapMatrix * vec4(PixelCoordinates, depthSample, 1.0)); // Undo the perspective calculations vec3 pixelPosition = (world.xyz / world.w) * 3; // How far the light should reach from it's point of origin float lightReach = LightColor.a / 2; // Vector in between light and pixel vec3 lightDir = (LightViewPosition.xyz - pixelPosition); float lightDistance = length(lightDir); vec3 lightDirN = normalize(lightDir); // Discard pixels too far from light source //if(lightReach < lightDistance) discard; // Get normal from texture vec3 normal = normalize((texture( NormalTexture, PixelCoordinates ).xyz * 2) - 1); // Half vector between the light direction and eye, used for specular component vec3 halfVector = normalize(lightDirN + normalize(-pixelPosition)); // Dot product of normal and light direction float NdotL = dot(normal, lightDirN); float attenuation = pow(lightReach / lightDistance, LightFalloff); // If pixel is lit by the light if(NdotL > 0) { // I have moved stuff from here to above so I can debug them. // Diffuse light color color += LightColor.rgb * NdotL * attenuation; // Specular light color color += LightColor.xyz * pow(max(dot(halfVector, normal), 0.0), 4.0) * attenuation; } RT0 = vec4(color, 1); //RT0 = vec4(pixelPosition, 1); //RT0 = vec4(depthSample, depthSample, depthSample, 1); //RT0 = vec4(NdotL, NdotL, NdotL, 1); RT0 = vec4(attenuation, attenuation, attenuation, 1); //RT0 = vec4(lightReach, lightReach, lightReach, 1); //RT0 = depthtexel; //RT0 = 100 / vec4(lightDistance, lightDistance, lightDistance, 1); //RT0 = vec4(lightDirN, 1); //RT0 = vec4(halfVector, 1); //RT0 = vec4(LightColor.xyz,1); //RT0 = vec4(LightViewPosition.xyz/100, 1); //RT0 = vec4(LightPosition.xyz, 1); //RT0 = vec4(normal,1); } What am I doing wrong here?

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• Fetching Partition Information

  - by Mike Femenella

  For a recent SSIS package at work I needed to determine the distinct values in a partition, the number of rows in each partition and the file group name on which each partition resided in order to come up with a grouping mechanism. Of course sys.partitions comes to mind for some of that but there are a few other tables you need to link to in order to grab the information required. The table I’m working on contains 8.8 billion rows. Finding the distinct partition keys from this table was not a fast operation. My original solution was to create  a temporary table, grab the distinct values for the partitioned column, then update via sys.partitions for the rows and the $partition function for the partitionid and finally look back to the sys.filegroups table for the filegroup names. It wasn’t pretty, it could take up to 15 minutes to return the results. The primary issue is pulling distinct values from the table. Queries for distinct against 8.8 billion rows don’t go quickly. A few beers into a conversation with a friend and we ended up talking about work which led to a conversation about the task described above. The solution was already built in SQL Server, just needed to pull it together. The first table I needed was sys.partition_range_values. This contains one row for each range boundary value for a partition function. In my case I have a partition function which uses dayid values. For example July 4th would be represented as an int, 20130704. This table lists out all of the dayid values which were defined in the function. This eliminated the need to query my source table for distinct dayid values, everything I needed was already built in here for me. The only caveat was that in my SSIS package I needed to create a bucket for any dayid values that were out of bounds for my function. For example if my function handled 20130501 through 20130704 and I had day values of 20130401 or 20130705 in my table, these would not be listed in sys.partition_range_values. I just created an “everything else” bucket in my ssis package just in case I had any dayid values unaccounted for. To get the number of rows for a partition is very easy. The sys.partitions table contains values for each partition. Easy enough to achieve by querying for the object_id and index value of 1 (the clustered index) The final piece of information was the filegroup name. There are 2 options available to get the filegroup name, sys.data_spaces or sys.filegroups. For my query I chose sys.filegroups but really it’s a matter of preference and data needs. In order to bridge between sys.partitions table and either sys.data_spaces or sys.filegroups you need to get the container_id. This can be done by joining sys.allocation_units.container_id to the sys.partitions.hobt_id. sys.allocation_units contains the field data_space_id which then lets you join in either sys.data_spaces or sys.file_groups. The end result is the query below, which typically executes for me in under 1 second. I’ve included the join to sys.filegroups and to sys.dataspaces, and I’ve  just commented out the join sys.filegroups. As I mentioned above, this shaves a good 10-15 minutes off of my original ssis package and is a really easy tweak to get a boost in my ETL time. Enjoy.

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• The gestures of Windows 8 (Consumer preview): part 2, More about Search

  - by Laurent Bugnion

  This is part 2 of a multipart blog post about the gestures and shortcuts in Windows 8 consumer preview. Part 1 can be found here! More about the Search charm In the first installment of this series, we talked about the charms and mentioned a few gestures to display the Search charm. Search is a very central and powerful feature in Windows 8, and allows you to search in Apps, Settings, Files and within Metro applications that support the Search contract. There are a few cool features around the Search, and especially the applications associated to it. I already mentioned the keyboard shortcuts you can use: Win-C shows the Charms bar (same as swiping from the right bevel towards the center of the screen). Win-Q open the Search fly out with Apps preselected. Win-W open the Search fly out with Settings preselected. Win-F open the Search fly out with Files preselected. Searching in Metro apps In addition to these three search domains, you can also search a Metro app, as long as it supports the Search contract (check this Build video to learn more about the Search contract). These apps show up in the Search flyout as shown here: Notice the list of apps below the Files button? That’s what we are talking about. First of all, the list order changes when you search in some applications. For instance, in the image above, I had used the Store with the Search charm. This is why the store shows up as the first app. I am not 100% what algorithm is used here (sorting according to number of searches is my guess), but try it out and try to figure it out Applications that have never been searched are sorted alphabetically. Does it mean we will see cool app names like ___AAA_MyCoolApp? I certainly hope not!! Pinning You can also pin often used apps to the Search flyout. To pin an app with the mouse, right click on it in the Search flyout and select Pin from the context menu. With the keyboard, use the arrow keys to go down to the selected app, and then open the context menu. With the finger, simply tap and hold until you see a semi transparent rectangle indicating that the context menu will be shown, then release. The context menu opens up and you can select Pin. Pin context menu Pinned apps Unpinning, Hiding Using the same technique as for pinning here above, you can also unpin a pinned application. Finally, you can also choose to hide an app from the Search flyout altogether. This is a convenient way to clean up and make it easy to find stuff. Note: At this point, I am not sure how to re-add a hidden app to the Search flyout. If anyone knows, please mention it in the comments, thanks! Reordering You can also reorder pinned apps. To do this, with the finger, tap, hold and pull the app to the side, then pull it vertically to reorder it. You can also reorder with the mouse, simply by clicking on an app and pulling it vertically to the place you want to put it. I don’t think there is a way to do that with the keyboard though. That’s it for now More gestures will follow in a next installment! Have fun with Windows 8   Laurent Bugnion (GalaSoft) Subscribe | Twitter | Facebook | Flickr | LinkedIn

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• MPI Cluster Debugger launch integration in VS2010

  Let's assume that you have all the HPC bits installed and that you have existing MPI code (or you created a "Hello World" project using the MPI project template). Of course, you create a single MPI application and at runtime it will correspond to multiple processes (of the same app) launched on multiple nodes (i.e. machines) on the cluster. So how do you debug such a situation by simply hitting the familiar "F5" keystroke (i.e. Debug - Start Debugging)?WATCH IT INSTEAD OF READING ABOUT ITIf you can't bear to read through all the details below, just watch this 19-minute screencast explaining this VS2010 feature. Alternatively, or even additionally, keep on reading.REQUIREMENTWhen you debug an MPI application, you would want the copying of resources from your client machine (where Visual Studio is installed) to each compute node (where Windows HPC Server is installed) to take place automatically for you. 'Resources' in the previous sentence includes your application binary, plus any binary or data dependencies it may have, plus PDBs if needed, plus the debug CRT of the correct bitness, plus msvsmon for remote debugging to work. You would also want, after copying is complete, to have your app and msvsmon launched and attached so that you can hit breakpoints back in Visual Studio on your client machine. All these thing that you would want are delivered in VS2010.STEPS TO F51. In your MPI project where you have placed a breakpoint go to Project Properties - Configuration Properties - Debugging. Ensure the "Debugger to launch" combo box value is set to MPI Cluster Debugger.2. There are a whole bunch of properties here and typically you can ignore all of them except one: Run Environment. By default it is set to run 1 process on your local machine and if you change the number after that to, for example, 4 it will launch 4 processes of your app on your local machine.You want this to run on your cluster though, so go to the dropdown arrow at the end of the Run Environment cell and open it to expose the "Edit Hpc node" menu which opens the Node Selector dialog:In this dialog you can enter (or pick from a list) the cluster head node name and then the number of processes you want to execute on the cluster and then hit OK and… you are done.3. Press F5 and watch your breakpoint get hit (after giving it some time for copying, remote execution, attachment and symbol resolution to take place).GOING DEEPERIn the MPI Cluster Debugger project properties above, you can see many additional properties to the Run Environment. They are all optional, but you may want to understand them in order to fine tune your cluster debugging. Read all about each one of these on the MSDN page Configuration Properties for the MPI Cluster Debugger.In the Node Selector dialog above you can see more options than just the Head Node name and Number of Process to run. They should be self-explanatory but I also cover them in depth in my screencast showing you an example of why you would choose to schedule processes per core versus per node. You can also read about these options on MSDN as part of the page How to: Configure and Launch the MPI Cluster Debugger.To read through an example that touches on MPI project creation, project properties, node selector, and also usage of MPI with OpenMP plus MPI with PPL, read the MSDN page Walkthrough: Launching the MPI Cluster Debugger in Visual Studio 2010.Happy MPI debugging! Comments about this post welcome at the original blog.

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• Make your TSQL easier to read during a presentation

  - by Jonathan Allen

  SQL Server Management Studio 2012 has some neat settings that you can use to help your presentations at a SQL event better for the attendees if you are willing to spend a few minutes making some settings changes. Historically, I have been reluctant to make changes to my SSMS settings as it is such a tedious process and it’s not 100% clear that what you think you are changing is actually what gets changed. With SSMS 2012 this has become a lot easier and a lot less risky. In any session that involves TSQL there is a trade off between the speaker having all the code on screen and the attendees being able to read any of what is on screen. You (the speaker) might be able to read this when you are working on the code but plenty of your audience wont be able to make head or tail of it. SSMS 2012 has a zoom facility that can help: but don’t go nuts … Having the font too big means you will be scrolling a lot and the code will again be rendered unreadable. There is more though but you need to take a deep breath and open the Tools menu and delve into the SSMS options. In previous versions of SSMS this is a deep, dark and scary place where changing values can be obscure and sometimes catastrophic to the UI when you get back to the code editor. First things first, we set out as a good DBA and save our current (and presumably acceptable) SSMS configuration. From the import and Export Settings you can set up a file to hold all of the settings that you currently have. The wizard will open and ask you to pick an option. This time around choose to export settings. hit next and next again and then name your settings profile in the final step of the wizard and then click Finish. Once this is done then you can change whatever you like and always get back to this configuration in a couple of clicks. So what can you change to make for a good experience? Well there are plenty of things that can be altered but don’t go too mad and change too many things without taking a look at the results for every item on the list above you can change font, size, weight, colour, background colour etc. etc. but consider what you are trying to achieve and take it slowly. I have seen presenters with their settings set to have a yellow highlight and black font rather than the default pale blue background and slightly darker font so to achieve that select Text Editor and then select “Selected Text” in the Display Items listbox. As you change things the Sample area give you an idea of what effect you are going to have. Black and yellow is the colour combination with the highest contrast – that’s why bees and wasps# are that colour. What next? how about increasing the default font for your demo scripts? This means that any script you open and any new ones that you start will take on this font. No more zooming (or forgetting to) in the middle of sessions. now don’t forget to save this profile – follow the same steps as above but give the profile a different name, something like PresentationBigFontHighContrast might be appropriate. Once you are done making changes, export the settings once more and then go into the Import Export wizard and import settings from the first profile you created. Everything will be back to normal. Now making changes to suit your environment can be done very easily and with confidence. * – and warning tape and safety signs and so forth – Health and Safety officers simply copy nature!

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• Converting .docx to pdf (or .doc to pdf, or .doc to odt, etc.) with libreoffice on a webserver on the fly using php

  - by robertphyatt

  Ok, so I needed to convert .docx files to .pdf files on the fly, but none of the free php libraries that were available let me do it on my server (a webservice was not good enough). Basically either I needed to pay for a library (and have it maybe suck) or just deal with the free ones that didn't convert the formatting well enough. Not good enough! I found that LibreOffice (OpenOffice's successor) allows command line conversion using the LibreOffice conversion engine (which DID preserve the formatting like I wanted and generally worked great). I loaded the latest version of Ubuntu (http://www.ubuntu.com/download/ubuntu/download) onto my Virtual Box (https://www.virtualbox.org/wiki/Downloads) on my computer and found that I was able to easily convert files using the commandline like this: libreoffice --headless -convert-to pdf fileToConvert.docx -outdir output/path/for/pdf I thought: sweet...but I don't have admin rights on my host's web server. I tried to use a "portable" version of LibreOffice that I obtained from http://portablelinuxapps.org/ but I was unable to get it to work on my host's webserver, because my host's webserver didn't have all the dependencies (Dependency Hell! http://en.wikipedia.org/wiki/Dependency_hell) I was at a loss of how to make it work, until I ran across a cool project made by a Ph.D. student (Philip J. Guo) at Stanford called CDE: http://www.stanford.edu/~pgbovine/cde.html I will let you look at his explanations of how it works (I followed what he did in http://www.youtube.com/watch?feature=player_embedded&v=6XdwHo1BWwY, starting at about 32:00 as well as the directions on his site), but in short, it allows one to avoid dependency hell by copying all the files used when you run certain commands, recreating the linux environment where the command worked. I was able to use this to run LibreOffice without having to resort to someone's portable version of it, and it worked just like it did when I did it on Ubuntu with the command above, with a tweak: I needed to run the wrapper of LibreOffice the CDE generated. So, below is my PHP code that calls it. In this code snippet, the filename to be copied is passed in as $_POST["filename"]. I copy the file to the same spot where I originally converted the file, convert it, copy it back and then delete all the files (so that it doesn't start growing exponentially). I did it this way because I wasn't able to make it work otherwise on the webserver. If there is a linux + webserver ninja out there that can figure out how to make it work without doing this, I would be interested to know what you did. Please post a comment or something if you did that. <?php //first copy the file to the magic place where we can convert it to a pdf on the fly copy($time.$_POST["filename"], "../LibreOffice/cde-package/cde-root/home/robert/Desktop/".$_POST["filename"]); //change to that directory chdir('../LibreOffice/cde-package/cde-root/home/robert'); //the magic command that does the conversion $myCommand = "./libreoffice.cde --headless -convert-to pdf Desktop/".$_POST["filename"]." -outdir Desktop/"; exec ($myCommand); //copy the file back copy("Desktop/".str_replace(".docx", ".pdf", $_POST["filename"]), "../../../../../documents/".str_replace(".docx", ".pdf", $_POST["filename"])); //delete all the files out of the magic place where we can convert it to a pdf on the fly $files1 = scandir('Desktop'); //my files that I generated all happened to start with a number. $pattern = '/^[0-9]/'; foreach ($files1 as $value) { preg_match($pattern, $value, $matches); if(count($matches) ?> 0) { unlink("Desktop/".$value); } } //changing the header to the location of the file makes it work well on androids header( 'Location: '.str_replace(".docx", ".pdf", $_POST["filename"]) ); ?> And here is the tar.gz file I generated I generated with CDE. To duplicate what I did exactly, put the tar.gz file in a folder somewhere. I will call that folder the "root". Make a new folder called "documents" in the "root" folder. Unpack the tar.gz and run the php script above from the "documents" folder. Success! I made a truly portable version of LibreOffice that can convert files on the fly on a webserver using 100% free, open source software!

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• AI to move custom-shaped spaceships (shape affecting movement behaviour)

  - by kaoD

  I'm designing a networked turn based 3D-6DOF space fleet combat strategy game which relies heavily on ship customization. Let me explain the game a bit, since you need to know a bit about it to set the question. What I aim for is the ability to create your own fleet of ships with custom shapes and attached modules (propellers, tractor beams...) which would give advantages and disadvantages to each ship, so you have lots of different fleet distributions. E.g., long ship with two propellers at the side would let the ship spin around that plane easily, bigger ships would move slowly unless you place lots of propellers at the back (therefore spending more "construction" points and energy when moving, and it will only move fast towards that direction.) I plan to balance all the game around this feature. The game would revolve around two phases: orders and combat phase. During the orders phase, you command the different ships. When all players finish the order phase, the combat phase begins and the ship orders get resolved in real-time for some time, then the action pauses and there's a new orders phase. The problem comes when I think about player input. To move a ship, you need to turn on or off different propellers if you want to steer, travel forward, brake, rotate in place... These propellers don't have to work at their whole power, so you can achieve more movement combinations with less propellers. I think this approach is a bit boring. The player doesn't want to fiddle with motors or anything, you just want to MOVE and KILL. The way I intend the player to give orders to these ships is by a destination and a rotation, and then the AI would calculate the correct propeller power to achive that movement and rotation. Propulsion doesn't have to be the same throught the entire turn calculation (after the orders have been given) so it would be cool if the ships reacted as they move, adjusting the power of the propellers for their needs dynamically, but it may be too hard to implement and it's not really needed for the game to work. In both cases, how would that AI decide which propellers to activate for the best (or at least not worst) trajectory to be achieved? I though about some approaches: Learning AI: The ship types would learn about their movement by trial and error, adjusting their behaviour with more uses, and finally becoming "smart". I don't want to get involved THAT far in AI coding, and I think it can be frustrating for the player (even if you can let it learn without playing.) Pre-calculated timestep movement: Upon ship creation, ALL possible movements are calculated for each propeller configuration and power for a given delta-time. Memory intensive, ugly, bad. Pre-calculated trajectories: The same as above but not for each delta-time but the whole trajectory, which would then be fitted as much as possible. Requires a fixed propeller configuration for the whole combat phase and is still memory intensive, ugly and bad. Continuous brute forcing: The AI continously checks ALL possible propeller configurations throughout the entire combat phase, precalculates a few time steps and decides which is the best one based on that. Con: what's good now might not be that good later, and it's too CPU intensive, ugly, and bad too. Single brute forcing: Same as above, but only brute forcing at the beginning of the simulation, so it needs constant propeller configuration throughout the entire combat phase. Coninuous angle check: This is not a full movement method, but maybe a way to discard "stupid" propeller configurations. Given the current propeller's normal vector and the final one, you can approximate the power needed for the propeller based on the angle. You must do this continuously throughout the whole combat phase. I figured this one out recently so I didn't put in too much thought. A priori, it has the "what's good now might not be that good later" drawback too, and it doesn't care about the other propellers which may act together to make a better propelling configuration. I'm really stuck here. Any ideas?

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• Rotation of viewplatform in Java3D

  - by user29163

  I have just started with Java3D programming. I thought I had built up some basic intuition about how the scene graph works, but something that should work, does not work. I made a simple program for rotating a pyramid around the y-axis. This was done just by adding a RotationInterpolator R to the TransformGroup above the pyramid. Then I thought hey, can I now remove the RotationInterpolator from this TransformGroup, then add it to the TransformGroup above my ViewPlatform leaf. This should work if I have understood how things work. Adding the RotationInterpolator to this TransformGroup, should make the children of this TransformGroup rotate, and the ViewingPlatform is a child of the TransformGroup. Any ideas on where my reasoning is flawed? Here is the code for setting up the universe, and the view branchgroup. import java.awt.*; import java.awt.event.*; import javax.media.j3d.*; import javax.vecmath.*; public class UniverseBuilder { // User-specified canvas Canvas3D canvas; // Scene graph elements to which the user may want access VirtualUniverse universe; Locale locale; TransformGroup vpTrans; View view; public UniverseBuilder(Canvas3D c) { this.canvas = c; // Establish a virtual universe that has a single // hi-res Locale universe = new VirtualUniverse(); locale = new Locale(universe); // Create a PhysicalBody and PhysicalEnvironment object PhysicalBody body = new PhysicalBody(); PhysicalEnvironment environment = new PhysicalEnvironment(); // Create a View and attach the Canvas3D and the physical // body and environment to the view. view = new View(); view.addCanvas3D(c); view.setPhysicalBody(body); view.setPhysicalEnvironment(environment); // Create a BranchGroup node for the view platform BranchGroup vpRoot = new BranchGroup(); // Create a ViewPlatform object, and its associated // TransformGroup object, and attach it to the root of the // subgraph. Attach the view to the view platform. Transform3D t = new Transform3D(); Transform3D s = new Transform3D(); t.set(new Vector3f(0.0f, 0.0f, 10.0f)); t.rotX(-Math.PI/4); s.set(new Vector3f(0.0f, 0.0f, 10.0f)); //forandre verdier her for å endre viewing position t.mul(s); ViewPlatform vp = new ViewPlatform(); vpTrans = new TransformGroup(t); vpTrans.setCapability(TransformGroup.ALLOW_TRANSFORM_WRITE); // Rotator stuff Transform3D yAxis = new Transform3D(); //yAxis.rotY(Math.PI/2); Alpha rotationAlpha = new Alpha( -1, Alpha.INCREASING_ENABLE, 0, 0,4000, 0, 0, 0, 0, 0); RotationInterpolator rotator = new RotationInterpolator( rotationAlpha, vpTrans, yAxis, 0.0f, (float) Math.PI*2.0f); RotationInterpolator rotator2 = new RotationInterpolator( rotationAlpha, vpTrans); BoundingSphere bounds = new BoundingSphere(new Point3d(0.0,0.0,0.0), 1000.0); rotator.setSchedulingBounds(bounds); vpTrans.addChild(rotator); vpTrans.addChild(vp); vpRoot.addChild(vpTrans); view.attachViewPlatform(vp); // Attach the branch graph to the universe, via the // Locale. The scene graph is now live! locale.addBranchGraph(vpRoot); } public void addBranchGraph(BranchGroup bg) { locale.addBranchGraph(bg); } }

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• Making AI jump on a spot effectively

  - by Pasquale Sada

  How to calculate, in 3D environment, the closest point, from which an AI character can jump onto a platform? Setup I have an initial velocity V(Vx,Vy,VZ) and a spot where the character stands still at S(Sx,Sy,Sz). What I'm trying to achieve is a successful jump on a spot E(Ex,Ey,Ez) where you have clicked on(only lower or higher spot, because I've in place a simple steering behavior for even terrains). There are no obstacles around. I've implemented a formula that can make him jump in a precise way on a spot but you need to declare an angle: the problem arise when the selected spot is straight above your head. It' pretty lame that the char hang there and can reach a thing that is 1cm above is head. I'll share the code I'm using: Vector3 dir = target - transform.position; // get target direction float h = dir.y; // get height difference dir.y = 0; // retain only the horizontal direction float dist = dir.magnitude ; // get horizontal distance float a = angle * Mathf.Deg2Rad; // convert angle to radians dir.y = dist * Mathf.Tan(a); // set dir to the elevation angle dist += h / Mathf.Tan(a); // correct for small height differences // calculate the velocity magnitude float vel = Mathf.Sqrt(dist * Physics.gravity.magnitude / Mathf.Sin(2 *a)); return vel * dir.normalized; Ended up using the lowest angle (20 degree) and checking for collision on the trajectory. If found any increase the angle. Here some code (to improve the code maybe must stop the check at the highest point of the curve): Vector3 BallisticVel(Vector3 target, float angle) { Vector3 dir = target - transform.position; // get target direction float h = dir.y; // get height difference dir.y = 0; // retain only the horizontal direction float dist = dir.magnitude ; // get horizontal distance float a = angle * Mathf.Deg2Rad; // convert angle to radians dir.y = dist * Mathf.Tan(a); // set dir to the elevation angle dist += h / Mathf.Tan(a); // correct for small height differences // calculate the velocity magnitude float vel = Mathf.Sqrt(dist * Physics.gravity.magnitude / Mathf.Sin(2 * a)); return vel * dir.normalized; } Vector3 TrajectoryPoint(Vector3 startingPosition, Vector3 startingVelocity, float n ) { float t = 1/60 ; // seconds per time step Vector3 stepVelocity = t * startingVelocity; // m/s Vector3 stepGravity = t * t * Physics.gravity; // m/s/s return startingPosition + n * stepVelocity + 0.5f * (n*n+n) * stepGravity; } bool CheckTrajectory(Vector3 startingPosition,Vector3 target, float angle_jump) { Debug.Log("checking"); if(angle_jump < 80f) { Debug.Log("if"); Vector3 startingVelocity = BallisticVel(target, angle_jump); for (int i = 0; i < 180; i++) { //Debug.Log(i); Vector3 trajectoryPosition = TrajectoryPoint( startingPosition, startingVelocity, i ); if(Physics.Raycast(trajectoryPosition,Vector3.forward,safeDistance)) { angle_jump += 10; break; // restart loop with the new angle } else continue; } return true; JumpVelocity = BallisticVel(target, angle_jump); } return false; }

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• ARTS Reference Model for Retail

  - by Sanjeev Sharma

  Consider a hypothetical scenario where you have been tasked to set up retail operations for a electronic goods or daily consumables or a luxury brand etc. It is very likely you will be faced with the following questions: What are the essential business capabilities that you must have in place?  What are the essential business activities under-pinning each of the business capabilities, identified in Step 1? What are the set of steps that you need to perform to execute each of the business activities, identified in Step 2? Answers to the above will drive your investments in software and hardware to enable the core retail operations. More importantly, the choices you make in responding to the above questions will several implications in the short-run and in the long-run. In the short-term, you will incur the time and cost of defining your technology requirements, procuring the software/hardware components and getting them up and running. In the long-term, as you grow in operations organically or through M&A, partnerships and franchiser business models  you will invariably need to make more technology investments to manage the greater complexity (scale and scope) of business operations.  "As new software applications, such as time & attendance, labor scheduling, and POS transactions, just to mention a few, are introduced into the store environment, it takes a disproportionate amount of time and effort to integrate them with existing store applications. These integration projects can add up to 50 percent to the time needed to implement a new software application and contribute significantly to the cost of the overall project, particularly if a systems integrator is called in. This has been the reality that all retailers have had to live with over the last two decades. The effect of the environment has not only been to increase costs, but also to limit retailers' ability to implement change and the speed with which they can do so." (excerpt taken from here) Now, one would think a lot of retailers would have already gone through the pain of finding answers to these questions, so why re-invent the wheel? Precisely so, a major effort began almost 17 years ago in the retail industry to make it less expensive and less difficult to deploy new technology in stores and at the retail enterprise level. This effort is called the Association for Retail Technology Standards (ARTS). Without standards such as those defined by ARTS, you would very likely end up experiencing the following: Increased Time and Cost due to resource wastage arising from re-inventing the wheel i.e. re-creating vanilla processes from scratch, and incurring, otherwise avoidable, mistakes and errors by ignoring experience of others Sub-optimal Process Efficiency due to narrow, isolated view of processes thereby ignoring process inter-dependencies i.e. optimizing parts but not the whole, and resulting in lack of transparency and inter-departmental finger-pointing Embracing ARTS standards as a blue-print for establishing or managing or streamlining your retail operations can benefit you in the following ways: Improved Time-to-Market from parity with industry best-practice processes e.g. ARTS, thus avoiding “reinventing the wheel” for common retail processes and focusing more on customizing processes for differentiations, and lowering integration complexity and risk with a standardized vocabulary for exchange between internal and external i.e. partner systems Lower Operating Costs by embracing the ARTS enterprise-wide process reference model for developing and streamlining retail operations holistically instead of a narrow, silo-ed view, and  procuring IT systems in compliance with ARTS thus avoiding IT budget marginalization While parity with industry standards such as ARTS business process model by itself does not create a differentiation, it does however provide a higher starting point for bridging the strategy-execution gap in setting up and improving retail operations.

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• Why won't my Broadcom BCM4312 LP-PHY work with the STA driver?

  - by Jackson Taylor

  I tried the steps here for a 4312: https://help.ubuntu.com/community/WifiDocs/Driver/bcm43xx Both of these: sudo modprobe -r b43 ssb wl sudo modprobe wl return: FATAL: Module wl not found. FATAL: Error running install command for wl (this one is only for the second one actually) I tried the broadcom-sta, didn't work. What's confusing is down below in the next steps for STA with internet access it says to use the bcmwl one. So I install that and it succeeds but with some errors: sudo apt-get install bcmwl-kernel-source Reading package lists... Done Building dependency tree Reading state information... Done The following package was automatically installed and is no longer required: module-assistant Use 'apt-get autoremove' to remove it. The following NEW packages will be installed: bcmwl-kernel-source 0 upgraded, 1 newly installed, 0 to remove and 0 not upgraded. Need to get 0 B/1,181 kB of archives. After this operation, 3,609 kB of additional disk space will be used. Selecting previously unselected package bcmwl-kernel-source. (Reading database ... 168005 files and directories currently installed.) Unpacking bcmwl-kernel-source (from .../bcmwl-kernel-source_5.100.82.112+bdcom-0ubuntu3_amd64.deb) ... Setting up bcmwl-kernel-source (5.100.82.112+bdcom-0ubuntu3) ... Loading new bcmwl-5.100.82.112+bdcom DKMS files... Building only for 3.5.0-21-generic Building for architecture x86_64 Module build for the currently running kernel was skipped since the kernel source for this kernel does not seem to be installed. ERROR: Module b43 does not exist in /proc/modules ERROR: Module b43legacy does not exist in /proc/modules ERROR: Module ssb does not exist in /proc/modules ERROR: Module bcm43xx does not exist in /proc/modules ERROR: Module brcm80211 does not exist in /proc/modules ERROR: Module brcmfmac does not exist in /proc/modules ERROR: Module brcmsmac does not exist in /proc/modules ERROR: Module bcma does not exist in /proc/modules FATAL: Module wl not found. FATAL: Error running install command for wl update-initramfs: deferring update (trigger activated) Processing triggers for initramfs-tools ... update-initramfs: Generating /boot/initrd.img-3.5.0-21-generic jtaylor991@jtaylor991-whiteHP:~$ sudo modprobe wl FATAL: Module wl not found. FATAL: Error running install command for wl Then I do the modprobe wl commands listed above and it gives the above listed errors. It didn't work with the broadcom-sta driver either. I installed the b43 ones but nothing happened, and I don't know why so those are still installed. firmware-b43legacy-installer, b43-fwcutter and firmware-b43-lpphy-installer (yes it is a LP-PHY) are currently installed. If I go into System Settings Software Sources Additional Drivers it says "Using Broadcom 802.11 Linux STA wireless driver source from bcmwl-kernel-source (proprietary) But bcmwl-kernel-source isn't installed. I could try again but I remember rebooting and it still said this. What's funny is it found wireless networks during the Ubuntu setup/installation, I don't remember if I got it to connect or not though. I think it kept asking for a password when I put it in (yes it was right I showed password and looked at it) so I just ignored it. But right now the Enable Wireless option in the top right is just gone, it's just Enable Networking and I'm on ethernet on this HP Pavilion dv4-1435dx right here. If I run rfkill list it shows: 0: hp-wifi: Wireless LAN Soft blocked: no Hard blocked: no It was hard blocked at the beginning but unblocking it makes no change. Also it's a touch sensitive button, and it appears to be always orange no matter if it's enabled or not because when I touch it the hard blocked changes between yes and no in rfkill list. I think it was blue for a minute at one point. What is going on?!?! Help me! Lol, thanks for any and all of your time guys. Oh yeah this is Ubuntu 12.10 fresh install.

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• Web Safe Area (optimal resolution) for web app design?

  - by M.A.X

  I'm in the process of designing a new web app and I'm wondering for what 'Web Safe Area' should I optimize the app layout and design. By Web Safe Area I mean the actual area available to display the website in the browser (which is influenced by monitor resolution as well as the space taken up by the browser and OS) I did some investigation and thinking on my own but wanted to share this to see what the general opinion is. Here is what I found: Optimal Display Resolution: w3schools web stats seems to be the most referenced source (however they state that these are results from their site and is biased towards tech savvy users) http://www.w3counter.com/globalstats.php (aggregate data from something like 15,000 different sites that use their tracking services) StatCounter Global Stats Display Resolution (Stats are based on aggregate data collected by StatCounter on a sample exceeding 15 billion pageviews per month collected from across the StatCounter network of more than 3 million websites) NetMarketShare Screen Resolutions (marketshare.hitslink.com) (a web analytics consulting firm, they get data from browsers of site visitors to their on-demand network of live stats customers. The data is compiled from approximately 160 million visitors per month) Display Resolution Summary: There is a bit of variation between the above sources but in general as of Jan 2011 looks like 1024x768 is about 20%, while ~85% have a higher resolution of at least 1280x768 (1280x800 is the most common of these with 15-20% of total web, depending on the source; 1280x1024 and 1366x768 follow behind with 9-14% of the share). My guess would be that the higher resolution values will be even more common if we filter on North America, and even higher if we filter on N.American corporate users (unfortunately I couldn't find any free geographically filtered statistics). Another point to note is that the 1024x768 desktop user population is likely lower than the aforementioned 20%, seeing as the iPad (1024x768 native display) is likely propping up those number (the app I'm designing is flash based, Apple mobile devices don't support flash so iPad support isn't a concern). My recommendation would be to optimize around the 1280x768 constraint (*note: 1280x768 is actually a relatively rare resolution, but I think it's a valid constraint range considering that 1366x768 is relatively common and 1280 is the most common horizontal resolution). Browser + OS Constraints: To further add to the constraints we have to subtract the space taken up by the browser (assuming IE, which is the most space consuming) and the OS (assuming WinXP-Win7): Win7 has the biggest taskbar footprint at a height of 40px (XP's and Vista's is 30px) The default IE8 view uses up 25px at the bottom of the screen with the status bar and a further 120px at the top of the screen with the windows title bar and the browser UI (assuming the default 'favorites' toolbar is present, it would instead be 91px without the favorites toolbar). Assuming no scrollbar, we also loose a total of 4px horizontally for the window outline. This means that we are left with 583px of vertical space and 1276px of horizontal. In other words, a Web Safe Area of 1276 x 583 Is this a correct line of thinking? I'm really surprised that I couldn't find this type of investigation anywhere on the web. Lots of websites talk about designing for 1024x768, but that's only half the equation! There is no mention of browser/OS influences on the actual area you have to display the site/app. Any help on this would be greatly appreciated! Thanks. EDIT Another caveat to my line of thinking above is that different browsers actually take up different amounts of pixels based on the OS they're running on. For example, under WinXP IE8 takes up 142px on top of the screen (instead the aforementioned 120px for Win7) because the file menu shows up by default on XP while in Win7 the file menu is hidden by default. So it looks like on WinXP + IE8 the Web Safe Area would be a mere 572px (768px-142-30-24=572)

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