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• Clustering for Mere Mortals (Pt3)

  - by Geoff N. Hiten

  The Controller Now we get to the meat of the matter.  You want a virtual cluster, the first thing you have to do is create your own portable domain.  IStart with a plain vanilla install of Windows 2003 R2 Standard on a semi-default VM. (1 GB RAM, 2 cores, 2 NICs, 128GB dynamically expanding VHD file).  I chose this because it had the smallest disk and memory footprint of any current supported Microsoft Server product.  I created the VM with a single dynamically expanding VHD, one fixed 16 GB VHD, and two NICs.  One NIC is connected to the outside world and the other one is part of an internal-only network.  The first NIC is set up as a DHCP client.  We will get to the other one later. I actually tried this with Windows 2008 R2, but it failed miserably.  Not sure whether it was 2008 R2 or the fact I tried to use cloned VMs in the cluster.  Clustering is one place where NewSID would really come in handy.  Too bad Microsoft bought and buried it. Load and Patch the OS (hence the need for the outside connection).This is a good time to go get dinner.  Maybe a movie too.  There are close to a hundred patches that need to be downloaded and applied.  Avoiding that mess was why I put so much time into trying to get the 2008 R2 version working.  Maybe next time.  Don’t forget to add the extensions for VMLite (or whatever virtualization product you prefer). Set a fixed IP address on the internal-only NIC.  Do not give it a gateway.  Put the same IP address for the NIC and for the DNS Server.  This IP should be in a range that is never available on your public network.  You will need all the addresses in the range available.  See the previous post for the exact settings I used. I chose 10.97.230.1 as the server.  The rest of the 10.97.230 range is what I will use later.  For the curious, those numbers are based on elements of my home address.  Not truly random, but good enough for this project. Do not bridge the network connections.  I never allowed the cluster nodes direct access to any public network. Format the fixed VHD and leave it alone for now. Promote the VM to a Domain Controller.  If you have never done this, don’t worry.  The only meaningful decision is what to call the new domain.  I prefer a bogus name that does not correspond to a real Top-Level Domain (TLD).  .com, .biz., .net, .org  are all TLDs that we know and love.  I chose .test as the TLD since it is descriptive AND it does not exist in the real world.  The domain is called MicroAD.  This gives me MicroAD.Test as my domain. During the promotion process, you will be prompted to install DNS as part of the Domain creation process.  You want to accept this option.  The installer will automatically assign this DNS server as the authoritative owner of the MicroAD.test DNS domain (not to be confused with the MicroAD.test Active Directory domain.) For the rest of the DCPROMO process, just accept the defaults. Now let’s make our IP address management easy.  Add the DHCP Role to the server.  Add the server (10.97.230.1 in this case) as the default gateway to assign to DHCP clients.  Here is where you have to be VERY careful and bind it ONLY to the Internal NIC.  Trust me, your network admin will NOT like an extra DHCP server “helping” out on her network.  Go ahead and create a range of 10-20 IP Addresses in your scope.  You might find other uses for a pocket domain controller <cough> Mirroring </cough> than just for building a cluster.  And Clustering in SQL 2008 and Windows 2008 R2 fully supports DHCP addresses. Now we have three of the five key roles ready.  Two more to go. Next comes file sharing.  Since your cluster node VMs will not have access to any outside, you have to have some way to get files into these VMs.  I simply go to the root of C: and create a “Shared” folder.  I then share it out and grant full control to “Everyone” to both the share and to the underlying NTFS folder.   This will be immensely useful for Service Packs, demo databases, and any other software that isn’t packaged as an ISO that we can mount to the VM. Finally we need to create a block-level multi-connect storage device.  The kind folks at Starwinds Software (http://www.starwindsoftware.com/) graciously gave me a non-expiring demo license for expressly this purpose.  Their iSCSI SAN software lets you create an iSCSI target from nearly any storage medium.  Refreshingly, their product does exactly what they say it does.  Thanks. Remember that 16 GB VHD file?  That is where we are going to carve into our LUNs.  I created an iSCSI folder off the root, just so I can keep everything organized.  I then carved 5 ea. 2 GB iSCSI targets from that folder.  I chose a fixed VHD for performance.  I tried this earlier with a dynamically expanding VHD, but too many layers of abstraction and sparseness combined to make it unusable even for a demo.  Stick with a fixed VHD so there is a one-to-one mapping between abstract and physical storage.  If you read the previous post, you know what I named these iSCSI LUNs and why.  Yes, I do have some left over space.  Always leave yourself room for future growth or options. This gets us up to where we can actually build the nodes and install SQL.  As with most clusters, the real work happens long before the individual nodes get installed and configured.  At least it does if you want the cluster to be a true high-availability platform.

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• Applications: How to create a custom dialog box for Windows Mobile 6 (native)

  - by TechTwaddle

  Ashraf, on the MSDN forum, asks, “Is there a way to make a default choice for the messagebox that happens after a period of time if the user doesn't choose (Clicked ) Yes or No buttons.” To elaborate, the requirement is to show a message box to the user with certain options to select, and if the user does not respond within a predefined time limit (say 8 seconds) then the message box must dismiss itself and select a default option. Now such a functionality is not available with the MessageBox() api, you will have to write your own custom dialog box. Surely, creating a dialog box is quite a simple task using the DialogBox() api, and we have been creating full screen dialog boxes all the while. So how will this custom message box be any different? It’s not much different from a regular dialog box except for a few changes in its properties. First, it has a title bar but no buttons on the title bar (no ‘x’ or ‘ok’ button on the title bar), it doesn’t occupy full screen and it contains the controls that you put into it, thus justifying the title ‘custom’. So in this post we create a custom dialog box with two buttons, ‘Black’ and ‘White’. The user is given 8 seconds to select one of those colours, if the user doesn’t make a selection in 8 seconds, the default option ‘Black’ is selected. Before going into the implementation here is a video of how the dialog box works; Custom dialog box To start off, add a new dialog resource into your application, size it appropriately and add whatever controls you need to the dialog. In my case, I added two static text labels and two buttons, as below; Now we need to write up the window procedure for this dialog, here is the complete function; BOOL CALLBACK CustomDialogProc(HWND hDlg, UINT uMessage, WPARAM wParam, LPARAM lParam) {     int wmID, wmEvent;     PAINTSTRUCT ps;     HDC hdc;     static int timeCount = 0;     switch(uMessage)     {         case WM_INITDIALOG:             {                 SHINITDLGINFO shidi;                 memset(&shidi, 0, sizeof(shidi));                 shidi.dwMask = SHIDIM_FLAGS;                 //shidi.dwFlags = SHIDIF_DONEBUTTON | SHIDIF_SIPDOWN | SHIDIF_SIZEDLGFULLSCREEN | SHIDIF_EMPTYMENU;                 shidi.dwFlags = SHIDIF_SIPDOWN | SHIDIF_EMPTYMENU;                 shidi.hDlg = hDlg;                 SHInitDialog(&shidi);                 SHDoneButton(hDlg, SHDB_HIDE);                 timeCount = 0;                 SetWindowText(GetDlgItem(hDlg, IDC_STATIC_TIME_REMAINING), L"Time remaining: 8 second(s)");                 SetTimer(hDlg, MY_TIMER, 1000, NULL);             }             return TRUE;         case WM_COMMAND:             {                 wmID = LOWORD(wParam);                 wmEvent = HIWORD(wParam);                 switch(wmID)                 {                     case IDC_BUTTON_BLACK:                         KillTimer(hDlg, MY_TIMER);                         EndDialog(hDlg, IDC_BUTTON_BLACK);                         break;                     case IDC_BUTTON_WHITE:                         KillTimer(hDlg, MY_TIMER);                         EndDialog(hDlg, IDC_BUTTON_WHITE);                         break;                 }             }             break;         case WM_TIMER:             {                 if (wParam == MY_TIMER)                 {                     WCHAR wszText[128];                     memset(&wszText, 0, sizeof(wszText));                     timeCount++;                     //8 seconds are over, dismiss the dialog, select def value                     if (timeCount >= 8)                     {                         KillTimer(hDlg, MY_TIMER);                         EndDialog(hDlg, IDC_BUTTON_BLACK_DEF);                     }                     wsprintf(wszText, L"Time remaining: %d second(s)", 8-timeCount);                     SetWindowText(GetDlgItem(hDlg, IDC_STATIC_TIME_REMAINING), wszText);                     UpdateWindow(GetDlgItem(hDlg, IDC_STATIC_TIME_REMAINING));                 }             }             break;         case WM_PAINT:             {                 hdc = BeginPaint(hDlg, &ps);                 EndPaint(hDlg, &ps);             }             break;     }     return FALSE; } The MSDN documentation mentions that you need to specify the flag WS_NONAVDONEBUTTON, but I got an error saying that the value could not be found, so we can ignore this for now. Next up, while calling SHInitDialog() for your custom dialog, make sure that you don’t specify SHDIF_DONEBUTTON in the dwFlags member of the SHINITDIALOG structure, this member makes the ‘ok’ button appear on the dialog title bar. Finally, we need to call SHDoneButton() with SHDB_HIDE flag to, well, hide the Done button. The ‘Done’ button is the same as the ‘ok’ button, so this step might seem redundant, and the dialog works fine without calling SHDoneButton() too, but it’s better to stick with the documentation (; So you can see that we have followed all these steps above, under WM_INITDIALOG. We also setup a few things like a variable to keep track of the time, and setting off a one second timer. Every time the timer fires, we receive a WM_TIMER message. We then update the static label displaying the amount of time left to the user. If 8 seconds go by without the user selecting any option, we kill the timer and end the dialog with IDC_BUTTON_BLACK_DEF. This is just a #define’d integer value, make sure it’s unique. You’ll see why this is important. If the user makes a selection, either Black or White, we kill the timer and end the dialog with corresponding selection the user made, that is, either IDC_BUTTON_BLACK or IDC_BUTTON_WHITE. Ok, so now our custom dialog is ready to be used. I invoke the custom dialog from a menu entry in the main windows as below, case IDM_MENU_CUSTOMDLG:     {         int ret = DialogBox(g_hInst, MAKEINTRESOURCE(IDD_CUSTOM_DIALOG), hWnd, CustomDialogProc);         switch(ret)         {             case IDC_BUTTON_BLACK_DEF:                 SetWindowText(g_hStaticSelection, L"You Selected: Black (default)");                 break;             case IDC_BUTTON_BLACK:                 SetWindowText(g_hStaticSelection, L"You Selected: Black");                 break;             case IDC_BUTTON_WHITE:                 SetWindowText(g_hStaticSelection, L"You Selected: White");                 break;         }         UpdateWindow(g_hStaticSelection);     }     break; So you see why ending the dialog with the corresponding value was important, that’s what the DialogBox() api returns with. And in the main window I update a static text label to show which option was selected. I cranked this out in about an hour, and unfortunately don’t have time for a managed C# version. That will have to be another post, if I manage to get it working that is (;

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• Working With Extended Events

  - by Fatherjack

  SQL Server 2012 has made working with Extended Events (XE) pretty simple when it comes to what sessions you have on your servers and what options you have selected and so forth but if you are like me then you still have some SQL Server instances that are 2008 or 2008 R2. For those servers there is no built-in way to view the Extended Event sessions in SSMS. I keep coming up against the same situations – Where are the xel log files? What events, actions or predicates are set for the events on the server? What sessions are there on the server already? I got tired of this being a perpetual question and wrote some TSQL to save as a snippet in SQL Prompt so that these details are permanently only a couple of clicks away. First, some history. If you just came here for the code skip down a few paragraphs and it’s all there. If you want a little time to reminisce about SQL Server then stick with me through the next paragraph or two. We are in a bit of a cross-over period currently, there are many versions of SQL Server but I would guess that SQL Server 2008, 2008 R2 and 2012 comprise the majority of installations. With each of these comes a set of management tools, of which SQL Server Management Studio (SSMS) is one. In 2008 and 2008 R2 Extended Events made their first appearance and there was no way to work with them in the SSMS interface. At some point the Extended Events guru Jonathan Kehayias (http://www.sqlskills.com/blogs/jonathan/) created the SQL Server 2008 Extended Events SSMS Addin which is really an excellent tool to ease XE session administration. This addin will install in SSMS 2008 or 2008R2 but not SSMS 2012. If you use a compatible version of SSMS then I wholly recommend downloading and using it to make your work with XE much easier. If you have SSMS 2012 installed, and there is no reason not to as it will let you work with all versions of SQL Server, then you cannot install this addin. If you are working with SQL Server 2012 then SSMS 2012 has built in functionality to manage XE sessions – this functionality does not apply for 2008 or 2008 R2 instances though. This means you are somewhat restricted and have to use TSQL to manage XE sessions on older versions of SQL Server. OK, those of you that skipped ahead for the code, you need to start from here: So, you are working with SSMS 2012 but have a SQL Server that is an earlier version that needs an XE session created or you think there is a session created but you aren’t sure, or you know it’s there but can’t remember if it is running and where the output is going. How do you find out? Well, none of the information is hidden as such but it is a bit of a wrangle to locate it and it isn’t a lot of code that is unlikely to remain in your memory. I have created two pieces of code. The first examines the SYS.Server_Event_… management views in combination with the SYS.DM_XE_… management views to give the name of all sessions that exist on the server, regardless of whether they are running or not and two pieces of TSQL code. One piece will alter the state of the session: if the session is running then the code will stop the session if executed and vice versa. The other piece of code will drop the selected session. If the session is running then the code will stop it first. Do not execute the DROP code unless you are sure you have the Create code to hand. It will be dropped from the server without a second chance to change your mind. /**************************************************************/ /***   To locate and describe event sessions on a server    ***/ /***                                                        ***/ /***   Generates TSQL to start/stop/drop sessions           ***/ /***                                                        ***/ /***        Jonathan Allen - @fatherjack                    ***/ /***                 June 2013                                ***/ /***                                                        ***/ /**************************************************************/ SELECT  [EES].[name] AS [Session Name - all sessions] ,         CASE WHEN [MXS].[name] IS NULL THEN ISNULL([MXS].[name], 'Stopped')              ELSE 'Running'         END AS SessionState ,         CASE WHEN [MXS].[name] IS NULL              THEN ISNULL([MXS].[name],                          'ALTER EVENT SESSION [' + [EES].[name]                          + '] ON SERVER STATE = START;')              ELSE 'ALTER EVENT SESSION [' + [EES].[name]                   + '] ON SERVER STATE = STOP;'         END AS ALTER_SessionState ,         CASE WHEN [MXS].[name] IS NULL              THEN ISNULL([MXS].[name],                          'DROP EVENT SESSION [' + [EES].[name]                          + '] ON SERVER; -- This WILL drop the session. It will no longer exist. Don't do it unless you are certain you can recreate it if you need it.')              ELSE 'ALTER EVENT SESSION [' + [EES].[name]                   + '] ON SERVER STATE = STOP; ' + CHAR(10)                   + '-- DROP EVENT SESSION [' + [EES].[name]                   + '] ON SERVER; -- This WILL stop and drop the session. It will no longer exist. Don't do it unless you are certain you can recreate it if you need it.'         END AS DROP_Session FROM    [sys].[server_event_sessions] AS EES         LEFT JOIN [sys].[dm_xe_sessions] AS MXS ON [EES].[name] = [MXS].[name] WHERE   [EES].[name] NOT IN ( 'system_health', 'AlwaysOn_health' ) ORDER BY SessionState GO I have excluded the system_health and AlwaysOn sessions as I don’t want to accidentally execute the drop script for these sessions that are created as part of the SQL Server installation. It is possible to recreate the sessions but that is a whole lot of aggravation I’d rather avoid. The second piece of code gathers details of running XE sessions only and provides information on the Events being collected, any predicates that are set on those events, the actions that are set to be collected, where the collected information is being logged and if that logging is to a file target, where that file is located. /**********************************************/ /***    Running Session summary                ***/ /***                                        ***/ /***    Details key values of XE sessions     ***/ /***    that are in a running state            ***/ /***                                        ***/ /***        Jonathan Allen - @fatherjack    ***/ /***        June 2013                        ***/ /***                                        ***/ /**********************************************/ SELECT  [EES].[name] AS [Session Name - running sessions] ,         [EESE].[name] AS [Event Name] ,         COALESCE([EESE].[predicate], 'unfiltered') AS [Event Predicate Filter(s)] ,         [EESA].[Action] AS [Event Action(s)] ,         [EEST].[Target] AS [Session Target(s)] ,         ISNULL([EESF].[value], 'No file target in use') AS [File_Target_UNC] -- select * FROM    [sys].[server_event_sessions] AS EES         INNER JOIN [sys].[dm_xe_sessions] AS MXS ON [EES].[name] = [MXS].[name]         INNER JOIN [sys].[server_event_session_events] AS [EESE] ON [EES].[event_session_id] = [EESE].[event_session_id]         LEFT JOIN [sys].[server_event_session_fields] AS EESF ON ( [EES].[event_session_id] = [EESF].[event_session_id]                                                               AND [EESF].[name] = 'filename'                                                               )         CROSS APPLY ( SELECT    STUFF(( SELECT  ', ' + sest.name                                         FROM    [sys].[server_event_session_targets]                                                 AS SEST                                         WHERE   [EES].[event_session_id] = [SEST].[event_session_id]                                       FOR                                         XML PATH('')                                       ), 1, 2, '') AS [Target]                     ) AS EEST         CROSS APPLY ( SELECT    STUFF(( SELECT  ', ' + [sesa].NAME                                         FROM    [sys].[server_event_session_actions]                                                 AS sesa                                         WHERE   [sesa].[event_session_id] = [EES].[event_session_id]                                       FOR                                         XML PATH('')                                       ), 1, 2, '') AS [Action]                     ) AS EESA WHERE   [EES].[name] NOT IN ( 'system_health', 'AlwaysOn_health' ) /*Optional to exclude 'out-of-the-box' traces*/ I hope that these scripts are useful to you and I would be obliged if you would keep my name in the script comments. I have no problem with you using it in production or personal circumstances, however it has no warranty or guarantee. Don’t use it unless you understand it and are happy with what it is going to do. I am not ever responsible for the consequences of executing this script on your servers.

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• Self-signed certificates for a known community

  - by costlow

  Recently announced changes scheduled for Java 7 update 51 (January 2014) have established that the default security slider will require code signatures and the Permissions Manifest attribute. Code signatures are a common practice recommended in the industry because they help determine that the code your computer will run is the same code that the publisher created. This post is written to help users that need to use self-signed certificates without involving a public Certificate Authority. The role of self-signed certificates within a known community You may still use self-signed certificates within a known community. The difference between self-signed and purchased-from-CA is that your users must import your self-signed certificate to indicate that it is valid, whereas Certificate Authorities are already trusted by default. This works for known communities where people will trust that my certificate is mine, but does not scale widely where I cannot actually contact or know the systems that will need to trust my certificate. Public Certificate Authorities are widely trusted already because they abide by many different requirements and frequent checks. An example would be students in a university class sharing their public certificates on a mailing list or web page, employees publishing on the intranet, or a system administrator rolling certificates out to end-users. Managed machines help this because you can automate the rollout, but they are not required -- the major point simply that people will trust and import your certificate. How to distribute self-signed certificates for a known community There are several steps required to distribute a self-signed certificate to users so that they will properly trust it. These steps are: Creating a public/private key pair for signing. Exporting your public certificate for others Importing your certificate onto machines that should trust you Verify work on a different machine Creating a public/private key pair for signing Having a public/private key pair will give you the ability both to sign items yourself and issue a Certificate Signing Request (CSR) to a certificate authority. Create your public/private key pair by following the instructions for creating key pairs.Every Certificate Authority that I looked at provided similar instructions, but for the sake of cohesiveness I will include the commands that I used here: Generate the key pair.keytool -genkeypair -alias erikcostlow -keyalg EC -keysize 571 -validity 730 -keystore javakeystore_keepsecret.jks Provide a good password for this file. The alias "erikcostlow" is my name and therefore easy to remember. Substitute your name of something like "mykey." The sigalg of EC (Elliptical Curve) and keysize of 571 will give your key a good strong lifetime. All keys are set to expire. Two years or 730 days is a reasonable compromise between not-long-enough and too-long. Most public Certificate Authorities will sign something for one to five years. You will be placing your keys in javakeystore_keepsecret.jks -- this file will contain private keys and therefore should not be shared. If someone else gets these private keys, they can impersonate your signature. Please be cautious about automated cloud backup systems and private key stores. Answer all the questions. It is important to provide good answers because you will stick with them for the "-validity" days that you specified above.What is your first and last name?  [Unknown]:  First LastWhat is the name of your organizational unit?  [Unknown]:  Line of BusinessWhat is the name of your organization?  [Unknown]:  MyCompanyWhat is the name of your City or Locality?  [Unknown]:  City NameWhat is the name of your State or Province?  [Unknown]:  CAWhat is the two-letter country code for this unit?  [Unknown]:  USIs CN=First Last, OU=Line of Business, O=MyCompany, L=City, ST=CA, C=US correct?  [no]:  yesEnter key password for <erikcostlow>        (RETURN if same as keystore password): Verify your work:keytool -list -keystore javakeystore_keepsecret.jksYou should see your new key pair. Exporting your public certificate for others Public Key Infrastructure relies on two simple concepts: the public key may be made public and the private key must be private. By exporting your public certificate, you are able to share it with others who can then import the certificate to trust you. keytool -exportcert -keystore javakeystore_keepsecret.jks -alias erikcostlow -file erikcostlow.cer To verify this, you can open the .cer file by double-clicking it on most operating systems. It should show the information that you entered during the creation prompts. This is the file that you will share with others. They will use this certificate to prove that artifacts signed by this certificate came from you. If you do not manage machines directly, place the certificate file on an area that people within the known community should trust, such as an intranet page. Import the certificate onto machines that should trust you In order to trust the certificate, people within your known network must import your certificate into their keystores. The first step is to verify that the certificate is actually yours, which can be done through any band: email, phone, in-person, etc. Known networks can usually do this Determine the right keystore: For an individual user looking to trust another, the correct file is within that user’s directory.e.g. USER_HOME\AppData\LocalLow\Sun\Java\Deployment\security\trusted.certs For system-wide installations, Java’s Certificate Authorities are in JAVA_HOMEe.g. C:\Program Files\Java\jre8\lib\security\cacerts File paths for Mac and Linux are included in the link above. Follow the instructions to import the certificate into the keystore. keytool -importcert -keystore THEKEYSTOREFROMABOVE -alias erikcostlow -file erikcostlow.cer In this case, I am still using my name for the alias because it’s easy for me to remember. You may also use an alias of your company name. Scaling distribution of the import The easiest way to apply your certificate across many machines is to just push the .certs or cacerts file onto them. When doing this, watch out for any changes that people would have made to this file on their machines. Trusted.certs: When publishing into user directories, your file will overwrite any keys that the user has added since last update. CACerts: It is best to re-run the import command with each installation rather than just overwriting the file. If you just keep the same cacerts file between upgrades, you will overwrite any CAs that have been added or removed. By re-importing, you stay up to date with changes. Verify work on a different machine Verification is a way of checking on the client machine to ensure that it properly trusts signed artifacts after you have added your signing certificate. Many people have started using deployment rule sets. You can validate the deployment rule set by: Create and sign the deployment rule set on the computer that holds the private key. Copy the deployment rule set on to the different machine where you have imported the signing certificate. Verify that the Java Control Panel’s security tab shows your deployment rule set. Verifying an individual JAR file or multiple JAR files You can test a certificate chain by using the jarsigner command. jarsigner -verify filename.jar If the output does not say "jar verified" then run the following command to see why: jarsigner -verify -verbose -certs filename.jar Check the output for the term “CertPath not validated.”

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• Oracle Tutor: Top 10 to Implement Sustainable Policies and Procedures

  - by emily.chorba(at)oracle.com

  Overview Your organization (executives, managers, and employees) understands the value of having written business process documents (process maps, procedures, instructions, reference documents, and form abstracts). Policies and procedures should be documented because they help to reduce the range of individual decisions and encourage management by exception: the manager only needs to give special attention to unusual problems, not covered by a specific policy or procedure. As more and more procedures are written to cover recurring situations, managers will begin to make decisions which will be consistent from one functional area to the next.Companies should take a project management approach when implementing an environment for a sustainable documentation program and do the following:1. Identify an Executive Champion2. Put together a winning team3. Assign ownership4. Centralize publishing5. Establish the Document Maintenance Process Up Front6. Document critical activities only7. Document actual practice8. Minimize documentation9. Support continuous improvement10. Keep it simple 1. Identify an Executive ChampionAppoint a top down driver. Select one key individual to be a mentor for the procedure planning team. The individual should be a senior manager, such as your company president, CIO, CFO, the vice-president of quality, manufacturing, or engineering. Written policies and procedures can be important supportive aids when known to express the thinking for the chief executive officer and / or the president and to have his or her full support. 2. Put Together a Winning TeamChoose a strong Project Management Leader and staff the procedure planning team with management members from cross functional groups. Make sure team members have the responsibility - and the authority - to make things happen.The winning team should consist of the Documentation Project Manager, Document Owners (one for each functional area), a Document Controller, and Document Specialists (as needed). The Tutor Implementation Guide has complete job descriptions for these roles. 3. Assign Ownership It is virtually impossible to keep process documentation simple and meaningful if employees who are far removed from the activity itself create it. It is impossible to keep documentation up-to-date when responsibility for the document is not clearly understood.Key to the Tutor methodology, therefore, is the concept of ownership. Each document has a single owner, who is responsible for ensuring that the document is necessary and that it reflects actual practice. The owner must be a person who is knowledgeable about the activity and who has the authority to build consensus among the persons who participate in the activity as well as the authority to define or change the way an activity is performed. The owner must be an advocate of the performers and negotiate, not dictate practices.In the Tutor environment, a document's owner is the only person with the authority to approve an update to that document. 4. Centralize Publishing Although it is tempting (especially in a networked environment and with document management software solutions) to decentralize the control of all documents -- with each owner updating and distributing his own -- Tutor promotes centralized publishing by assigning the Document Administrator (gate keeper) to manage the updates and distribution of the procedures library. 5. Establish a Document Maintenance Process Up Front (and stick to it) Everyone in your organization should know they are invited to suggest changes to procedures and should understand exactly what steps to take to do so. Tutor provides a set of procedures to help your company set up a healthy document control system. There are many document management products available to automate some of the document change and maintenance steps. Depending on the size of your organization, a simple document management system can reduce the effort it takes to track and distribute document changes and updates. Whether your company decides to store the written policies and procedures on a file server or in a database, the essential tasks for maintaining documents are the same, though some tasks are automated. 6. Document Critical Activities Only The best way to keep your documentation simple is to reduce the number of process documents to a bare minimum and to include in those documents only as much detail as is absolutely necessary. The first step to reducing process documentation is to document only those activities that are deemed critical. Not all activities require documentation. In fact, some critical activities cannot and should not be standardized. Others may be sufficiently documented with an instruction or a checklist and may not require a procedure. A document should only be created when it enhances the performance of the employee performing the activity. If it does not help the employee, then there is no reason to maintain the document. Activities that represent little risk (such as project status), activities that cannot be defined in terms of specific tasks (such as product research), and activities that can be performed in a variety of ways (such as advertising) often do not require documentation. Sometimes, an activity will evolve to the point where documentation is necessary. For example, an activity performed by single employee may be straightforward and uncomplicated -- that is, until the activity is performed by multiple employees. Sometimes, it is the interaction between co-workers that necessitates documentation; sometimes, it is the complexity or the diversity of the activity.7. Document Actual Practices The only reason to maintain process documentation is to enhance the performance of the employee performing the activity. And documentation can only enhance performance if it reflects reality -- that is, current best practice. Documentation that reflects an unattainable ideal or outdated practices will end up on the shelf, unused and forgotten.Documenting actual practice means (1) auditing the activity to understand how the work is really performed, (2) identifying best practices with employees who are involved in the activity, (3) building consensus so that everyone agrees on a common method, and (4) recording that consensus.8. Minimize Documentation One way to keep it simple is to document at the highest level possible. That is, include in your documents only as much detail as is absolutely necessary.When writing a document, you should ask yourself, What is the purpose of this document? That is, what problem will it solve?By focusing on this question, you can target the critical information.• What questions are the end users likely to have?• What level of detail is required?• Is any of this information extraneous to the document's purpose? Short, concise documents are user friendly and they are easier to keep up to date. 9. Support Continuous Improvement Employees who perform an activity are often in the best position to identify improvements to the process. In other words, continuous improvement is a natural byproduct of the work itself -- but only if the improvements are communicated to all employees who are involved in the process, and only if there is consensus among those employees.Traditionally, process documentation has been used to dictate performance, to limit employees' actions. In the Tutor environment, process documents are used to communicate improvements identified by employees. How does this work? The Tutor methodology requires a process document to reflect actual practice, so the owner of a document must routinely audit its content -- does the document match what the employees are doing? If it doesn't, the owner has the responsibility to evaluate the process, to build consensus among the employees, to identify "best practices," and to communicate these improvements via a document update. Continuous improvement can also be an outgrowth of corrective action -- but only if the solutions to problems are communicated effectively. The goal should be to solve a problem once and only once, which means not only identifying the solution, but ensuring that the solution becomes part of the process. The Tutor system provides the method through which improvements and solutions are documented and communicated to all affected employees in a cost-effective, timely manner; it ensures that improvements are not lost or confined to a single employee. 10. Keep it Simple Process documents don't have to be complex and unfriendly. In fact, the simpler the format and organization, the more likely the documents will be used. And the simpler the method of maintenance, the more likely the documents will be kept up-to-date. Keep it simply by:• Minimizing skills and training required• Following the established Tutor document format and layout• Avoiding technology just for technology's sake No other rule has as major an impact on the success of your internal documentation as -- keep it simple. Learn More For more information about Tutor, visit Oracle.Com or the Tutor Blog. Post your questions at the Tutor Forum.   Emily Chorba Principle Product Manager Oracle Tutor & BPM 

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• Developing a Cost Model for Cloud Applications

  - by BuckWoody

  Note - please pay attention to the date of this post. As much as I attempt to make the information below accurate, the nature of distributed computing means that components, units and pricing will change over time. The definitive costs for Microsoft Windows Azure and SQL Azure are located here, and are more accurate than anything you will see in this post: http://www.microsoft.com/windowsazure/offers/  When writing software that is run on a Platform-as-a-Service (PaaS) offering like Windows Azure / SQL Azure, one of the questions you must answer is how much the system will cost. I will not discuss the comparisons between on-premise costs (which are nigh impossible to calculate accurately) versus cloud costs, but instead focus on creating a general model for estimating costs for a given application. You should be aware that there are (at this writing) two billing mechanisms for Windows and SQL Azure: “Pay-as-you-go” or consumption, and “Subscription” or commitment. Conceptually, you can consider the former a pay-as-you-go cell phone plan, where you pay by the unit used (at a slightly higher rate) and the latter as a standard cell phone plan where you commit to a contract and thus pay lower rates. In this post I’ll stick with the pay-as-you-go mechanism for simplicity, which should be the maximum cost you would pay. From there you may be able to get a lower cost if you use the other mechanism. In any case, the model you create should hold. Developing a good cost model is essential. As a developer or architect, you’ll most certainly be asked how much something will cost, and you need to have a reliable way to estimate that. Businesses and Organizations have been used to paying for servers, software licenses, and other infrastructure as an up-front cost, and power, people to the systems and so on as an ongoing (and sometimes not factored) cost. When presented with a new paradigm like distributed computing, they may not understand the true cost/value proposition, and that’s where the architect and developer can guide the conversation to make a choice based on features of the application versus the true costs. The two big buckets of use-types for these applications are customer-based and steady-state. In the customer-based use type, each successful use of the program results in a sale or income for your organization. Perhaps you’ve written an application that provides the spot-price of foo, and your customer pays for the use of that application. In that case, once you’ve estimated your cost for a successful traversal of the application, you can build that into the price you charge the user. It’s a standard restaurant model, where the price of the meal is determined by the cost of making it, plus any profit you can make. In the second use-type, the application will be used by a more-or-less constant number of processes or users and no direct revenue is attached to the system. A typical example is a customer-tracking system used by the employees within your company. In this case, the cost model is often created “in reverse” - meaning that you pilot the application, monitor the use (and costs) and that cost is held steady. This is where the comparison with an on-premise system becomes necessary, even though it is more difficult to estimate those on-premise true costs. For instance, do you know exactly how much cost the air conditioning is because you have a team of system administrators? This may sound trivial, but that, along with the insurance for the building, the wiring, and every other part of the system is in fact a cost to the business. There are three primary methods that I’ve been successful with in estimating the cost. None are perfect, all are demand-driven. The general process is to lay out a matrix of: components units cost per unit and then multiply that times the usage of the system, based on which components you use in the program. That sounds a bit simplistic, but using those metrics in a calculation becomes more detailed. In all of the methods that follow, you need to know your application. The components for a PaaS include computing instances, storage, transactions, bandwidth and in the case of SQL Azure, database size. In most cases, architects start with the first model and progress through the other methods to gain accuracy. Simple Estimation The simplest way to calculate costs is to architect the application (even UML or on-paper, no coding involved) and then estimate which of the components you’ll use, and how much of each will be used. Microsoft provides two tools to do this - one is a simple slider-application located here: http://www.microsoft.com/windowsazure/pricing-calculator/  The other is a tool you download to create an “Return on Investment” (ROI) spreadsheet, which has the advantage of leading you through various questions to estimate what you plan to use, located here: https://roianalyst.alinean.com/msft/AutoLogin.do?d=176318219048082115  You can also just create a spreadsheet yourself with a structure like this: Program Element Azure Component Unit of Measure Cost Per Unit Estimated Use of Component Total Cost Per Component Cumulative Cost               Of course, the consideration with this model is that it is difficult to predict a system that is not running or hasn’t even been developed. Which brings us to the next model type. Measure and Project A more accurate model is to actually write the code for the application, using the Software Development Kit (SDK) which can run entirely disconnected from Azure. The code should be instrumented to estimate the use of the application components, logging to a local file on the development system. A series of unit and integration tests should be run, which will create load on the test system. You can use standard development concepts to track this usage, and even use Windows Performance Monitor counters. The best place to start with this method is to use the Windows Azure Diagnostics subsystem in your code, which you can read more about here: http://blogs.msdn.com/b/sumitm/archive/2009/11/18/introducing-windows-azure-diagnostics.aspx This set of API’s greatly simplifies tracking the application, and in fact you can use this information for more than just a cost model. After you have the tracking logs, you can plug the numbers into ay of the tools above, which should give a representative cost or in some cases a unit cost. The consideration with this model is that the SDK fabric is not a one-to-one comparison with performance on the actual Windows Azure fabric. Those differences are usually smaller, but they do need to be considered. Also, you may not be able to accurately predict the load on the system, which might lead to an architectural change, which changes the model. This leads us to the next, most accurate method for a cost model. Sample and Estimate Using standard statistical and other predictive math, once the application is deployed you will get a bill each month from Microsoft for your Azure usage. The bill is quite detailed, and you can export the data from it to do analysis, and using methods like regression and so on project out into the future what the costs will be. I normally advise that the architect also extrapolate a unit cost from those metrics as well. This is the information that should be reported back to the executives that pay the bills: the past cost, future projected costs, and unit cost “per click” or “per transaction”, as your case warrants. The challenge here is in the model itself - statistical methods are not foolproof, and the larger the sample (in this case I recommend the entire population, not a smaller sample) is key. References and Tools Articles: http://blogs.msdn.com/b/patrick_butler_monterde/archive/2010/02/10/windows-azure-billing-overview.aspx http://technet.microsoft.com/en-us/magazine/gg213848.aspx http://blog.codingoutloud.com/2011/06/05/azure-faq-how-much-will-it-cost-me-to-run-my-application-on-windows-azure/ http://blogs.msdn.com/b/johnalioto/archive/2010/08/25/10054193.aspx http://geekswithblogs.net/iupdateable/archive/2010/02/08/qampa-how-can-i-calculate-the-tco-and-roi-when.aspx   Other Tools: http://cloud-assessment.com/ http://communities.quest.com/community/cloud_tools

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• Scheduling thread tiles with C++ AMP

  - by Daniel Moth

  This post assumes you are totally comfortable with, what some of us call, the simple model of C++ AMP, i.e. you could write your own matrix multiplication. We are now ready to explore the tiled model, which builds on top of the non-tiled one. Tiling the extent We know that when we pass a grid (which is just an extent under the covers) to the parallel_for_each call, it determines the number of threads to schedule and their index values (including dimensionality). For the single-, two-, and three- dimensional cases you can go a step further and subdivide the threads into what we call tiles of threads (others may call them thread groups). So here is a single-dimensional example: extent<1> e(20); // 20 units in a single dimension with indices from 0-19 grid<1> g(e);      // same as extent tiled_grid<4> tg = g.tile<4>(); …on the 3rd line we subdivided the single-dimensional space into 5 single-dimensional tiles each having 4 elements, and we captured that result in a concurrency::tiled_grid (a new class in amp.h). Let's move on swiftly to another example, in pictures, this time 2-dimensional: So we start on the left with a grid of a 2-dimensional extent which has 8*6=48 threads. We then have two different examples of tiling. In the first case, in the middle, we subdivide the 48 threads into tiles where each has 4*3=12 threads, hence we have 2*2=4 tiles. In the second example, on the right, we subdivide the original input into tiles where each has 2*2=4 threads, hence we have 4*3=12 tiles. Notice how you can play with the tile size and achieve different number of tiles. The numbers you pick must be such that the original total number of threads (in our example 48), remains the same, and every tile must have the same size. Of course, you still have no clue why you would do that, but stick with me. First, we should see how we can use this tiled_grid, since the parallel_for_each function that we know expects a grid. Tiled parallel_for_each and tiled_index It turns out that we have additional overloads of parallel_for_each that accept a tiled_grid instead of a grid. However, those overloads, also expect that the lambda you pass in accepts a concurrency::tiled_index (new in amp.h), not an index<N>. So how is a tiled_index different to an index? A tiled_index object, can have only 1 or 2 or 3 dimensions (matching exactly the tiled_grid), and consists of 4 index objects that are accessible via properties: global, local, tile_origin, and tile. The global index is the same as the index we know and love: the global thread ID. The local index is the local thread ID within the tile. The tile_origin index returns the global index of the thread that is at position 0,0 of this tile, and the tile index is the position of the tile in relation to the overall grid. Confused? Here is an example accompanied by a picture that hopefully clarifies things: array_view<int, 2> data(8, 6, p_my_data); parallel_for_each(data.grid.tile<2,2>(), [=] (tiled_index<2,2> t_idx) restrict(direct3d) { /* todo */ }); Given the code above and the picture on the right, what are the values of each of the 4 index objects that the t_idx variables exposes, when the lambda is executed by T (highlighted in the picture on the right)? If you can't work it out yourselves, the solution follows: t_idx.global       = index<2> (6,3) t_idx.local          = index<2> (0,1) t_idx.tile_origin = index<2> (6,2) t_idx.tile             = index<2> (3,1) Don't move on until you are comfortable with this… the picture really helps, so use it. Tiled Matrix Multiplication Example – part 1 Let's paste here the C++ AMP matrix multiplication example, bolding the lines we are going to change (can you guess what the changes will be?) 01: void MatrixMultiplyTiled_Part1(vector<float>& vC, const vector<float>& vA, const vector<float>& vB, int M, int N, int W) 02: { 03: 04: array_view<const float,2> a(M, W, vA); 05: array_view<const float,2> b(W, N, vB); 06: array_view<writeonly<float>,2> c(M, N, vC); 07: parallel_for_each(c.grid, 08: [=](index<2> idx) restrict(direct3d) { 09: 10: int row = idx[0]; int col = idx[1]; 11: float sum = 0.0f; 12: for(int i = 0; i < W; i++) 13: sum += a(row, i) * b(i, col); 14: c[idx] = sum; 15: }); 16: } To turn this into a tiled example, first we need to decide our tile size. Let's say we want each tile to be 16*16 (which assumes that we'll have at least 256 threads to process, and that c.grid.extent.size() is divisible by 256, and moreover that c.grid.extent[0] and c.grid.extent[1] are divisible by 16). So we insert at line 03 the tile size (which must be a compile time constant). 03: static const int TS = 16; ...then we need to tile the grid to have tiles where each one has 16*16 threads, so we change line 07 to be as follows 07: parallel_for_each(c.grid.tile<TS,TS>(), ...that means that our index now has to be a tiled_index with the same characteristics as the tiled_grid, so we change line 08 08: [=](tiled_index<TS, TS> t_idx) restrict(direct3d) { ...which means, without changing our core algorithm, we need to be using the global index that the tiled_index gives us access to, so we insert line 09 as follows 09: index<2> idx = t_idx.global; ...and now this code just works and it is tiled! Closing thoughts on part 1 The process we followed just shows the mechanical transformation that can take place from the simple model to the tiled model (think of this as step 1). In fact, when we wrote the matrix multiplication example originally, the compiler was doing this mechanical transformation under the covers for us (and it has additional smarts to deal with the cases where the total number of threads scheduled cannot be divisible by the tile size). The point is that the thread scheduling is always tiled, even when you use the non-tiled model. But with this mechanical transformation, we haven't gained anything… Hint: our goal with explicitly using the tiled model is to gain even more performance. In the next post, we'll evolve this further (beyond what the compiler can automatically do for us, in this first release), so you can see the full usage of the tiled model and its benefits… Comments about this post by Daniel Moth welcome at the original blog.

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• DBA Best Practices - A Blog Series: Episode 1 - Backups

  - by Argenis

    This blog post is part of the DBA Best Practices series, on which various topics of concern for daily database operations are discussed. Your feedback and comments are very much welcome, so please drop by the comments section and be sure to leave your thoughts on the subject. Morning Coffee When I was a DBA, the first thing I did when I sat down at my desk at work was checking that all backups had completed successfully. It really was more of a ritual, since I had a dual system in place to check for backup completion: 1) the scheduled agent jobs to back up the databases were set to alert the NOC in failure, and 2) I had a script run from a central server every so often to check for any backup failures. Why the redundancy, you might ask. Well, for one I was once bitten by the fact that database mail doesn't work 100% of the time. Potential causes for failure include issues on the SMTP box that relays your server email, firewall problems, DNS issues, etc. And so to be sure that my backups completed fine, I needed to rely on a mechanism other than having the servers do the taking - I needed to interrogate the servers and ask each one if an issue had occurred. This is why I had a script run every so often. Some of you might have monitoring tools in place like Microsoft System Center Operations Manager (SCOM) or similar 3rd party products that would track all these things for you. But at that moment, we had no resort but to write our own Powershell scripts to do it. Now it goes without saying that if you don't have backups in place, you might as well find another career. Your most sacred job as a DBA is to protect the data from a disaster, and only properly safeguarded backups can offer you peace of mind here. "But, we have a cluster...we don't need backups" Sadly I've heard this line more than I would have liked to. You need to understand that a cluster is comprised of shared storage, and that is precisely your single point of failure. A cluster will protect you from an issue at the Operating System level, and also under an outage of any SQL-related service or dependent devices. But it will most definitely NOT protect you against corruption, nor will it protect you against somebody deleting data from a table - accidentally or otherwise. Backup, fine. How often do I take a backup? The answer to this is something you will hear frequently when working with databases: it depends. What does it depend on? For one, you need to understand how much data your business is willing to lose. This is what's called Recovery Point Objective, or RPO. If you don't know how much data your business is willing to lose, you need to have an honest and realistic conversation about data loss expectations with your customers, internal or external. From my experience, their first answer to the question "how much data loss can you withstand?" will be "zero". In that case, you will need to explain how zero data loss is very difficult and very costly to achieve, even in today's computing environments. Do you want to go ahead and take full backups of all your databases every hour, or even every day? Probably not, because of the impact that taking a full backup can have on a system. That's what differential and transaction log backups are for. Have I answered the question of how often to take a backup? No, and I did that on purpose. You need to think about how much time you have to recover from any event that requires you to restore your databases. This is what's called Recovery Time Objective. Again, if you go ask your customer how long of an outage they can withstand, at first you will get a completely unrealistic number - and that will be your starting point for discussing a solution that is cost effective. The point that I'm trying to get across is that you need to have a plan. This plan needs to be practiced, and tested. Like a football playbook, you need to rehearse the moves you'll perform when the time comes. How often is up to you, and the objective is that you feel better about yourself and the steps you need to follow when emergency strikes. A backup is nothing more than an untested restore Backups are files. Files are prone to corruption. Put those two together and realize how you feel about those backups sitting on that network drive. When was the last time you restored any of those? Restoring your backups on another box - that, by the way, doesn't have to match the specs of your production server - will give you two things: 1) peace of mind, because now you know that your backups are good and 2) a place to offload your consistency checks with DBCC CHECKDB or any of the other DBCC commands like CHECKTABLE or CHECKCATALOG. This is a great strategy for VLDBs that cannot withstand the additional load created by the consistency checks. If you choose to offload your consistency checks to another server though, be sure to run DBCC CHECKDB WITH PHYSICALONLY on the production server, and if you're using SQL Server 2008 R2 SP1 CU4 and above, be sure to enable traceflags 2562 and/or 2549, which will speed up the PHYSICALONLY checks further - you can read more about this enhancement here. Back to the "How Often" question for a second. If you have the disk, and the network latency, and the system resources to do so, why not backup the transaction log often? As in, every 5 minutes, or even less than that? There's not much downside to doing it, as you will have to clear the log with a backup sooner than later, lest you risk running out space on your tlog, or even your drive. The one drawback to this approach is that you will have more files to deal with at restore time, and processing each file will add a bit of extra time to the entire process. But it might be worth that time knowing that you minimized the amount of data lost. Again, test your plan to make sure that it matches your particular needs. Where to back up to? Network share? Locally? SAN volume? This is another topic where everybody has a favorite choice. So, I'll stick to mentioning what I like to do and what I consider to be the best practice in this regard. I like to backup to a SAN volume, i.e., a drive that actually lives in the SAN, and can be easily attached to another server in a pinch, saving you valuable time - you wouldn't need to restore files on the network (slow) or pull out drives out a dead server (been there, done that, it’s also slow!). The key is to have a copy of those backup files made quickly, and, if at all possible, to a remote target on a different datacenter - or even the cloud. There are plenty of solutions out there that can help you put such a solution together. That right there is the first step towards a practical Disaster Recovery plan. But there's much more to DR, and that's material for a different blog post in this series.

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• Need personal advice on how to get out of a company..

  - by SOfan

  Hi, I am an SO user since past 6 months and this is the first time I am turning to SO for personal help. I have asked technical questions before with my real ID. I am stuck inside a service based IT company for the past one year and haven't been able to decide if to leave it, when to leave it and how to leave it. I had taken 2 weeks LWP on medical reason roughly at end of 1 year and then soon after reporting, I applied for 2 months more LWP (on medical/personal ground) with the intention of working on my health,take up a hobby class to ward off depression,pessimism, to have some fun in life, and to look for a job which I really would be excited about - that interests me and which matches with my strength. My leave starts from this Monday. So in any case, I had hard set in mind that I will leave the company after I join them back hopefully with some job offer already in hand (after figuring out what I want do). Neither I can stand the past project,past colleagues,company, HR, pathetically low salary. But if I really listen to my heart, I don't want to have to go back to that office after my sabbatical and again have to see those people. I will have to resign it after my sabbatical ends. Then HR people perhaps wont like it, may even accuse me on face or behind back that primary purpose of my leave must have been to hunt for a better job and I lied about medical and person reasons. Also, if they get nasty and force me to serve 2 months notice period. There is no way I see myself after sabbatical resuming in old project or starting new work. It will be a pain. Since they have already approved 2 months leave and stuff, ideally if they want, they should be just able to relieve me right on the next day after I join back. But, I don't know if they want to get nasty, will they mention about my 2 months sabbatical leave in my experience letter or more scary, the term medical/personal reason. I have hard earned my experience here, have worked against my will, mostly it has been painful and slogged like anything, because I realize the importance of work experience in IT industry. I don't have greed to have those 2 months included extra in my experience letter, but I don't want to mess up with my experience letter in a way which makes my next employer ask question, get suspicious, or be wary if I have any medical reason going on. Being an emotional,moody person or somebody who can't be in an environment, once my mind and heart starts hating it. I think it perhaps is best, if I resign on Monday itself telling them (in polite manner) something that look I took sabbatical for some reason but I don't want to resume working in the company after my sabbatical ends. So please accept my resignation. Now tell me what you want to do about my leave request, my notice period and when you are willing to relieve me. What should I write and how? Some background: I am working in an IT company in India.I am overqualified in the company. It is grossly underpaying me. My education qualifications far exceed anyone's in the whole company being a CS undergrad as well as a CS grad. I joined this company after finishing the grad. I had self-doubts about my skills and interest as a programmer. I like doing research oriented work, though didn't have any particular success during grad. My life here has been very hectic. The project containing many many sub-projects has kept me on my toes and I have never really liked the work. I have been playing against my strength. Also the company strict internet usage policy (you can't read gmails, can't browse any non-work related sites not even news). When working for a client, from the machine we can't even check company related emails.For this one has to go to kiosk like 5 machines in a small room etc. Most of the times those machines are not available, so it was not unusual to keep making rounds to these kiosk machines to check company emails, browse company related emails etc.So it was not so easy to keep in touch with company related basic affairs for a not particular careful person. Things like this which are new to me, make me feel restricted. I am an undecisive person with a sense of failure, self-doubt, not meeting up unrealistic expectation. Somewhere at back of mind, I envy my classmates who make a smooth transition from company to company without causing any gap in their resume. I on other hand have gaps in resume. I get tired after working in a place for sometime. problem with colleagues in general. I am not particular great with people, have few friends, not known for a fun nature, rather serious, scholar. I am not a typical conventional female. I think females are usually more disciplined. But I am not so. I reach office late (though after informing manager). I don't want to blame them entirely, because from my past, it is not unusual for me to get undecisive on things. Also I had doubts about my ability as researched and to succeed there. of building a relationship in a group, to have something to talk about, newspaper. I get cut-off from people. peer pressure. I make blunders in coding, lose patience. Consciously or unconsciously I feel contempt for people here, work here, environment here. I have doubts that either I go to a place which does innovation, does research oriented work, product biggies, have great motivated people, have competent people passionate about products they are building. But then I also doubt my ability to survive there. I have identified that an idea job for me would be 4 days a week, a high salary job. When among people in company/team, I can't think much. I need some time at home to read good authentic books written in good style on what work I am doing.So that I am comfortable with my understanding of work. I get into pressure easily under deadline and need 5th day to cool myself off. I took for 2 weeks leave, because each day was hell for me. May be the depression phase of bipolar is on and also partially it could be that being a work centered person, who derives happiness,self-esteem from work, haven't been enjoying work and have been working for the sole person of proving stability, and ability to stick, against all odds, and facing what challenges I see, bonding with people, identifying opportunities to learn in given task etc.have been averaging one day LWP in 1 week or 10 days. or may be because of my nature,ADD,not being able to switch context,out of touch with news, don't have a circle of friends with who I enjoy. less knowledge in general to talk about, just some technical stuff.anyway, so due to emotional reason, some practical reason etc, I wanted to be very sure before leaving. So my leave starts from Monday and I should feel happy about it. I have taken the leave to for a few purposes - to take care of my health by regular yoga/exercise (with project on, I just can't do anything regular), reassess myself to see what I want to try next which work I might like, look for next job, take up a hobby which I like say singing. I am not clear on my career,job aspiration. I have tried my hands on research. During this year appraisal yesterday, I even had some conflict with my last manager. In meeting with me one on one, he would say all nice things about me, but in feedback to new manager, he hasn't given any excellent feedback. It is all only good. I am angry at this old Manager. Also new manager also scolded me as I didn't agree to his appraisal and waited to hear myself from old Manager. He kind of scolded me for wasting his time. Am I being unethical somewhere? I am always very conscious of if I am cheating anywhere. What advice I am seeking? How to resign and what to write in resignation letter

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• Advice for a distracted, unhappy, recently graduated programmer? [closed]

  - by Re-Invent

  I graduated 4 months ago. I had offers from a few good places to work at. At the same time I wanted to stick to building a small software business of my own, still have some ideas with good potential, some half done projects frozen in my github. But due to social pressures, I chose a job, the pay is great, but I am half-passionate about it. A small team of smart folks building useful product, working out contracts across the world. I've started finding it extremely boring. Boring to the extent that I skip 2-3 days a week together not doing work. Neither do I spend that time progressing any of my own projects. Yes, I feel stupid at the way I'm wasting time, but I don't understand exactly why is it happening. It's as if all the excitement has been drained. What can I do about it? Long version: School - I was in third standard. Only students, 6th grade had access to computer labs. I once peeked into the lab from the little door opening. No hard-disks, MS DOS on 5 1/2 inch floppies. I asked a senior student to play some sound in BASIC. He used PLAY to compose a tune. Boy! I was so excited, I was jumping from within. Back home, asked my brother to teach me some programming. We bought a book "MODERN All About GW-BASIC for Schools & Colleges". The book had everything, right from printing, to taking input, file i/o, game programming, machine level support, etc. I was in 6th standard, wrote my first game - a wheel of fortune, rotated the wheel by manipulating 16 color palette's definition. Got internet soon, got hooked to QuickBasic programming community. Made some more games "007 in Danger", "Car Crush 2" for submission to allbasiccode archives. I was extremely excited about all this. My interests now swayed into "hacking" (computer security). Taught myself some perl, found it annoying, learnt PHP and a bit of SQL. Also taught myself Visual Basic one of the winters and wrote a pacman clone with Direct X. By the time I was in 10th standard, I created some evil tools using visual basic, php and mysql and eventually landed myself into an unpaid side-job at a government facility, building evil tools for them. It was a dream come true for crackers of that time. And so was I, still very excited. Things changed soon, last two years of school were not so great as I was balancing preps for college, work at govt. and studies for school at same time. College - College was opposite of all I had wished it to be. I imagined it to be a place where I'd spend my 4 years building something awesome. It was rather an epitome of rote learning, attendance, rules, busy schedules, ban on personal laptops, hardly any hackers surrounding you and shit like that. We had to take permissions to even introduce some cultural/creative activities in our annual schedule. The labs won't be open on weekends because the lab employees had to have their leaves. Yes, a horrible place for someone like me. I still managed to pull out a project with a friend over 2 months. Showed it to people high in the academia hierarchy. They were immensely impressed, we proposed to allow personal computers for students. They made up half-assed reasons and didn't agree. We felt frustrated. And so on, I still managed to teach myself new languages, do new projects of my own, do an intern at the same govt. facility, start a small business for sometime, give a talk at a conference I'm passionate about, win game-dev and hacking contest at most respected colleges, solve good deal of programming contest problems, etc. At the same time I was not content with all these restrictions, great emphasis on rote learning, and sheer wastage of time due to college. I never felt I was overdoing, but now I feel I burnt myself out. During my last days at college, I did an intern at a bigco. While I spent my time building prototypes for certain LBS, the other interns around me, even a good friend, was just skipping time. I thought maybe, in a few weeks he would put in some serious efforts at work assigned to him, but all he did was to find creative ways to skip work, hide his face from manager, engage people in talks if they try to question his progress, etc. I tried a few time to get him on track, but it seems all he wanted was to "not to work hard at all and still reap the fruits". I don't know how others take such people, but I find their vicinity very very poisonous to one's own motivation and productivity. Over that, the place where I come from, HRs don't give much value to what have you done past 4 years. So towards the end of out intern, we all were offered work at the bigco, but the slacker, even after not writing more than 200 lines of code was made a much better offer. I felt enraged instantly - "Is this how the corp world treats someone who does fruitful, if not extra-ordinary work form them for past 6 months?". Yes, I did try to negotiate and debate. The bigcos seem blind due to departmentalization of responsibilities and many layers of management. I decided not to be in touch with any characters of that depressing play. Probably the busy time I had at college, ignoring friends, ignoring fun and squeezing every bit of free time for myself is also responsible. Probably this is what has drained all my willingness to work for anyone. I find my day job boring, at the same time I with to maintain it for financial reasons. I feel a bit burnt out, unsatisfied and at the same time an urge to quit working for someone else and start finishing my frozen side-projects (which may be profitable). Though I haven't got much to support myself with food, office, internet bills, etc in savings. I still have my day job, but I don't find it very interesting, even though the pay is higher than the slacker, I don't find money to be a great motivator here. I keep comparing myself to my past version. I wonder how to get rid of this and reboot myself back to the way I was in school days - excited about it, tinkering, building, learning new things daily, and NOT BORED?

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• DBA Best Practices - A Blog Series: Episode 1 - Backups

  - by Argenis

    This blog post is part of the DBA Best Practices series, on which various topics of concern for daily database operations are discussed. Your feedback and comments are very much welcome, so please drop by the comments section and be sure to leave your thoughts on the subject. Morning Coffee When I was a DBA, the first thing I did when I sat down at my desk at work was checking that all backups have completed successfully. It really was more of a ritual, since I had a dual system in place to check for backup completion: 1) the scheduled agent jobs to back up the databases were set to alert the NOC in failure, and 2) I had a script run from a central server every so often to check for any backup failures. Why the redundancy, you might ask. Well, for one I was once bitten by the fact that database mail doesn't work 100% of the time. Potential causes for failure include issues on the SMTP box that relays your server email, firewall problems, DNS issues, etc. And so to be sure that my backups completed fine, I needed to rely on a mechanism other than having the servers do the taking - I needed to interrogate the servers and ask each one if an issue had occurred. This is why I had a script run every so often. Some of you might have monitoring tools in place like Microsoft System Center Operations Manager (SCOM) or similar 3rd party products that would track all these things for you. But at that moment, we had no resort but to write our own Powershell scripts to do it. Now it goes without saying that if you don't have backups in place, you might as well find another career. Your most sacred job as a DBA is to protect the data from a disaster, and only properly safeguarded backups can offer you peace of mind here. "But, we have a cluster...we don't need backups" Sadly I've heard this line more than I would have liked to. You need to understand that a cluster is comprised of shared storage, and that is precisely your single point of failure. A cluster will protect you from an issue at the Operating System level, and also under an outage of any SQL-related service or dependent devices. But it will most definitely NOT protect you against corruption, nor will it protect you against somebody deleting data from a table - accidentally or otherwise. Backup, fine. How often do I take a backup? The answer to this is something you will hear frequently when working with databases: it depends. What does it depend on? For one, you need to understand how much data your business is willing to lose. This is what's called Recovery Point Objective, or RPO. If you don't know how much data your business is willing to lose, you need to have an honest and realistic conversation about data loss expectations with your customers, internal or external. From my experience, their first answer to the question "how much data loss can you withstand?" will be "zero". In that case, you will need to explain how zero data loss is very difficult and very costly to achieve, even in today's computing environments. Do you want to go ahead and take full backups of all your databases every hour, or even every day? Probably not, because of the impact that taking a full backup can have on a system. That's what differential and transaction log backups are for. Have I answered the question of how often to take a backup? No, and I did that on purpose. You need to think about how much time you have to recover from any event that requires you to restore your databases. This is what's called Recovery Time Objective. Again, if you go ask your customer how long of an outage they can withstand, at first you will get a completely unrealistic number - and that will be your starting point for discussing a solution that is cost effective. The point that I'm trying to get across is that you need to have a plan. This plan needs to be practiced, and tested. Like a football playbook, you need to rehearse the moves you'll perform when the time comes. How often is up to you, and the objective is that you feel better about yourself and the steps you need to follow when emergency strikes. A backup is nothing more than an untested restore Backups are files. Files are prone to corruption. Put those two together and realize how you feel about those backups sitting on that network drive. When was the last time you restored any of those? Restoring your backups on another box - that, by the way, doesn't have to match the specs of your production server - will give you two things: 1) peace of mind, because now you know that your backups are good and 2) a place to offload your consistency checks with DBCC CHECKDB or any of the other DBCC commands like CHECKTABLE or CHECKCATALOG. This is a great strategy for VLDBs that cannot withstand the additional load created by the consistency checks. If you choose to offload your consistency checks to another server though, be sure to run DBCC CHECKDB WITH PHYSICALONLY on the production server, and if you're using SQL Server 2008 R2 SP1 CU4 and above, be sure to enable traceflags 2562 and/or 2549, which will speed up the PHYSICALONLY checks further - you can read more about this enhancement here. Back to the "How Often" question for a second. If you have the disk, and the network latency, and the system resources to do so, why not backup the transaction log often? As in, every 5 minutes, or even less than that? There's not much downside to doing it, as you will have to clear the log with a backup sooner than later, lest you risk running out space on your tlog, or even your drive. The one drawback to this approach is that you will have more files to deal with at restore time, and processing each file will add a bit of extra time to the entire process. But it might be worth that time knowing that you minimized the amount of data lost. Again, test your plan to make sure that it matches your particular needs. Where to back up to? Network share? Locally? SAN volume? This is another topic where everybody has a favorite choice. So, I'll stick to mentioning what I like to do and what I consider to be the best practice in this regard. I like to backup to a SAN volume, i.e., a drive that actually lives in the SAN, and can be easily attached to another server in a pinch, saving you valuable time - you wouldn't need to restore files on the network (slow) or pull out drives out a dead server (been there, done that, it’s also slow!). The key is to have a copy of those backup files made quickly, and, if at all possible, to a remote target on a different datacenter - or even the cloud. There are plenty of solutions out there that can help you put such a solution together. That right there is the first step towards a practical Disaster Recovery plan. But there's much more to DR, and that's material for a different blog post in this series.

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• OData &ndash; The easiest service I can create: now with updates

  - by Jon Dalberg

  The other day I created a simple NastyWord service exposed via OData. It was read-only and used an in-memory backing store for the words. Today I’ll modify it to use a file instead of a list and I’ll accept new nasty words by implementing IUpdatable directly. The first thing to do is enable the service to accept new entries. This is done at configuration time by adding the “WriteAppend” access rule: 1: public class NastyWords : DataService<NastyWordsDataSource> 2: { 3: // This method is called only once to initialize service-wide policies. 4: public static void InitializeService(DataServiceConfiguration config) 5: { 6: config.SetEntitySetAccessRule("*", EntitySetRights.AllRead | EntitySetRights.WriteAppend); 7: config.DataServiceBehavior.MaxProtocolVersion = DataServiceProtocolVersion.V2; 8: } 9: }   Next I placed a file, NastyWords.txt, in the “App_Data” folder and added a few *choice* words to start. This required one simple change to our NastyWordDataSource.cs file: 1: public NastyWordsDataSource() 2: { 3: UpdateFromSource(); 4: } 5:   6: private void UpdateFromSource() 7: { 8: var words = File.ReadAllLines(pathToFile); 9: NastyWords = (from w in words 10: select new NastyWord { Word = w }).AsQueryable(); 11: }   Nothing too shocking here, just reading each line from the NastyWords.txt file and exposing them. Next, I implemented IUpdatable which comes with a boat-load of methods. We don’t need all of them for now since we are only concerned with allowing new values. Here are the methods we must implement, all the others throw a NotImplementedException: 1: public object CreateResource(string containerName, string fullTypeName) 2: { 3: var nastyWord = new NastyWord(); 4: pendingUpdates.Add(nastyWord); 5: return nastyWord; 6: } 7:   8: public object ResolveResource(object resource) 9: { 10: return resource; 11: } 12:   13: public void SaveChanges() 14: { 15: var intersect = (from w in pendingUpdates 16: select w.Word).Intersect(from n in NastyWords 17: select n.Word); 18:   19: if (intersect.Count() > 0) 20: throw new DataServiceException(500, "duplicate entry"); 21:   22: var lines = from w in pendingUpdates 23: select w.Word; 24:   25: File.AppendAllLines(pathToFile, 26: lines, 27: Encoding.UTF8); 28:   29: pendingUpdates.Clear(); 30:   31: UpdateFromSource(); 32: } 33:   34: public void SetValue(object targetResource, string propertyName, object propertyValue) 35: { 36: targetResource.GetType().GetProperty(propertyName).SetValue(targetResource, propertyValue, null); 37: }   I use a simple list to contain the pending updates and only commit them when the “SaveChanges” method is called. Here’s the order these methods are called in our service during an insert: CreateResource – here we just instantiate a new NastyWord and stick a reference to it in our pending updates list. SetValue – this is where the “Word” property of the NastyWord instance is set. SaveChanges – get the list of pending updates, barfing on duplicates, write them to the file and clear our pending list. ResolveResource – the newly created resource will be returned directly here since we aren’t dealing with “handles” to objects but the actual objects themselves. Not too bad, eh? I didn’t find this documented anywhere but a little bit of digging in the OData spec and use of Fiddler made it pretty easy to figure out. Here is some client code which would add a new nasty word: 1: static void Main(string[] args) 2: { 3: var svc = new ServiceReference1.NastyWordsDataSource(new Uri("http://localhost.:60921/NastyWords.svc")); 4: svc.AddToNastyWords(new ServiceReference1.NastyWord() { Word = "shat" }); 5:   6: svc.SaveChanges(); 7: }   Here’s all of the code so far for to implement the service: 1: using System; 2: using System.Collections.Generic; 3: using System.Data.Services; 4: using System.Data.Services.Common; 5: using System.Linq; 6: using System.ServiceModel.Web; 7: using System.Web; 8: using System.IO; 9: using System.Text; 10:   11: namespace ONasty 12: { 13: [DataServiceKey("Word")] 14: public class NastyWord 15: { 16: public string Word { get; set; } 17: } 18:   19: public class NastyWordsDataSource : IUpdatable 20: { 21: private List<NastyWord> pendingUpdates = new List<NastyWord>(); 22: private string pathToFile = @"path to your\App_Data\NastyWords.txt"; 23:   24: public NastyWordsDataSource() 25: { 26: UpdateFromSource(); 27: } 28:   29: private void UpdateFromSource() 30: { 31: var words = File.ReadAllLines(pathToFile); 32: NastyWords = (from w in words 33: select new NastyWord { Word = w }).AsQueryable(); 34: } 35:   36: public IQueryable<NastyWord> NastyWords { get; private set; } 37:   38: public void AddReferenceToCollection(object targetResource, string propertyName, object resourceToBeAdded) 39: { 40: throw new NotImplementedException(); 41: } 42:   43: public void ClearChanges() 44: { 45: pendingUpdates.Clear(); 46: } 47:   48: public object CreateResource(string containerName, string fullTypeName) 49: { 50: var nastyWord = new NastyWord(); 51: pendingUpdates.Add(nastyWord); 52: return nastyWord; 53: } 54:   55: public void DeleteResource(object targetResource) 56: { 57: throw new NotImplementedException(); 58: } 59:   60: public object GetResource(IQueryable query, string fullTypeName) 61: { 62: throw new NotImplementedException(); 63: } 64:   65: public object GetValue(object targetResource, string propertyName) 66: { 67: throw new NotImplementedException(); 68: } 69:   70: public void RemoveReferenceFromCollection(object targetResource, string propertyName, object resourceToBeRemoved) 71: { 72: throw new NotImplementedException(); 73: } 74:   75: public object ResetResource(object resource) 76: { 77: throw new NotImplementedException(); 78: } 79:   80: public object ResolveResource(object resource) 81: { 82: return resource; 83: } 84:   85: public void SaveChanges() 86: { 87: var intersect = (from w in pendingUpdates 88: select w.Word).Intersect(from n in NastyWords 89: select n.Word); 90:   91: if (intersect.Count() > 0) 92: throw new DataServiceException(500, "duplicate entry"); 93:   94: var lines = from w in pendingUpdates 95: select w.Word; 96:   97: File.AppendAllLines(pathToFile, 98: lines, 99: Encoding.UTF8); 100:   101: pendingUpdates.Clear(); 102:   103: UpdateFromSource(); 104: } 105:   106: public void SetReference(object targetResource, string propertyName, object propertyValue) 107: { 108: throw new NotImplementedException(); 109: } 110:   111: public void SetValue(object targetResource, string propertyName, object propertyValue) 112: { 113: targetResource.GetType().GetProperty(propertyName).SetValue(targetResource, propertyValue, null); 114: } 115: } 116:   117: public class NastyWords : DataService<NastyWordsDataSource> 118: { 119: // This method is called only once to initialize service-wide policies. 120: public static void InitializeService(DataServiceConfiguration config) 121: { 122: config.SetEntitySetAccessRule("*", EntitySetRights.AllRead | EntitySetRights.WriteAppend); 123: config.DataServiceBehavior.MaxProtocolVersion = DataServiceProtocolVersion.V2; 124: } 125: } 126: } Next time we’ll allow removing nasty words. Enjoy!

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• What's new in EJB 3.2 ? - Java EE 7 chugging along!

  - by arungupta

  EJB 3.1 added a whole ton of features for simplicity and ease-of-use such as @Singleton, @Asynchronous, @Schedule, Portable JNDI name, EJBContainer.createEJBContainer, EJB 3.1 Lite, and many others. As part of Java EE 7, EJB 3.2 (JSR 345) is making progress and this blog will provide highlights from the work done so far. This release has been particularly kept small but include several minor improvements and tweaks for usability. More features in EJB.Lite Asynchronous session bean Non-persistent EJB Timer service This also means these features can be used in embeddable EJB container and there by improving testability of your application. Pruning - The following features were made Proposed Optional in Java EE 6 and are now made optional. EJB 2.1 and earlier Entity Bean Component Contract for CMP and BMP Client View of an EJB 2.1 and earlier Entity Bean EJB QL: Query Language for CMP Query Methods JAX-RPC-based Web Service Endpoints and Client View The optional features are moved to a separate document and as a result EJB specification is now split into Core and Optional documents. This allows the specification to be more readable and better organized. Updates and Improvements Transactional lifecycle callbacks in Stateful Session Beans, only for CMT. In EJB 3.1, the transaction context for lifecyle callback methods (@PostConstruct, @PreDestroy, @PostActivate, @PrePassivate) are defined as shown. @PostConstruct @PreDestroy @PrePassivate @PostActivate Stateless Unspecified Unspecified N/A N/A Stateful Unspecified Unspecified Unspecified Unspecified Singleton Bean's transaction management type Bean's transaction management type N/A N/A In EJB 3.2, stateful session bean lifecycle callback methods can opt-in to be transactional. These methods are then executed in a transaction context as shown. @PostConstruct @PreDestroy @PrePassivate @PostActivate Stateless Unspecified Unspecified N/A N/A Stateful Bean's transaction management type Bean's transaction management type Bean's transaction management type Bean's transaction management type Singleton Bean's transaction management type Bean's transaction management type N/A N/A For example, the following stateful session bean require a new transaction to be started for @PostConstruct and @PreDestroy lifecycle callback methods. @Statefulpublic class HelloBean {   @PersistenceContext(type=PersistenceContextType.EXTENDED)   private EntityManager em;    @TransactionAttribute(TransactionAttributeType.REQUIRES_NEW)   @PostConstruct   public void init() {        myEntity = em.find(...);   }   @TransactionAttribute(TransactionAttributeType.REQUIRES_NEW)    @PostConstruct    public void destroy() {        em.flush();    }} Notice, by default the lifecycle callback methods are not transactional for backwards compatibility. They need to be explicitly opt-in to be made transactional. Opt-out of passivation for stateful session bean - If your stateful session bean needs to stick around or it has non-serializable field then the bean can be opt-out of passivation as shown. @Stateful(passivationCapable=false)public class HelloBean {    private NonSerializableType ref = ... . . .} Simplified the rules to define all local/remote views of the bean. For example, if the bean is defined as: @Statelesspublic class Bean implements Foo, Bar {    . . .} where Foo and Bar have no annotations of their own, then Foo and Bar are exposed as local views of the bean. The bean may be explicitly marked @Local as @Local@Statelesspublic class Bean implements Foo, Bar {    . . .} then this is the same behavior as explained above, i.e. Foo and Bar are local views. If the bean is marked @Remote as: @Remote@Statelesspublic class Bean implements Foo, Bar {    . . .} then Foo and Bar are remote views. If an interface is marked @Local or @Remote then each interface need to be explicitly marked explicitly to be exposed as a view. For example: @Remotepublic interface Foo { . . . }@Statelesspublic class Bean implements Foo, Bar {    . . .} only exposes one remote interface Foo. Section 4.9.7 from the specification provide more details about this feature. TimerService.getAllTimers is a newly added convenience API that returns all timers in the same bean. This is only for displaying the list of timers as the timer can only be canceled by its owner. Removed restriction to obtain the current class loader, and allow to use java.io package. This is handy if you want to do file access within your beans. JMS 2.0 alignment - A standard list of activation-config properties is now defined destinationLookup connectionFactoryLookup clientId subscriptionName shareSubscriptions Tons of other clarifications through out the spec. Appendix A provide a comprehensive list of changes since EJB 3.1. ThreadContext in Singleton is guaranteed to be thread-safe. Embeddable container implement Autocloseable. A complete replay of Enterprise JavaBeans Today and Tomorrow from JavaOne 2012 can be seen here (click on CON4654_mp4_4654_001 in Media). The specification is still evolving so the actual property or method names or their actual behavior may be different from the currently proposed ones. Are there any improvements that you'd like to see in EJB 3.2 ? The EJB 3.2 Expert Group would love to hear your feedback. An Early Draft of the specification is available. The latest version of the specification can always be downloaded from here. Java EE 7 Specification Status EJB Specification Project JIRA of EJB Specification JSR Expert Group Discussion Archive These features will start showing up in GlassFish 4 Promoted Builds soon.

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• CentOS 5.4 NFS v4 client file permissions differ from original files & NFS Share file contents

  - by p4guru

  Having a strange problem with NFS share and file permissions on the 1 out of the 2 NFS clients, web1 has file permissions issues but web2 is fine. web1 and web2 are load balanced web servers. So questions are: how do I ensure NFS share file contents retain the same permissions for user/group as the original files on web1 server like they do on web2 server ? how do I reverse what I did on web1, i tried unmount command and said command not found ? Information: I'm using 3 dedicated server setup. All 3 servers CentOS 5.4 64bit based. servers are as follows: web1 - nfs client with file permissions issues web2 - nfs client file permissions are OKAY db1 - nfs share at /nfsroot web2 nfs client was setup by my web host, while web1 was setup by me. I did the following commands on web1 and it worked with updating db1 nfsroot share at /nfsroot/site_css with latest files on web1 but the file permissions don't stick even if i use tar with -p command to perserve file permissions ? cd /home/username/public_html/forums/script/ tar -zcp site_css/ > site_css.tar.gz mount -t nfs4 nfsshareipaddress:/site_css /home/username/public_html/forums/scripts/site_css/ -o rw,soft cd /home/username/public_html/forums/script/ tar -zxf site_css.tar.gz But checking on web1 file permissions no longer username user/group but owned by nobody ? but web2 file permissions correct ? This is only a problem for web1 while web2 is correct ? Looks like numeric ids aren't the same ? Not sure how to correct this ? web1 with incorrect user/group of nobody ls -alh /home/username/public_html/forums/scripts/site_css total 48K drwxrwxrwx 2 nobody nobody 4.0K Feb 22 02:37 ./ drwxr-xr-x 3 username username 4.0K Feb 22 02:43 ../ -rw-r--r-- 1 nobody nobody 1 Nov 30 2006 index.html -rw-r--r-- 1 nobody nobody 5.8K Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 nobody nobody 5.8K Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 nobody nobody 5.8K Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 nobody nobody 5.8K Feb 18 05:37 style-cc2f96c9-00011.css web1 numeric ids ls -n /home/username/public_html/forums/scripts/site_css total 48 drwxrwxrwx 2 99 99 4096 Feb 22 02:37 ./ drwxr-xr-x 3 503 500 4096 Feb 22 02:43 ../ -rw-r--r-- 1 99 99 1 Nov 30 2006 index.html -rw-r--r-- 1 99 99 5876 Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 99 99 5877 Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 99 99 5877 Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 99 99 5876 Feb 18 05:37 style-cc2f96c9-00011.css web2 correct username user/group permissions ls -alh /home/username/public_html/forums/scripts/site_css total 48K drwxrwxrwx 2 root root 4.0K Feb 22 02:37 ./ drwxr-xr-x 3 username username 4.0K Dec 2 14:51 ../ -rw-r--r-- 1 username username 1 Nov 30 2006 index.html -rw-r--r-- 1 username username 5.8K Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 username username 5.8K Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 username username 5.8K Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 username username 5.8K Feb 18 05:37 style-cc2f96c9-00011.css web2 numeric ids ls -n /home/username/public_html/forums/scripts/site_css total 48 drwxrwxrwx 2 503 500 4096 Feb 22 02:37 ./ drwxr-xr-x 3 503 500 4096 Dec 2 14:51 ../ -rw-r--r-- 1 503 500 1 Nov 30 2006 index.html -rw-r--r-- 1 503 500 5876 Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 503 500 5877 Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 503 500 5877 Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 503 500 5876 Feb 18 05:37 style-cc2f96c9-00011.css I checked db1 /nfsroot/site_css and user/group ownership was incorrect for newer files dated feb22 owned by root and not username ? on db1 originally incorrect root assigned user/group for new feb22 dated files ls -alh /nfsroot/site_css total 44K drwxrwxrwx 2 root root 4.0K Feb 22 02:37 . drwxr-xr-x 17 root root 4.0K Feb 17 12:06 .. -rw-r--r-- 1 root root 1 Nov 30 2006 index.html -rw-r--r-- 1 root root 5.8K Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 root root 5.8K Feb 22 02:37 style-95001864-00002.css -rw------- 1 username nfs 5.8K Feb 18 05:37 style-b1879ba7-00002.css -rw------- 1 username nfs 5.8K Feb 18 05:37 style-cc2f96c9-00011.css Then I chmod them all on db1 and chown to set to right ownership on db1 so it looks like below on db1 once corrected the newer feb22 dated files ls -alh /nfsroot/site_css total 44K drwxrwxrwx 2 root root 4.0K Feb 22 02:37 . drwxr-xr-x 17 root root 4.0K Feb 17 12:06 .. -rw-r--r-- 1 username username 1 Nov 30 2006 index.html -rw-r--r-- 1 username username 5.8K Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 username username 5.8K Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 username username 5.8K Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 username username 5.8K Feb 18 05:37 style-cc2f96c9-00011.css but still web1 shows owned by nobody ? while web2 shows correct permissions ? web1 still with incorrect user/group of nobody not matching what web2 and db1 are set to ? ls -alh /home/username/public_html/forums/scripts/site_css total 48K drwxrwxrwx 2 nobody nobody 4.0K Feb 22 02:37 ./ drwxr-xr-x 3 username username 4.0K Feb 22 02:43 ../ -rw-r--r-- 1 nobody nobody 1 Nov 30 2006 index.html -rw-r--r-- 1 nobody nobody 5.8K Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 nobody nobody 5.8K Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 nobody nobody 5.8K Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 nobody nobody 5.8K Feb 18 05:37 style-cc2f96c9-00011.css Just so confusing so any help is very very much appreciated! thanks

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• CentOS 5.4 NFS v4 client file permissions differ from original files & NFS Share file contents

  - by p4guru

  Having a strange problem with NFS share and file permissions on the 1 out of the 2 NFS clients, web1 has file permissions issues but web2 is fine. web1 and web2 are load balanced web servers. So questions are: how do I ensure NFS share file contents retain the same permissions for user/group as the original files on web1 server like they do on web2 server ? how do I reverse what I did on web1, i tried unmount command and said command not found ? Information: I'm using 3 dedicated server setup. All 3 servers CentOS 5.4 64bit based. servers are as follows: web1 - nfs client with file permissions issues web2 - nfs client file permissions are OKAY db1 - nfs share at /nfsroot web2 nfs client was setup by my web host, while web1 was setup by me. I did the following commands on web1 and it worked with updating db1 nfsroot share at /nfsroot/site_css with latest files on web1 but the file permissions don't stick even if i use tar with -p command to perserve file permissions ? cd /home/username/public_html/forums/script/ tar -zcp site_css/ > site_css.tar.gz mount -t nfs4 nfsshareipaddress:/site_css /home/username/public_html/forums/scripts/site_css/ -o rw,soft cd /home/username/public_html/forums/script/ tar -zxf site_css.tar.gz But checking on web1 file permissions no longer username user/group but owned by nobody ? but web2 file permissions correct ? This is only a problem for web1 while web2 is correct ? Looks like numeric ids aren't the same ? Not sure how to correct this ? web1 with incorrect user/group of nobody ls -alh /home/username/public_html/forums/scripts/site_css total 48K drwxrwxrwx 2 nobody nobody 4.0K Feb 22 02:37 ./ drwxr-xr-x 3 username username 4.0K Feb 22 02:43 ../ -rw-r--r-- 1 nobody nobody 1 Nov 30 2006 index.html -rw-r--r-- 1 nobody nobody 5.8K Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 nobody nobody 5.8K Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 nobody nobody 5.8K Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 nobody nobody 5.8K Feb 18 05:37 style-cc2f96c9-00011.css web1 numeric ids ls -n /home/username/public_html/forums/scripts/site_css total 48 drwxrwxrwx 2 99 99 4096 Feb 22 02:37 ./ drwxr-xr-x 3 503 500 4096 Feb 22 02:43 ../ -rw-r--r-- 1 99 99 1 Nov 30 2006 index.html -rw-r--r-- 1 99 99 5876 Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 99 99 5877 Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 99 99 5877 Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 99 99 5876 Feb 18 05:37 style-cc2f96c9-00011.css web2 correct username user/group permissions ls -alh /home/username/public_html/forums/scripts/site_css total 48K drwxrwxrwx 2 root root 4.0K Feb 22 02:37 ./ drwxr-xr-x 3 username username 4.0K Dec 2 14:51 ../ -rw-r--r-- 1 username username 1 Nov 30 2006 index.html -rw-r--r-- 1 username username 5.8K Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 username username 5.8K Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 username username 5.8K Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 username username 5.8K Feb 18 05:37 style-cc2f96c9-00011.css web2 numeric ids ls -n /home/username/public_html/forums/scripts/site_css total 48 drwxrwxrwx 2 503 500 4096 Feb 22 02:37 ./ drwxr-xr-x 3 503 500 4096 Dec 2 14:51 ../ -rw-r--r-- 1 503 500 1 Nov 30 2006 index.html -rw-r--r-- 1 503 500 5876 Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 503 500 5877 Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 503 500 5877 Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 503 500 5876 Feb 18 05:37 style-cc2f96c9-00011.css I checked db1 /nfsroot/site_css and user/group ownership was incorrect for newer files dated feb22 owned by root and not username ? on db1 originally incorrect root assigned user/group for new feb22 dated files ls -alh /nfsroot/site_css total 44K drwxrwxrwx 2 root root 4.0K Feb 22 02:37 . drwxr-xr-x 17 root root 4.0K Feb 17 12:06 .. -rw-r--r-- 1 root root 1 Nov 30 2006 index.html -rw-r--r-- 1 root root 5.8K Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 root root 5.8K Feb 22 02:37 style-95001864-00002.css -rw------- 1 username nfs 5.8K Feb 18 05:37 style-b1879ba7-00002.css -rw------- 1 username nfs 5.8K Feb 18 05:37 style-cc2f96c9-00011.css Then I chmod them all on db1 and chown to set to right ownership on db1 so it looks like below on db1 once corrected the newer feb22 dated files ls -alh /nfsroot/site_css total 44K drwxrwxrwx 2 root root 4.0K Feb 22 02:37 . drwxr-xr-x 17 root root 4.0K Feb 17 12:06 .. -rw-r--r-- 1 username username 1 Nov 30 2006 index.html -rw-r--r-- 1 username username 5.8K Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 username username 5.8K Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 username username 5.8K Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 username username 5.8K Feb 18 05:37 style-cc2f96c9-00011.css but still web1 shows owned by nobody ? while web2 shows correct permissions ? web1 still with incorrect user/group of nobody not matching what web2 and db1 are set to ? ls -alh /home/username/public_html/forums/scripts/site_css total 48K drwxrwxrwx 2 nobody nobody 4.0K Feb 22 02:37 ./ drwxr-xr-x 3 username username 4.0K Feb 22 02:43 ../ -rw-r--r-- 1 nobody nobody 1 Nov 30 2006 index.html -rw-r--r-- 1 nobody nobody 5.8K Feb 22 02:37 style-057c3df0-00011.css -rw-r--r-- 1 nobody nobody 5.8K Feb 22 02:37 style-95001864-00002.css -rw-r--r-- 1 nobody nobody 5.8K Feb 18 05:37 style-b1879ba7-00002.css -rw-r--r-- 1 nobody nobody 5.8K Feb 18 05:37 style-cc2f96c9-00011.css Just so confusing so any help is very very much appreciated! thanks

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• Problem installing Ubuntu 10.04 64 bit side by side with Vista by using a bootable USB drive. What n

  - by Adam Siddhi

  What happened I decided to install Ubuntu 10.04 64 bit side by side with Vista Home Premium (I guess on another partition) with a USB stick. I found instructions on how to do this here: https://help.ubuntu.com/community/Installation/FromUSBStick To create the bootable USB drive I had to download a program called Unetbootin. That process was simple enough. All I had to do was just choose the disk image option, select the ubuntu-10.04-desktop-amd64.iso image, make sure it recognizes my USB drive and then press OK. It takes only like a few minutes to create a working bootable USB drive. Then I have to restart my computer, enter the BIOS, select my USB drive as the first boot drive, save options and continue with booting up. After this Ubuntu actually loads up. I think this is known as the Live version of Ubuntu so you can try it out before fully installing it. Any ways, on the Ubuntu 10.04 desktop I saw an installer. I click it and begin the installation process. Just so you know, I tried installing it 2 times. I will explain what happened each time: The first time I tried installing Ubuntu 10.04 I got stuck at step 4 of 7. I remember selecting the last option in the window which was Specify Partitions Manually (Advanced) I made my partition for Ubuntu like 52 gigs. I clicked forward and a little pop up window appeared saying Please Wait. So the installation process stalled on this window so I closed out of it and quit the installation process. So at this point I was worried because I had already selected the partition size and assumed it started making it. Since it stalled I had to quit out though. Anyways, once again I reached step 4 of 7 a decided to select the first option which is Install them side by side choosing between them each startup. I figured this was the safe way to go. I did that and the pop up window saying Please Wait popped up again but lasted only like 10 seconds. Then I got to I guess step 6 where it asks you to enter your desired name and password. Did that and clicked forward. The Ubuntu 10.04 installation load screen appeared and the loading bar at the bottom started filling up. So I got to 83% and stalled during the Importing other profile information (I think it was called this. I had the option to do this during I think step 6) process. So at this point I decided to get stop the installation process. I was getting very nervous. I tried to restart the computer but all that happened was that Ubuntu restarted. I finally got the computer to restart. I was pretty sure I had screwed something up big time by this point. As my computer was restarting I entered BIOS again and switched back to it booting from my main hard drive containing Vista. Saved it and continued the boot process. My worst fears were confirmed as Vista would not boot up. I mean I saw the little Microsoft Windows choppy animated green loading bar at the bottom of the screen and then boom! It decided to restart. When it restarted I had the option to run a memory test check to see if there was anything that needed to be repaired. That took like 20 minutes and at the end I saw that I did indeed have to repair something. I had to go through 2 repair processes. After each I had to restart the computer. The 2nd time it went through the repair process it said that it could not fully repair the damage. I was scared and restarted but Vista did load up. I got to my desktop and saw a message saying something like Repairs have been made, Please restart for changes to take effect I noticed that some Notification icons were missing and I could not hear volume in a video. Things were a bit funky. So I did restart and here I am. Now what?! So since I got back into Vista and thankfully have a working Internet connection I am trying to find answers to my problem (that is why I am writing this post). I am scared that I have partioned my hard drive 2 times after researching Installing Ubuntu 10.04 and seeing this post http://techie-buzz.com/foss/ubuntu-10-04-lts-installation-guide.html The author shows screen shots of installing Ubuntu 10.04. He shows the image of step 4 of 7 with a caption at the bottom. I will recreate it below: Select a partitioning option. Unless you want to format all the hard drive and install Ubuntu afresh, select the last option and proceed. Questions If I have indeed partitioned my HD 2 times (which I am sure it is), how do I get to a point where I can see all my bad, unfinished Ubuntu partitions and get rid of them? How do I clean this big mess up? & How can I ensure that this mess will not happen next time I try installing Ubuntu 10.04? Thank you Adam

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• Cross-platform distributed fault-tolerant (disconnected operation/local cache) filesystem

  - by Adrian Frühwirth

  We are facing a design "challenge" where we are required to set up a storage solution with the following properties: What we need HA a scalable storage backend offline/disconnected operation on the client to account for network outages cross-platform access client-side access from certainly Windows (probably XP upwards), possibly Linux backend integrates with AD/LDAP (permission management (user/group management, ...)) should work reasonably well over slow WAN-links Another problem is that we don't really know all possible use cases here, if people need to be able to have concurrent access to shared files or if they will only be accessing their own files, so a possible solution needs to account for concurrent access and how conflict management would look in this case from a user's point of view. This two years old blog posts sums up the impression that I have been getting during the last couple of days of research, that there are lots of current übercool projects implementing (non-Windows) clustered petabyte-capable blob-storage solutions but that there is none that supports disconnected operation nicely and natively, but I am hoping that we have missed an obvious solution. What we have tried OpenAFS We figured that we want a distributed network filesystem with a local cache and tested OpenAFS (which, as the only currently "stable" DFS supporting disconnected operation, seemed the way to go) for a week but there are several problems with it: it's a real pain to set up there are no official RHEL/CentOS packages the package of the current stable version 1.6.5.1 from elrepo randomly kernel panics on fresh installs, this is an absolute no-go Windows support (including the required Kerberos packages) is mystical. The current client for the 1.6 branch does not run on Windows 8, the current client for the 1.7 does but it just randomly crashes. After that experience we didn't even bother testing on XP and Windows 7. Suffice to say, we couldn't get it working and the whole setup has been so unstable and complicated to setup that it's just not an option for production. Samba + Unison Since OpenAFS was a complete disaster and no other DFS seems to support disconnected operation we went for a simpler idea that would sync files against a Samba server using Unison. This has the following advantages: Samba integrates with ADs; it's a pain but can be done. Samba solves the problem of remotely accessing the storage from Windows but introduces another SPOF and does not address the actual storage problem. We could probably stick any clustered FS underneath Samba, but that means we need a HA Samba setup on top of that to maintain HA which probably adds a lot of additional complexity. I vaguely remember trying to implement redundancy with Samba before and I could not silently failover between servers. Even when online, you are working with local files which will result in more conflicts than would be necessary if a local cache were only touched when disconnected It's not automatic. We cannot expect users to manually sync their files using the (functional, but not-so-pretty) GTK GUI on a regular basis. I attempted to semi-automate the process using the Windows task scheduler, but you cannot really do it in a satisfactory way. On top of that, the way Unison works makes syncing against Samba a costly operation, so I am afraid that it just doesn't scale very well or even at all. Samba + "Offline Files" After that we became a little desparate and gave Windows "offline files" a chance. We figured that having something that is inbuilt into the OS would reduce administrative efforts, helps blaming someone else when it's not working properly and should just work since people have been using this for years. Right? Wrong. We really wanted it to work, but it just doesn't. 30 minutes of copying files around and unplugging network cables/disabling network interfaces left us with (silent! there is only a tiny notification in Windows explorer in the statusbar, which doesn't even open Sync Center if you click on it!) undeletable files on the server (!) and conflicts that should not even be conflicts. In the end, we had one successful sync of a tiny text file, everything else just exploded horribly. Beyond that, there are other problems: Microsoft admits that "offline files" in Windows XP cannot cope with "large files" and therefore does not cache/sync them at all which would mean those files become unavailable if the connection drop In Windows 7 the feature is only available in the Professional/Ultimate/Enterprise editions. Summary Unless there is another fault-tolerant DFS that supports Windows natively I assume that stacking a HA Samba cluster on top of something like GlusterFS/Lustre/whatnot is the only option, but I hope that I am wrong here. How do other companies allow fault-tolerant network access to redundant storage in a heterogeneous environment with Windows?

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• Add collection or array to wpf resource dictionary

  - by Chris Cap

  I've search high and low and can't find an answer to this. I have two questions How do you create an array or collection in XAML. I've got an array I want to stick in there and bind to a combo box. My first idea was to put an ItemsControl in a resource dictionary, but the ItemsSource of a combo box expects IEnumerable so that didn't work. Here's what I've tried in my resource dictionary and neither works <ItemsControl x:Key="stateList"> <sys:String>AL</sys:String> <sys:String>CA</sys:String> <sys:String>CN</sys:String> </ItemsControl> <ItemsControl x:Key="stateList2"> <ComboBoxItem>AL</ComboBoxItem> <ComboBoxItem>CA</ComboBoxItem> <ComboBoxItem>CN</ComboBoxItem> </ItemsControl> and here's how I bind to it <ComboBox SelectedValue="{Binding Path=State}" ItemsSource="{Binding Source={StaticResource stateList2}}" > </ComboBox> EDIT: UPDATED I got this first part to work this way <col:ArrayList x:Key="stateList3"> <sys:String>AL</sys:String> <sys:String>CA</sys:String> <sys:String>CN</sys:String> </col:ArrayList> However, I'd rather not use an array list, I'd like to use a generic list so if anyone knows how please let me know. EDIT UPDATE: I guess XAML has very limited support for generics so maybe an array list is the best I can do for now, but I would still like help on the second question if anyone has an anser 2nd. I've tried referencing a merged resource dictionary in my XAML and had problems because under Window.resources I had more than just the dictionary so it required me to add x:Key. Once I add the key, the system can no longer find the items in my resource dictionary. I had to move the merged dictionary to Grid.Resources instead. Ideally I'd like to reference the merged dictionary in the app.xaml but I have the same problem Here's some sample code. This first part required an x:key to compile because I have converter and it complained that every item must have a key if there is more than one <UserControl.Resources> <win:BooleanToVisibilityConverter x:Key="VisibilityConverter" /> <ResourceDictionary> <ResourceDictionary.MergedDictionaries> <ResourceDictionary Source="/ResourcesD.xaml" /> </ResourceDictionary.MergedDictionaries> </ResourceDictionary> </UserControl.Resources> I had to change it to this <UI:BaseStep.Resources> <win:BooleanToVisibilityConverter x:Key="VisibilityConverter" /> </UI:BaseStep.Resources> <Grid> <Grid.Resources> <ResourceDictionary> <ResourceDictionary.MergedDictionaries> <ResourceDictionary Source="/ResourcesD.xaml" /> </ResourceDictionary.MergedDictionaries> </ResourceDictionary> </Grid.Resources> </Grid> Thank you

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• Parallax backgrounds in OpenGL ES on the iPhone

  - by Scott

  I've got basically a 2d game on the iPhone and I'm trying to set up multiple backgrounds that scroll at different speeds (known as parallax backgrounds). So my thought was to just stick the backgrounds BEHIND the foreground using different z-coordinate planes, and just make them bigger than the foreground (in size) to accommodate, so that the whole thing can be scrolled (just at a different speed). And (as far as I know) I basically implemented that. The only problem is that it seems to entirely ignore whatever z-value I give it, or rather it just zeroes all of them. I see the background (I've only tested ONE background so far, to keep it simple...so for now I just have a foreground and I want one background scrolling at a different speed), but it scrolls 1:1 with my foreground, so it obviously doesn't look right, and most of it is cut off (cause it's bigger). And I've tried various z-values for the background and various near/far clipping planes...it's always the same. I'm probably just doing one simple thing wrong, but I can't figure it out. I'm wondering if it has to do with me using only 2 coordinates in glVertexPointer for the foreground? (Of course for the background I AM passing in 3) I'll post some code: This is some initial setup: glMatrixMode(GL_PROJECTION); glLoadIdentity(); glOrthof(-1.0f, 1.0f, -1.5f, 1.5f, -10.0f, 10.0f); glMatrixMode(GL_MODELVIEW); glLoadIdentity(); glEnableClientState(GL_VERTEX_ARRAY); //glEnableClientState(GL_COLOR_ARRAY); glEnableClientState(GL_TEXTURE_COORD_ARRAY); //transparency glEnable (GL_BLEND); glBlendFunc (GL_ONE, GL_ONE_MINUS_SRC_ALPHA); A little bit about my foreground's float array....it's interleaved. For my foreground it goes vertex x, vertex y, texture x, texture y, repeat. This all works just fine. This is my FOREGROUND rendering: glVertexPointer(2, GL_FLOAT, 4*sizeof(GLfloat), texes); <br> glTexCoordPointer(2, GL_FLOAT, 4*sizeof(GLfloat), (GLvoid*)texes + 2*sizeof(GLfloat)); <br> glDrawArrays(GL_TRIANGLES, 0, indexCount / 4); BACKGROUND rendering: Same drill here except this time it goes vertex x, vertex y, vertex z, texture x, texture y, repeat. Note the z value this time. I did make sure the data in this array was correct while debugging (getting the right z values). And again, it shows up...it's just not going far back in the distance like it should. glVertexPointer(3, GL_FLOAT, 5*sizeof(GLfloat), b1Texes); glTexCoordPointer(2, GL_FLOAT, 5*sizeof(GLfloat), (GLvoid*)b1Texes + 3*sizeof(GLfloat)); glDrawArrays(GL_TRIANGLES, 0, b1IndexCount / 5); And to move my camera, I just do a simple glTranslatef(x, y, 0.0f); I'm not understanding what I'm doing wrong cause this seems like the most basic 3D function imaginable...things further away are smaller and don't move as fast when the camera moves. Not the case for me. Seems like it should be pretty basic and not even really be affected by my projection and all that (though I've even tried doing glFrustum just for fun, no success). Please help, I feel like it's just one dumb thing. I will post more code if necessary.

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• HttpWebRequest: How to find a postal code at Canada Post through a WebRequest with x-www-form-enclos

  - by Will Marcouiller

  I'm currently writing some tests so that I may improve my skills with the Internet interaction through Windows Forms. One of those tests is to find a postal code which should be returned by Canada Post website. My default URL setting is set to: http://www.canadapost.ca/cpotools/apps/fpc/personal/findByCity?execution=e4s1 The required form fields are: streetNumber, streetName, city, province The contentType is "application/x-www-form-enclosed" EDIT: Please consider the value "application/x-www-form-encoded" instead of point 3 value as the contentType. (Thanks EricLaw-MSFT!) The result I get is not the result expected. I get the HTML source code of the page where I could manually enter the information to find the postal code, but not the HTML source code with the found postal code. Any idea of what I'm doing wrong? Shall I consider going the XML way? Is it first of all possible to search on Canada Post anonymously? Here's a code sample for better description: public static string FindPostalCode(ICanadadianAddress address) { var postData = string.Concat(string.Format("&streetNumber={0}", address.StreetNumber) , string.Format("&streetName={0}", address.StreetName) , string.Format("&city={0}", address.City) , string.Format("&province={0}", address.Province)); var encoding = new ASCIIEncoding(); byte[] postDataBytes = encoding.GetBytes(postData); request = (HttpWebRequest)WebRequest.Create(DefaultUrlSettings); request.ImpersonationLevel = System.Security.Principal.TokenImpersonationLevel.Anonymous; request.Container = new CookieContainer(); request.Timeout = 10000; request.ContentType = contentType; request.ContentLength = postDataBytes.LongLength; request.Method = @"post"; var senderStream = new StreamWriter(request.GetRequestStream()); senderStream.Write(postDataBytes, 0, postDataBytes.Length); senderStream.Close(); string htmlResponse = new StreamReader(request.GetResponse().GetResponseStream()).ReadToEnd(); return processedResult(htmlResponse); // Processing the HTML source code parsing, etc. } I seem stuck in a bottle neck in my point of view. I find no way out to the desired result. EDIT: There seems to have to parameters as for the ContentType of this site. Let me explain. There's one with the "meta"-variables which stipulates the following: meta http-equiv="Content-Type" content="application/xhtml+xml, text/xml, text/html; charset=utf-8" And another one later down the code that is read as: form id="fpcByAdvancedSearch:fpcSearch" name="fpcByAdvancedSearch:fpcSearch" method="post" action="/cpotools/apps/fpc/personal/findByCity?execution=e1s1" enctype="application/x-www-form-urlencoded" My question is the following: With which one do I have to stick? Let me guess, the first ContentType is to be considered as the second is only for another request to a function or so when the data is posted? EDIT: As per request, the closer to the solution I am is listed under this question: WebRequest: How to find a postal code using a WebRequest against this ContentType=”application/xhtml+xml, text/xml, text/html; charset=utf-8”? Thanks for any help! :-)

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• Silverlight Image Loading Question

  - by Matt

  I'm playing around with Silverlight Images and a listbox. Here's the scenario. Using WCF I grab some images out of my database and, using a custom class, add items to a listbox. It's working great right now. The images load and appear in the listbox, just like I want them to. I want to refine and improve my control just a little more so here's what I've done. <ListBox x:Name="lbMedia" Background="Transparent" ScrollViewer.HorizontalScrollBarVisibility="Disabled" ScrollViewer.VerticalScrollBarVisibility="Auto"> <ItemsControl.ItemsPanel> <ItemsPanelTemplate> <c:WrapPanel></c:WrapPanel> </ItemsPanelTemplate> </ItemsControl.ItemsPanel> <ItemsControl.ItemTemplate> <DataTemplate> <im:MediaManagerItem></im:MediaManagerItem> </DataTemplate> </ItemsControl.ItemTemplate> </ListBox> Just a simple listbox. The datatemplate is a custom control and literally it contains a contentpresenter, nothing more. Now the class that I use as the ItemSource has a Source property. Here's what it looks like. private UIElement _LoadingSource; private UIElement _Source; public UIElement Source { get { if( _Source == null ) { LoadMedia(); return new LoadingElement(); } return _Source; } set { if( !( value is Image ) && !( value is MediaElement ) ) throw new Exception( "Media Source must be an Image or MediaElement" ); _Source = value; NotifyPropertyChanged( "Source" ); } } Essentially, on the get I check if the image/video has been loaded from the server. If it hasn't I return a loading control, then I proceed to load my image. Here's the code for my LoadMedia method. private void LoadMedia() { if( _Media != null && _Media.MediaId > 0 ) { //load the media BackgroundWorker mediaLoader = new BackgroundWorker(); mediaLoader.DoWork += mediaLoader_DoWork; mediaLoader.RunWorkerCompleted += mediaLoader_RunWorkerCompleted; mediaLoader.RunWorkerAsync(); } } void mediaLoader_RunWorkerCompleted( object sender, RunWorkerCompletedEventArgs e ) { if(_LoadingSource != null) Source = _LoadingSource; } void mediaLoader_DoWork( object sender, DoWorkEventArgs e ) { string url = App.siteUrl + "download.ashx?MediaId=" + _Media.MediaId; SmartDispatcher.BeginInvoke( () => { Image img = new Image(); img.Source = new BitmapImage( new Uri( url, UriKind.Absolute ) ); _LoadingSource = img; } ); } So as the code goes, I create a new image element, and set the Uri. The images that I'm downloading take about 2-5 seconds to download. Now for the problem / fine tuning. Right now my code will check if the source is null and if it is, return a loading element, and run the background worker to get the image. Once the background worker finishes, set the source to the new downloaded image. I want to be able to set the Source property AFTER the image has fully downloaded. Right now my loading element appears for a brief second, then there's nothing for 2-5 seconds until the image finishes downloading. I want the loading elements to stick around until the image is completely ready but I'm having troubles doing this. I've tried adding a a listener to the ImageOpened event and update the Source property then, but it doesn't work. Thanks in advance.

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• How do I evaluate my skillset against the current market to see what needs improvement and where my

  - by baijajusav

  First of all, this question may be out of bounds for this site. If so, remove it. I say this because this site seems to be a place for more concrete questions that are not so relative in nature. And before I begin, for those of you whom just prefer a question and not this sort of dialog, here is my question: How can I assess my current skills as a programmer and decide where and what areas to improve upon? That said, here's what I'm asking/talking about, in essence. The market is always in constant flux. As programmers we're always having to learn new things, update our skills, push ourselves into that next project. There's not a very good litmus test that I know of for us to get an idea of where we stand as programmers. I came across this blog post by Jeff Atwood talking about why can't programmers code. Instinctively (and as the post goes on to state) I rushed through the program in about 4 minutes (most of that time was b/c I was hand writing it out. Still, this doesn't really answer the question of where do my skills need to be to succeed in today's world. I real blogs, listen to podcasts, try to keep up on the latest things coming out. It has only been in the past couple of months that I made a decision to pick a focus area for my learning as I can't learn everything and trying to do so is to spread myself too thin. I chose ASP.NET MVC & C#. I plan to stick with Microsoft technologies, not out of some sense of loyalty or stubbornness, but rather because they seem to stream together and have a unifying connection between them. With Windows Phone 7 coming out, it seems that now is the obvious time to pick up WPF and Silverlight as well. Still, if you asked me to code something apart from intellisense and the internet, I probably couldn't get the syntax right. I don't have libraries memorized or know precisely where the classes I use exist within the .Net framework, namely because I haven't had to pull that knowledge out of the air. In a way, I suppose Visual Studio has insulated me, which isn't a good thing, but, at the same time, I've still been able to be productive. I'm working on my own side project to try and help my learning. In doing so, I'm trying to make use of best practices and 3rd party frameworks where I can. I'm using automapper and EF 1.0. I know everyone in the .net community seems to cry foul at the sound of EF 1.0, but I can't say why because I've never used it. There's no lazy loading and that has proven rather annoying; however, aside from that, I haven't had that much of an issue. Granted this is probably because I'm not writing tests as I go (which I'm not doing because I don't know how to test EF in tests and don't really have a clue how to write tests for ASP.NET MVC 1.0). I'm also using a custom membership provider; granted, it's a barebone implementation, but I'm using it still. My thinking in all of this is, while I am neglecting a great many important technologies that are in the mainstream, I'll have a working project in the end. I can come back and add those things after I finish. Doing it all now and at once seems like too much. I know how I work and I don't think I'd ever get it done that way. I've elected to make this a community wiki as I think this question might fight better there. If a moderator disagrees with that choice or the decision to post this here, the just delete the question. I'm not trying to make undue work for anyone. I'm just a programmer trying to assess my where his skills are now and where I should be improving.

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• In Ruby, how to I read memory values from an external process?

  - by grg-n-sox

  So all I simply want to do is make a Ruby program that reads some values from known memory address in another process's virtual memory. Through my research and basic knowledge of hex editing a running process's x86 assembly in memory, I have found the base address and offsets for the values in memory I want. I do not want to change them; I just want to read them. I asked a developer of a memory editor how to approach this abstract of language and assuming a Windows platform. He told me the Win32API calls for OpenProcess, CreateProcess, ReadProcessMemory, and WriteProcessMemory were the way to go using either C or C++. I think that the way to go would be just using the Win32API class and mapping two instances of it; One for either OpenProcess or CreateProcess, depending on if the user already has th process running or not, and another instance will be mapped to ReadProcessMemory. I probably still need to find the function for getting the list of running processes so I know which running process is the one I want if it is running already. This would take some work to put all together, but I am figuring it wouldn't be too bad to code up. It is just a new area of programming for me since I have never worked this low level from a high level language (well, higher level than C anyways). I am just wondering of the ways to approach this. I could just use a bunch or Win32API calls, but that means having to deal with a bunch of string and array pack and unpacking that is system dependant I want to eventually make this work cross-platform since the process I am reading from is produced from an executable that has multiple platform builds, (I know the memory address changes from system to system. The idea is to have a flat file that contains all memory mappings so the Ruby program can just match the current platform environment to the matching memory mapping.) but from the looks of things I'll just have to make a class that wraps whatever is the current platform's system shared library memory related function calls. For all I know, there could already exist a Ruby gem that takes care of all of this for me that I am just not finding. I could also possibly try editing the executables for each build to make it so whenever the memory values I want to read from are written to by the process, it also writes a copy of the new value to a space in shared memory that I somehow have Ruby make an instance of a class that is a pointer under the hood to that shared memory address and somehow signal to the Ruby program that the value was updated and should be reloaded. Basically a interrupt based system would be nice, but since the purpose of reading these values is just to send to a scoreboard broadcasted from a central server, I could just stick to a polling based system that sends updates at fixed time intervals. I also could just abandon Ruby altogether and go for C or C++ but I do not know those nearly as well. I actually know more x86 than C++ and I only know C as far as system independent ANSI C and have never dealt with shared system libraries before. So is there a gem or lesser known module available that has already done this? If not, then any additional information as to how to accomplish this would be nice. I guess, long story short, how do I do all this? Thanks in advance, Grg PS: Also a confirmation that those Win32API calls should be aimed at the kernel32.dll library would be nice.

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• Both tab & hover triggered popups problem

  - by carpenter

  I am trying to display divs when hovering over thumb-nails and/or both when tabbing onto them. If I stick to my mouse, the popups seem to work OK - if I start with a tab press I can show the popops also (foward only - no shift + tab yet). Any help getting them to play well together? <script type="text/javascript"> // Note: the below is being run from an onmouseover on a asp:HyperLink at the moment function onhovering_and_tabbingon2() { var active_hover = 0; var num_of_thumb; // set the default focus onto the first thumb-nail and make its popup display document.getElementById('link_no' + active_hover).focus(); // set focus on the first thumb $('#pop' + active_hover).toggleClass('popup'); // show its popup as it is hidden // for when hovering over the thumbs $(".box img").hover( // so as to effect only images/thumb-nails within divs of class=box when hovering over them function () { // test for if the image is a thumb-entry and not a popup image - of class=thumbs2 thumb = $(this).attr('class'); if (thumb != "thumbs2") { // I need to add/toggle the class here to a "div" and not to the image being hovered on, a div with text that corrosponds to the hovered on image though // so grab the number of the thumb_entry - to use to id the div. num_of_thumb = $(this).attr('id').replace('thumb_entry_No', ''); // find the div with id 'pop' + num_of_thumb, and toggleClass on it $('#pop' + num_of_thumb).toggleClass('popup'); // shows the hovered on pic's popup // move the focus to the hovered on pic's a tag ?????? document.getElementById('link_no' + num_of_thumb).focus(); // if the previous popup that was showing was in box2.. if (active_hover == 1 || active_hover% 2 == 1) { $('#pop' + active_hover).toggleClass('popup4_line2'); } else { // remove/toggle the previous active popup's visibility $('#pop' + active_hover).toggleClass('popup'); } // set the new active_hover to num_of_thumb active_hover = num_of_thumb; } }, function () { } ); // same thing again - but for my second row/line of entries/thumb-nails... $(".box2 img").hover( // so as to effect only images/thumbs within divs of class=box2 function () { // test if the image is a thumb-entry and not a popup image thumb = $(this).attr('class'); if (thumb != "thumbs2") { // I need to add the class here to a "div" and not to the image being hovered on, a div that corrosponds to the hovered on image though // so grab the number of the thumb_entry being hovered on, so as to id the div. num_of_thumb = $(this).attr('id').replace('thumb_entry_No', ''); // find the div with id='pop' + num_of_thumb, and toggleClass on it $('#pop' + num_of_thumb).toggleClass('popup4_line2'); // move the focus to the hovered on pic's a tag ?? document.getElementById('link_no' + num_of_thumb).focus(); // if the previous popup that was showing was in box.. // or if the active_hover is even (modulus) if (active_hover == 0 || active_hover % 2 == 0) { $('#pop' + active_hover).toggleClass('popup'); } else { // remove the previous active visible popup $('#pop' + active_hover).toggleClass('popup4_line2'); } // set the new active_hover to num_of_thumb active_hover = num_of_thumb; } }, function () { } ); // todo: I would like to try to show the popups when tabbing through the thumb-nails also // but am lost... document.onkeyup = keypress; // ???? function keypress() { // alert("The key pressed was: " + window.event.keyCode); if (window.event.keyCode == "9") { //alert("The tab key was pressed!"); active_hover = active_hover + 1; // for tabbing into box 2 (odd numbers) if (active_hover == 1 || active_hover % 2 == 1) { // toggle visibility of previous popup $('#pop' + (active_hover - 1)).toggleClass('popup'); // toggle visibility of current popup $('#pop' + active_hover).toggleClass('popup4_line2'); // } else { // for tabbing into box from box2 // toggle visibility of previous popup $('#pop' + (active_hover - 1)).toggleClass('popup4_line2'); // toggle visibility of current popup $('#pop' + active_hover).toggleClass('popup'); // } // ?????? // // if (window.event.keyCode == "shift&9") { } } } } </script>

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• Clear data at serial port in Linux in C?

  - by ipkiss

  Hello guys, I am testing the sending and receiving programs with the code as The main() function is below: include include include include include include include "read_write.h" int fd; int initport(int fd) { struct termios options; // Get the current options for the port... tcgetattr(fd, &options); // Set the baud rates to 19200... cfsetispeed(&options, B9600); cfsetospeed(&options, B9600); // Enable the receiver and set local mode... options.c_cflag |= (CLOCAL | CREAD); options.c_cflag &= ~PARENB; options.c_cflag &= ~CSTOPB; options.c_cflag &= ~CSIZE; options.c_cflag |= CS8; // Set the new options for the port... tcsetattr(fd, TCSANOW, &options); return 1; } int main(int argc, char **argv) { fd = open("/dev/pts/2", O_RDWR | O_NOCTTY | O_NDELAY); if (fd == -1) { perror("open_port: Unable to open /dev/pts/1 - "); return 1; } else { fcntl(fd, F_SETFL, 0); } printf("baud=%d\n", getbaud(fd)); initport(fd); printf("baud=%d\n", getbaud(fd)); char sCmd[254]; sCmd[0] = 0x41; sCmd[1] = 0x42; sCmd[2] = 0x43; sCmd[3] = 0x00; if (!writeport(fd, sCmd)) { printf("write failed\n"); close(fd); return 1; } printf("written:%s\n", sCmd); usleep(500000); char sResult[254]; fcntl(fd, F_SETFL, FNDELAY); if (!readport(fd,sResult)) { printf("read failed\n"); close(fd); return 1; } printf("readport=%s\n", sResult); close(fd); return 0; } read_write.h: #include <stdio.h> /* Standard input/output definitions */ include /* String function definitions */ include /* UNIX standard function definitions */ include /* File control definitions */ include /* Error number definitions */ include /* POSIX terminal control definitions */ int writeport(int fd, char *chars) { int len = strlen(chars); chars[len] = 0x0d; // stick a after the command chars[len+1] = 0x00; // terminate the string properly int n = write(fd, chars, strlen(chars)); if (n < 0) { fputs("write failed!\n", stderr); return 0; } return 1; } int readport(int fd, char *result) { int iIn = read(fd, result, 254); result[iIn-1] = 0x00; if (iIn < 0) { if (errno == EAGAIN) { printf("SERIAL EAGAIN ERROR\n"); return 0; } else { printf("SERIAL read error %d %s\n", errno, strerror(errno)); return 0; } } return 1; } and got the issue: In order to test with serial port, I used the socat (https://help.ubuntu.com/community/VirtualSerialPort ) to create a pair serial ports on Linux and test my program with these port. The first time the program sends the data and the program receives data is ok. However, if I read again or even re-write the new data into the serial port, the return data is always null until I stop the virtual serial port and start it again, then the write and read data is ok, but still, only one time. (In the real case, the sending part will be done by another device, I am just taking care of the reading data from the serial port. I wrote both parts just to test my reading code.) Does anyone have any ideas? Thanks a lot.

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