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  • Windows Mobile 6.5 GPS Device - WaitForMultipleObjects returns 258 (timeout)

    - by wizmagister
    I’ve created a GPS program that track positions in realtime in the background for Windows mobile 6.1 in 2008-2009. It ran fine on these devices for many years. For some reason, the same code never worked perfectly on Windows Mobile 6.5. After many hour of operations (mostly when nobody use the device), I receive a “Timeout” (code 258) from the function "WaitForMultipleObjects": this.GPSEvent_WaitValue = WaitForMultipleObjects(2, this.GPSEvent_Handles, 0, 45000); Again, this can work for hours and suddenly, it's just impossible to get another position without : UPDATE: - Restarting the device (GoogleMap confirms that there's no GPS device present!) It has something to do with Windows Mobile going to sleep and slowing up my thread. Here's the core code (adapted from Microsoft SDK Sample): /// <summary> /// When "WindowsMobile" wake up the program to check for a new position /// </summary> private void OnNextGPSEvent_Callback() { int SecondsToNextWakeUp = ETL.Mobile.Device.ScheduledCallback.MINIMUM_SECONDTONEXTWAKEUP; switch (this.SleepingState) { case SleepingStateType.SleepingForNextPosition: // Get position this.GPSEvent_WaitValue = (WaitForEventThreadResultType)WaitForMultipleObjects(2, this.GPSEvent_Handles, 0, 45000); switch (this.GPSEvent_WaitValue) { case WaitForEventThreadResultType.Event_LocationChanged: // Got a new position this.FireLocationChanged(this.GetCurrentPosition()); // Manage device shutdown (save battery) if (this.PositionFrequency > MIN_SECONDS_FREQUENCY_FORDEVICE_SHUTDOWN) { // Close device this.CloseDevice(); SecondsToNextWakeUp = (this.PositionFrequency - GPSDEVICE_LOAD_SECONDS_LOAD_TIME); this.SleepingState = SleepingStateType.SleepingBeforeDeviceWakeUp; } else { // Default Wait Time this.SleepingState = SleepingStateType.SleepingForNextPosition; } break; case WaitForEventThreadResultType.Event_StateChanged: break; case WaitForEventThreadResultType.Timeout: case WaitForEventThreadResultType.Failed: case WaitForEventThreadResultType.Stop: // >>>>>>>>>>>>>> This is where the error happens <<<<<<<<<<<<<<<<<<<<<<<<<<< // >>>>>>>>>>>>>> This is where the error happens <<<<<<<<<<<<<<<<<<<<<<<<<<< // >>>>>>>>>>>>>> This is where the error happens <<<<<<<<<<<<<<<<<<<<<<<<<<< // Too many errors this.ConsecutiveErrorReadingDevice++; if (this.ConsecutiveErrorReadingDevice > MAX_ERRORREADINGDEVICE) { this.CloseDevice(); SecondsToNextWakeUp = (this.PositionFrequency - GPSDEVICE_LOAD_SECONDS_LOAD_TIME); this.SleepingState = SleepingStateType.SleepingBeforeDeviceWakeUp; } else { // Default Wait Time this.SleepingState = SleepingStateType.SleepingForNextPosition; } break; } #endregion break; case SleepingStateType.SleepingBeforeDeviceWakeUp: this.OpenDevice(); SecondsToNextWakeUp = GPSDEVICE_LOAD_SECONDS_LOAD_TIME; this.SleepingState = SleepingStateType.SleepingForNextPosition; break; } if (this.IsListeningGPSEvent) { // Ajustement du prochain rappel this.NextGPSEvent_Callback.SecondToNextWakeUp = SecondsToNextWakeUp; this.NextGPSEvent_Callback.RequestWakeUpCallback(); } } /// <summary> ///Create Thread /// </summary> private void StartListeningThreadForGPSEvent() { // We only want to create the thread if we don't have one created already and we have opened the gps device if (this._GPSEventThread == null) { // Create and start thread to listen for GPS events this._GPSEventThread = new System.Threading.Thread(new System.Threading.ThreadStart(this.ListeningThreadForGPSEvent)); this._GPSEventThread.Start(); } } private void ListeningThreadForGPSEvent() { this.GPSEvent_WaitValue = WaitForEventThreadResultType.Stop; this.IsListeningGPSEvent = true; // Allocate handles worth of memory to pass to WaitForMultipleObjects this.GPSEvent_Handles = Helpers.LocalAlloc(12); Marshal.WriteInt32(this.GPSEvent_Handles, 0, this._StopHandle.ToInt32()); Marshal.WriteInt32(this.GPSEvent_Handles, 4, this._NewLocationHandle.ToInt32()); Marshal.WriteInt32(this.GPSEvent_Handles, 8, this._GPSDeviceStateChanged.ToInt32()); this.Start_NextGPSEvent_Timer(this.PositionFrequency); this.SleepingState = SleepingStateType.SleepingBeforeDeviceWakeUp; this.OnNextGPSEvent_Callback(); }

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  • Quick guide to Oracle IRM 11g: Server configuration

    - by Simon Thorpe
    Quick guide to Oracle IRM 11g index Welcome to the second article in this quick quide to Oracle IRM 11g. Hopefully you've just finished the first article which takes you through deploying the software onto a Linux server. This article walks you through the configuration of this new service and contains a subset of information from the official documentation and is focused on installing the server on Oracle Enterprise Linux. If you are planning to deploy on a non-Linux platform, you will need to reference the documentation for platform specific information. Contents Introduction Create IRM WebLogic Domain Starting the Admin Server and initial configuration Introduction In the previous article the database was prepared, the WebLogic Application Server installed and the files required for an IRM server installed. But we don't actually have a configured system yet. We need to now create a WebLogic Domain in which the IRM server will run, then configure some of the settings and crypography so that we can create a context and be ready to seal some content and test it all works. This article doesn't cover the configuration of SSL communication from client to server. This is quite a big topic and a separate article has been dedicated for this area. In these articles I also use the hostname, irm.company.internal to reference the IRM server and later on use the hostname irm.company.com in reference to the public facing service. Create IRM WebLogic Domain First step is creating the WebLogic domain, in a console switch to the newly created IRM installation folder as shown below and we will run the domain configuration wizard. [oracle@irm /]$ cd /oracle/middleware/Oracle_IRM/common/bin [oracle@irm bin]$ ./config.sh First thing the wizard will ask is if you wish to create a new or extend an existing domain. This guide is creating a standalone system so you should select to create a new domain. Next step is to choose what technologies from the Oracle ECM Suite you wish this domain to host. You are only interested in selecting the option "Oracle Information Rights Management". When you select this check box you will notice that it also selects "Oracle Enterprise Manager" and "Oracle JRF" as these are dependencies of the IRM server. You then need to specify where you wish to place the domain files. I usually just change the domain name from base_domain or irm_domain and leave the others with their defaults. Now the domain will have a single user initially and by default this user is called "weblogic". I usually change this account name to "sysadmin" or "administrator", but in this guide lets just accept the default. With respects to the next dialog, again for eval or dev reasons, leave the server startup mode as development. The JDK should also be automatically detected. We now need to provide details of the database. This guide is using the Oracle 11gR2 database and the settings I used can be seen in the image to the right. There is a lot of configuration that can now be done for the admin server, any managed servers and where the deployments reside. In this guide I am leaving all of these to their defaults so do not check any of the boxes. However I will on this blog be detailing later how you can go back and setup things such as automated startup of an IRM server which require changes to these default settings. But for now, lets leave it all alone and just click next. Now we are ready to install. Note that from this dialog you can scroll the left window and see there are going to be two servers created from the defaults. The AdminServer which is where you modify settings for the WebLogic Server and also hosts the Oracle Enterprise Manager for IRM which allows to monitor the IRM service performance and also make service related settings (which we shortly do below) and the IRM_server1 which hosts the actual IRM services themselves. So go right ahead and hit create, the process is pretty quick and usually under 10 minutes. When the domain creation ends, it will give you the URL to the admin server. It's worth noting this down and the URL is usually; http://irm.company.internal:7001 Starting the Admin Server and initial configuration First thing to do is to start the WebLogic Admin server and review the initial IRM server settings. In this guide we are going to run the Admin server and IRM server in console windows, in another article I will discuss running these as background services. So for now, start a console and run the Admin server by doing the following. cd /oracle/middleware/user_projects/domains/irm_domain/ ./startWebLogic.sh Wait for the server to start, you are looking for the following line to be reported in the console window. <BEA-00360><Server started in RUNNING mode> First step is configuring the IRM service via Enterprise Manager. Now that the Admin server is running you can point a browser at http://irm.company.internal:7001/em. Login with the username and password you supplied when you created the domain. In Enterprise Manager the IRM service administrator is able to make server wide configuration. However finding where to access the pages with these settings can be a bit of a challenge. After logging in on the left you'll see a tree containing elements of the Enterprise Manager farm Farm_irm_domain. Open up Content Management, then Information Rights Management and finally select the IRM node. On the right then select the IRM menu item, navigate to the Administration section and now we have four options, for now, we are just going to look at General Settings. The image on the right proves that a picture is worth a thousand words (or 113 in this case). The General Settings page allows you to set the cryptographic algorithms used for protecting sealed content. Unless you have a burning need to increase the key lengths or you need to comply to a regulation or government mandate, AES192 is a good start. You can change this later on without worry. The most important setting here we need to make is the Server URL. In this blog article I go over why this URL is so important, basically every single piece of content you protect with Oracle IRM is going to have this URL embedded in it, so if it's wrong or unresolvable, then nobody can open the secured documents. Note that in our environment we have yet to do any SSL configuration of the service. If you intend to build a server without SSL, then use http as the protocol instead of https. But I would recommend using SSL and setting this up is described in the next article. I would also probably up the device count from 1 to 3. This means that any user can retrieve rights to access content onto 3 computers at any one time. The default of 1 doesn't really make sense in development, evaluation nor even production environments and my experience is that 3 is a better number. Next step is to create the keystore for the IRM server. When a classification (called a context) is created, Oracle IRM generates a unique set of symmetric keys which are used to secure the content itself. These keys are then encrypted with a set of "wrapper" asymmetric cryptography keys which are stored externally to the server either in a Java Key Store or a HSM. These keys need to be generated and the following shows my commands and the resulting output. I have greyed out the responses from the commands so you can see the input a little easier. [oracle@irmsrv ~]$ cd /oracle/middleware/wlserver_10.3/server/bin/ [oracle@irmsrv bin]$ ./setWLSEnv.sh CLASSPATH=/oracle/middleware/patch_wls1033/profiles/default/sys_manifest_classpath/weblogic_patch.jar:/oracle/middleware/patch_ocp353/profiles/default/sys_manifest_classpath/weblogic_patch.jar:/usr/java/jdk1.6.0_18/lib/tools.jar:/oracle/middleware/wlserver_10.3/server/lib/weblogic_sp.jar:/oracle/middleware/wlserver_10.3/server/lib/weblogic.jar:/oracle/middleware/modules/features/weblogic.server.modules_10.3.3.0.jar:/oracle/middleware/wlserver_10.3/server/lib/webservices.jar:/oracle/middleware/modules/org.apache.ant_1.7.1/lib/ant-all.jar:/oracle/middleware/modules/net.sf.antcontrib_1.1.0.0_1-0b2/lib/ant-contrib.jar: PATH=/oracle/middleware/wlserver_10.3/server/bin:/oracle/middleware/modules/org.apache.ant_1.7.1/bin:/usr/java/jdk1.6.0_18/jre/bin:/usr/java/jdk1.6.0_18/bin:/usr/kerberos/bin:/usr/local/bin:/bin:/usr/bin:/home/oracle/bin Your environment has been set. [oracle@irmsrv bin]$ cd /oracle/middleware/user_projects/domains/irm_domain/config/fmwconfig/ [oracle@irmsrv fmwconfig]$ keytool -genkeypair -alias oracle.irm.wrap -keyalg RSA -keysize 2048 -keystore irm.jks Enter keystore password: Re-enter new password: What is your first and last name? [Unknown]: Simon Thorpe What is the name of your organizational unit? [Unknown]: Oracle What is the name of your organization? [Unknown]: Oracle What is the name of your City or Locality? [Unknown]: San Francisco What is the name of your State or Province? [Unknown]: CA What is the two-letter country code for this unit? [Unknown]: US Is CN=Simon Thorpe, OU=Oracle, O=Oracle, L=San Francisco, ST=CA, C=US correct? [no]: yes Enter key password for (RETURN if same as keystore password): At this point we now have an irm.jks in the directory /oracle/middleware/user_projects/domains/irm_domain/config/fmwconfig. The reason we store it here is this folder would be backed up as part of a domain backup. As with any cryptographic technology, DO NOT LOSE THESE KEYS OR THIS KEY STORE. Once you've sealed content against a context, the keys will be wrapped with these keys, lose these keys, and you can't get access to any secured content, pretty important. Now we've got the keys created, we need to go back to the IRM Enterprise Manager and set the location of the key store. Going back to the General Settings page in Enterprise Manager scroll down to Keystore Settings. Leave the type as JKS but change the location to; /oracle/Middleware/user_projects/domains/irm_domain/config/fmwconfig/irm.jks and hit Apply. The final step with regards to the key store is we need to tell the server what the password is for the Java Key Store so that it can be opened and the keys accessed. Once more fire up a console window and run these commands (again i've greyed out the clutter to see the commands easier). You will see dummy passed into the commands, this is because the command asks for a username, but in this instance we don't use one, hence the value dummy is passed and it isn't used. [oracle@irmsrv fmwconfig]$ cd /oracle/middleware/Oracle_IRM/common/bin/ [oracle@irmsrv bin]$ ./wlst.sh ... lots of settings fly by... Welcome to WebLogic Server Administration Scripting Shell Type help() for help on available commands wls:/offline>connect('weblogic','password','t3://irmsrv.us.oracle.com:7001') Connecting to t3://irmsrv.us.oracle.com:7001 with userid weblogic ... Successfully connected to Admin Server 'AdminServer' that belongs to domain 'irm_domain'. Warning: An insecure protocol was used to connect to the server. To ensure on-the-wire security, the SSL port or Admin port should be used instead. wls:/irm_domain/serverConfig>createCred("IRM","keystore:irm.jks","dummy","password") Location changed to domainRuntime tree. This is a read-only tree with DomainMBean as the root. For more help, use help(domainRuntime)wls:/irm_domain/serverConfig>createCred("IRM","key:irm.jks:oracle.irm.wrap","dummy","password") Already in Domain Runtime Tree wls:/irm_domain/serverConfig> At last we are now ready to fire up the IRM server itself. The domain creation created a managed server called IRM_server1 and we need to start this, use the following commands in a new console window. cd /oracle/middleware/user_projects/domains/irm_domain/bin/ ./startManagedWebLogic.sh IRM_server1 This will start up the server in the console, unlike the Admin server, you need to provide the username and password for the service to start. Enter in your weblogic username and password when prompted. You can change this behavior by putting the password into a boot.properties file, read more about this in the WebLogic Server documentation. Once running, wait until you see the line; <Notice><WebLogicServer><BEA-000360><Server started in RUNNING mode> At this point we can now login to the Oracle IRM Management Website at the URL. http://irm.company.internal:1600/irm_rights/ The server is just configured for HTTP at the moment, no SSL involved. Just want to ensure we can get a working system up and running. You should now see a login like the image on the right and you can now login using your weblogic username and password. The next article in this guide goes over adding SSL and now testing your server by actually adding a few users, sealing some content and opening this content as a user.

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  • Continuous Integration for SQL Server Part II – Integration Testing

    - by Ben Rees
    My previous post, on setting up Continuous Integration for SQL Server databases using GitHub, Bamboo and Red Gate’s tools, covered the first two parts of a simple Database Continuous Delivery process: Putting your database in to a source control system, and, Running a continuous integration process, each time changes are checked in. However there is, of course, a lot more to to Continuous Delivery than that. Specifically, in addition to the above: Putting some actual integration tests in to the CI process (otherwise, they don’t really do much, do they!?), Deploying the database changes with a managed, automated approach, Monitoring what you’ve just put live, to make sure you haven’t broken anything. This post will detail how to set up a very simple pipeline for implementing the first of these (continuous integration testing). NB: A lot of the setup in this post is built on top of the configuration from before, so it might be difficult to implement this post without running through part I first. There’ll then be a third post on automated database deployment followed by a final post dealing with the last item – monitoring changes on the live system. In the previous post, I used a mixture of Red Gate products and other 3rd party software – GitHub and Atlassian Bamboo specifically. This was partly because I believe most people work in an heterogeneous environment, using software from different vendors to suit their purposes and I wanted to show how this could work for this process. For example, you could easily substitute Atlassian’s BitBucket or Stash for GitHub, depending on your needs, or use an alternative CI server such as TeamCity, TFS or Jenkins. However, in this, post, I’ll be mostly using Red Gate products only (other than tSQLt). I would do this, firstly because I work for Red Gate. However, I also think that in the area of Database Delivery processes, nobody else has the offerings to implement this process fully – so I didn’t have any choice!   Background on Continuous Delivery For me, a great source of information on what makes a proper Continuous Delivery process is the Jez Humble and David Farley classic: Continuous Delivery – Reliable Software Releases through Build, Test, and Deployment Automation This book is not of course, primarily about databases, and the process I outline here and in the previous article is a gross simplification of what Jez and David describe (not least because it’s that much harder for databases!). However, a lot of the principles that they describe can be equally applied to database development and, I would argue, should be. As I say however, what I describe here is a very simple version of what would be required for a full production process. A couple of useful resources on handling some of these complexities can be found in the following two references: Refactoring Databases – Evolutionary Database Design, by Scott J Ambler and Pramod J. Sadalage Versioning Databases – Branching and Merging, by Scott Allen In particular, I don’t deal at all with the issues of multiple branches and merging of those branches, an issue made particularly acute by the use of GitHub. The other point worth making is that, in the words of Martin Fowler: Continuous Delivery is about keeping your application in a state where it is always able to deploy into production.   I.e. we are not talking about continuously delivery updates to the production database every time someone checks in an amendment to a stored procedure. That is possible (and what Martin calls Continuous Deployment). However, again, that’s more than I describe in this article. And I doubt I need to remind DBAs or Developers to Proceed with Caution!   Integration Testing Back to something practical. The next stage, building on our set up from the previous article, is to add in some integration tests to the process. As I say, the CI process, though interesting, isn’t enormously useful without some sort of test process running. For this we’ll use the tSQLt framework, an open source framework designed specifically for running SQL Server tests. tSQLt is part of Red Gate’s SQL Test found on http://www.red-gate.com/products/sql-development/sql-test/ or can be downloaded separately from www.tsqlt.org - though I’ll provide a step-by-step guide below for setting this up. Getting tSQLt set up via SQL Test Click on the link http://www.red-gate.com/products/sql-development/sql-test/ and click on the blue Download button to download the Red Gate SQL Test product, if not already installed. Follow the install process for SQL Test to install the SQL Server Management Studio (SSMS) plugin on to your machine, if not already installed. Open SSMS. You should now see SQL Test under the Tools menu:   Clicking this link will give you the basic SQL Test dialogue: As yet, though we’ve installed the SQL Test product we haven’t yet installed the tSQLt test framework on to any particular database. To do this, we need to add our RedGateApp database using this dialogue, by clicking on the + Add Database to SQL Test… link, selecting the RedGateApp database and clicking the Add Database link:   In the next screen, SQL Test describes what will be installed on the database for the tSQLt framework. Also in this dialogue, uncheck the “Add SQL Cop tests” option (shown below). SQL Cop is a great set of pre-defined tests that work within the tSQLt framework to check the general health of your SQL Server database. However, we won’t be using them in this particular simple example: Once you’ve clicked on the OK button, the changes described in the dialogue will be made to your database. Some of these are shown in the left-hand-side below: We’ve now installed the framework. However, we haven’t actually created any tests, so this will be the next step. But, before we proceed, we’ve made an update to our database so should, again check this in to source control, adding comments as required:   Also worth a quick check that your build still runs with the new additions!: (And a quick check of the RedGateAppCI database shows that the changes have been made).   Creating and Testing a Unit Test There are, of course, a lot of very interesting unit tests that you could and should set up for a database. The great thing about the tSQLt framework is that you can write these in SQL. The example I’m going to use here is pretty Mickey Mouse – our database table is going to include some email addresses as reference data and I want to check whether these are all in a correct email format. Nothing clever but it illustrates the process and hopefully shows the method by which more interesting tests could be set up. Adding Reference Data to our Database To start, I want to add some reference data to my database, and have this source controlled (as well as the schema). First of all I need to add some data in to my solitary table – this can be done a number of ways, but I’ll do this in SSMS for simplicity: I then add some reference data to my table: Currently this reference data just exists in the database. For proper integration testing, this needs to form part of the source-controlled version of the database – and so needs to be added to the Git repository. This can be done via SQL Source Control, though first a Primary Key needs to be added to the table. Right click the table, select Design, then right-click on the first “id” row. Then click on “Set Primary Key”: NB: once this change is made, click Save to save the change to the table. Then, to source control this reference data, right click on the table (dbo.Email) and selecting the following option:   In the next screen, link the data in the Email table, by selecting it from the list and clicking “save and close”: We should at this point re-commit the changes (both the addition of the Primary Key, and the data) to the Git repo. NB: From here on, I won’t show screenshots for the GitHub side of things – it’s the same each time: whenever a change is made in SQL Source Control and committed to your local folder, you then need to sync this in the GitHub Windows client (as this is where the build server, Bamboo is taking it from). An interesting point to note here, when these changes are committed in SQL Source Control (right-click database and select “Commit Changes to Source Control..”): The display gives a warning about possibly needing a migration script for the “Add Primary Key” step of the changes. This isn’t actually necessary in this case, but this mechanism would allow you to create override scripts to replace the default change scripts created by the SQL Compare engine (which runs underneath SQL Source Control). Ignoring this message (!), we add a comment and commit the changes to Git. I then sync these, run a build (or the build gets run automatically), and check that the data is being deployed over to the target RedGateAppCI database:   Creating and Running the Test As I mention, the test I’m going to use here is a very simple one - are the email addresses in my reference table valid? This isn’t of course, a full test of email validation (I expect the email addresses I’ve chosen here aren’t really the those of the Fab Four) – but just a very basic check of format used. I’ve taken the relevant SQL from this Stack Overflow article. In SSMS select “SQL Test” from the Tools menu, then click on + New Test: In the next screen, give your new test a name, and also enter a name in the Test Class box (test classes are schemas that help you keep things organised). Also check that the database in which the test is going to be created is correct – RedGateApp in this example: Click “Create Test”. After closing a couple of subsequent dialogues, you’ll see a dummy script for the test, that needs filling in:   We now need to define the SQL for our test. As mentioned before, tSQLt allows you to write your unit tests in T-SQL, and the code I’m going to use here is as below. This needs to be copied and pasted in to the query window, to replace the default given by tSQLt: –  Basic email check test ALTER PROCEDURE [MyChecks].[test Check Email Addresses] AS BEGIN SET NOCOUNT ON         Declare @Output VarChar(max)     Set @Output = ”       SELECT  @Output = @Output + Email +Char(13) + Char(10) FROM dbo.Email WHERE email NOT LIKE ‘%_@__%.__%’       If @Output > ”         Begin             Set @Output = Char(13) + Char(10)                           + @Output             EXEC tSQLt.Fail@Output         End   END;   Once this script is entered, hit execute to add the Stored Procedure to the database. Before committing the test to source control,  it’s worth just checking that it works! For a positive test, click on “SQL Test” from the Tools menu, then click Run Tests. You should see output like the following: - a green tick to indicate success! But of course, what we also need to do is test that this is actually doing something by showing a failed test. Edit one of the email addresses in your table to an incorrect format: Now, re-run the same SQL Test as before and you’ll see the following: Great – we now know that our test is really doing something! You’ll also see a useful error message at the bottom of SSMS: (leave the email address as invalid for now, for the next steps). The next stage is to check this new test in to source control again, by right-clicking on the database and checking in the changes with a commit message (and not forgetting to sync in the GitHub client):   Checking that the Tests are Running as Integration Tests After the changes above are made, and after a build has run on Bamboo (manual or automatic), looking at the Stored Procedures for the RedGateAppCI, the SPROC for the new test has been moved over to the database. However this is not exactly what we were after. We didn’t want to just copy objects from one database to another, but actually run the tests as part of the build/integration test process. I.e. we’re continuously checking any changes we make (in this case, to the reference data emails), to ensure we’re not breaking a test that we’ve set up. The behaviour we want to see is that, if we check in static data that is incorrect (as we did in step 9 above) and we have the tSQLt test set up, then our build in Bamboo should fail. However, re-running the build shows the following: - sadly, a successful build! To make sure the tSQLt tests are run as part of the integration test, we need to amend a switch in the Red Gate CI config file. First, navigate to file sqlCI.targets in your working folder: Edit this document, make the following change, save the document, then commit and sync this change in the GitHub client: <!-- tSQLt tests --> <!-- Optional --> <!-- To run tSQLt tests in source control for the database, enter true. --> <enableTsqlt>true</enableTsqlt> Now, if we re-run the build in Bamboo (NB: I’ve moved to a new server here, hence different address and build number): - superb, a broken build!! The error message isn’t great here, so to get more detailed info, click on the full build log link on this page (below the fold). The interesting part of the log shown is towards the bottom. Pulling out this part:   21-Jun-2013 11:35:19 Build FAILED. 21-Jun-2013 11:35:19 21-Jun-2013 11:35:19 "C:\Users\Administrator\bamboo-home\xml-data\build-dir\RGA-RGP-JOB1\sqlCI.proj" (default target) (1) -> 21-Jun-2013 11:35:19 (sqlCI target) -> 21-Jun-2013 11:35:19 EXEC : sqlCI error occurred: RedGate.Deploy.SqlServerDbPackage.Shared.Exceptions.InvalidSqlException: Test Case Summary: 1 test case(s) executed, 0 succeeded, 1 failed, 0 errored. [C:\Users\Administrator\bamboo-home\xml-data\build-dir\RGA-RGP-JOB1\sqlCI.proj] 21-Jun-2013 11:35:19 EXEC : sqlCI error occurred: [MyChecks].[test Check Email Addresses] failed: [C:\Users\Administrator\bamboo-home\xml-data\build-dir\RGA-RGP-JOB1\sqlCI.proj] 21-Jun-2013 11:35:19 EXEC : sqlCI error occurred: ringo.starr@beatles [C:\Users\Administrator\bamboo-home\xml-data\build-dir\RGA-RGP-JOB1\sqlCI.proj] 21-Jun-2013 11:35:19 EXEC : sqlCI error occurred: [C:\Users\Administrator\bamboo-home\xml-data\build-dir\RGA-RGP-JOB1\sqlCI.proj] 21-Jun-2013 11:35:19 EXEC : sqlCI error occurred: +----------------------+ [C:\Users\Administrator\bamboo-home\xml-data\build-dir\RGA-RGP-JOB1\sqlCI.proj] 21-Jun-2013 11:35:19 EXEC : sqlCI error occurred: |Test Execution Summary| [C:\Users\Administrator\bamboo-home\xml-data\build-dir\RGA-RGP-JOB1\sqlCI.proj]   As a final check, we should make sure that, if we now fix this error, the build succeeds. So in SSMS, I’m going to correct the invalid email address, then check this change in to SQL Source Control (with a comment), commit to GitHub, and re-run the build:   This should have fixed the build: It worked! Summary This has been a very quick run through the implementation of CI for databases, including tSQLt tests to test whether your database updates are working. The next post in this series will focus on automated deployment – we’ve tested our database changes, how can we now deploy these to target sites?  

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  • Another Marketing Conference, part one – the best morning sessions.

    - by Roger Hart
    Yesterday I went to Another Marketing Conference. I honestly can’t tell if the title is just tipping over into smug, but in the balance of things that doesn’t matter, because it was a good conference. There was an enjoyable blend of theoretical and practical, and enough inter-disciplinary spread to keep my inner dilettante grinning from ear to ear. Sure, there was a bumpy bit in the middle, with two back-to-back sales pitches and a rather thin overview of the state of the web. But the signal:noise ratio at AMC2012 was impressively high. Here’s the first part of my write-up of the sessions. It’s a bit of a mammoth. It’s also a bit of a mash-up of what was said and what I thought about it. I’ll add links to the videos and slides from the sessions as they become available. Although it was in the morning session, I’ve not included Vanessa Northam’s session on the power of internal comms to build brand ambassadors. It’ll be in the next roundup, as this is already pushing 2.5k words. First, the important stuff. I was keeping a tally, and nobody said “synergy” or “leverage”. I did, however, hear the term “marketeers” six times. Shame on you – you know who you are. 1 – Branding in a post-digital world, Graham Hales This initially looked like being a sales presentation for Interbrand, but Graham pulled it out of the bag a few minutes in. He introduced a model for brand management that was essentially Plan >> Do >> Check >> Act, with Do and Check rolled up together, and went on to stress that this looks like on overall business management model for a reason. Brand has to be part of your overall business strategy and metrics if you’re going to care about it at all. This was the first iteration of what proved to be one of the event’s emergent themes: do it throughout the stack or don’t bother. Graham went on to remind us that brands, in so far as they are owned at all, are owned by and co-created with our customers. Advertising can offer a message to customers, but they provide the expression of a brand. This was a preface to talking about an increasingly chaotic marketplace, with increasingly hard-to-manage purchase processes. Services like Amazon reviews and TripAdvisor (four presenters would make this point) saturate customers with information, and give them a kind of vigilante power to comment on and define brands. Consequentially, they experience a number of “moments of deflection” in our sales funnels. Our control is lessened, and failure to engage can negatively-impact buying decisions increasingly poorly. The clearest example given was the failure of NatWest’s “caring bank” campaign, where staff in branches, customer support, and online presences didn’t align. A discontinuity of experience basically made the campaign worthless, and disgruntled customers talked about it loudly on social media. This in turn presented an opportunity to engage and show caring, but that wasn’t taken. What I took away was that brand (co)creation is ongoing and needs monitoring and metrics. But reciprocally, given you get what you measure, strategy and metrics must include brand if any kind of branding is to work at all. Campaigns and messages must permeate product and service design. What that doesn’t mean (and Graham didn’t say it did) is putting Marketing at the top of the pyramid, and having them bawl demands at Product Management, Support, and Development like an entitled toddler. It’s going to have to be collaborative, and session 6 on internal comms handled this really well. The main thing missing here was substantiating data, and the main question I found myself chewing on was: if we’re building brands collaboratively and in the open, what about the cultural politics of trolling? 2 – Challenging our core beliefs about human behaviour, Mark Earls This was definitely the best show of the day. It was also some of the best content. Mark talked us through nudging, behavioural economics, and some key misconceptions around decision making. Basically, people aren’t rational, they’re petty, reactive, emotional sacks of meat, and they’ll go where they’re led. Comforting stuff. Examples given were the spread of the London Riots and the “discovery” of the mountains of Kong, and the popularity of Susan Boyle, which, in turn made me think about Per Mollerup’s concept of “social wayshowing”. Mark boiled his thoughts down into four key points which I completely failed to write down word for word: People do, then think – Changing minds to change behaviour doesn’t work. Post-rationalization rules the day. See also: mere exposure effects. Spock < Kirk - Emotional/intuitive comes first, then we rationalize impulses. The non-thinking, emotive, reactive processes run much faster than the deliberative ones. People are not really rational decision makers, so  intervening with information may not be appropriate. Maximisers or satisficers? – Related to the last point. People do not consistently, rationally, maximise. When faced with an abundance of choice, they prefer to satisfice than evaluate, and will often follow social leads rather than think. Things tend to converge – Behaviour trends to a consensus normal. When faced with choices people overwhelmingly just do what they see others doing. Humans are extraordinarily good at mirroring behaviours and receiving influence. People “outsource the cognitive load” of choices to the crowd. Mark’s headline quote was probably “the real influence happens at the table next to you”. Reference examples, word of mouth, and social influence are tremendously important, and so talking about product experiences may be more important than talking about products. This reminded me of Kathy Sierra’s “creating bad-ass users” concept of designing to make people more awesome rather than products they like. If we can expose user-awesome, and make sharing easy, we can normalise the behaviours we want. If we normalize the behaviours we want, people should make and post-rationalize the buying decisions we want.  Where we need to be: “A bigger boy made me do it” Where we are: “a wizard did it and ran away” However, it’s worth bearing in mind that some purchasing decisions are personal and informed rather than social and reactive. There’s a quadrant diagram, in fact. What was really interesting, though, towards the end of the talk, was some advice for working out how social your products might be. The standard technology adoption lifecycle graph is essentially about social product diffusion. So this idea isn’t really new. Geoffrey Moore’s “chasm” idea may not strictly apply. However, his concepts of beachheads and reference segments are exactly what is required to normalize and thus enable purchase decisions (behaviour change). The final thing is that in only very few categories does a better product actually affect purchase decision. Where the choice is personal and informed, this is true. But where it’s personal and impulsive, or in any way social, “better” is trumped by popularity, endorsement, or “point of sale salience”. UX, UCD, and e-commerce know this to be true. A better (and easier) experience will always beat “more features”. Easy to use, and easy to observe being used will beat “what the user says they want”. This made me think about the astounding stickiness of rational fallacies, “common sense” and the pathological willful simplifications of the media. Rational fallacies seem like they’re basically the heuristics we use for post-rationalization. If I were profoundly grimy and cynical, I’d suggest deploying a boat-load in our messaging, to see if they’re really as sticky and appealing as they look. 4 – Changing behaviour through communication, Stephen Donajgrodzki This was a fantastic follow up to Mark’s session. Stephen basically talked us through some tactics used in public information/health comms that implement the kind of behavioural theory Mark introduced. The session was largely about how to get people to do (good) things they’re predisposed not to do, and how communication can (and can’t) make positive interventions. A couple of things stood out, in particular “implementation intentions” and how they can be linked to goals. For example, in order to get people to check and test their smoke alarms (a goal intention, rarely actualized  an information campaign will attempt to link this activity to the clocks going back or forward (a strong implementation intention, well-actualized). The talk reinforced the idea that making behaviour changes easy and visible normalizes them and makes them more likely to succeed. To do this, they have to be embodied throughout a product and service cycle. Experiential disconnects undermine the normalization. So campaigns, products, and customer interactions must be aligned. This is underscored by the second section of the presentation, which talked about interventions and pre-conditions for change. Taking the examples of drug addiction and stopping smoking, Stephen showed us a framework for attempting (and succeeding or failing in) behaviour change. He noted that when the change is something people fundamentally want to do, and that is easy, this gets a to simpler. Coordinated, easily-observed environmental pressures create preconditions for change and build motivation. (price, pub smoking ban, ad campaigns, friend quitting, declining social acceptability) A triggering even leads to a change attempt. (getting a cold and panicking about how bad the cough is) Interventions can be made to enable an attempt (NHS services, public information, nicotine patches) If it succeeds – yay. If it fails, there’s strong negative enforcement. Triggering events seem largely personal, but messaging can intervene in the creation of preconditions and in supporting decisions. Stephen talked more about systems of thinking and “bounded rationality”. The idea being that to enable change you need to break through “automatic” thinking into “reflective” thinking. Disruption and emotion are great tools for this, but that is only the start of the process. It occurs to me that a great deal of market research is focused on determining triggers rather than analysing necessary preconditions. Although they are presumably related. The final section talked about setting goals. Marketing goals are often seen as deriving directly from business goals. However, marketing may be unable to deliver on these directly where decision and behaviour-change processes are involved. In those cases, marketing and communication goals should be to create preconditions. They should also consider priming and norms. Content marketing and brand awareness are good first steps here, as brands can be heuristics in decision making for choice-saturated consumers, or those seeking education. 5 – The power of engaged communities and how to build them, Harriet Minter (the Guardian) The meat of this was that you need to let communities define and establish themselves, and be quick to react to their needs. Harriet had been in charge of building the Guardian’s community sites, and learned a lot about how they come together, stabilize  grow, and react. Crucially, they can’t be about sales or push messaging. A community is not just an audience. It’s essential to start with what this particular segment or tribe are interested in, then what they want to hear. Eventually you can consider – in light of this – what they might want to buy, but you can’t start with the product. A community won’t cohere around one you’re pushing. Her tips for community building were (again, sorry, not verbatim): Set goals Have some targets. Community building sounds vague and fluffy, but you can have (and adjust) concrete goals. Think like a start-up This is the “lean” stuff. Try things, fail quickly, respond. Don’t restrict platforms Let the audience choose them, and be aware of their differences. For example, LinkedIn is very different to Twitter. Track your stats Related to the first point. Keeping an eye on the numbers lets you respond. They should be qualified, however. If you want a community of enterprise decision makers, headcount alone may be a bad metric – have you got CIOs, or just people who want to get jobs by mingling with CIOs? Build brand advocates Do things to involve people and make them awesome, and they’ll cheer-lead for you. The last part really got my attention. Little bits of drive-by kindness go a long way. But more than that, genuinely helping people turns them into powerful advocates. Harriet gave an example of the Guardian engaging with an aspiring journalist on its Q&A forums. Through a series of serendipitous encounters he became a BBC producer, and now enthusiastically speaks up for the Guardian community sites. Cultivating many small, authentic, influential voices may have a better pay-off than schmoozing the big guys. This could be particularly important in the context of Mark and Stephen’s models of social, endorsement-led, and example-led decision making. There’s a lot here I haven’t covered, and it may be worth some follow-up on community building. Thoughts I was quite sceptical of nudge theory and behavioural economics. First off it sounds too good to be true, and second it sounds too sinister to permit. But I haven’t done the background reading. So I’m going to, and if it seems to hold real water, and if it’s possible to do it ethically (Stephen’s presentations suggests it may be) then it’s probably worth exploring. The message seemed to be: change what people do, and they’ll work out why afterwards. Moreover, the people around them will do it too. Make the things you want them to do extraordinarily easy and very, very visible. Normalize and support the decisions you want them to make, and they’ll make them. In practice this means not talking about the thing, but showing the user-awesome. Glib? Perhaps. But it feels worth considering. Also, if I ever run a marketing conference, I’m going to ban speakers from using examples from Apple. Quite apart from not being consistently generalizable, it’s becoming an irritating cliché.

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  • The Incremental Architect&acute;s Napkin - #2 - Balancing the forces

    - by Ralf Westphal
    Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/06/02/the-incremental-architectacutes-napkin---2---balancing-the-forces.aspxCategorizing requirements is the prerequisite for ecconomic architectural decisions. Not all requirements are created equal. However, to truely understand and describe the requirement forces pulling on software development, I think further examination of the requirements aspects is varranted. Aspects of Functionality There are two sides to Functionality requirements. It´s about what a software should do. I call that the Operations it implements. Operations are defined by expressions and control structures or calls to frameworks of some sort, i.e. (business) logic statements. Operations calculate, transform, aggregate, validate, send, receive, load, store etc. Operations are about behavior; they take input and produce output by considering state. I´m not using the term “function” here, because functions - or methods or sub-programs - are not necessary to implement Operations. Functions belong to a different sub-aspect of requirements (see below). Operations alone are not enough, though, to make a customer happy with regard to his/her Functionality requirements. Only correctly implemented Operations provide full value. This should make clear, why testing is so important. And not just manual tests during development of some operational feature, but automated tests. Because only automated tests scale when over time the number of operations increases. Without automated tests there is no guarantee formerly correct operations are still correct after more got added. To retest all previous operations manually is infeasible. So whoever relies just on manual tests is not really balancing the two forces Operations and Correctness. With manual tests more weight is put on the side of the scale of Operations. That might be ok for a short period of time - but in the long run it will bite you. You need to plan for Correctness in the long run from the first day of your project on. Aspects of Quality As important as Functionality is, it´s not the driver for software development. No software has ever been written to just implement some operation in code. We don´t need computers just to do something. All computers can do with software we can do without them. Well, at least given enough time and resources. We could calculate the most complex formulas without computers. We could do auctions with millions of people without computers. The only reason we want computers to help us with this and a million other Operations is… We don´t want to wait for the results very long. Or we want less errors. Or we want easier accessability to complicated solutions. So the main reason for customers to buy/order software is some Quality. They want some Functionality with a higher Quality (e.g. performance, scalability, usability, security…) than without the software. But Qualities come in at least two flavors: Most important are Primary Qualities. That´s the Qualities software truely is written for. Take an online auction website for example. Its Primary Qualities are performance, scalability, and usability, I´d say. Auctions should come within reach of millions of people; setting up an auction should be very easy; finding a suitable auction and bidding on it should be as fast as possible. Only if those Qualities have been implemented does security become relevant. A secure auction website is important - but not as important as a fast auction website. Nobody would want to use the most secure auction website if it was unbearably slow. But there would be people willing to use the fastest auction website even it was lacking security. That´s why security - with regard to online auction software - is not a Primary Quality, but just a Secondary Quality. It´s a supporting quality, so to speak. It does not deliver value by itself. With a password manager software this might be different. There security might be a Primary Quality. Please get me right: I don´t want to denigrate any Quality. There´s a long list of non-functional requirements at Wikipedia. They are all created equal - but that does not mean they are equally important for all software projects. When confronted with Quality requirements check with the customer which are primary and which are secondary. That will help to make good economical decisions when in a crunch. Resources are always limited - but requirements are a bottomless ocean. Aspects of Security of Investment Functionality and Quality are traditionally the requirement aspects cared for most - by customers and developers alike. Even today, when pressure rises in a project, tunnel vision will focus on them. Any measures to create and hold up Security of Investment (SoI) will be out of the window pretty quickly. Resistance to customers and/or management is futile. As long as SoI is not placed on equal footing with Functionality and Quality it´s bound to suffer under pressure. To look closer at what SoI means will help to become more conscious about it and make customers and management aware of the risks of neglecting it. SoI to me has two facets: Production Efficiency (PE) is about speed of delivering value. Customers like short response times. Short response times mean less money spent. So whatever makes software development faster supports this requirement. This must not lead to duct tape programming and banging out features by the dozen, though. Because customers don´t just want Operations and Quality, but also Correctness. So if Correctness gets compromised by focussing too much on Production Efficiency it will fire back. Customers want PE not just today, but over the whole course of a software´s lifecycle. That means, it´s not just about coding speed, but equally about code quality. If code quality leads to rework the PE is on an unsatisfactory level. Also if code production leads to waste it´s unsatisfactory. Because the effort which went into waste could have been used to produce value. Rework and waste cost money. Rework and waste abound, however, as long as PE is not addressed explicitly with management and customers. Thanks to the Agile and Lean movements that´s increasingly the case. Nevertheless more could and should be done in many teams. Each and every developer should keep in mind that Production Efficiency is as important to the customer as Functionality and Quality - whether he/she states it or not. Making software development more efficient is important - but still sooner or later even agile projects are going to hit a glas ceiling. At least as long as they neglect the second SoI facet: Evolvability. Delivering correct high quality functionality in short cycles today is good. But not just any software structure will allow this to happen for an indefinite amount of time.[1] The less explicitly software was designed the sooner it´s going to get stuck. Big ball of mud, monolith, brownfield, legacy code, technical debt… there are many names for software structures that have lost the ability to evolve, to be easily changed to accomodate new requirements. An evolvable code base is the opposite of a brownfield. It´s code which can be easily understood (by developers with sufficient domain expertise) and then easily changed to accomodate new requirements. Ideally the costs of adding feature X to an evolvable code base is independent of when it is requested - or at least the costs should only increase linearly, not exponentially.[2] Clean Code, Agile Architecture, and even traditional Software Engineering are concerned with Evolvability. However, it seems no systematic way of achieving it has been layed out yet. TDD + SOLID help - but still… When I look at the design ability reality in teams I see much room for improvement. As stated previously, SoI - or to be more precise: Evolvability - can hardly be measured. Plus the customer rarely states an explicit expectation with regard to it. That´s why I think, special care must be taken to not neglect it. Postponing it to some large refactorings should not be an option. Rather Evolvability needs to be a core concern for every single developer day. This should not mean Evolvability is more important than any of the other requirement aspects. But neither is it less important. That´s why more effort needs to be invested into it, to bring it on par with the other aspects, which usually are much more in focus. In closing As you see, requirements are of quite different kinds. To not take that into account will make it harder to understand the customer, and to make economic decisions. Those sub-aspects of requirements are forces pulling in different directions. To improve performance might have an impact on Evolvability. To increase Production Efficiency might have an impact on security etc. No requirement aspect should go unchecked when deciding how to allocate resources. Balancing should be explicit. And it should be possible to trace back each decision to a requirement. Why is there a null-check on parameters at the start of the method? Why are there 5000 LOC in this method? Why are there interfaces on those classes? Why is this functionality running on the threadpool? Why is this function defined on that class? Why is this class depending on three other classes? These and a thousand more questions are not to mean anything should be different in a code base. But it´s important to know the reason behind all of these decisions. Because not knowing the reason possibly means waste and having decided suboptimally. And how do we ensure to balance all requirement aspects? That needs practices and transparency. Practices means doing things a certain way and not another, even though that might be possible. We´re dealing with dangerous tools here. Like a knife is a dangerous tool. Harm can be done if we use our tools in just any way at the whim of the moment. Over the centuries rules and practices have been established how to use knifes. You don´t put them in peoples´ legs just because you´re feeling like it. You hand over a knife with the handle towards the receiver. You might not even be allowed to cut round food like potatos or eggs with it. The same should be the case for dangerous tools like object-orientation, remote communication, threads etc. We need practices to use them in a way so requirements are balanced almost automatically. In addition, to be able to work on software as a team we need transparency. We need means to share our thoughts, to work jointly on mental models. So far our tools are focused on working with code. Testing frameworks, build servers, DI containers, intellisense, refactoring support… That´s all nice and well. I don´t want to miss any of that. But I think it´s not enough. We´re missing mental tools, tools for making thinking and talking about software (independently of code) easier. You might think, enough of such tools already exist like all those UML diagram types or Flow Charts. But then, isn´t it strange, hardly any team is using them to design software? Or is that just due to a lack of education? I don´t think so. It´s a matter value/weight ratio: the current mental tools are too heavy weight compared to the value they deliver. So my conclusion is, we need lightweight tools to really be able to balance requirements. Software development is complex. We need guidance not to forget important aspects. That´s like with flying an airplane. Pilots don´t just jump in and take off for their destination. Yes, there are times when they are “flying by the seats of their pants”, when they are just experts doing thing intuitively. But most of the time they are going through honed practices called checklist. See “The Checklist Manifesto” for very enlightening details on this. Maybe then I should say it like this: We need more checklists for the complex businss of software development.[3] But that´s what software development mostly is about: changing software over an unknown period of time. It needs to be corrected in order to finally provide promised operations. It needs to be enhanced to provide ever more operations and qualities. All this without knowing when it´s going to stop. Probably never - until “maintainability” hits a wall when the technical debt is too large, the brownfield too deep. Software development is not a sprint, is not a marathon, not even an ultra marathon. Because to all this there is a foreseeable end. Software development is like continuously and foreever running… ? And sometimes I dare to think that costs could even decrease over time. Think of it: With each feature a software becomes richer in functionality. So with each additional feature the chance of there being already functionality helping its implementation increases. That should lead to less costs of feature X if it´s requested later than sooner. X requested later could stand on the shoulders of previous features. Alas, reality seems to be far from this despite 20+ years of admonishing developers to think in terms of reusability.[1] ? Please don´t get me wrong: I don´t want to bog down the “art” of software development with heavyweight practices and heaps of rules to follow. The framework we need should be lightweight. It should not stand in the way of delivering value to the customer. It´s purpose is even to make that easier by helping us to focus and decreasing waste and rework. ?

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  • .NET Code Evolution

    - by Alois Kraus
    Originally posted on: http://geekswithblogs.net/akraus1/archive/2013/07/24/153504.aspxAt my day job I do look at a lot of code written by other people. Most of the code is quite good and some is even a masterpiece. And there is also code which makes you think WTF… oh it was written by me. Hm not so bad after all. There are many excuses reasons for bad code. Most often it is time pressure followed by not enough ambition (who cares) or insufficient training. Normally I do care about code quality quite a lot which makes me a (perceived) slow worker who does write many tests and refines the code quite a lot because of the design deficiencies. Most of the deficiencies I do find by putting my design under stress while checking for invariants. It does also help a lot to step into the code with a debugger (sometimes also Windbg). I do this much more often when my tests are red. That way I do get a much better understanding what my code really does and not what I think it should be doing. This time I do want to show you how code can evolve over the years with different .NET Framework versions. Once there was  time where .NET 1.1 was new and many C++ programmers did switch over to get rid of not initialized pointers and memory leaks. There were also nice new data structures available such as the Hashtable which is fast lookup table with O(1) time complexity. All was good and much code was written since then. At 2005 a new version of the .NET Framework did arrive which did bring many new things like generics and new data structures. The “old” fashioned way of Hashtable were coming to an end and everyone used the new Dictionary<xx,xx> type instead which was type safe and faster because the object to type conversion (aka boxing) was no longer necessary. I think 95% of all Hashtables and dictionaries use string as key. Often it is convenient to ignore casing to make it easy to look up values which the user did enter. An often followed route is to convert the string to upper case before putting it into the Hashtable. Hashtable Table = new Hashtable(); void Add(string key, string value) { Table.Add(key.ToUpper(), value); } This is valid and working code but it has problems. First we can pass to the Hashtable a custom IEqualityComparer to do the string matching case insensitive. Second we can switch over to the now also old Dictionary type to become a little faster and we can keep the the original keys (not upper cased) in the dictionary. Dictionary<string, string> DictTable = new Dictionary<string, string>(StringComparer.OrdinalIgnoreCase); void AddDict(string key, string value) { DictTable.Add(key, value); } Many people do not user the other ctors of Dictionary because they do shy away from the overhead of writing their own comparer. They do not know that .NET has for strings already predefined comparers at hand which you can directly use. Today in the many core area we do use threads all over the place. Sometimes things break in subtle ways but most of the time it is sufficient to place a lock around the offender. Threading has become so mainstream that it may sound weird that in the year 2000 some guy got a huge incentive for the idea to reduce the time to process calibration data from 12 hours to 6 hours by using two threads on a dual core machine. Threading does make it easy to become faster at the expense of correctness. Correct and scalable multithreading can be arbitrarily hard to achieve depending on the problem you are trying to solve. Lets suppose we want to process millions of items with two threads and count the processed items processed by all threads. A typical beginners code might look like this: int Counter; void IJustLearnedToUseThreads() { var t1 = new Thread(ThreadWorkMethod); t1.Start(); var t2 = new Thread(ThreadWorkMethod); t2.Start(); t1.Join(); t2.Join(); if (Counter != 2 * Increments) throw new Exception("Hmm " + Counter + " != " + 2 * Increments); } const int Increments = 10 * 1000 * 1000; void ThreadWorkMethod() { for (int i = 0; i < Increments; i++) { Counter++; } } It does throw an exception with the message e.g. “Hmm 10.222.287 != 20.000.000” and does never finish. The code does fail because the assumption that Counter++ is an atomic operation is wrong. The ++ operator is just a shortcut for Counter = Counter + 1 This does involve reading the counter from a memory location into the CPU, incrementing value on the CPU and writing the new value back to the memory location. When we do look at the generated assembly code we will see only inc dword ptr [ecx+10h] which is only one instruction. Yes it is one instruction but it is not atomic. All modern CPUs have several layers of caches (L1,L2,L3) which try to hide the fact how slow actual main memory accesses are. Since cache is just another word for redundant copy it can happen that one CPU does read a value from main memory into the cache, modifies it and write it back to the main memory. The problem is that at least the L1 cache is not shared between CPUs so it can happen that one CPU does make changes to values which did change in meantime in the main memory. From the exception you can see we did increment the value 20 million times but half of the changes were lost because we did overwrite the already changed value from the other thread. This is a very common case and people do learn to protect their  data with proper locking.   void Intermediate() { var time = Stopwatch.StartNew(); Action acc = ThreadWorkMethod_Intermediate; var ar1 = acc.BeginInvoke(null, null); var ar2 = acc.BeginInvoke(null, null); ar1.AsyncWaitHandle.WaitOne(); ar2.AsyncWaitHandle.WaitOne(); if (Counter != 2 * Increments) throw new Exception(String.Format("Hmm {0:N0} != {1:N0}", Counter, 2 * Increments)); Console.WriteLine("Intermediate did take: {0:F1}s", time.Elapsed.TotalSeconds); } void ThreadWorkMethod_Intermediate() { for (int i = 0; i < Increments; i++) { lock (this) { Counter++; } } } This is better and does use the .NET Threadpool to get rid of manual thread management. It does give the expected result but it can result in deadlocks because you do lock on this. This is in general a bad idea since it can lead to deadlocks when other threads use your class instance as lock object. It is therefore recommended to create a private object as lock object to ensure that nobody else can lock your lock object. When you read more about threading you will read about lock free algorithms. They are nice and can improve performance quite a lot but you need to pay close attention to the CLR memory model. It does make quite weak guarantees in general but it can still work because your CPU architecture does give you more invariants than the CLR memory model. For a simple counter there is an easy lock free alternative present with the Interlocked class in .NET. As a general rule you should not try to write lock free algos since most likely you will fail to get it right on all CPU architectures. void Experienced() { var time = Stopwatch.StartNew(); Task t1 = Task.Factory.StartNew(ThreadWorkMethod_Experienced); Task t2 = Task.Factory.StartNew(ThreadWorkMethod_Experienced); t1.Wait(); t2.Wait(); if (Counter != 2 * Increments) throw new Exception(String.Format("Hmm {0:N0} != {1:N0}", Counter, 2 * Increments)); Console.WriteLine("Experienced did take: {0:F1}s", time.Elapsed.TotalSeconds); } void ThreadWorkMethod_Experienced() { for (int i = 0; i < Increments; i++) { Interlocked.Increment(ref Counter); } } Since time does move forward we do not use threads explicitly anymore but the much nicer Task abstraction which was introduced with .NET 4 at 2010. It is educational to look at the generated assembly code. The Interlocked.Increment method must be called which does wondrous things right? Lets see: lock inc dword ptr [eax] The first thing to note that there is no method call at all. Why? Because the JIT compiler does know very well about CPU intrinsic functions. Atomic operations which do lock the memory bus to prevent other processors to read stale values are such things. Second: This is the same increment call prefixed with a lock instruction. The only reason for the existence of the Interlocked class is that the JIT compiler can compile it to the matching CPU intrinsic functions which can not only increment by one but can also do an add, exchange and a combined compare and exchange operation. But be warned that the correct usage of its methods can be tricky. If you try to be clever and look a the generated IL code and try to reason about its efficiency you will fail. Only the generated machine code counts. Is this the best code we can write? Perhaps. It is nice and clean. But can we make it any faster? Lets see how good we are doing currently. Level Time in s IJustLearnedToUseThreads Flawed Code Intermediate 1,5 (lock) Experienced 0,3 (Interlocked.Increment) Master 0,1 (1,0 for int[2]) That lock free thing is really a nice thing. But if you read more about CPU cache, cache coherency, false sharing you can do even better. int[] Counters = new int[12]; // Cache line size is 64 bytes on my machine with an 8 way associative cache try for yourself e.g. 64 on more modern CPUs void Master() { var time = Stopwatch.StartNew(); Task t1 = Task.Factory.StartNew(ThreadWorkMethod_Master, 0); Task t2 = Task.Factory.StartNew(ThreadWorkMethod_Master, Counters.Length - 1); t1.Wait(); t2.Wait(); Counter = Counters[0] + Counters[Counters.Length - 1]; if (Counter != 2 * Increments) throw new Exception(String.Format("Hmm {0:N0} != {1:N0}", Counter, 2 * Increments)); Console.WriteLine("Master did take: {0:F1}s", time.Elapsed.TotalSeconds); } void ThreadWorkMethod_Master(object number) { int index = (int) number; for (int i = 0; i < Increments; i++) { Counters[index]++; } } The key insight here is to use for each core its own value. But if you simply use simply an integer array of two items, one for each core and add the items at the end you will be much slower than the lock free version (factor 3). Each CPU core has its own cache line size which is something in the range of 16-256 bytes. When you do access a value from one location the CPU does not only fetch one value from main memory but a complete cache line (e.g. 16 bytes). This means that you do not pay for the next 15 bytes when you access them. This can lead to dramatic performance improvements and non obvious code which is faster although it does have many more memory reads than another algorithm. So what have we done here? We have started with correct code but it was lacking knowledge how to use the .NET Base Class Libraries optimally. Then we did try to get fancy and used threads for the first time and failed. Our next try was better but it still had non obvious issues (lock object exposed to the outside). Knowledge has increased further and we have found a lock free version of our counter which is a nice and clean way which is a perfectly valid solution. The last example is only here to show you how you can get most out of threading by paying close attention to your used data structures and CPU cache coherency. Although we are working in a virtual execution environment in a high level language with automatic memory management it does pay off to know the details down to the assembly level. Only if you continue to learn and to dig deeper you can come up with solutions no one else was even considering. I have studied particle physics which does help at the digging deeper part. Have you ever tried to solve Quantum Chromodynamics equations? Compared to that the rest must be easy ;-). Although I am no longer working in the Science field I take pride in discovering non obvious things. This can be a very hard to find bug or a new way to restructure data to make something 10 times faster. Now I need to get some sleep ….

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  • Why is the upgrade manager freezing when upgrading to 13.10

    - by Nil
    Earlier today I started to upgrade to 13.10 only to return much much later and notice that the update manager was still running. It seems to be frozen and I am reluctant to hit ctrl+C. I can't launch nautilus using the icon on the launcher. When I try to run it via the terminal, this is what happens: $ nautilus Could not register the application: GDBus.Error:org.freedesktop.DBus.Error.UnknownMethod: No such interface `org.gtk.Actions' on object at path /org/gnome/Nautilus My printers aren't showing up when I attempt to print. I don't know whether these a symptoms of the same problem. Should I let the update manage continue to run, or should I shut it down? Here are the processes running if it is any help: $ ps aux USER PID %CPU %MEM VSZ RSS TTY STAT START TIME COMMAND root 1 0.0 0.2 5920 4072 ? Ss Oct18 0:02 /sbin/init root 2 0.0 0.0 0 0 ? S Oct18 0:00 [kthreadd] root 3 0.0 0.0 0 0 ? S Oct18 0:31 [ksoftirqd/0] root 5 0.0 0.0 0 0 ? S< Oct18 0:00 [kworker/0:0H] root 6 0.0 0.0 0 0 ? S Oct18 0:00 [kworker/u:0] root 7 0.0 0.0 0 0 ? S< Oct18 0:00 [kworker/u:0H] root 8 0.0 0.0 0 0 ? S Oct18 0:00 [migration/0] root 9 0.0 0.0 0 0 ? S Oct18 0:00 [rcu_bh] root 10 0.0 0.0 0 0 ? S Oct18 0:54 [rcu_sched] root 11 0.0 0.0 0 0 ? S Oct18 0:00 [watchdog/0] root 12 0.0 0.0 0 0 ? S Oct18 0:00 [watchdog/1] root 13 0.0 0.0 0 0 ? S Oct18 0:43 [ksoftirqd/1] root 14 0.0 0.0 0 0 ? S Oct18 0:00 [migration/1] root 16 0.0 0.0 0 0 ? S< Oct18 0:00 [kworker/1:0H] root 17 0.0 0.0 0 0 ? S< Oct18 0:00 [cpuset] root 18 0.0 0.0 0 0 ? S< Oct18 0:00 [khelper] root 19 0.0 0.0 0 0 ? S Oct18 0:00 [kdevtmpfs] root 20 0.0 0.0 0 0 ? S< Oct18 0:00 [netns] root 21 0.0 0.0 0 0 ? S Oct18 0:00 [bdi-default] root 22 0.0 0.0 0 0 ? S< Oct18 0:00 [kintegrityd] root 23 0.0 0.0 0 0 ? S< Oct18 0:00 [kblockd] root 24 0.0 0.0 0 0 ? S< Oct18 0:00 [ata_sff] root 25 0.0 0.0 0 0 ? S Oct18 0:00 [khubd] root 26 0.0 0.0 0 0 ? S< Oct18 0:00 [md] root 27 0.0 0.0 0 0 ? S< Oct18 0:00 [devfreq_wq] root 29 0.0 0.0 0 0 ? S Oct18 0:00 [khungtaskd] root 30 0.0 0.0 0 0 ? S Oct18 0:09 [kswapd0] root 31 0.0 0.0 0 0 ? SN Oct18 0:00 [ksmd] root 32 0.0 0.0 0 0 ? SN Oct18 0:00 [khugepaged] root 33 0.0 0.0 0 0 ? S Oct18 0:00 [fsnotify_mark] root 34 0.0 0.0 0 0 ? S Oct18 0:00 [ecryptfs-kthre root 35 0.0 0.0 0 0 ? S< Oct18 0:00 [crypto] root 46 0.0 0.0 0 0 ? S< Oct18 0:00 [kthrotld] root 49 0.0 0.0 0 0 ? S< Oct18 0:00 [binder] root 69 0.0 0.0 0 0 ? S< Oct18 0:00 [deferwq] root 70 0.0 0.0 0 0 ? S< Oct18 0:00 [charger_manage root 166 0.0 0.0 0 0 ? S Oct18 0:00 [scsi_eh_0] root 167 0.0 0.0 0 0 ? S Oct18 0:00 [scsi_eh_1] root 188 0.0 0.0 0 0 ? S Oct18 0:00 [scsi_eh_2] root 244 0.0 0.0 0 0 ? S Oct18 0:00 [kworker/u:4] root 245 0.0 0.0 0 0 ? S< Oct18 0:00 [ttm_swap] root 260 0.0 0.0 0 0 ? S Oct18 0:00 [scsi_eh_3] root 266 0.0 0.0 0 0 ? S Oct18 0:00 [scsi_eh_4] root 267 0.0 0.0 0 0 ? S Oct18 1:08 [usb-storage] root 268 0.0 0.0 0 0 ? S Oct18 0:00 [scsi_eh_5] root 269 0.0 0.0 0 0 ? S Oct18 0:06 [usb-storage] root 302 0.0 0.0 2904 504 ? S 14:11 0:00 upstart-udev-br root 305 0.0 0.0 12080 1632 ? Ss 14:11 0:00 /lib/systemd/sy root 329 0.0 0.0 0 0 ? S Oct18 0:24 [jbd2/sda2-8] root 330 0.0 0.0 0 0 ? S< Oct18 0:00 [ext4-dio-unwri root 352 0.0 0.0 2944 4 ? S Oct18 0:00 /sbin/ureadahea root 440 0.0 0.0 0 0 ? S Oct18 0:05 [flush-8:0] root 734 0.0 0.0 0 0 ? S< Oct18 0:00 [cfg80211] root 761 0.0 0.0 0 0 ? S< Oct18 0:00 [kpsmoused] root 780 0.0 0.0 0 0 ? S Oct18 0:00 [pccardd] root 784 0.0 0.0 0 0 ? S< Oct18 0:00 [kvm-irqfd-clea root 902 0.0 0.0 0 0 ? S< Oct18 0:00 [hd-audio0] syslog 916 0.0 0.0 31120 680 ? Sl Oct18 0:13 rsyslogd -c5 102 1010 0.0 0.1 4344 1988 ? Ss Oct18 0:04 dbus-daemon --s root 1061 0.0 0.0 4844 924 ? Ss Oct18 0:00 /usr/sbin/bluet root 1077 0.0 0.0 2268 388 ? Ss Oct18 0:00 /bin/sh /etc/in root 1079 0.0 0.0 4664 484 tty4 Ss+ Oct18 0:00 /sbin/getty -8 root 1087 0.0 0.0 4664 484 tty5 Ss+ Oct18 0:00 /sbin/getty -8 root 1089 0.0 0.0 0 0 ? S< Oct18 0:00 [krfcommd] root 1098 0.0 0.0 4664 484 tty2 Ss+ Oct18 0:00 /sbin/getty -8 root 1099 0.0 0.1 4408 2076 tty3 Ss Oct18 0:00 /bin/login -- root 1101 0.0 0.0 4664 484 tty6 Ss+ Oct18 0:00 /sbin/getty -8 root 1168 0.0 0.0 2780 524 ? Ss Oct18 0:00 cron daemon 1169 0.0 0.0 2636 212 ? Ss Oct18 0:00 atd root 1183 0.0 0.0 34872 1448 ? SLsl Oct18 0:00 lightdm root 1249 0.0 0.0 3536 468 ? S Oct18 0:00 /bin/bash /etc/ root 1254 4.2 2.2 125832 40040 tty7 Rsl+ Oct18 81:27 /usr/bin/X :0 - root 1261 0.0 0.0 2268 344 ? S Oct18 0:00 /bin/sh /etc/ac root 1265 0.0 0.1 42004 2836 ? Ssl Oct18 0:01 NetworkManager root 1272 0.0 0.0 2268 376 ? S Oct18 0:00 /bin/sh /usr/sb root 1286 0.0 0.3 30616 5824 ? Sl Oct18 0:05 /usr/lib/policy root 1304 0.0 0.0 2268 372 ? D Oct18 0:00 /bin/sh /usr/li root 1360 0.0 0.0 5532 560 ? S Oct18 0:00 /sbin/dhclient nobody 1368 0.0 0.0 5476 784 ? S Oct18 0:00 /usr/sbin/dnsma root 1514 0.0 0.1 34036 1932 ? Sl Oct18 0:01 /usr/lib/accoun root 1530 0.0 0.0 0 0 ? S Oct18 0:00 [kauditd] root 1536 0.0 0.1 30480 2260 ? Sl Oct18 0:01 /usr/sbin/conso root 1653 0.0 0.1 28908 2104 ? Sl Oct18 0:00 /usr/lib/upower root 1698 0.0 0.0 17464 1388 ? Sl Oct18 0:00 lightdm --sessi rtkit 1750 0.0 0.0 21368 696 ? SNl Oct18 0:00 /usr/lib/rtkit/ 1000 1844 0.0 0.1 88116 2320 ? SLl Oct18 0:00 /usr/bin/gnome- 1000 1855 0.0 0.3 73076 5884 ? Ssl Oct18 0:00 gnome-session - 1000 1901 0.0 0.0 4128 24 ? Ss Oct18 0:00 /usr/bin/ssh-ag 1000 1904 0.0 0.0 3880 192 ? S Oct18 0:00 /usr/bin/dbus-l 1000 1905 0.0 0.1 5520 2500 ? Ss Oct18 0:23 //bin/dbus-daem 1000 1915 0.0 0.0 43348 1420 ? Sl Oct18 0:00 /usr/lib/at-spi 1000 1919 0.0 0.0 3412 1252 ? S Oct18 0:01 /bin/dbus-daemo 1000 1922 0.0 0.0 17176 1624 ? Sl Oct18 0:00 /usr/lib/at-spi 1000 1932 0.0 0.5 165916 9124 ? Sl Oct18 0:21 /usr/lib/gnome- 1000 1947 1.9 0.2 100716 4024 ? S<l Oct18 37:48 /usr/bin/pulsea 1000 1949 0.0 0.0 27568 1616 ? Sl Oct18 0:00 /usr/lib/gvfs/g 1000 1953 0.0 0.0 42628 1184 ? Sl Oct18 0:00 /usr/lib/gvfs// 1000 1962 0.0 0.0 14472 916 ? S Oct18 0:00 /usr/lib/pulsea 1000 1964 0.0 0.1 9548 2480 ? S Oct18 0:00 /usr/lib/i386-l 1000 1980 0.0 0.0 3764 364 ? S Oct18 0:43 syndaemon -i 1. 1000 1987 0.0 0.0 24476 1668 ? Sl Oct18 0:00 /usr/lib/dconf/ 1000 1990 0.0 0.4 122968 8844 ? Sl Oct18 0:00 /usr/lib/policy 1000 1991 0.0 0.2 80480 5392 ? Sl Oct18 0:00 /usr/lib/gnome- 1000 1992 0.0 1.2 167532 22776 ? Sl Oct18 0:07 nautilus -n 1000 1998 0.0 0.4 181444 7744 ? Sl Oct18 0:00 nm-applet 1000 2002 0.0 0.1 38020 2892 ? Sl Oct18 0:00 /usr/lib/gvfs/g root 2012 0.0 0.1 59908 2664 ? Sl Oct18 0:24 /usr/lib/udisks 1000 2024 0.0 0.0 26456 1540 ? Sl Oct18 0:00 /usr/lib/gvfs/g 1000 2028 0.0 0.0 27684 1536 ? Sl Oct18 0:00 /usr/lib/gvfs/g 1000 2036 0.0 0.0 38964 1452 ? Sl Oct18 0:00 /usr/lib/gvfs/g root 2049 0.0 0.0 3328 588 ? Ss Oct18 0:00 /sbin/mount.ntf 1000 2053 0.0 0.0 36792 1284 ? Sl Oct18 0:00 /usr/lib/gvfs/g 1000 2058 0.0 0.1 53664 2364 ? Sl Oct18 0:00 /usr/lib/gvfs/g 1000 2069 0.0 0.4 82816 8112 ? Sl Oct18 0:07 /usr/lib/i386-l 1000 2084 0.0 0.1 17984 2048 ? Sl Oct18 0:00 /usr/lib/gvfs/g 1000 2086 0.0 0.0 2268 392 ? Ss Oct18 0:00 /bin/sh -c /usr 1000 2087 0.0 0.7 68100 12856 ? Sl Oct18 0:13 /usr/bin/gtk-wi 1000 2089 0.0 0.9 98508 17756 ? Sl Oct18 0:13 /usr/lib/unity/ 1000 2091 0.0 0.3 65380 6692 ? Sl Oct18 0:01 /usr/lib/i386-l 1000 2117 0.0 0.2 98024 3888 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2125 0.0 0.1 86644 3408 ? Sl Oct18 0:00 /usr/lib/indica 1000 2126 0.0 0.3 84272 6664 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2127 0.0 0.1 94384 2752 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2128 0.0 0.1 83968 2828 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2129 0.0 0.2 150020 4684 ? Sl Oct18 0:01 /usr/lib/i386-l 1000 2130 0.0 0.2 86572 3884 ? Sl Oct18 0:00 /usr/lib/indica 1000 2131 0.0 0.1 69352 2524 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2144 0.0 0.1 74192 3152 ? Sl Oct18 0:00 /usr/lib/evolut 1000 2182 0.0 0.2 101120 4420 ? Sl Oct18 0:02 /usr/lib/gnome- 1000 2193 0.0 0.3 77752 6448 ? Sl Oct18 0:00 telepathy-indic 1000 2200 0.0 0.1 44032 2708 ? Sl Oct18 0:00 /usr/lib/telepa 1000 2209 0.0 0.2 77664 3860 ? Sl Oct18 0:02 zeitgeist-datah 1000 2216 0.0 0.2 44464 4180 ? Sl Oct18 0:01 /usr/bin/zeitge root 2234 0.0 0.0 0 0 ? S< Oct18 0:00 [kworker/1:1H] 1000 2246 0.0 1.1 93428 21256 ? Sl Oct18 0:02 /usr/bin/python 1000 2284 0.0 0.6 110040 11656 ? Sl Oct18 0:14 /usr/lib/i386-l 1000 2289 0.0 0.2 85632 3728 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2296 0.0 0.1 77900 3388 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2298 0.0 0.6 120356 11992 ? Sl Oct18 0:00 /usr/bin/python 1000 2300 0.0 0.1 87560 2408 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2301 0.0 0.2 91764 4404 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2303 0.0 0.2 78224 4592 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2370 0.0 0.2 74976 4908 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2372 0.0 0.4 106760 8972 ? Sl Oct18 0:00 /usr/bin/python 1000 2394 0.0 0.1 95624 2736 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2433 0.0 0.1 46640 2124 ? Sl Oct18 0:00 /usr/lib/i386-l 1000 2457 0.0 0.0 34496 1648 ? Sl Oct18 0:00 /usr/lib/libuni root 2513 0.0 0.0 0 0 ? S< Oct18 0:00 [kworker/0:1H] 1000 3361 0.0 0.0 2268 396 ? SN 07:54 0:20 /bin/sh -c /usr root 4919 1.8 2.1 201196 38760 ? SNl 13:29 11:25 /usr/bin/python root 4957 0.0 0.0 3880 400 ? SN 13:29 0:00 dbus-launch --a root 4958 0.0 0.0 3424 1196 ? SNs 13:29 0:05 //bin/dbus-daem root 5128 0.0 0.0 2268 416 ? SN 13:50 0:00 /bin/sh -c whil root 5141 0.0 0.0 2436 508 ? SN 13:50 0:00 gnome-pty-helpe root 5145 0.0 1.7 245280 30872 pts/1 SNs+ 13:50 0:05 /usr/bin/python root 5159 0.0 0.4 64200 7432 ? SNl 13:50 0:05 /usr/bin/gnome- root 5163 0.0 0.0 27440 1552 ? SNl 13:50 0:00 /usr/lib/gvfs/g root 5167 0.0 0.0 42628 1648 ? SNl 13:50 0:00 /usr/lib/gvfs// root 9236 0.0 0.1 19112 2680 ? Ss 14:33 0:00 /usr/sbin/winbi root 9243 0.0 0.0 19112 1448 ? S 14:33 0:00 /usr/sbin/winbi whoopsie 9409 0.0 0.2 53608 4264 ? Ssl 14:33 0:00 whoopsie root 20087 0.0 0.0 0 0 ? S< 14:34 0:00 [xfsalloc] root 20088 0.0 0.0 0 0 ? S< 14:34 0:00 [xfs_mru_cache] root 20089 0.0 0.0 0 0 ? S< 14:34 0:00 [xfslogd] root 20092 0.0 0.0 0 0 ? S 14:34 0:00 [jfsIO] root 20093 0.0 0.0 0 0 ? S 14:34 0:00 [jfsCommit] root 20094 0.0 0.0 0 0 ? S 14:34 0:00 [jfsCommit] root 20095 0.0 0.0 0 0 ? S 14:34 0:00 [jfsSync] root 20845 0.0 0.3 7980 6048 pts/2 SNs+ 14:29 0:04 /usr/bin/dpkg - root 23330 0.0 0.0 2896 568 ? S 14:09 0:00 upstart-file-br root 23332 0.0 0.0 2884 572 ? S 14:09 0:00 upstart-socket- root 24577 0.2 0.0 0 0 ? S 23:09 0:04 [kworker/1:2] root 24656 0.1 0.0 0 0 ? S 23:10 0:02 [kworker/0:0] 1000 24758 2.8 4.7 243692 85516 ? Sl 23:11 0:50 compiz root 25774 0.0 0.0 0 0 ? S< 14:39 0:00 [iprt] 1000 26128 5.5 10.3 641628 187420 ? Sl 23:27 0:46 /usr/lib/firefo root 26374 0.0 0.0 3964 720 ? Ss 14:39 0:02 /usr/sbin/irqba root 26534 0.0 0.0 0 0 ? S 23:34 0:00 [kworker/0:1] root 26564 0.0 0.0 0 0 ? S 23:35 0:00 [kworker/1:1] 1000 26664 0.0 0.1 6784 3068 tty3 S+ 23:36 0:00 -bash 1000 26936 15.2 1.3 67520 23672 ? Sl 23:39 0:21 gnome-system-mo root 26992 0.0 0.0 0 0 ? S 23:40 0:00 [kworker/1:0] root 27049 0.0 0.0 4248 288 ? SN 23:41 0:00 sleep 30 1000 27057 9.5 0.8 68624 16140 ? Rl 23:41 0:00 gnome-terminal 1000 27064 0.0 0.0 2440 704 ? S 23:41 0:00 gnome-pty-helpe 1000 27065 2.6 0.1 6344 2608 pts/3 Ss 23:41 0:00 bash 1000 27113 0.0 0.0 5240 1144 pts/3 R+ 23:41 0:00 ps aux root 28267 0.0 0.0 2216 632 ? Ss 14:39 0:00 acpid -c /etc/a root 28333 0.0 0.0 2272 552 pts/2 SN+ 14:39 0:00 /bin/sh /var/li root 29699 0.0 0.2 8384 4608 pts/2 SN+ 14:40 0:00 modprobe wl

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  • openvpn WARNING: No server certificate verification method has been enabled

    - by tmedtcom
    I tried to install openvpn on debian squeez (server) and connect from my fedora 17 as (client). Here is my configuration: server configuration ###cat server.conf # Serveur TCP ** proto tcp** port 1194 dev tun # Cles et certificats ca /etc/openvpn/easy-rsa/keys/ca.crt cert /etc/openvpn/easy-rsa/keys/server.crt key /etc/openvpn/easy-rsa/keys/server.key dh /etc/openvpn/easy-rsa/keys/dh1024.pem # Reseau #Adresse virtuel du reseau vpn server 192.170.70.0 255.255.255.0 #Cette ligne ajoute sur le client la route du reseau vers le serveur push "route 192.168.1.0 255.255.255.0" #Creer une route du server vers l'interface tun. #route 192.170.70.0 255.255.255.0 # Securite keepalive 10 120 #type d'encryptage des données **cipher AES-128-CBC** #activation de la compression comp-lzo #nombre maximum de clients autorisés max-clients 10 #pas d'utilisateur et groupe particuliers pour l'utilisation du VPN user nobody group nogroup #pour rendre la connexion persistante persist-key persist-tun #Log d'etat d'OpenVPN status /var/log/openvpn-status.log #logs openvpnlog /var/log/openvpn.log log-append /var/log/openvpn.log #niveau de verbosité verb 5 ###cat client.conf # Client client dev tun [COLOR="Red"]proto tcp-client[/COLOR] remote <my server wan IP> 1194 resolv-retry infinite **cipher AES-128-CBC** # Cles ca ca.crt cert client.crt key client.key # Securite nobind persist-key persist-tun comp-lzo verb 3 Message from the host client (fedora 17) in the log file / var / log / messages: Dec 6 21:56:00 GlobalTIC NetworkManager[691]: <info> Starting VPN service 'openvpn'... Dec 6 21:56:00 GlobalTIC NetworkManager[691]: <info> VPN service 'openvpn' started (org.freedesktop.NetworkManager.openvpn), PID 7470 Dec 6 21:56:00 GlobalTIC NetworkManager[691]: <info> VPN service 'openvpn' appeared; activating connections Dec 6 21:56:00 GlobalTIC NetworkManager[691]: <info> VPN plugin state changed: starting (3) Dec 6 21:56:01 GlobalTIC NetworkManager[691]: <info> VPN connection 'Connexion VPN 1' (Connect) reply received. Dec 6 21:56:01 GlobalTIC nm-openvpn[7472]: OpenVPN 2.2.2 x86_64-redhat-linux-gnu [SSL] [LZO2] [EPOLL] [PKCS11] [eurephia] built on Sep 5 2012 Dec 6 21:56:01 GlobalTIC nm-openvpn[7472]:[COLOR="Red"][U][B] WARNING: No server certificate verification method has been enabled.[/B][/U][/COLOR] See http://openvpn.net/howto.html#mitm for more info. Dec 6 21:56:01 GlobalTIC nm-openvpn[7472]: NOTE: the current --script-security setting may allow this configuration to call user-defined scripts Dec 6 21:56:01 GlobalTIC nm-openvpn[7472]:[COLOR="Red"] WARNING: file '/home/login/client/client.key' is group or others accessible[/COLOR] Dec 6 21:56:01 GlobalTIC nm-openvpn[7472]: UDPv4 link local: [undef] Dec 6 21:56:01 GlobalTIC nm-openvpn[7472]: UDPv4 link remote: [COLOR="Red"]<my server wan IP>[/COLOR]:1194 Dec 6 21:56:01 GlobalTIC nm-openvpn[7472]: [COLOR="Red"]read UDPv4 [ECONNREFUSED]: Connection refused (code=111)[/COLOR] Dec 6 21:56:03 GlobalTIC nm-openvpn[7472]: [COLOR="Red"]read UDPv4[/COLOR] [ECONNREFUSED]: Connection refused (code=111) Dec 6 21:56:07 GlobalTIC nm-openvpn[7472]: read UDPv4 [ECONNREFUSED]: Connection refused (code=111) Dec 6 21:56:15 GlobalTIC nm-openvpn[7472]: read UDPv4 [ECONNREFUSED]: Connection refused (code=111) Dec 6 21:56:31 GlobalTIC nm-openvpn[7472]: read UDPv4 [ECONNREFUSED]: Connection refused (code=111) Dec 6 21:56:41 GlobalTIC NetworkManager[691]: <warn> VPN connection 'Connexion VPN 1' (IP Conf[/CODE] ifconfig on server host(debian): ifconfig eth0 Link encap:Ethernet HWaddr 08:00:27:16:21:ac inet addr:192.168.1.6 Bcast:192.168.1.255 Mask:255.255.255.0 inet6 addr: fe80::a00:27ff:fe16:21ac/64 Scope:Link UP BROADCAST RUNNING MULTICAST MTU:1500 Metric:1 RX packets:9059 errors:0 dropped:0 overruns:0 frame:0 TX packets:5660 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:1000 RX bytes:919427 (897.8 KiB) TX bytes:1273891 (1.2 MiB) tun0 Link encap:UNSPEC HWaddr 00-00-00-00-00-00-00-00-00-00-00-00-00-00-00-00 inet addr:192.170.70.1 P-t-P:192.170.70.2 Mask:255.255.255.255 UP POINTOPOINT RUNNING NOARP MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:0 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:100 RX bytes:0 (0.0 B) TX bytes:0 (0.0 B) ifconfig on the client host (fedora 17) as0t0: flags=4305<UP,POINTOPOINT,RUNNING,NOARP,MULTICAST> mtu 1500 inet 5.5.0.1 netmask 255.255.252.0 destination 5.5.0.1 unspec 00-00-00-00-00-00-00-00-00-00-00-00-00-00-00-00 txqueuelen 200 (UNSPEC) RX packets 0 bytes 0 (0.0 B) RX errors 0 dropped 0 overruns 0 frame 0 TX packets 2 bytes 321 (321.0 B) TX errors 0 dropped 0 overruns 0 carrier 0 collisions 0 as0t1: flags=4305<UP,POINTOPOINT,RUNNING,NOARP,MULTICAST> mtu 1500 inet 5.5.4.1 netmask 255.255.252.0 destination 5.5.4.1 unspec 00-00-00-00-00-00-00-00-00-00-00-00-00-00-00-00 txqueuelen 200 (UNSPEC) RX packets 0 bytes 0 (0.0 B) RX errors 0 dropped 0 overruns 0 frame 0 TX packets 2 bytes 321 (321.0 B) TX errors 0 dropped 0 overruns 0 carrier 0 collisions 0 as0t2: flags=4305<UP,POINTOPOINT,RUNNING,NOARP,MULTICAST> mtu 1500 inet 5.5.8.1 netmask 255.255.252.0 destination 5.5.8.1 unspec 00-00-00-00-00-00-00-00-00-00-00-00-00-00-00-00 txqueuelen 200 (UNSPEC) RX packets 0 bytes 0 (0.0 B) RX errors 0 dropped 0 overruns 0 frame 0 TX packets 2 bytes 321 (321.0 B) TX errors 0 dropped 0 overruns 0 carrier 0 collisions 0 as0t3: flags=4305<UP,POINTOPOINT,RUNNING,NOARP,MULTICAST> mtu 1500 inet 5.5.12.1 netmask 255.255.252.0 destination 5.5.12.1 unspec 00-00-00-00-00-00-00-00-00-00-00-00-00-00-00-00 txqueuelen 200 (UNSPEC) RX packets 0 bytes 0 (0.0 B) RX errors 0 dropped 0 overruns 0 frame 0 TX packets 2 bytes 321 (321.0 B) TX errors 0 dropped 0 overruns 0 carrier 0 collisions 0 **p255p1**: flags=4163<UP,BROADCAST,RUNNING,MULTICAST> mtu 1500 inet 192.168.1.2 netmask 255.255.255.0 broadcast 192.168.1.255 inet6 fe80::21d:baff:fe20:b7e6 prefixlen 64 scopeid 0x20<link> ether 00:1d:ba:20:b7:e6 txqueuelen 1000 (Ethernet) RX packets 4842070 bytes 3579798184 (3.3 GiB) RX errors 0 dropped 0 overruns 0 frame 0 TX packets 3996158 bytes 2436442882 (2.2 GiB) TX errors 0 dropped 0 overruns 0 carrier 0 collisions 0 device interrupt 16 p255p1 is label for eth0 interface and on the server : root@hoteserver:/etc/openvpn# tree . +-- client ¦** +-- ca.crt ¦** +-- client.conf ¦** +-- client.crt ¦** +-- client.csr ¦** +-- client.key ¦** +-- client.ovpn ¦* ¦** +-- easy-rsa ¦** +-- build-ca ¦** +-- build-dh ¦** +-- build-inter ¦** +-- build-key ¦** +-- build-key-pass ¦** +-- build-key-pkcs12 ¦** +-- build-key-server ¦** +-- build-req ¦** +-- build-req-pass ¦** +-- clean-all ¦** +-- inherit-inter ¦** +-- keys ¦** ¦** +-- 01.pem ¦** ¦** +-- 02.pem ¦** ¦** +-- ca.crt ¦** ¦** +-- ca.key ¦** ¦** +-- client.crt ¦** ¦** +-- client.csr ¦** ¦** +-- client.key ¦** ¦** +-- dh1024.pem ¦** ¦** +-- index.txt ¦** ¦** +-- index.txt.attr ¦** ¦** +-- index.txt.attr.old ¦** ¦** +-- index.txt.old ¦** ¦** +-- serial ¦** ¦** +-- serial.old ¦** ¦** +-- server.crt ¦** ¦** +-- server.csr ¦** ¦** +-- server.key ¦** +-- list-crl ¦** +-- Makefile ¦** +-- openssl-0.9.6.cnf.gz ¦** +-- openssl.cnf ¦** +-- pkitool ¦** +-- README.gz ¦** +-- revoke-full ¦** +-- sign-req ¦** +-- vars ¦** +-- whichopensslcnf +-- openvpn.log +-- openvpn-status.log +-- server.conf +-- update-resolv-conf on the client: [login@hoteclient openvpn]$ tree . |-- easy-rsa | |-- 1.0 | | |-- build-ca | | |-- build-dh | | |-- build-inter | | |-- build-key | | |-- build-key-pass | | |-- build-key-pkcs12 | | |-- build-key-server | | |-- build-req | | |-- build-req-pass | | |-- clean-all | | |-- list-crl | | |-- make-crl | | |-- openssl.cnf | | |-- README | | |-- revoke-crt | | |-- revoke-full | | |-- sign-req | | `-- vars | `-- 2.0 | |-- build-ca | |-- build-dh | |-- build-inter | |-- build-key | |-- build-key-pass | |-- build-key-pkcs12 | |-- build-key-server | |-- build-req | |-- build-req-pass | |-- clean-all | |-- inherit-inter | |-- keys [error opening dir] | |-- list-crl | |-- Makefile | |-- openssl-0.9.6.cnf | |-- openssl-0.9.8.cnf | |-- openssl-1.0.0.cnf | |-- pkitool | |-- README | |-- revoke-full | |-- sign-req | |-- vars | `-- whichopensslcnf |-- keys -> ./easy-rsa/2.0/keys/ `-- server.conf the problem source is cipher AES-128-CBC ,proto tcp-client or UDP or the interface p255p1 on fedora17 or file authentification ta.key is not found ????

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  • openvpn: after changing to server mode, client does not create TUN device

    - by lurscher
    i had a previously working configuration with the config files used in a previous question However, i've changed this now to the following configuration using server mode, everything on the logs seem fine, however the client doesn't create any tun interface, so i don't have anything to connect to, presumably, i need to add or push some route commands, but i don't have any idea at this point what i need to do. I am posting all my relevant configuration files server.conf: dev tun server 10.8.117.0 255.255.255.0 ifconfig-pool-persist ipp.txt tls-server dh /home/lurscher/keys/dh1024.pem ca /home/lurscher/keys/ca.crt cert /home/lurscher/keys/vpnCh8TestServer.crt key /home/lurscher/keys/vpnCh8TestServer.key status openvpn-status.log log openvpn.log comp-lzo verb 3 and client.conf: dev tun remote my.server.com tls-client ca /home/chuckq/keys/ca.crt cert /home/chuckq/keys/vpnCh8TestClient.crt key /home/chuckq/keys/vpnCh8TestClient.key ns-cert-type server ; port 1194 ; user nobody ; group nogroup status openvpn-status.log log openvpn.log comp-lzo verb 3 the server ifconfig shows a tun device: tun0 Link encap:UNSPEC HWaddr 00-00-00-00-00-00-00-00-00-00-00-00-00-00-00-00 inet addr:10.8.117.1 P-t-P:10.8.117.2 Mask:255.255.255.255 UP POINTOPOINT RUNNING NOARP MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:0 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:100 RX bytes:0 (0.0 B) TX bytes:0 (0.0 B) However the client ifconfig does not show any tun interface! $ ifconfig tun0 tun0 Link encap:UNSPEC HWaddr 00-00-00-00-00-00-00-00-00-00-00-00-00-00-00-00 POINTOPOINT NOARP MULTICAST MTU:1500 Metric:1 RX packets:0 errors:0 dropped:0 overruns:0 frame:0 TX packets:0 errors:0 dropped:0 overruns:0 carrier:0 collisions:0 txqueuelen:100 RX bytes:0 (0.0 B) TX bytes:0 (0.0 B) the client log says: Tue May 17 23:27:09 2011 OpenVPN 2.1.0 i686-pc-linux-gnu [SSL] [LZO2] [EPOLL] [PKCS11] [MH] [PF_INET6] [eurephia] built on Jul 12 2010 Tue May 17 23:27:09 2011 IMPORTANT: OpenVPN's default port number is now 1194, based on an official port number assignment by IANA. OpenVPN 2.0-beta16 and earlier used 5000 as the default port. Tue May 17 23:27:09 2011 NOTE: the current --script-security setting may allow this configuration to call user-defined scripts Tue May 17 23:27:09 2011 /usr/bin/openssl-vulnkey -q -b 1024 -m <modulus omitted> Tue May 17 23:27:09 2011 LZO compression initialized Tue May 17 23:27:09 2011 Control Channel MTU parms [ L:1542 D:138 EF:38 EB:0 ET:0 EL:0 ] Tue May 17 23:27:09 2011 TUN/TAP device tun0 opened Tue May 17 23:27:09 2011 TUN/TAP TX queue length set to 100 Tue May 17 23:27:09 2011 Data Channel MTU parms [ L:1542 D:1450 EF:42 EB:135 ET:0 EL:0 AF:3/1 ] Tue May 17 23:27:09 2011 Local Options hash (VER=V4): '41690919' Tue May 17 23:27:09 2011 Expected Remote Options hash (VER=V4): '530fdded' Tue May 17 23:27:09 2011 Socket Buffers: R=[114688->131072] S=[114688->131072] Tue May 17 23:27:09 2011 UDPv4 link local (bound): [undef] Tue May 17 23:27:09 2011 UDPv4 link remote: [AF_INET]192.168.0.101:1194 Tue May 17 23:27:09 2011 TLS: Initial packet from [AF_INET]192.168.0.101:1194, sid=8e8bdc33 f4275407 Tue May 17 23:27:09 2011 VERIFY OK: depth=1, /C=CA/ST=Out/L=There/O=Ubuntu/OU=Home/CN=Ubuntu_CA/name=lurscher/[email protected] Tue May 17 23:27:09 2011 VERIFY OK: nsCertType=SERVER Tue May 17 23:27:09 2011 VERIFY OK: depth=0, /C=CA/ST=Out/L=There/O=Ubuntu/OU=Home/CN=vpnCh8TestServer/name=lurscher/[email protected] Tue May 17 23:27:09 2011 Data Channel Encrypt: Cipher 'BF-CBC' initialized with 128 bit key Tue May 17 23:27:09 2011 Data Channel Encrypt: Using 160 bit message hash 'SHA1' for HMAC authentication Tue May 17 23:27:09 2011 Data Channel Decrypt: Cipher 'BF-CBC' initialized with 128 bit key Tue May 17 23:27:09 2011 Data Channel Decrypt: Using 160 bit message hash 'SHA1' for HMAC authentication Tue May 17 23:27:09 2011 Control Channel: TLSv1, cipher TLSv1/SSLv3 DHE-RSA-AES256-SHA, 1024 bit RSA Tue May 17 23:27:09 2011 [vpnCh8TestServer] Peer Connection Initiated with [AF_INET]192.168.0.101:1194 Tue May 17 23:27:10 2011 Initialization Sequence Completed the client status log: OpenVPN STATISTICS Updated,Tue May 17 23:30:09 2011 TUN/TAP read bytes,0 TUN/TAP write bytes,0 TCP/UDP read bytes,5604 TCP/UDP write bytes,4244 Auth read bytes,0 pre-compress bytes,0 post-compress bytes,0 pre-decompress bytes,0 post-decompress bytes,0 END and the server log says: Tue May 17 23:18:25 2011 OpenVPN 2.1.0 x86_64-pc-linux-gnu [SSL] [LZO2] [EPOLL] [PKCS11] [MH] [PF_INET6] [eurephia] built on Jul 12 2010 Tue May 17 23:18:25 2011 IMPORTANT: OpenVPN's default port number is now 1194, based on an official port number assignment by IANA. OpenVPN 2.0-beta16 and earlier used 5000 as the default port. Tue May 17 23:18:25 2011 WARNING: --keepalive option is missing from server config Tue May 17 23:18:25 2011 NOTE: your local LAN uses the extremely common subnet address 192.168.0.x or 192.168.1.x. Be aware that this might create routing conflicts if you connect to the VPN server from public locations such as internet cafes that use the same subnet. Tue May 17 23:18:25 2011 NOTE: the current --script-security setting may allow this configuration to call user-defined scripts Tue May 17 23:18:25 2011 Diffie-Hellman initialized with 1024 bit key Tue May 17 23:18:25 2011 /usr/bin/openssl-vulnkey -q -b 1024 -m <modulus omitted> Tue May 17 23:18:25 2011 TLS-Auth MTU parms [ L:1542 D:138 EF:38 EB:0 ET:0 EL:0 ] Tue May 17 23:18:25 2011 ROUTE default_gateway=192.168.0.1 Tue May 17 23:18:25 2011 TUN/TAP device tun0 opened Tue May 17 23:18:25 2011 TUN/TAP TX queue length set to 100 Tue May 17 23:18:25 2011 /sbin/ifconfig tun0 10.8.117.1 pointopoint 10.8.117.2 mtu 1500 Tue May 17 23:18:25 2011 /sbin/route add -net 10.8.117.0 netmask 255.255.255.0 gw 10.8.117.2 Tue May 17 23:18:25 2011 Data Channel MTU parms [ L:1542 D:1450 EF:42 EB:135 ET:0 EL:0 AF:3/1 ] Tue May 17 23:18:25 2011 Socket Buffers: R=[126976->131072] S=[126976->131072] Tue May 17 23:18:25 2011 UDPv4 link local (bound): [undef] Tue May 17 23:18:25 2011 UDPv4 link remote: [undef] Tue May 17 23:18:25 2011 MULTI: multi_init called, r=256 v=256 Tue May 17 23:18:25 2011 IFCONFIG POOL: base=10.8.117.4 size=62 Tue May 17 23:18:25 2011 IFCONFIG POOL LIST Tue May 17 23:18:25 2011 vpnCh8TestClient,10.8.117.4 Tue May 17 23:18:25 2011 Initialization Sequence Completed Tue May 17 23:27:22 2011 MULTI: multi_create_instance called Tue May 17 23:27:22 2011 192.168.0.104:1194 Re-using SSL/TLS context Tue May 17 23:27:22 2011 192.168.0.104:1194 LZO compression initialized Tue May 17 23:27:22 2011 192.168.0.104:1194 Control Channel MTU parms [ L:1542 D:138 EF:38 EB:0 ET:0 EL:0 ] Tue May 17 23:27:22 2011 192.168.0.104:1194 Data Channel MTU parms [ L:1542 D:1450 EF:42 EB:135 ET:0 EL:0 AF:3/1 ] Tue May 17 23:27:22 2011 192.168.0.104:1194 Local Options hash (VER=V4): '530fdded' Tue May 17 23:27:22 2011 192.168.0.104:1194 Expected Remote Options hash (VER=V4): '41690919' Tue May 17 23:27:22 2011 192.168.0.104:1194 TLS: Initial packet from [AF_INET]192.168.0.104:1194, sid=8972b565 79323f68 Tue May 17 23:27:22 2011 192.168.0.104:1194 VERIFY OK: depth=1, /C=CA/ST=Out/L=There/O=Ubuntu/OU=Home/CN=Ubuntu_CA/name=lurscher/[email protected] Tue May 17 23:27:22 2011 192.168.0.104:1194 VERIFY OK: depth=0, /C=CA/ST=Out/L=There/O=Ubuntu/OU=Home/CN=Ubuntu_CA/name=lurscher/[email protected] Tue May 17 23:27:22 2011 192.168.0.104:1194 Data Channel Encrypt: Cipher 'BF-CBC' initialized with 128 bit key Tue May 17 23:27:22 2011 192.168.0.104:1194 Data Channel Encrypt: Using 160 bit message hash 'SHA1' for HMAC authentication Tue May 17 23:27:22 2011 192.168.0.104:1194 Data Channel Decrypt: Cipher 'BF-CBC' initialized with 128 bit key Tue May 17 23:27:22 2011 192.168.0.104:1194 Data Channel Decrypt: Using 160 bit message hash 'SHA1' for HMAC authentication Tue May 17 23:27:22 2011 192.168.0.104:1194 Control Channel: TLSv1, cipher TLSv1/SSLv3 DHE-RSA-AES256-SHA, 1024 bit RSA Tue May 17 23:27:22 2011 192.168.0.104:1194 [vpnCh8TestClient] Peer Connection Initiated with [AF_INET]192.168.0.104:1194 Tue May 17 23:27:22 2011 vpnCh8TestClient/192.168.0.104:1194 MULTI: Learn: 10.8.117.6 -> vpnCh8TestClient/192.168.0.104:1194 Tue May 17 23:27:22 2011 vpnCh8TestClient/192.168.0.104:1194 MULTI: primary virtual IP for vpnCh8TestClient/192.168.0.104:1194: 10.8.117.6 finally, the server status log: OpenVPN CLIENT LIST Updated,Tue May 17 23:36:25 2011 Common Name,Real Address,Bytes Received,Bytes Sent,Connected Since vpnCh8TestClient,192.168.0.104:1194,4244,5604,Tue May 17 23:27:22 2011 ROUTING TABLE Virtual Address,Common Name,Real Address,Last Ref 10.8.117.6,vpnCh8TestClient,192.168.0.104:1194,Tue May 17 23:27:22 2011 GLOBAL STATS Max bcast/mcast queue length,0 END

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  • Does anyone really understand how HFSC scheduling in Linux/BSD works?

    - by Mecki
    I read the original SIGCOMM '97 PostScript paper about HFSC, it is very technically, but I understand the basic concept. Instead of giving a linear service curve (as with pretty much every other scheduling algorithm), you can specify a convex or concave service curve and thus it is possible to decouple bandwidth and delay. However, even though this paper mentions to kind of scheduling algorithms being used (real-time and link-share), it always only mentions ONE curve per scheduling class (the decoupling is done by specifying this curve, only one curve is needed for that). Now HFSC has been implemented for BSD (OpenBSD, FreeBSD, etc.) using the ALTQ scheduling framework and it has been implemented Linux using the TC scheduling framework (part of iproute2). Both implementations added two additional service curves, that were NOT in the original paper! A real-time service curve and an upper-limit service curve. Again, please note that the original paper mentions two scheduling algorithms (real-time and link-share), but in that paper both work with one single service curve. There never have been two independent service curves for either one as you currently find in BSD and Linux. Even worse, some version of ALTQ seems to add an additional queue priority to HSFC (there is no such thing as priority in the original paper either). I found several BSD HowTo's mentioning this priority setting (even though the man page of the latest ALTQ release knows no such parameter for HSFC, so officially it does not even exist). This all makes the HFSC scheduling even more complex than the algorithm described in the original paper and there are tons of tutorials on the Internet that often contradict each other, one claiming the opposite of the other one. This is probably the main reason why nobody really seems to understand how HFSC scheduling really works. Before I can ask my questions, we need a sample setup of some kind. I'll use a very simple one as seen in the image below: Here are some questions I cannot answer because the tutorials contradict each other: What for do I need a real-time curve at all? Assuming A1, A2, B1, B2 are all 128 kbit/s link-share (no real-time curve for either one), then each of those will get 128 kbit/s if the root has 512 kbit/s to distribute (and A and B are both 256 kbit/s of course), right? Why would I additionally give A1 and B1 a real-time curve with 128 kbit/s? What would this be good for? To give those two a higher priority? According to original paper I can give them a higher priority by using a curve, that's what HFSC is all about after all. By giving both classes a curve of [256kbit/s 20ms 128kbit/s] both have twice the priority than A2 and B2 automatically (still only getting 128 kbit/s on average) Does the real-time bandwidth count towards the link-share bandwidth? E.g. if A1 and B1 both only have 64kbit/s real-time and 64kbit/s link-share bandwidth, does that mean once they are served 64kbit/s via real-time, their link-share requirement is satisfied as well (they might get excess bandwidth, but lets ignore that for a second) or does that mean they get another 64 kbit/s via link-share? So does each class has a bandwidth "requirement" of real-time plus link-share? Or does a class only have a higher requirement than the real-time curve if the link-share curve is higher than the real-time curve (current link-share requirement equals specified link-share requirement minus real-time bandwidth already provided to this class)? Is upper limit curve applied to real-time as well, only to link-share, or maybe to both? Some tutorials say one way, some say the other way. Some even claim upper-limit is the maximum for real-time bandwidth + link-share bandwidth? What is the truth? Assuming A2 and B2 are both 128 kbit/s, does it make any difference if A1 and B1 are 128 kbit/s link-share only, or 64 kbit/s real-time and 128 kbit/s link-share, and if so, what difference? If I use the seperate real-time curve to increase priorities of classes, why would I need "curves" at all? Why is not real-time a flat value and link-share also a flat value? Why are both curves? The need for curves is clear in the original paper, because there is only one attribute of that kind per class. But now, having three attributes (real-time, link-share, and upper-limit) what for do I still need curves on each one? Why would I want the curves shape (not average bandwidth, but their slopes) to be different for real-time and link-share traffic? According to the little documentation available, real-time curve values are totally ignored for inner classes (class A and B), they are only applied to leaf classes (A1, A2, B1, B2). If that is true, why does the ALTQ HFSC sample configuration (search for 3.3 Sample configuration) set real-time curves on inner classes and claims that those set the guaranteed rate of those inner classes? Isn't that completely pointless? (note: pshare sets the link-share curve in ALTQ and grate the real-time curve; you can see this in the paragraph above the sample configuration). Some tutorials say the sum of all real-time curves may not be higher than 80% of the line speed, others say it must not be higher than 70% of the line speed. Which one is right or are they maybe both wrong? One tutorial said you shall forget all the theory. No matter how things really work (schedulers and bandwidth distribution), imagine the three curves according to the following "simplified mind model": real-time is the guaranteed bandwidth that this class will always get. link-share is the bandwidth that this class wants to become fully satisfied, but satisfaction cannot be guaranteed. In case there is excess bandwidth, the class might even get offered more bandwidth than necessary to become satisfied, but it may never use more than upper-limit says. For all this to work, the sum of all real-time bandwidths may not be above xx% of the line speed (see question above, the percentage varies). Question: Is this more or less accurate or a total misunderstanding of HSFC? And if assumption above is really accurate, where is prioritization in that model? E.g. every class might have a real-time bandwidth (guaranteed), a link-share bandwidth (not guaranteed) and an maybe an upper-limit, but still some classes have higher priority needs than other classes. In that case I must still prioritize somehow, even among real-time traffic of those classes. Would I prioritize by the slope of the curves? And if so, which curve? The real-time curve? The link-share curve? The upper-limit curve? All of them? Would I give all of them the same slope or each a different one and how to find out the right slope? I still haven't lost hope that there exists at least a hand full of people in this world that really understood HFSC and are able to answer all these questions accurately. And doing so without contradicting each other in the answers would be really nice ;-)

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  • Can't connect to samba

    - by Rick
    Windows 7, connecting to Samba shares I have a follow up question from the link above. I am running Samba 3.0.23d on FreeBSD is release 7.1 I changed the policies as described above but still cannot connect to the samba server with the windows 7 or a server 2008. I feel it is a problem with recognizing the new machines on the network. the windows machines can see the samba server, but cannot connect to it or view any of the files. After changing the security policies the samba server asked for network id and password but would not allow the machine to connect, said they were unknown username or bad password. Here is my current config file. there is no sign of encryption anywhere, should I just add the line? not sure what that would do elsewhere. Workgroup = WWOFFSET server string = WWO File Server (%v) security = server username map = /usr/local/etc/smb.users hosts allow = 10. 127. # If you want to automatically load your printer list rather # than setting them up individually then you'll need this ; load printers = yes # you may wish to override the location of the printcap file ; printcap name = /etc/printcap # on SystemV system setting printcap name to lpstat should allow # you to automatically obtain a printer list from the SystemV spool # system ; printcap name = lpstat # It should not be necessary to specify the print system type unless # it is non-standard. Currently supported print systems include: # bsd, cups, sysv, plp, lprng, aix, hpux, qnx ; printing = cups # Uncomment this if you want a guest account, you must add this to /etc/passwd # otherwise the user "nobody" is used ; guest account = pcguest # this tells Samba to use a separate log file for each machine # that connects log file = /var/log/samba/log.%m # Put a capping on the size of the log files (in Kb). max log size = 50 # Use password server option only with security = server # The argument list may include: # password server = My_PDC_Name [My_BDC_Name] [My_Next_BDC_Name] # or to auto-locate the domain controller/s # password server = * ; password server = <NT-Server-Name> password server = SERVER0 # Use the realm option only with security = ads # Specifies the Active Directory realm the host is part of ; realm = MY_REALM # Backend to store user information in. New installations should # use either tdbsam or ldapsam. smbpasswd is available for backwards # compatibility. tdbsam requires no further configuration. ; passdb backend = tdbsam ; passdb backend = smbpasswd # Using the following line enables you to customise your configuration # on a per machine basis. The %m gets replaced with the netbios name # of the machine that is connecting. # Note: Consider carefully the location in the configuration file of # this line. The included file is read at that point. ; include = /usr/local/etc/smb.conf.%m # Most people will find that this option gives better performance. # See the chapter 'Samba performance issues' in the Samba HOWTO Collection # and the manual pages for details. # You may want to add the following on a Linux system: # SO_RCVBUF=8192 SO_SNDBUF=8192 socket options = TCP_NODELAY # Configure Samba to use multiple interfaces # If you have multiple network interfaces then you must list them # here. See the man page for details. ; interfaces = 192.168.12.2/24 192.168.13.2/24 # Browser Control Options: # set local master to no if you don't want Samba to become a master # browser on your network. Otherwise the normal election rules apply ; local master = no # OS Level determines the precedence of this server in master browser # elections. The default value should be reasonable ; os level = 33 # Domain Master specifies Samba to be the Domain Master Browser. This # allows Samba to collate browse lists between subnets. Don't use this # if you already have a Windows NT domain controller doing this job ; domain master = yes # Preferred Master causes Samba to force a local browser election on startup # and gives it a slightly higher chance of winning the election ; preferred master = yes # Enable this if you want Samba to be a domain logon server for # Windows95 workstations. ; domain logons = yes # if you enable domain logons then you may want a per-machine or # per user logon script # run a specific logon batch file per workstation (machine) ; logon script = %m.bat # run a specific logon batch file per username ; logon script = %U.bat # Where to store roving profiles (only for Win95 and WinNT) # %L substitutes for this servers netbios name, %U is username # You must uncomment the [Profiles] share below ; logon path = \\%L\Profiles\%U # Windows Internet Name Serving Support Section: # WINS Support - Tells the NMBD component of Samba to enable it's WINS Server ; wins support = yes # WINS Server - Tells the NMBD components of Samba to be a WINS Client # Note: Samba can be either a WINS Server, or a WINS Client, but NOT both ; wins server = w.x.y.z # WINS Proxy - Tells Samba to answer name resolution queries on # behalf of a non WINS capable client, for this to work there must be # at least one WINS Server on the network. The default is NO. ; wins proxy = yes # DNS Proxy - tells Samba whether or not to try to resolve NetBIOS names # via DNS nslookups. The default is NO. dns proxy = no # charset settings ; display charset = ASCII ; unix charset = ASCII ; dos charset = ASCII # These scripts are used on a domain controller or stand-alone # machine to add or delete corresponding unix accounts ; add user script = /usr/sbin/useradd %u ; add group script = /usr/sbin/groupadd %g ; add machine script = /usr/sbin/adduser -n -g machines -c Machine -d /dev/null -s /bin/false %u ; delete user script = /usr/sbin/userdel %u ; delete user from group script = /usr/sbin/deluser %u %g ; delete group script = /usr/sbin/groupdel %g unix extensions = no

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  • IRQ problem with 2.6.32/2.6.39 kernel on Debian Squeeze x86_64

    - by MasterM
    I recently assembled a new computer so that all hardware is pretty new. Since then I've been experiencing some problem with IRQs when running Debian 6.0. On random occasions, usually after an hour or so of running I hear a beep and this shows up in dmesg: [ 3537.762795] irq 16: nobody cared (try booting with the "irqpoll" option) [ 3537.762797] Pid: 0, comm: swapper Tainted: P W O 2.6.39-2-amd64 #1 [ 3537.762798] Call Trace: [ 3537.762799] <IRQ> [<ffffffff810924d4>] ? __report_bad_irq+0x3a/0xa2 [ 3537.762803] [<ffffffff810926a4>] ? note_interrupt+0x168/0x1da [ 3537.762805] [<ffffffff81090dd4>] ? handle_irq_event_percpu+0x171/0x18f [ 3537.762807] [<ffffffff8100e0e2>] ? read_tsc+0x5/0x16 [ 3537.762809] [<ffffffff8106b8a2>] ? update_ts_time_stats+0x32/0x6b [ 3537.762810] [<ffffffff81090e26>] ? handle_irq_event+0x34/0x52 [ 3537.762812] [<ffffffff81063fb7>] ? sched_clock_idle_wakeup_event+0x12/0x1c [ 3537.762813] [<ffffffff81092df2>] ? handle_fasteoi_irq+0x82/0xa4 [ 3537.762815] [<ffffffff8100aadb>] ? handle_irq+0x1a/0x23 [ 3537.762816] [<ffffffff8100a384>] ? do_IRQ+0x45/0xaa [ 3537.762818] [<ffffffff81332c93>] ? common_interrupt+0x13/0x13 [ 3537.762818] <EOI> [<ffffffff81332c8e>] ? common_interrupt+0xe/0x13 [ 3537.762821] [<ffffffff81026800>] ? native_safe_halt+0x2/0x3 [ 3537.762829] [<ffffffffa016ed58>] ? acpi_idle_do_entry+0x39/0x62 [processor] [ 3537.762831] [<ffffffffa016edde>] ? acpi_idle_enter_c1+0x5d/0xad [processor] [ 3537.762834] [<ffffffff81261033>] ? cpuidle_idle_call+0x11f/0x1cc [ 3537.762835] [<ffffffff81008dd2>] ? cpu_idle+0xab/0xe1 [ 3537.762837] [<ffffffff8169fc60>] ? start_kernel+0x3e0/0x3eb [ 3537.762838] [<ffffffff8169f3c8>] ? x86_64_start_kernel+0x102/0x10f [ 3537.762839] handlers: [ 3537.762840] [<ffffffffa0358d5a>] (rtl8169_interrupt+0x0/0x2d7 [r8169]) [ 3537.762842] [<ffffffffa08ff2ca>] (nv_kern_isr+0x0/0x54 [nvidia]) [ 3537.762902] Disabling IRQ #16 After that Xorg either hogs on CPU or is unstable (up to hanging the system completely). When I restart Xorg everything is fine again and the problem doesn't occur until next reboot. I tried to upgrade the kernel from stock 2.6.32 to 2.6.39 from unstable repository but that didn't help. Booting with irqpoll option only seems to prolong the initial time period after which the problem occurs. I'm using latest NVIDIA drivers and Realtek firmware from firmware-realtek package. I have two GTX 560Ti that run in SLI. Disabling SLI or taking out one card completely doesn't solve the problem either. Output of uname -a is: Linux whitestar 2.6.39-2-amd64 #1 SMP Wed Jun 8 11:01:04 UTC 2011 x86_64 GNU/Linux Output of lspci is: 00:00.0 Host bridge: Intel Corporation Sandy Bridge DRAM Controller (rev 09) 00:01.0 PCI bridge: Intel Corporation Sandy Bridge PCI Express Root Port (rev 09) 00:01.1 PCI bridge: Intel Corporation Sandy Bridge PCI Express Root Port (rev 09) 00:16.0 Communication controller: Intel Corporation Cougar Point HECI Controller #1 (rev 04) 00:19.0 Ethernet controller: Intel Corporation 82579V Gigabit Network Connection (rev 05) 00:1a.0 USB Controller: Intel Corporation Cougar Point USB Enhanced Host Controller #2 (rev 05) 00:1b.0 Audio device: Intel Corporation Cougar Point High Definition Audio Controller (rev 05) 00:1c.0 PCI bridge: Intel Corporation Cougar Point PCI Express Root Port 1 (rev b5) 00:1c.1 PCI bridge: Intel Corporation Cougar Point PCI Express Root Port 2 (rev b5) 00:1c.2 PCI bridge: Intel Corporation Cougar Point PCI Express Root Port 3 (rev b5) 00:1c.4 PCI bridge: Intel Corporation Cougar Point PCI Express Root Port 5 (rev b5) 00:1c.6 PCI bridge: Intel Corporation 82801 PCI Bridge (rev b5) 00:1d.0 USB Controller: Intel Corporation Cougar Point USB Enhanced Host Controller #1 (rev 05) 00:1f.0 ISA bridge: Intel Corporation Cougar Point LPC Controller (rev 05) 00:1f.2 SATA controller: Intel Corporation Cougar Point 6 port SATA AHCI Controller (rev 05) 00:1f.3 SMBus: Intel Corporation Cougar Point SMBus Controller (rev 05) 01:00.0 VGA compatible controller: nVidia Corporation Device 1200 (rev a1) 01:00.1 Audio device: nVidia Corporation Device 0e0c (rev a1) 02:00.0 VGA compatible controller: nVidia Corporation Device 1200 (rev a1) 02:00.1 Audio device: nVidia Corporation Device 0e0c (rev a1) 04:00.0 USB Controller: NEC Corporation uPD720200 USB 3.0 Host Controller (rev 04) 06:00.0 USB Controller: NEC Corporation uPD720200 USB 3.0 Host Controller (rev 04) 07:00.0 PCI bridge: Device 1b21:1080 (rev 01) 08:02.0 Ethernet controller: Realtek Semiconductor Co., Ltd. RTL-8110SC/8169SC Gigabit Ethernet (rev 10) 08:03.0 FireWire (IEEE 1394): VIA Technologies, Inc. VT6306/7/8 [Fire II(M)] IEEE 1394 OHCI Controller (rev c0) Contents of /proc/interrupts: CPU0 CPU1 CPU2 CPU3 CPU4 CPU5 CPU6 CPU7 0: 77 0 0 0 0 0 0 0 IO-APIC-edge timer 1: 2 0 0 0 0 0 0 0 IO-APIC-edge i8042 8: 1 0 0 0 0 0 0 0 IO-APIC-edge rtc0 9: 0 0 0 0 0 0 0 0 IO-APIC-fasteoi acpi 12: 4 0 0 0 0 0 0 0 IO-APIC-edge i8042 16: 699083 0 0 0 0 0 0 0 IO-APIC-fasteoi nvidia, eth0 17: 87810 0 0 0 0 0 0 0 IO-APIC-fasteoi firewire_ohci, hda_intel, nvidia 18: 242 0 0 0 0 0 0 0 IO-APIC-fasteoi hda_intel 23: 85925 0 0 0 0 0 0 0 IO-APIC-fasteoi ehci_hcd:usb5, ehci_hcd:usb6 40: 0 0 0 0 0 0 0 0 PCI-MSI-edge PCIe PME 41: 0 0 0 0 0 0 0 0 PCI-MSI-edge PCIe PME 42: 0 0 0 0 0 0 0 0 PCI-MSI-edge PCIe PME 43: 0 0 0 0 0 0 0 0 PCI-MSI-edge PCIe PME 44: 0 0 0 0 0 0 0 0 PCI-MSI-edge PCIe PME 45: 0 0 0 0 0 0 0 0 PCI-MSI-edge PCIe PME 46: 79853 0 0 0 0 0 0 0 PCI-MSI-edge ahci 48: 1 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 49: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 50: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 51: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 52: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 53: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 54: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 55: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 56: 1 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 57: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 58: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 59: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 60: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 61: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 62: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 63: 0 0 0 0 0 0 0 0 PCI-MSI-edge xhci_hcd 64: 173506 0 0 0 0 0 0 0 PCI-MSI-edge hda_intel NMI: 482 89 25 13 277 24 11 10 Non-maskable interrupts LOC: 783857 194752 114133 70577 372438 179065 117179 162016 Local timer interrupts SPU: 0 0 0 0 0 0 0 0 Spurious interrupts PMI: 482 89 25 13 277 24 11 10 Performance monitoring interrupts IWI: 0 0 0 0 0 0 0 0 IRQ work interrupts RES: 131917 46750 7432 3291 150003 9576 3435 3067 Rescheduling interrupts CAL: 2759 6563 7150 6997 5387 7140 7269 6678 Function call interrupts TLB: 4396 2038 1336 492 5434 1896 1121 606 TLB shootdowns TRM: 0 0 0 0 0 0 0 0 Thermal event interrupts THR: 0 0 0 0 0 0 0 0 Threshold APIC interrupts MCE: 0 0 0 0 0 0 0 0 Machine check exceptions MCP: 37 37 37 37 37 37 37 37 Machine check polls ERR: 0 MIS: 0 Last but not least, right after boot-up those lines are usually present in dmesg: [ 18.367094] hda-intel: IRQ timing workaround is activated for card #1. Suggest a bigger bdl_pos_adj. [ 18.458859] hda-intel: IRQ timing workaround is activated for card #2. Suggest a bigger bdl_pos_adj. I'm not sure if it's related or a symptom of a bigger problem so I'm posting it just in case. I don't really know what other information might be of relevance here. Don't hesitate to ask for more in the comments.

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  • Does anyone really understand how HFSC scheduling in Linux/BSD works?

    - by Mecki
    I read the original SIGCOMM '97 PostScript paper about HFSC, it is very technically, but I understand the basic concept. Instead of giving a linear service curve (as with pretty much every other scheduling algorithm), you can specify a convex or concave service curve and thus it is possible to decouple bandwidth and delay. However, even though this paper mentions to kind of scheduling algorithms being used (real-time and link-share), it always only mentions ONE curve per scheduling class (the decoupling is done by specifying this curve, only one curve is needed for that). Now HFSC has been implemented for BSD (OpenBSD, FreeBSD, etc.) using the ALTQ scheduling framework and it has been implemented Linux using the TC scheduling framework (part of iproute2). Both implementations added two additional service curves, that were NOT in the original paper! A real-time service curve and an upper-limit service curve. Again, please note that the original paper mentions two scheduling algorithms (real-time and link-share), but in that paper both work with one single service curve. There never have been two independent service curves for either one as you currently find in BSD and Linux. Even worse, some version of ALTQ seems to add an additional queue priority to HSFC (there is no such thing as priority in the original paper either). I found several BSD HowTo's mentioning this priority setting (even though the man page of the latest ALTQ release knows no such parameter for HSFC, so officially it does not even exist). This all makes the HFSC scheduling even more complex than the algorithm described in the original paper and there are tons of tutorials on the Internet that often contradict each other, one claiming the opposite of the other one. This is probably the main reason why nobody really seems to understand how HFSC scheduling really works. Before I can ask my questions, we need a sample setup of some kind. I'll use a very simple one as seen in the image below: Here are some questions I cannot answer because the tutorials contradict each other: What for do I need a real-time curve at all? Assuming A1, A2, B1, B2 are all 128 kbit/s link-share (no real-time curve for either one), then each of those will get 128 kbit/s if the root has 512 kbit/s to distribute (and A and B are both 256 kbit/s of course), right? Why would I additionally give A1 and B1 a real-time curve with 128 kbit/s? What would this be good for? To give those two a higher priority? According to original paper I can give them a higher priority by using a curve, that's what HFSC is all about after all. By giving both classes a curve of [256kbit/s 20ms 128kbit/s] both have twice the priority than A2 and B2 automatically (still only getting 128 kbit/s on average) Does the real-time bandwidth count towards the link-share bandwidth? E.g. if A1 and B1 both only have 64kbit/s real-time and 64kbit/s link-share bandwidth, does that mean once they are served 64kbit/s via real-time, their link-share requirement is satisfied as well (they might get excess bandwidth, but lets ignore that for a second) or does that mean they get another 64 kbit/s via link-share? So does each class has a bandwidth "requirement" of real-time plus link-share? Or does a class only have a higher requirement than the real-time curve if the link-share curve is higher than the real-time curve (current link-share requirement equals specified link-share requirement minus real-time bandwidth already provided to this class)? Is upper limit curve applied to real-time as well, only to link-share, or maybe to both? Some tutorials say one way, some say the other way. Some even claim upper-limit is the maximum for real-time bandwidth + link-share bandwidth? What is the truth? Assuming A2 and B2 are both 128 kbit/s, does it make any difference if A1 and B1 are 128 kbit/s link-share only, or 64 kbit/s real-time and 128 kbit/s link-share, and if so, what difference? If I use the seperate real-time curve to increase priorities of classes, why would I need "curves" at all? Why is not real-time a flat value and link-share also a flat value? Why are both curves? The need for curves is clear in the original paper, because there is only one attribute of that kind per class. But now, having three attributes (real-time, link-share, and upper-limit) what for do I still need curves on each one? Why would I want the curves shape (not average bandwidth, but their slopes) to be different for real-time and link-share traffic? According to the little documentation available, real-time curve values are totally ignored for inner classes (class A and B), they are only applied to leaf classes (A1, A2, B1, B2). If that is true, why does the ALTQ HFSC sample configuration (search for 3.3 Sample configuration) set real-time curves on inner classes and claims that those set the guaranteed rate of those inner classes? Isn't that completely pointless? (note: pshare sets the link-share curve in ALTQ and grate the real-time curve; you can see this in the paragraph above the sample configuration). Some tutorials say the sum of all real-time curves may not be higher than 80% of the line speed, others say it must not be higher than 70% of the line speed. Which one is right or are they maybe both wrong? One tutorial said you shall forget all the theory. No matter how things really work (schedulers and bandwidth distribution), imagine the three curves according to the following "simplified mind model": real-time is the guaranteed bandwidth that this class will always get. link-share is the bandwidth that this class wants to become fully satisfied, but satisfaction cannot be guaranteed. In case there is excess bandwidth, the class might even get offered more bandwidth than necessary to become satisfied, but it may never use more than upper-limit says. For all this to work, the sum of all real-time bandwidths may not be above xx% of the line speed (see question above, the percentage varies). Question: Is this more or less accurate or a total misunderstanding of HSFC? And if assumption above is really accurate, where is prioritization in that model? E.g. every class might have a real-time bandwidth (guaranteed), a link-share bandwidth (not guaranteed) and an maybe an upper-limit, but still some classes have higher priority needs than other classes. In that case I must still prioritize somehow, even among real-time traffic of those classes. Would I prioritize by the slope of the curves? And if so, which curve? The real-time curve? The link-share curve? The upper-limit curve? All of them? Would I give all of them the same slope or each a different one and how to find out the right slope? I still haven't lost hope that there exists at least a hand full of people in this world that really understood HFSC and are able to answer all these questions accurately. And doing so without contradicting each other in the answers would be really nice ;-)

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  • ProFTPd server on Ubuntu getting access denied message when successfully authenticated?

    - by exxoid
    I have a Ubuntu box with a ProFTPD 1.3.4a Server, when I try to log in via my FTP Client I cannot do anything as it does not allow me to list directories; I have tried logging in as root and as a regular user and tried accessing different paths within the FTP Server. The error I get in my FTP Client is: Status: Retrieving directory listing... Command: CDUP Response: 250 CDUP command successful Command: PWD Response: 257 "/var" is the current directory Command: PASV Response: 227 Entering Passive Mode (172,16,4,22,237,205). Command: MLSD Response: 550 Access is denied. Error: Failed to retrieve directory listing Any idea? Here is the config of my proftpd: # # /etc/proftpd/proftpd.conf -- This is a basic ProFTPD configuration file. # To really apply changes, reload proftpd after modifications, if # it runs in daemon mode. It is not required in inetd/xinetd mode. # # Includes DSO modules Include /etc/proftpd/modules.conf # Set off to disable IPv6 support which is annoying on IPv4 only boxes. UseIPv6 off # If set on you can experience a longer connection delay in many cases. IdentLookups off ServerName "Drupal Intranet" ServerType standalone ServerIdent on "FTP Server ready" DeferWelcome on # Set the user and group that the server runs as User nobody Group nogroup MultilineRFC2228 on DefaultServer on ShowSymlinks on TimeoutNoTransfer 600 TimeoutStalled 600 TimeoutIdle 1200 DisplayLogin welcome.msg DisplayChdir .message true ListOptions "-l" DenyFilter \*.*/ # Use this to jail all users in their homes # DefaultRoot ~ # Users require a valid shell listed in /etc/shells to login. # Use this directive to release that constrain. # RequireValidShell off # Port 21 is the standard FTP port. Port 21 # In some cases you have to specify passive ports range to by-pass # firewall limitations. Ephemeral ports can be used for that, but # feel free to use a more narrow range. # PassivePorts 49152 65534 # If your host was NATted, this option is useful in order to # allow passive tranfers to work. You have to use your public # address and opening the passive ports used on your firewall as well. # MasqueradeAddress 1.2.3.4 # This is useful for masquerading address with dynamic IPs: # refresh any configured MasqueradeAddress directives every 8 hours <IfModule mod_dynmasq.c> # DynMasqRefresh 28800 </IfModule> # To prevent DoS attacks, set the maximum number of child processes # to 30. If you need to allow more than 30 concurrent connections # at once, simply increase this value. Note that this ONLY works # in standalone mode, in inetd mode you should use an inetd server # that allows you to limit maximum number of processes per service # (such as xinetd) MaxInstances 30 # Set the user and group that the server normally runs at. # Umask 022 is a good standard umask to prevent new files and dirs # (second parm) from being group and world writable. Umask 022 022 # Normally, we want files to be overwriteable. AllowOverwrite on # Uncomment this if you are using NIS or LDAP via NSS to retrieve passwords: # PersistentPasswd off # This is required to use both PAM-based authentication and local passwords AuthPAMConfig proftpd AuthOrder mod_auth_pam.c* mod_auth_unix.c # Be warned: use of this directive impacts CPU average load! # Uncomment this if you like to see progress and transfer rate with ftpwho # in downloads. That is not needed for uploads rates. # UseSendFile off TransferLog /var/log/proftpd/xferlog SystemLog /var/log/proftpd/proftpd.log # Logging onto /var/log/lastlog is enabled but set to off by default #UseLastlog on # In order to keep log file dates consistent after chroot, use timezone info # from /etc/localtime. If this is not set, and proftpd is configured to # chroot (e.g. DefaultRoot or <Anonymous>), it will use the non-daylight # savings timezone regardless of whether DST is in effect. #SetEnv TZ :/etc/localtime <IfModule mod_quotatab.c> QuotaEngine off </IfModule> <IfModule mod_ratio.c> Ratios off </IfModule> # Delay engine reduces impact of the so-called Timing Attack described in # http://www.securityfocus.com/bid/11430/discuss # It is on by default. <IfModule mod_delay.c> DelayEngine on </IfModule> <IfModule mod_ctrls.c> ControlsEngine off ControlsMaxClients 2 ControlsLog /var/log/proftpd/controls.log ControlsInterval 5 ControlsSocket /var/run/proftpd/proftpd.sock </IfModule> <IfModule mod_ctrls_admin.c> AdminControlsEngine off </IfModule> # # Alternative authentication frameworks # #Include /etc/proftpd/ldap.conf #Include /etc/proftpd/sql.conf # # This is used for FTPS connections # #Include /etc/proftpd/tls.conf # # Useful to keep VirtualHost/VirtualRoot directives separated # #Include /etc/proftpd/virtuals.con # A basic anonymous configuration, no upload directories. # <Anonymous ~ftp> # User ftp # Group nogroup # # We want clients to be able to login with "anonymous" as well as "ftp" # UserAlias anonymous ftp # # Cosmetic changes, all files belongs to ftp user # DirFakeUser on ftp # DirFakeGroup on ftp # # RequireValidShell off # # # Limit the maximum number of anonymous logins # MaxClients 10 # # # We want 'welcome.msg' displayed at login, and '.message' displayed # # in each newly chdired directory. # DisplayLogin welcome.msg # DisplayChdir .message # # # Limit WRITE everywhere in the anonymous chroot # <Directory *> # <Limit WRITE> # DenyAll # </Limit> # </Directory> # # # Uncomment this if you're brave. # # <Directory incoming> # # # Umask 022 is a good standard umask to prevent new files and dirs # # # (second parm) from being group and world writable. # # Umask 022 022 # # <Limit READ WRITE> # # DenyAll # # </Limit> # # <Limit STOR> # # AllowAll # # </Limit> # # </Directory> # # </Anonymous> # Include other custom configuration files Include /etc/proftpd/conf.d/ UseReverseDNS off <Global> RootLogin on UseFtpUsers on ServerIdent on DefaultChdir /var/www DeleteAbortedStores on LoginPasswordPrompt on AccessGrantMsg "You have been authenticated successfully." </Global> Any idea what could be wrong? Thanks for your help!

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  • Fibre channel long distance woes

    - by Marki
    I need a fresh pair of eyes. We're using a 15km fibre optic line across which fibrechannel and 10GbE is multiplexed (passive optical CWDM). For FC we have long distance lasers suitable up to 40km (Skylane SFCxx0404F0D). The multiplexer is limited by the SFPs which can do max. 4Gb fibrechannel. The FC switch is a Brocade 5000 series. The respective wavelengths are 1550,1570,1590 and 1610nm for FC and 1530nm for 10GbE. The problem is the 4GbFC fabrics are almost never clean. Sometimes they are for a while even with a lot of traffic on them. Then they may suddenly start producing errors (RX CRC, RX encoding, RX disparity, ...) even with only marginal traffic on them. I am attaching some error and traffic graphs. Errors are currently in the order of 50-100 errors per 5 minutes when with 1Gb/s traffic. Optics Here is the power output of one port summarized (collected using sfpshow on different switches) SITE-A units=uW (microwatt) SITE-B ********************************************** FAB1 SW1 TX 1234.3 RX 49.1 SW3 1550nm (ko) RX 95.2 TX 1175.6 FAB2 SW2 TX 1422.0 RX 104.6 SW4 1610nm (ok) RX 54.3 TX 1468.4 What I find curious at this point is the asymmetry in the power levels. While SW2 transmits with 1422uW which SW4 receives with 104uW, SW2 only receives the SW4 signal with similar original power only with 54uW. Vice versa for SW1-3. Anyway the SFPs have RX sensitivity down to -18dBm (ca. 20uW) so in any case it should be fine... But nothing is. Some SFPs have been diagnosed as malfunctioning by the manufacturer (the 1550nm ones shown above with "ko"). The 1610nm ones apparently are ok, they have been tested using a traffic generator. The leased line has also been tested more than once. All is within tolerances. I'm awaiting the replacements but for some reason I don't believe it will make things better as the apparently good ones don't produce ZERO errors either. Earlier there was active equipment involved (some kind of 4GFC retimer) before putting the signal on the line. No idea why. That equipment was eliminated because of the problems so we now only have: the long distance laser in the switch, (new) 10m LC-SC monomode cable to the mux (for each fabric), the leased line, the same thing but reversed on the other side of the link. FC switches Here is a port config from the Brocade portcfgshow (it's like that on both sides, obviously) Area Number: 0 Speed Level: 4G Fill Word(On Active) 0(Idle-Idle) Fill Word(Current) 0(Idle-Idle) AL_PA Offset 13: OFF Trunk Port ON Long Distance LS VC Link Init OFF Desired Distance 32 Km Reserved Buffers 70 Locked L_Port OFF Locked G_Port OFF Disabled E_Port OFF Locked E_Port OFF ISL R_RDY Mode OFF RSCN Suppressed OFF Persistent Disable OFF LOS TOV enable OFF NPIV capability ON QOS E_Port OFF Port Auto Disable: OFF Rate Limit OFF EX Port OFF Mirror Port OFF Credit Recovery ON F_Port Buffers OFF Fault Delay: 0(R_A_TOV) NPIV PP Limit: 126 CSCTL mode: OFF Forcing the links to 2GbFC produces no errors, but we bought 4GbFC and we want 4GbFC. I don't know where to look anymore. Any ideas what to try next or how to proceed? If we can't make 4GbFC work reliably I wonder what the people working with 8 or 16 do... I don't assume that "a few errors here and there" are acceptable. Oh and BTW we are in contact with everyone of the manufacturers (FC switch, MUX, SFPs, ...) Except for the SFPs to be changed (some have been changed before) nobody has a clue. Brocade SAN Health says the fabric is ok. MUX, well, it's passive, it's only a prism, nature at it's best. Any shots in the dark? APPENDIX: Answers to your questions @Chopper3: This is the second generation of Brocades exhibiting the problem. Before we had 5000s, now we have 5100s. In the beginning when we still had the active MUX we rented a longdistance laser once to put it into the switch directly in order to make tests for a day, during that day of course it was clean. But as I said, sometimes it's clean just like that. And sometimes it's not. Alternative switches would mean to rebuild the entire SAN with those only to test. Alternative SFPs, well they're hard to come by just like that. @longneck: The line is rented. It's a dark fibre (9um monomode) so there's noone else on it. Sure there are splices. I can't go and look but I have to trust they have been done correctly. As I said the line has been checked and rechecked (using an optical time-domain reflectometer). Obviously you don't have all this equipment yourself because it's way too expensive. @mdpc: What would be the "wrong" type of cable according to you? Up to the switch everything is monomode, yes. The connectors are the correct ones too. Yeah I know there are the green ones where the fibre is cut off at a certain angle etc. But we have the correct ones for all that I know. Progress Report #1 We have had two fabrics (=2x2 switches) with Brocade 5100s with FabricOS 6.4.1 and two fabrics (another 2x4 switches) on FabricOS 7.0.2. On the longdistance ISLs (one in each fabric) it turned out that with FOS 6.4.1 setting it to long distance issues warnings about the VC Init setting and consequently the fill word. But those are only warnings. FOS 7.0.2 requires you to do modifications to VCI and the fillword for long distance links. Setting FOS 6.4.1 to the LS (long-distance static distance) setting with wrong VCI and fillword setting made the whole fabric inoperational (stuck in an SCN loop, use fabriclog -s to see, you don't see it anywhere else, no port error counters or anything increasing). Currently I'm giving the one fabric with the IMHO more correct settings a beating and it seems to do fine, whereas the other one without much traffic still has errors here and there. In short: We have eliminated the active part of the MUX (the FC retimer). We are putting the long distance SFPs into the end equipment themselves. Just to be sure we bought new monomode cables to connect the end equipment to the remaining passive part of the MUX. We are now trying out several long distance configs. It's almost black magic. Everything that happens is mostly empirical, noone seems to have a clue what are the exact reasons to do something. ("We have tried this, and it didn't work, then we tried that and it worked, so we stuck with that." But noone really seems to know why.) I'll keep you updated. Progress Report #2 We got the new lasers for one of the fabrics on warranty. It's ultra clean even on 4GbFC. They're transmitting with roughly 2mW (3dBm) whereas the others are only at 1.5mW (1.5dBm) although that should really be enough. The other fabric (where the lasers are apparently ok) still produces one or two CRCs infrequently. Using sfpshow the SFP producing the actual RX errors shows Status/Ctrl: 0x82 Alarm flags[0,1] = 0x5, 0x40 Warn Flags[0,1] = 0x5, 0x40 Now I'll have to find out what that means. Not sure if it was there before. Well I'll first clear my head with a week of vacation. 8-)

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  • pre-commit hook in svn: could not be translated from the native locale to UTF-8

    - by Alexandre Moraes
    Hi everybody, I have a problem with my pre-commit hook. This hook test if a file is locked when the user commits. When a bad condition happens, it should output that the another user is locking this file or if nobody is locking, it should show "you are not locking this file message (file´s name)". The error happens when the file´s name has some latin character like "ç" and tortoise show me this in the output. Commit failed (details follow): Commit blocked by pre-commit hook (exit code 1) with output: [Erro output could not be translated from the native locale to UTF-8.] Do you know how can I solve this? Thanks, Alexandre My shell script is here: #!/bin/sh REPOS="$1" TXN="$2" export LANG="en_US.UTF-8" /app/svn/hooks/ensure-has-need-lock.pl "$REPOS" "$TXN" if [ $? -ne 0 ]; then exit 1; fi exit 0 And my perl is here: !/usr/bin/env perl #Turn on warnings the best way depending on the Perl version. BEGIN { if ( $] >= 5.006_000) { require warnings; import warnings; } else { $^W = 1; } } use strict; use Carp; &usage unless @ARGV == 2; my $repos = shift; my $txn = shift; my $svnlook = "/usr/local/bin/svnlook"; my $user; my $ok = 1; foreach my $program ($svnlook) { if (-e $program) { unless (-x $program) { warn "$0: required program $program' is not executable, ", "edit $0.\n"; $ok = 0; } } else { warn "$0: required program $program' does not exist, edit $0.\n"; $ok = 0; } } exit 1 unless $ok; unless (-e $repos){ &usage("$0: repository directory $repos' does not exist."); } unless (-d $repos){ &usage("$0: repository directory $repos' is not a directory."); } foreach my $user_tmp (&read_from_process($svnlook, 'author', $repos, '-t', $txn)) { $user = $user_tmp; } my @errors; foreach my $transaction (&read_from_process($svnlook, 'changed', $repos, '-t', $txn)){ if ($transaction =~ /^U. (.*[^\/])$/){ my $file = $1; my $err = 0; foreach my $locks (&read_from_process($svnlook, 'lock', $repos, $file)){ $err = 1; if($locks=~ /Owner: (.*)/){ if($1 != $user){ push @errors, "$file : You are not locking this file!"; } } } if($err==0){ push @errors, "$file : You are not locking this file!"; } } elsif($transaction =~ /^D. (.*[^\/])$/){ my $file = $1; my $tchan = &read_from_process($svnlook, 'lock', $repos, $file); foreach my $locks (&read_from_process($svnlook, 'lock', $repos, $file)){ push @errors, "$1 : cannot delete locked Files"; } } elsif($transaction =~ /^A. (.*[^\/])$/){ my $needs_lock; my $path = $1; foreach my $prop (&read_from_process($svnlook, 'proplist', $repos, '-t', $txn, '--verbose', $path)){ if ($prop =~ /^\s*svn:needs-lock : (\S+)/){ $needs_lock = $1; } } if (not $needs_lock){ push @errors, "$path : svn:needs-lock is not set. Pleas ask TCC for support."; } } } if (@errors) { warn "$0:\n\n", join("\n", @errors), "\n\n"; exit 1; } else { exit 0; } sub usage { warn "@_\n" if @_; die "usage: $0 REPOS TXN-NAME\n"; } sub safe_read_from_pipe { unless (@_) { croak "$0: safe_read_from_pipe passed no arguments.\n"; } print "Running @_\n"; my $pid = open(SAFE_READ, '-|'); unless (defined $pid) { die "$0: cannot fork: $!\n"; } unless ($pid) { open(STDERR, ">&STDOUT") or die "$0: cannot dup STDOUT: $!\n"; exec(@_) or die "$0: cannot exec @_': $!\n"; } my @output; while (<SAFE_READ>) { chomp; push(@output, $_); } close(SAFE_READ); my $result = $?; my $exit = $result >> 8; my $signal = $result & 127; my $cd = $result & 128 ? "with core dump" : ""; if ($signal or $cd) { warn "$0: pipe from @_' failed $cd: exit=$exit signal=$signal\n"; } if (wantarray) { return ($result, @output); } else { return $result; } } sub read_from_process { unless (@_) { croak "$0: read_from_process passed no arguments.\n"; } my ($status, @output) = &safe_read_from_pipe(@_); if ($status) { if (@output) { die "$0: @_' failed with this output:\n", join("\n", @output), "\n"; } else { die "$0: @_' failed with no output.\n"; } } else { return @output; } }

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  • Using WeakReference to resolve issue with .NET unregistered event handlers causing memory leaks.

    - by Eric
    The problem: Registered event handlers create a reference from the event to the event handler's instance. If that instance fails to unregister the event handler (via Dispose, presumably), then the instance memory will not be freed by the garbage collector. Example: class Foo { public event Action AnEvent; public void DoEvent() { if (AnEvent != null) AnEvent(); } } class Bar { public Bar(Foo l) { l.AnEvent += l_AnEvent; } void l_AnEvent() { } } If I instantiate a Foo, and pass this to a new Bar constructor, then let go of the Bar object, it will not be freed by the garbage collector because of the AnEvent registration. I consider this a memory leak, and seems just like my old C++ days. I can, of course, make Bar IDisposable, unregister the event in the Dispose() method, and make sure to call Dispose() on instances of it, but why should I have to do this? I first question why events are implemented with strong references? Why not use weak references? An event is used to abstractly notify an object of changes in another object. It seems to me that if the event handler's instance is no longer in use (i.e., there are no non-event references to the object), then any events that it is registered with should automatically be unregistered. What am I missing? I have looked at WeakEventManager. Wow, what a pain. Not only is it very difficult to use, but its documentation is inadequate (see http://msdn.microsoft.com/en-us/library/system.windows.weakeventmanager.aspx -- noticing the "Notes to Inheritors" section that has 6 vaguely described bullets). I have seen other discussions in various places, but nothing I felt I could use. I propose a simpler solution based on WeakReference, as described here. My question is: Does this not meet the requirements with significantly less complexity? To use the solution, the above code is modified as follows: class Foo { public WeakReferenceEvent AnEvent = new WeakReferenceEvent(); internal void DoEvent() { AnEvent.Invoke(); } } class Bar { public Bar(Foo l) { l.AnEvent += l_AnEvent; } void l_AnEvent() { } } Notice two things: 1. The Foo class is modified in two ways: The event is replaced with an instance of WeakReferenceEvent, shown below; and the invocation of the event is changed. 2. The Bar class is UNCHANGED. No need to subclass WeakEventManager, implement IWeakEventListener, etc. OK, so on to the implementation of WeakReferenceEvent. This is shown here. Note that it uses the generic WeakReference that I borrowed from here: http://damieng.com/blog/2006/08/01/implementingweakreferencet I had to add Equals() and GetHashCode() to his class, which I include below for reference. class WeakReferenceEvent { public static WeakReferenceEvent operator +(WeakReferenceEvent wre, Action handler) { wre._delegates.Add(new WeakReference<Action>(handler)); return wre; } public static WeakReferenceEvent operator -(WeakReferenceEvent wre, Action handler) { foreach (var del in wre._delegates) if (del.Target == handler) { wre._delegates.Remove(del); return wre; } return wre; } HashSet<WeakReference<Action>> _delegates = new HashSet<WeakReference<Action>>(); internal void Invoke() { HashSet<WeakReference<Action>> toRemove = null; foreach (var del in _delegates) { if (del.IsAlive) del.Target(); else { if (toRemove == null) toRemove = new HashSet<WeakReference<Action>>(); toRemove.Add(del); } } if (toRemove != null) foreach (var del in toRemove) _delegates.Remove(del); } } public class WeakReference<T> : IDisposable { private GCHandle handle; private bool trackResurrection; public WeakReference(T target) : this(target, false) { } public WeakReference(T target, bool trackResurrection) { this.trackResurrection = trackResurrection; this.Target = target; } ~WeakReference() { Dispose(); } public void Dispose() { handle.Free(); GC.SuppressFinalize(this); } public virtual bool IsAlive { get { return (handle.Target != null); } } public virtual bool TrackResurrection { get { return this.trackResurrection; } } public virtual T Target { get { object o = handle.Target; if ((o == null) || (!(o is T))) return default(T); else return (T)o; } set { handle = GCHandle.Alloc(value, this.trackResurrection ? GCHandleType.WeakTrackResurrection : GCHandleType.Weak); } } public override bool Equals(object obj) { var other = obj as WeakReference<T>; return other != null && Target.Equals(other.Target); } public override int GetHashCode() { return Target.GetHashCode(); } } It's functionality is trivial. I override operator + and - to get the += and -= syntactic sugar matching events. These create WeakReferences to the Action delegate. This allows the garbage collector to free the event target object (Bar in this example) when nobody else is holding on to it. In the Invoke() method, simply run through the weak references and call their Target Action. If any dead (i.e., garbage collected) references are found, remove them from the list. Of course, this only works with delegates of type Action. I tried making this generic, but ran into the missing where T : delegate in C#! As an alternative, simply modify class WeakReferenceEvent to be a WeakReferenceEvent, and replace the Action with Action. Fix the compiler errors and you have a class that can be used like so: class Foo { public WeakReferenceEvent<int> AnEvent = new WeakReferenceEvent<int>(); internal void DoEvent() { AnEvent.Invoke(5); } } Hopefully this will help someone else when they run into the mystery .NET event memory leak!

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  • Job conditions conflicting with personal principles on software-development - how much is too much?

    - by Baelnorn
    Sorry for the incoming wall'o'text (and for my probably bad English) but I just need to get this off somehow. I also accept that this question will be probably closed as subjective and argumentative, but I need to know one thing: "how much BS are programmers supposed to put up with before breaking?" My background I'm 27 years old and have a B.Sc. in Computer engineering with a graduation grade of 1.8 from a university of applied science. I went looking for a job right after graduation. I got three offers right away, with two offers paying vastly more than the last one, but that last one seemed more interesting so I went for that. My situation I've been working for the company now for 17 months now, but it feels like a drag more and more each day. Primarily because the company (which has only 5 other developers but me, and of these I work with 4) turned out to be pretty much the anti-thesis of what I expected (and was taught in university) from a modern software company. I agreed to accept less than half of the usual payment appropriate for my qualification for the first year because I was promised a trainee program. However, the trainee program turned out to be "here you got a computer, there's some links on the stuff we use, and now do what you colleagues tell you". Further, during my whole time there (trainee or not) I haven't been given the grace of even a single code-review - apparently nobody's interested in my work as long as it "just works". I was told in the job interview that "Microsoft technology played a central role in the company" yet I've been slowly eroding my congnitive functions with Flex/Actionscript/Cairngorm ever since I started (despite having applied as a C#/.NET developer). Actually, the company's primary projects are based on Java/XSLT and Flex/Actionscript (with some SAP/ABAP stuff here and there but I'm not involved in that) and they've been working on these before I even applied. Having had no experience either with that particular technology nor the framework nor the field (RIA) nor in developing business scale applications I obviously made several mistakes. However, my boss told me that he let me make those mistakes (which ate at least 2 months of development time on their own) on purpose to provide some "learning experience". Even when I was still a trainee I was already tasked with working on a business-critical application. On my own. Without supervision. Without code-reviews. My boss thinks agile methods are a waste of time/money and deems putting more than one developer on any project not efficient. Documentation is not necessary and each developer should only document what he himself needs for his work. Recently he wanted us to do bug tracking with Excel and Email instead of using an already existing Bugzilla, overriding an unanimous decision made by all developers and testers involved in the process - only after another senior developer had another hour-long private discussion with him he agreed to let us use the bugtracker. Project management is basically not present, there are only a few Excel sheets floating around where the senior developer lists some things (not all, mind you) with a time estimate ranging from days to months, trying to at least somehow organize the whole mess. A development process is also basically not present, each developer just works on his own however he wants. There are not even coding conventions in the company. Testing is done manually with a single tester (sometimes two testers) per project because automated testing wasn't given the least thought when the whole project was started. I guess it's not a big surprise when I say that each developer also has his own share of hundreds of overhours (which are, of course, unpaid). Each developer is tasked with working on his own project(s) which in turn leads to a very extensive knowledge monopolization - if one developer was to have an accident or become ill there would be absolutely no one who could even hope to do his work. Considering that each developer has his own business-critical application to work on, I guess that's a pretty bad situation. I've been trying to change things for the better. I tried to introduce a development process, but my first attempt was pretty much shot down by my boss with "I don't want to discuss agile methods". After that I put together a process that at least resembled how most of the developers were already working and then include stuff like automated (or at least organized) testing, coding conventions, etc. However, this was also shot down because it wasn't "simple" enought to be shown on a business slide (actually, I wasn't even given the 15 minutes I'd have needed to present the process in the meeting). My problem I can't stand working there any longer. Seriously, I consider to resign on monday, which still leaves me with 3 months to work there due to the cancelation period. My primary goal since I started studying computer science was being a good computer scientist, working with modern technologies and adhering to modern and proven principles and methods. However, the company I'm working for seems to make that impossible. Some days I feel as if was living in a perverted real-life version of the Dilbert comics. My question Am I overreacting? Is this the reality each graduate from university has to face? Should I betray my sound principles and just accept these working conditions? Or should I gtfo of there? What's the opinion of other developers on this matter. Would you put up with all that stuff?

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  • Modelling boost::Lockable with semaphore rather than mutex (previously titled: Unlocking a mutex fr

    - by dan
    I'm using the C++ boost::thread library, which in my case means I'm using pthreads. Officially, a mutex must be unlocked from the same thread which locks it, and I want the effect of being able to lock in one thread and then unlock in another. There are many ways to accomplish this. One possibility would be to write a new mutex class which allows this behavior. For example: class inter_thread_mutex{ bool locked; boost::mutex mx; boost::condition_variable cv; public: void lock(){ boost::unique_lock<boost::mutex> lck(mx); while(locked) cv.wait(lck); locked=true; } void unlock(){ { boost::lock_guard<boost::mutex> lck(mx); if(!locked) error(); locked=false; } cv.notify_one(); } // bool try_lock(); void error(); etc. } I should point out that the above code doesn't guarantee FIFO access, since if one thread calls lock() while another calls unlock(), this first thread may acquire the lock ahead of other threads which are waiting. (Come to think of it, the boost::thread documentation doesn't appear to make any explicit scheduling guarantees for either mutexes or condition variables). But let's just ignore that (and any other bugs) for now. My question is, if I decide to go this route, would I be able to use such a mutex as a model for the boost Lockable concept. For example, would anything go wrong if I use a boost::unique_lock< inter_thread_mutex for RAII-style access, and then pass this lock to boost::condition_variable_any.wait(), etc. On one hand I don't see why not. On the other hand, "I don't see why not" is usually a very bad way of determining whether something will work. The reason I ask is that if it turns out that I have to write wrapper classes for RAII locks and condition variables and whatever else, then I'd rather just find some other way to achieve the same effect. EDIT: The kind of behavior I want is basically as follows. I have an object, and it needs to be locked whenever it is modified. I want to lock the object from one thread, and do some work on it. Then I want to keep the object locked while I tell another worker thread to complete the work. So the first thread can go on and do something else while the worker thread finishes up. When the worker thread gets done, it unlocks the mutex. And I want the transition to be seemless so nobody else can get the mutex lock in between when thread 1 starts the work and thread 2 completes it. Something like inter_thread_mutex seems like it would work, and it would also allow the program to interact with it as if it were an ordinary mutex. So it seems like a clean solution. If there's a better solution, I'd be happy to hear that also. EDIT AGAIN: The reason I need locks to begin with is that there are multiple master threads, and the locks are there to prevent them from accessing shared objects concurrently in invalid ways. So the code already uses loop-level lock-free sequencing of operations at the master thread level. Also, in the original implementation, there were no worker threads, and the mutexes were ordinary kosher mutexes. The inter_thread_thingy came up as an optimization, primarily to improve response time. In many cases, it was sufficient to guarantee that the "first part" of operation A, occurs before the "first part" of operation B. As a dumb example, say I punch object 1 and give it a black eye. Then I tell object 1 to change it's internal structure to reflect all the tissue damage. I don't want to wait around for the tissue damage before I move on to punch object 2. However, I do want the tissue damage to occur as part of the same operation; for example, in the interim, I don't want any other thread to reconfigure the object in such a way that would make tissue damage an invalid operation. (yes, this example is imperfect in many ways, and no I'm not working on a game) So we made the change to a model where ownership of an object can be passed to a worker thread to complete an operation, and it actually works quite nicely; each master thread is able to get a lot more operations done because it doesn't need to wait for them all to complete. And, since the event sequencing at the master thread level is still loop-based, it is easy to write high-level master-thread operations, as they can be based on the assumption that an operation is complete when the corresponding function call returns. Finally, I thought it would be nice to use inter_thread mutex/semaphore thingies using RAII with boost locks to encapsulate the necessary synchronization that is required to make the whole thing work.

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  • Polymorphic associations in CakePHP2

    - by Joseph
    I have 3 models, Page , Course and Content Page and Course contain meta data and Content contains HTML content. Page and Course both hasMany Content Content belongsTo Page and Course To avoid having page_id and course_id fields in Content (because I want this to scale to more than just 2 models) I am looking at using Polymorphic Associations. I started by using the Polymorphic Behavior in the Bakery but it is generating waaay too many SQL queries for my liking and it's also throwing an "Illegal Offset" error which I don't know how to fix (it was written in 2008 and nobody seems to have referred to it recently so perhaps the error is due to it not having been designed for Cake 2?) Anyway, I've found that I can almost do everything I need by hardcoding the associations in the models as such: Page Model CREATE TABLE `pages` ( `id` int(11) NOT NULL AUTO_INCREMENT, `title` varchar(255) COLLATE utf8_unicode_ci NOT NULL, `slug` varchar(255) COLLATE utf8_unicode_ci NOT NULL, `created` datetime NOT NULL, `updated` datetime NOT NULL, PRIMARY KEY (`id`) ) <?php class Page extends AppModel { var $name = 'Page'; var $hasMany = array( 'Content' => array( 'className' => 'Content', 'foreignKey' => 'foreign_id', 'conditions' => array('Content.class' => 'Page'), ) ); } ?> Course Model CREATE TABLE `courses` ( `id` int(11) NOT NULL AUTO_INCREMENT, `title` varchar(255) COLLATE utf8_unicode_ci NOT NULL, `slug` varchar(255) COLLATE utf8_unicode_ci NOT NULL, `created` datetime NOT NULL, `updated` datetime NOT NULL, PRIMARY KEY (`id`) ) <?php class Course extends AppModel { var $name = 'Course'; var $hasMany = array( 'Content' => array( 'className' => 'Content', 'foreignKey' => 'foreign_id', 'conditions' => array('Content.class' => 'Course'), ) ); } ?> Content model CREATE TABLE IF NOT EXISTS `contents` ( `id` int(11) unsigned NOT NULL AUTO_INCREMENT, `class` varchar(30) COLLATE utf8_unicode_ci NOT NULL, `foreign_id` int(11) unsigned NOT NULL, `title` varchar(100) COLLATE utf8_unicode_ci NOT NULL, `content` text COLLATE utf8_unicode_ci NOT NULL, `created` datetime DEFAULT NULL, `modified` datetime DEFAULT NULL, PRIMARY KEY (`id`) ) <?php class Content extends AppModel { var $name = 'Content'; var $belongsTo = array( 'Page' => array( 'foreignKey' => 'foreign_id', 'conditions' => array('Content.class' => 'Page') ), 'Course' => array( 'foreignKey' => 'foreign_id', 'conditions' => array('Content.class' => 'Course') ) ); } ?> The good thing is that $this->Content->find('first') only generates a single SQL query instead of 3 (as was the case with the Polymorphic Behavior) but the problem is that the dataset returned includes both of the belongsTo models, whereas it should only really return the one that exists. Here's how the returned data looks: array( 'Content' => array( 'id' => '1', 'class' => 'Course', 'foreign_id' => '1', 'title' => 'something about this course', 'content' => 'The content here', 'created' => null, 'modified' => null ), 'Page' => array( 'id' => null, 'title' => null, 'slug' => null, 'created' => null, 'updated' => null ), 'Course' => array( 'id' => '1', 'title' => 'Course name', 'slug' => 'name-of-the-course', 'created' => '2012-10-11 00:00:00', 'updated' => '2012-10-11 00:00:00' ) ) I only want it to return one of either Page or Course depending on which one is specified in Content.class UPDATE: Combining the Page and Course models would seem like the obvious solution to this problem but the schemas I have shown above are just shown for the purpose of this question. The actual schemas are actually very different in terms of their fields and the each have a different number of associations with other models too. UPDATE 2 Here is the query that results from running $this->Content->find('first'); : SELECT `Content`.`id`, `Content`.`class`, `Content`.`foreign_id`, `Content`.`title`, `Content`.`slug`, `Content`.`content`, `Content`.`created`, `Content`.`modified`, `Page`.`id`, `Page`.`title`, `Page`.`slug`, `Page`.`created`, `Page`.`updated`, `Course`.`id`, `Course`.`title`, `Course`.`slug`, `Course`.`created`, `Course`.`updated` FROM `cakedb`.`contents` AS `Content` LEFT JOIN `cakedb`.`pages` AS `Page` ON (`Content`.`foreign_id` = `Page`.`id` AND `Content`.`class` = 'Page') LEFT JOIN `cakedb`.`courses` AS `Course` ON (`Content`.`foreign_id` = `Course`.`id` AND `Content`.`class` = 'Course') WHERE 1 = 1 LIMIT 1

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  • What Every Developer Should Know About MSI Components

    - by Alois Kraus
    Hopefully nothing. But if you have to do more than simple XCopy deployment and you need to support updates, upgrades and perhaps side by side scenarios there is no way around MSI. You can create Msi files with a Visual Studio Setup project which is severely limited or you can use the Windows Installer Toolset. I cannot talk about WIX with my German colleagues because WIX has a very special meaning. It is funny to always use the long name when I talk about deployment possibilities. Alternatively you can buy commercial tools which help you to author Msi files but I am not sure how good they are. Given enough pain with existing solutions you can also learn the MSI Apis and create your own packaging solution. If I were you I would use either a commercial visual tool when you do easy deployments or use the free Windows Installer Toolset. Once you know the WIX schema you can create well formed wix xml files easily with any editor. Then you can “compile” from the wxs files your Msi package. Recently I had the “pleasure” to get my hands dirty with C++ (again) and the MSI technology. Installation is a complex topic but after several month of digging into arcane MSI issues I can safely say that there should exist an easier way to install and update files as today. I am not alone with this statement as John Robbins (creator of the cool tool Paraffin) states: “.. It's a brittle and scary API in Windows …”. To help other people struggling with installation issues I present you the advice I (and others) found useful and what will happen if you ignore this advice. What is a MSI file? A MSI file is basically a database with tables which reference each other to control how your un/installation should work. The basic idea is that you declare via these tables what you want to install and MSI controls the how to get your stuff onto or off your machine. Your “stuff” consists usually of files, registry keys, shortcuts and environment variables. Therefore the most important tables are File, Registry, Environment and Shortcut table which define what will be un/installed. The key to master MSI is that every resource (file, registry key ,…) is associated with a MSI component. The actual payload consists of compressed files in the CAB format which can either be embedded into the MSI file or reside beside the MSI file or in a subdirectory below it. To examine MSI files you need Orca a free MSI editor provided by MS. There is also another free editor called Super Orca which does support diffs between MSI and it does not lock the MSI files. But since Orca comes with a shell extension I tend to use only Orca because it is so easy to right click on a MSI file and open it with this tool. How Do I Install It? Double click it. This does work for fresh installations as well as major upgrades. Updates need to be installed via the command line via msiexec /i <msi> REINSTALL=ALL REINSTALLMODE=vomus   This tells the installer to reinstall all already installed features (new features will NOT be installed). The reinstallmode letters do force an overwrite of the old cached package in the %WINDIR%\Installer folder. All files, shortcuts and registry keys are redeployed if they are missing or need to be replaced with a newer version. When things did go really wrong and you want to overwrite everything unconditionally use REINSTALLMODE=vamus. How To Enable MSI Logs? You can download a MSI from Microsoft which installs some registry keys to enable full MSI logging. The log files can be found in your %TEMP% folder and are called MSIxxxx.log. Alternatively you can add to your msiexec command line the option msiexec …. /l*vx <LogFileName> Personally I find it rather strange that * does not mean full logging. To really get all logs I need to add v and x which is documented in the msiexec help but I still find this behavior unintuitive. What are MSI components? The whole MSI logic is bound to the concept of MSI components. Nearly every msi table has a Component column which binds an installable resource to a component. Below are the screenshots of the FeatureComponents and Component table of an example MSI. The Feature table defines basically the feature hierarchy.  To find out what belongs to a feature you need to look at the FeatureComponents table where for each feature the components are listed which will be installed when a feature is installed. The MSI components are defined in the  Component table. This table has as first column the component name and as second column the component id which is a GUID. All resources you want to install belong to a MSI component. Therefore nearly all MSI tables have a Component_ column which contains the component name. If you look e.g. a the File table you see that every file belongs to a component which is true for all other tables which install resources. The component table is the glue between all other tables which contain the resources you want to install. So far so easy. Why is MSI then so complex? Most MSI problems arise from the fact that you did violate a MSI component rule in one or the other way. When you install a feature the reference count for all components belonging to this feature will increase by one. If your component is installed by more than one feature it will get a higher refcount. When you uninstall a feature its refcount will drop by one. Interesting things happen if the component reference count reaches zero: Then all associated resources will be deleted. That looks like a reasonable thing and it is. What it makes complex are the strange component rules you have to follow. Below are some important component rules from the Tao of the Windows Installer … Rule 16: Follow Component Rules Components are a very important part of the Installer technology. They are the means whereby the Installer manages the resources that make up your application. The SDK provides the following guidelines for creating components in your package: Never create two components that install a resource under the same name and target location. If a resource must be duplicated in multiple components, change its name or target location in each component. This rule should be applied across applications, products, product versions, and companies. Two components must not have the same key path file. This is a consequence of the previous rule. The key path value points to a particular file or folder belonging to the component that the installer uses to detect the component. If two components had the same key path file, the installer would be unable to distinguish which component is installed. Two components however may share a key path folder. Do not create a version of a component that is incompatible with all previous versions of the component. This rule should be applied across applications, products, product versions, and companies. Do not create components containing resources that will need to be installed into more than one directory on the user’s system. The installer installs all of the resources in a component into the same directory. It is not possible to install some resources into subdirectories. Do not include more than one COM server per component. If a component contains a COM server, this must be the key path for the component. Do not specify more than one file per component as a target for the Start menu or a Desktop shortcut. … And these rules do not even talk about component ids, update packages and upgrades which you need to understand as well. Lets suppose you install two MSIs (MSI1 and MSI2) which have the same ComponentId but different component names. Both do install the same file. What will happen when you uninstall MSI2?   Hm the file should stay there. But the component names are different. Yes and yes. But MSI uses not use the component name as key for the refcount. Instead the ComponentId column of the Component table which contains a GUID is used as identifier under which the refcount is stored. The components Comp1 and Comp2 are identical from the MSI perspective. After the installation of both MSIs the Component with the Id {100000….} has a refcount of two. After uninstallation of one MSI there is still a refcount of one which drops to zero just as expected when we uninstall the last msi. Then the file which was the same for both MSIs is deleted. You should remember that MSI keeps a refcount across MSIs for components with the same component id. MSI does manage components not the resources you did install. The resources associated with a component are then and only then deleted when the refcount of the component reaches zero.   The dependencies between features, components and resources can be described as relations. m,k are numbers >= 1, n can be 0. Inside a MSI the following relations are valid Feature    1  –> n Components Component    1 –> m Features Component      1  –>  k Resources These relations express that one feature can install several components and features can share components between them. Every (meaningful) component will install at least one resource which means that its name (primary key to stay in database speak) does occur in some other table in the Component column as value which installs some resource. Lets make it clear with an example. We want to install with the feature MainFeature some files a registry key and a shortcut. We can then create components Comp1..3 which are referenced by the resources defined in the corresponding tables.   Feature Component Registry File Shortcuts MainFeature Comp1 RegistryKey1     MainFeature Comp2   File.txt   MainFeature Comp3   File2.txt Shortcut to File2.txt   It is illegal that the same resource is part of more than one component since this would break the refcount mechanism. Lets illustrate this:            Feature ComponentId Resource Reference Count Feature1 {1000-…} File1.txt 1 Feature2 {2000-….} File1.txt 1 The installation part works well but what happens when you uninstall Feature2? Component {20000…} gets a refcount of zero where MSI deletes all resources belonging to this component. In this case File1.txt will be deleted. But Feature1 still has another component {10000…} with a refcount of one which means that the file was deleted too early. You just have ruined your installation. To fix it you then need to click on the Repair button under Add/Remove Programs to let MSI reinstall any missing registry keys, files or shortcuts. The vigilant reader might has noticed that there is more in the Component table. Beside its name and GUID it has also an installation directory, attributes and a KeyPath. The KeyPath is a reference to a file or registry key which is used to detect if the component is already installed. This becomes important when you repair or uninstall a component. To find out if the component is already installed MSI checks if the registry key or file referenced by the KeyPath property does exist. When it does not exist it assumes that it was either already uninstalled (can lead to problems during uninstall) or that it is already installed and all is fine. Why is this detail so important? Lets put all files into one component. The KeyPath should be then one of the files of your component to check if it was installed or not. When your installation becomes corrupt because a file was deleted you cannot repair it with the Repair button under Add/Remove Programs because MSI checks the component integrity via the Resource referenced by its KeyPath. As long as you did not delete the KeyPath file MSI thinks all resources with your component are installed and never executes any repair action. You get even more trouble when you try to remove files during an upgrade (you cannot remove files during an update) from your super component which contains all files. The only way out and therefore best practice is to assign for every resource you want to install an extra component. This ensures painless updatability and repairs and you have much less effort to remove specific files during an upgrade. In effect you get this best practice relation Feature 1  –> n Components Component   1  –>  1 Resources MSI Component Rules Rule 1 – One component per resource Every resource you want to install (file, registry key, value, environment value, shortcut, directory, …) must get its own component which does never change between versions as long as the install location is the same. Penalty If you add more than one resources to a component you will break the repair capability of MSI because the KeyPath is used to check if the component needs repair. MSI ComponentId Files MSI 1.0 {1000} File1-5 MSI 2.0 {2000} File2-5 You want to remove File1 in version 2.0 of your MSI. Since you want to keep the other files you create a new component and add them there. MSI will delete all files if the component refcount of {1000} drops to zero. The files you want to keep are added to the new component {2000}. Ok that does work if your upgrade does uninstall the old MSI first. This will cause the refcount of all previously installed components to reach zero which means that all files present in version 1.0 are deleted. But there is a faster way to perform your upgrade by first installing your new MSI and then remove the old one.  If you choose this upgrade path then you will loose File1-5 after your upgrade and not only File1 as intended by your new component design.   Rule 2 – Only add, never remove resources from a component If you did follow rule 1 you will not need Rule 2. You can add in a patch more resources to one component. That is ok. But you can never remove anything from it. There are tricky ways around that but I do not want to encourage bad component design. Penalty Lets assume you have 2 MSI files which install under the same component one file   MSI1 MSI2 {1000} - ComponentId {1000} – ComponentId File1.txt File2.txt   When you install and uninstall both MSIs you will end up with an installation where either File1 or File2 will be left. Why? It seems that MSI does not store the resources associated with each component in its internal database. Instead Windows will simply query the MSI that is currently uninstalled for all resources belonging to this component. Since it will find only one file and not two it will only uninstall one file. That is the main reason why you never can remove resources from a component!   Rule 3 Never Remove A Component From an Update MSI. This is the same as if you change the GUID of a component by accident for your new update package. The resulting update package will not contain all components from the previously installed package. Penalty When you remove a component from a feature MSI will set the feature state during update to Advertised and log a warning message into its log file when you did enable MSI logging. SELMGR: ComponentId '{2DCEA1BA-3E27-E222-484C-D0D66AEA4F62}' is registered to feature 'xxxxxxx, but is not present in the Component table.  Removal of components from a feature is not supported! MSI (c) (24:44) [07:53:13:436]: SELMGR: Removal of a component from a feature is not supported Advertised means that MSI treats all components of this feature as not installed. As a consequence during uninstall nothing will be removed since it is not installed! This is not only bad because uninstall does no longer work but this feature will also not get the required patches. All other features which have followed component versioning rules for update packages will be updated but the one faulty feature will not. This results in very hard to find bugs why an update was only partially successful. Things got better with Windows Installer 4.5 but you cannot rely on that nobody will use an older installer. It is a good idea to add to your update msiexec call MSIENFORCEUPGRADECOMPONENTRULES=1 which will abort the installation if you did violate this rule.

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  • Red Gate Coder interviews: Robin Hellen

    - by Michael Williamson
    Robin Hellen is a test engineer here at Red Gate, and is also the latest coder I’ve interviewed. We chatted about debugging code, the roles of software engineers and testers, and why Vala is currently his favourite programming language. How did you get started with programming?It started when I was about six. My dad’s a professional programmer, and he gave me and my sister one of his old computers and taught us a bit about programming. It was an old Amiga 500 with a variant of BASIC. I don’t think I ever successfully completed anything! It was just faffing around. I didn’t really get anywhere with it.But then presumably you did get somewhere with it at some point.At some point. The PC emerged as the dominant platform, and I learnt a bit of Visual Basic. I didn’t really do much, just a couple of quick hacky things. A bit of demo animation. Took me a long time to get anywhere with programming, really.When did you feel like you did start to get somewhere?I think it was when I started doing things for someone else, which was my sister’s final year of university project. She called up my dad two days before she was due to submit, saying “We need something to display a graph!”. Dad says, “I’m too busy, go talk to your brother”. So I hacked up this ugly piece of code, sent it off and they won a prize for that project. Apparently, the graph, the bit that I wrote, was the reason they won a prize! That was when I first felt that I’d actually done something that was worthwhile. That was my first real bit of code, and the ugliest code I’ve ever written. It’s basically an array of pre-drawn line elements that I shifted round the screen to draw a very spikey graph.When did you decide that programming might actually be something that you wanted to do as a career?It’s not really a decision I took, I always wanted to do something with computers. And I had to take a gap year for uni, so I was looking for twelve month internships. I applied to Red Gate, and they gave me a job as a tester. And that’s where I really started having to write code well. To a better standard that I had been up to that point.How did you find coming to Red Gate and working with other coders?I thought it was really nice. I learnt so much just from other people around. I think one of the things that’s really great is that people are just willing to help you learn. Instead of “Don’t you know that, you’re so stupid”, it’s “You can just do it this way”.If you could go back to the very start of that internship, is there something that you would tell yourself?Write shorter code. I have a tendency to write massive, many-thousand line files that I break out of right at the end. And then half-way through a project I’m doing something, I think “Where did I write that bit that does that thing?”, and it’s almost impossible to find. I wrote some horrendous code when I started. Just that principle, just keep things short. Even if looks a bit crazy to be jumping around all over the place all of the time, it’s actually a lot more understandable.And how do you hold yourself to that?Generally, if a function’s going off my screen, it’s probably too long. That’s what I tell myself, and within the team here we have code reviews, so the guys I’m with at the moment are pretty good at pulling me up on, “Doesn’t that look like it’s getting a bit long?”. It’s more just the subjective standard of readability than anything.So you’re an advocate of code review?Yes, definitely. Both to spot errors that you might have made, and to improve your knowledge. The person you’re reviewing will say “Oh, you could have done it that way”. That’s how we learn, by talking to others, and also just sharing knowledge of how your project works around the team, or even outside the team. Definitely a very firm advocate of code reviews.Do you think there’s more we could do with them?I don’t know. We’re struggling with how to add them as part of the process without it becoming too cumbersome. We’ve experimented with a few different ways, and we’ve not found anything that just works.To get more into the nitty gritty: how do you like to debug code?The first thing is to do it in my head. I’ll actually think what piece of code is likely to have caused that error, and take a quick look at it, just to see if there’s anything glaringly obvious there. The next thing I’ll probably do is throw in print statements, or throw some exceptions from various points, just to check: is it going through the code path I expect it to? A last resort is to actually debug code using a debugger.Why is the debugger the last resort?Probably because of the environments I learnt programming in. VB and early BASIC didn’t have much of a debugger, the only way to find out what your program was doing was to add print statements. Also, because a lot of the stuff I tend to work with is non-interactive, if it’s something that takes a long time to run, I can throw in the print statements, set a run off, go and do something else, and look at it again later, rather than trying to remember what happened at that point when I was debugging through it. So it also gives me the record of what happens. I hate just sitting there pressing F5, F5, continually. If you’re having to find out what your code is doing at each line, you’ve probably got a very wrong mental model of what your code’s doing, and you can find that out just as easily by inspecting a couple of values through the print statements.If I were on some codebase that you were also working on, what should I do to make it as easy as possible to understand?I’d say short and well-named methods. The one thing I like to do when I’m looking at code is to find out where a value comes from, and the more layers of indirection there are, particularly DI [dependency injection] frameworks, the harder it is to find out where something’s come from. I really hate that. I want to know if the value come from the user here or is a constant here, and if I can’t find that out, that makes code very hard to understand for me.As a tester, where do you think the split should lie between software engineers and testers?I think the split is less on areas of the code you write and more what you’re designing and creating. The developers put a structure on the code, while my major role is to say which tests we should have, whether we should test that, or it’s not worth testing that because it’s a tiny function in code that nobody’s ever actually going to see. So it’s not a split in the code, it’s a split in what you’re thinking about. Saying what code we should write, but alternatively what code we should take out.In your experience, do the software engineers tend to do much testing themselves?They tend to control the lowest layer of tests. And, depending on how the balance of people is in the team, they might write some of the higher levels of test. Or that might go to the testers. I’m the only tester on my team with three other developers, so they’ll be writing quite a lot of the actual test code, with input from me as to whether we should test that functionality, whereas on other teams, where it’s been more equal numbers, the testers have written pretty much all of the high level tests, just because that’s the best use of resource.If you could shuffle resources around however you liked, do you think that the developers should be writing those high-level tests?I think they should be writing them occasionally. It helps when they have an understanding of how testing code works and possibly what assumptions we’ve made in tests, and they can say “actually, it doesn’t work like that under the hood so you’ve missed this whole area”. It’s one of those agile things that everyone on the team should be at least comfortable doing the various jobs. So if the developers can write test code then I think that’s a very good thing.So you think testers should be able to write production code?Yes, although given most testers skills at coding, I wouldn’t advise it too much! I have written a few things, and I did make a few changes that have actually gone into our production code base. They’re not necessarily running every time but they are there. I think having that mix of skill sets is really useful. In some ways we’re using our own product to test itself, so being able to make those changes where it’s not working saves me a round-trip through the developers. It can be really annoying if the developers have no time to make a change, and I can’t touch the code.If the software engineers are consistently writing tests at all levels, what role do you think the role of a tester is?I think on a team like that, those distinctions aren’t quite so useful. There’ll be two cases. There’s either the case where the developers think they’ve written good tests, but you still need someone with a test engineer mind-set to go through the tests and validate that it’s a useful set, or the correct set for that code. Or they won’t actually be pure developers, they’ll have that mix of test ability in there.I think having slightly more distinct roles is useful. When it starts to blur, then you lose that view of the tests as a whole. The tester job is not to create tests, it’s to validate the quality of the product, and you don’t do that just by writing tests. There’s more things you’ve got to keep in your mind. And I think when you blur the roles, you start to lose that end of the tester.So because you’re working on those features, you lose that holistic view of the whole system?Yeah, and anyone who’s worked on the feature shouldn’t be testing it. You always need to have it tested it by someone who didn’t write it. Otherwise you’re a bit too close and you assume “yes, people will only use it that way”, but the tester will come along and go “how do people use this? How would our most idiotic user use this?”. I might not test that because it might be completely irrelevant. But it’s coming in and trying to have a different set of assumptions.Are you a believer that it should all be automated if possible?Not entirely. So an automated test is always better than a manual test for the long-term, but there’s still nothing that beats a human sitting in front of the application and thinking “What could I do at this point?”. The automated test is very good but they follow that strict path, and they never check anything off the path. The human tester will look at things that they weren’t expecting, whereas the automated test can only ever go “Is that value correct?” in many respects, and it won’t notice that on the other side of the screen you’re showing something completely wrong. And that value might have been checked independently, but you always find a few odd interactions when you’re going through something manually, and you always need to go through something manually to start with anyway, otherwise you won’t know where the important bits to write your automation are.When you’re doing that manual testing, do you think it’s important to do that across the entire product, or just the bits that you’ve touched recently?I think it’s important to do it mostly on the bits you’ve touched, but you can’t ignore the rest of the product. Unless you’re dealing with a very, very self-contained bit, you’re almost always encounter other bits of the product along the way. Most testers I know, even if they are looking at just one path, they’ll keep open and move around a bit anyway, just because they want to find something that’s broken. If we find that your path is right, we’ll go out and hunt something else.How do you think this fits into the idea of continuously deploying, so long as the tests pass?With deploying a website it’s a bit different because you can always pull it back. If you’re deploying an application to customers, when you’ve released it, it’s out there, you can’t pull it back. Someone’s going to keep it, no matter how hard you try there will be a few installations that stay around. So I’d always have at least a human element on that path. With websites, you could probably automate straight out, or at least straight out to an internal environment or a single server in a cloud of fifty that will serve some people. But I don’t think you should release to everyone just on automated tests passing.You’ve already mentioned using BASIC and C# — are there any other languages that you’ve used?I’ve used a few. That’s something that has changed more recently, I’ve become familiar with more languages. Before I started at Red Gate I learnt a bit of C. Then last year, I taught myself Python which I actually really enjoyed using. I’ve also come across another language called Vala, which is sort of a C#-like language. It’s basically a pre-processor for C, but it has very nice syntax. I think that’s currently my favourite language.Any particular reason for trying Vala?I have a completely Linux environment at home, and I’ve been looking for a nice language, and C# just doesn’t cut it because I won’t touch Mono. So, I was looking for something like C# but that was useable in an open source environment, and Vala’s what I found. C#’s got a few features that Vala doesn’t, and Vala’s got a few features where I think “It would be awesome if C# had that”.What are some of the features that it’s missing?Extension methods. And I think that’s the only one that really bugs me. I like to use them when I’m writing C# because it makes some things really easy, especially with libraries that you can’t touch the internals of. It doesn’t have method overloading, which is sometimes annoying.Where it does win over C#?Everything is non-nullable by default, you never have to check that something’s unexpectedly null.Also, Vala has code contracts. This is starting to come in C# 4, but the way it works in Vala is that you specify requirements in short phrases as part of your function signature and they stick to the signature, so that when you inherit it, it has exactly the same code contract as the base one, or when you inherit from an interface, you have to match the signature exactly. Just using those makes you think a bit more about how you’re writing your method, it’s not an afterthought when you’ve got contracts from base classes given to you, you can’t change it. Which I think is a lot nicer than the way C# handles it. When are those actually checked?They’re checked both at compile and run-time. The compile-time checking isn’t very strong yet, it’s quite a new feature in the compiler, and because it compiles down to C, you can write C code and interface with your methods, so you can bypass that compile-time check anyway. So there’s an extra runtime check, and if you violate one of the contracts at runtime, it’s game over for your program, there’s no exception to catch, it’s just goodbye!One thing I dislike about C# is the exceptions. You write a bit of code and fifty exceptions could come from any point in your ten lines, and you can’t mentally model how those exceptions are going to come out, and you can’t even predict them based on the functions you’re calling, because if you’ve accidentally got a derived class there instead of a base class, that can throw a completely different set of exceptions. So I’ve got no way of mentally modelling those, whereas in Vala they’re checked like Java, so you know only these exceptions can come out. You know in advance the error conditions.I think Raymond Chen on Old New Thing says “the only thing you know when you throw an exception is that you’re in an invalid state somewhere in your program, so just kill it and be done with it!”You said you’ve also learnt bits of Python. How did you find that compared to Vala and C#?Very different because of the dynamic typing. I’ve been writing a website for my own use. I’m quite into photography, so I take photos off my camera, post-process them, dump them in a file, and I get a webpage with all my thumbnails. So sort of like Picassa, but written by myself because I wanted something to learn Python with. There are some things that are really nice, I just found it really difficult to cope with the fact that I’m not quite sure what this object type that I’m passed is, I might not ever be sure, so it can randomly blow up on me. But once I train myself to ignore that and just say “well, I’m fairly sure it’s going to be something that looks like this, so I’ll use it like this”, then it’s quite nice.Any particular features that you’ve appreciated?I don’t like any particular feature, it’s just very straightforward to work with. It’s very quick to write something in, particularly as you don’t have to worry that you’ve changed something that affects a different part of the program. If you have, then that part blows up, but I can get this part working right now.If you were doing a big project, would you be willing to do it in Python rather than C# or Vala?I think I might be willing to try something bigger or long term with Python. We’re currently doing an ASP.NET MVC project on C#, and I don’t like the amount of reflection. There’s a lot of magic that pulls values out, and it’s all done under the scenes. It’s almost managed to put a dynamic type system on top of C#, which in many ways destroys the language to me, whereas if you’re already in a dynamic language, having things done dynamically is much more natural. In many ways, you get the worst of both worlds. I think for web projects, I would go with Python again, whereas for anything desktop, command-line or GUI-based, I’d probably go for C# or Vala, depending on what environment I’m in.It’s the fact that you can gain from the strong typing in ways that you can’t so much on the web app. Or, in a web app, you have to use dynamic typing at some point, or you have to write a hell of a lot of boilerplate, and I’d rather use the dynamic typing than write the boilerplate.What do you think separates great programmers from everyone else?Probably design choices. Choosing to write it a piece of code one way or another. For any given program you ask me to write, I could probably do it five thousand ways. A programmer who is capable will see four or five of them, and choose one of the better ones. The excellent programmer will see the largest proportion and manage to pick the best one very quickly without having to think too much about it. I think that’s probably what separates, is the speed at which they can see what’s the best path to write the program in. More Red Gater Coder interviews

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  • Microsoft Introduces WebMatrix

    - by Rick Strahl
    originally published in CoDe Magazine Editorial Microsoft recently released the first CTP of a new development environment called WebMatrix, which along with some of its supporting technologies are squarely aimed at making the Microsoft Web Platform more approachable for first-time developers and hobbyists. But in the process, it also provides some updated technologies that can make life easier for existing .NET developers. Let’s face it: ASP.NET development isn’t exactly trivial unless you already have a fair bit of familiarity with sophisticated development practices. Stick a non-developer in front of Visual Studio .NET or even the Visual Web Developer Express edition and it’s not likely that the person in front of the screen will be very productive or feel inspired. Yet other technologies like PHP and even classic ASP did provide the ability for non-developers and hobbyists to become reasonably proficient in creating basic web content quickly and efficiently. WebMatrix appears to be Microsoft’s attempt to bring back some of that simplicity with a number of technologies and tools. The key is to provide a friendly and fully self-contained development environment that provides all the tools needed to build an application in one place, as well as tools that allow publishing of content and databases easily to the web server. WebMatrix is made up of several components and technologies: IIS Developer Express IIS Developer Express is a new, self-contained development web server that is fully compatible with IIS 7.5 and based on the same codebase that IIS 7.5 uses. This new development server replaces the much less compatible Cassini web server that’s been used in Visual Studio and the Express editions. IIS Express addresses a few shortcomings of the Cassini server such as the inability to serve custom ISAPI extensions (i.e., things like PHP or ASP classic for example), as well as not supporting advanced authentication. IIS Developer Express provides most of the IIS 7.5 feature set providing much better compatibility between development and live deployment scenarios. SQL Server Compact 4.0 Database access is a key component for most web-driven applications, but on the Microsoft stack this has mostly meant you have to use SQL Server or SQL Server Express. SQL Server Compact is not new-it’s been around for a few years, but it’s been severely hobbled in the past by terrible tool support and the inability to support more than a single connection in Microsoft’s attempt to avoid losing SQL Server licensing. The new release of SQL Server Compact 4.0 supports multiple connections and you can run it in ASP.NET web applications simply by installing an assembly into the bin folder of the web application. In effect, you don’t have to install a special system configuration to run SQL Compact as it is a drop-in database engine: Copy the small assembly into your BIN folder (or from the GAC if installed fully), create a connection string against a local file-based database file, and then start firing SQL requests. Additionally WebMatrix includes nice tools to edit the database tables and files, along with tools to easily upsize (and hopefully downsize in the future) to full SQL Server. This is a big win, pending compatibility and performance limits. In my simple testing the data engine performed well enough for small data sets. This is not only useful for web applications, but also for desktop applications for which a fully installed SQL engine like SQL Server would be overkill. Having a local data store in those applications that can potentially be accessed by multiple users is a welcome feature. ASP.NET Razor View Engine What? Yet another native ASP.NET view engine? We already have Web Forms and various different flavors of using that view engine with Web Forms and MVC. Do we really need another? Microsoft thinks so, and Razor is an implementation of a lightweight, script-only view engine. Unlike the Web Forms view engine, Razor works only with inline code, snippets, and markup; therefore, it is more in line with current thinking of what a view engine should represent. There’s no support for a “page model” or any of the other Web Forms features of the full-page framework, but just a lightweight scripting engine that works with plain markup plus embedded expressions and code. The markup syntax for Razor is geared for minimal typing, plus some progressive detection of where a script block/expression starts and ends. This results in a much leaner syntax than the typical ASP.NET Web Forms alligator (<% %>) tags. Razor uses the @ sign plus standard C# (or Visual Basic) block syntax to delineate code snippets and expressions. Here’s a very simple example of what Razor markup looks like along with some comment annotations: <!DOCTYPE html> <html>     <head>         <title></title>     </head>     <body>     <h1>Razor Test</h1>          <!-- simple expressions -->     @DateTime.Now     <hr />     <!-- method expressions -->     @DateTime.Now.ToString("T")          <!-- code blocks -->     @{         List<string> names = new List<string>();         names.Add("Rick");         names.Add("Markus");         names.Add("Claudio");         names.Add("Kevin");     }          <!-- structured block statements -->     <ul>     @foreach(string name in names){             <li>@name</li>     }     </ul>           <!-- Conditional code -->        @if(true) {                        <!-- Literal Text embedding in code -->        <text>         true        </text>;    }    else    {        <!-- Literal Text embedding in code -->       <text>       false       </text>;    }    </body> </html> Like the Web Forms view engine, Razor parses pages into code, and then executes that run-time compiled code. Effectively a “page” becomes a code file with markup becoming literal text written into the Response stream, code snippets becoming raw code, and expressions being written out with Response.Write(). The code generated from Razor doesn’t look much different from similar Web Forms code that only uses script tags; so although the syntax may look different, the operational model is fairly similar to the Web Forms engine minus the overhead of the large Page object model. However, there are differences: -Razor pages are based on a new base class, Microsoft.WebPages.WebPage, which is hosted in the Microsoft.WebPages assembly that houses all the Razor engine parsing and processing logic. Browsing through the assembly (in the generated ASP.NET Temporary Files folder or GAC) will give you a good idea of the functionality that Razor provides. If you look closely, a lot of the feature set matches ASP.NET MVC’s view implementation as well as many of the helper classes found in MVC. It’s not hard to guess the motivation for this sort of view engine: For beginning developers the simple markup syntax is easier to work with, although you obviously still need to have some understanding of the .NET Framework in order to create dynamic content. The syntax is easier to read and grok and much shorter to type than ASP.NET alligator tags (<% %>) and also easier to understand aesthetically what’s happening in the markup code. Razor also is a better fit for Microsoft’s vision of ASP.NET MVC: It’s a new view engine without the baggage of Web Forms attached to it. The engine is more lightweight since it doesn’t carry all the features and object model of Web Forms with it and it can be instantiated directly outside of the HTTP environment, which has been rather tricky to do for the Web Forms view engine. Having a standalone script parser is a huge win for other applications as well – it makes it much easier to create script or meta driven output generators for many types of applications from code/screen generators, to simple form letters to data merging applications with user customizability. For me personally this is very useful side effect and who knows maybe Microsoft will actually standardize they’re scripting engines (die T4 die!) on this engine. Razor also better fits the “view-based” approach where the view is supposed to be mostly a visual representation that doesn’t hold much, if any, code. While you can still use code, the code you do write has to be self-contained. Overall I wouldn’t be surprised if Razor will become the new standard view engine for MVC in the future – and in fact there have been announcements recently that Razor will become the default script engine in ASP.NET MVC 3.0. Razor can also be used in existing Web Forms and MVC applications, although that’s not working currently unless you manually configure the script mappings and add the appropriate assemblies. It’s possible to do it, but it’s probably better to wait until Microsoft releases official support for Razor scripts in Visual Studio. Once that happens, you can simply drop .cshtml and .vbhtml pages into an existing ASP.NET project and they will work side by side with classic ASP.NET pages. WebMatrix Development Environment To tie all of these three technologies together, Microsoft is shipping WebMatrix with an integrated development environment. An integrated gallery manager makes it easy to download and load existing projects, and then extend them with custom functionality. It seems to be a prominent goal to provide community-oriented content that can act as a starting point, be it via a custom templates or a complete standard application. The IDE includes a project manager that works with a single project and provides an integrated IDE/editor for editing the .cshtml and .vbhtml pages. A run button allows you to quickly run pages in the project manager in a variety of browsers. There’s no debugging support for code at this time. Note that Razor pages don’t require explicit compilation, so making a change, saving, and then refreshing your page in the browser is all that’s needed to see changes while testing an application locally. It’s essentially using the auto-compiling Web Project that was introduced with .NET 2.0. All code is compiled during run time into dynamically created assemblies in the ASP.NET temp folder. WebMatrix also has PHP Editing support with syntax highlighting. You can load various PHP-based applications from the WebMatrix Web Gallery directly into the IDE. Most of the Web Gallery applications are ready to install and run without further configuration, with Wizards taking you through installation of tools, dependencies, and configuration of the database as needed. WebMatrix leverages the Web Platform installer to pull the pieces down from websites in a tight integration of tools that worked nicely for the four or five applications I tried this out on. Click a couple of check boxes and fill in a few simple configuration options and you end up with a running application that’s ready to be customized. Nice! You can easily deploy completed applications via WebDeploy (to an IIS server) or FTP directly from within the development environment. The deploy tool also can handle automatically uploading and installing the database and all related assemblies required, making deployment a simple one-click install step. Simplified Database Access The IDE contains a database editor that can edit SQL Compact and SQL Server databases. There is also a Database helper class that facilitates database access by providing easy-to-use, high-level query execution and iteration methods: @{       var db = Database.OpenFile("FirstApp.sdf");     string sql = "select * from customers where Id > @0"; } <ul> @foreach(var row in db.Query(sql,1)){         <li>@row.FirstName @row.LastName</li> } </ul> The query function takes a SQL statement plus any number of positional (@0,@1 etc.) SQL parameters by simple values. The result is returned as a collection of rows which in turn have a row object with dynamic properties for each of the columns giving easy (though untyped) access to each of the fields. Likewise Execute and ExecuteNonQuery allow execution of more complex queries using similar parameter passing schemes. Note these queries use string-based queries rather than LINQ or Entity Framework’s strongly typed LINQ queries. While this may seem like a step back, it’s also in line with the expectations of non .NET script developers who are quite used to writing and using SQL strings in code rather than using OR/M frameworks. The only question is why was something not included from the beginning in .NET and Microsoft made developers build custom implementations of these basic building blocks. The implementation looks a lot like a DataTable-style data access mechanism, but to be fair, this is a common approach in scripting languages. This type of syntax that uses simple, static, data object methods to perform simple data tasks with one line of code are common in scripting languages and are a good match for folks working in PHP/Python, etc. Seems like Microsoft has taken great advantage of .NET 4.0’s dynamic typing to provide this sort of interface for row iteration where each row has properties for each field. FWIW, all the examples demonstrate using local SQL Compact files - I was unable to get a SQL Server connection string to work with the Database class (the connection string wasn’t accepted). However, since the code in the page is still plain old .NET, you can easily use standard ADO.NET code or even LINQ or Entity Framework models that are created outside of WebMatrix in separate assemblies as required. The good the bad the obnoxious - It’s still .NET The beauty (or curse depending on how you look at it :)) of Razor and the compilation model is that, behind it all, it’s still .NET. Although the syntax may look foreign, it’s still all .NET behind the scenes. You can easily access existing tools, helpers, and utilities simply by adding them to the project as references or to the bin folder. Razor automatically recognizes any assembly reference from assemblies in the bin folder. In the default configuration, Microsoft provides a host of helper functions in a Microsoft.WebPages assembly (check it out in the ASP.NET temp folder for your application), which includes a host of HTML Helpers. If you’ve used ASP.NET MVC before, a lot of the helpers should look familiar. Documentation at the moment is sketchy-there’s a very rough API reference you can check out here: http://www.asp.net/webmatrix/tutorials/asp-net-web-pages-api-reference Who needs WebMatrix? Uhm… good Question Clearly Microsoft is trying hard to create an environment with WebMatrix that is easy to use for newbie developers. The goal seems to be simplicity in providing a minimal development environment and an easy-to-use script engine/language that makes it easy to get started with. There’s also some focus on community features that can be used as starting points, such as Web Gallery applications and templates. The community features in particular are very nice and something that would be nice to eventually see in Visual Studio as well. The question is whether this is too little too late. Developers who have been clamoring for a simpler development environment on the .NET stack have mostly left for other simpler platforms like PHP or Python which are catering to the down and dirty developer. Microsoft will be hard pressed to win those folks-and other hardcore PHP developers-back. Regardless of how much you dress up a script engine fronted by the .NET Framework, it’s still the .NET Framework and all the complexity that drives it. While .NET is a fine solution in its breadth and features once you get a basic handle on the core features, the bar of entry to being productive with the .NET Framework is still pretty high. The MVC style helpers Microsoft provides are a good step in the right direction, but I suspect it’s not enough to shield new developers from having to delve much deeper into the Framework to get even basic applications built. Razor and its helpers is trying to make .NET more accessible but the reality is that in order to do useful stuff that goes beyond the handful of simple helpers you still are going to have to write some C# or VB or other .NET code. If the target is a hobby/amateur/non-programmer the learning curve isn’t made any easier by WebMatrix it’s just been shifted a tad bit further along in your development endeavor when you run out of canned components that are supplied either by Microsoft or the community. The database helpers are interesting and actually I’ve heard a lot of discussion from various developers who’ve been resisting .NET for a really long time perking up at the prospect of easier data access in .NET than the ridiculous amount of code it takes to do even simple data access with raw ADO.NET. It seems sad that such a simple concept and implementation should trigger this sort of response (especially since it’s practically trivial to create helpers like these or pick them up from countless libraries available), but there it is. It also shows that there are plenty of developers out there who are more interested in ‘getting stuff done’ easily than necessarily following the latest and greatest practices which are overkill for many development scenarios. Sometimes it seems that all of .NET is focused on the big life changing issues of development, rather than the bread and butter scenarios that many developers are interested in to get their work accomplished. And that in the end may be WebMatrix’s main raison d'être: To bring some focus back at Microsoft that simpler and more high level solutions are actually needed to appeal to the non-high end developers as well as providing the necessary tools for the high end developers who want to follow the latest and greatest trends. The current version of WebMatrix hits many sweet spots, but it also feels like it has a long way to go before it really can be a tool that a beginning developer or an accomplished developer can feel comfortable with. Although there are some really good ideas in the environment (like the gallery for downloading apps and components) which would be a great addition for Visual Studio as well, the rest of the development environment just feels like crippleware with required functionality missing especially debugging and Intellisense, but also general editor support. It’s not clear whether these are because the product is still in an early alpha release or whether it’s simply designed that way to be a really limited development environment. While simple can be good, nobody wants to feel left out when it comes to necessary tool support and WebMatrix just has that left out feeling to it. If anything WebMatrix’s technology pieces (which are really independent of the WebMatrix product) are what are interesting to developers in general. The compact IIS implementation is a nice improvement for development scenarios and SQL Compact 4.0 seems to address a lot of concerns that people have had and have complained about for some time with previous SQL Compact implementations. By far the most interesting and useful technology though seems to be the Razor view engine for its light weight implementation and it’s decoupling from the ASP.NET/HTTP pipeline to provide a standalone scripting/view engine that is pluggable. The first winner of this is going to be ASP.NET MVC which can now have a cleaner view model that isn’t inconsistent due to the baggage of non-implemented WebForms features that don’t work in MVC. But I expect that Razor will end up in many other applications as a scripting and code generation engine eventually. Visual Studio integration for Razor is currently missing, but is promised for a later release. The ASP.NET MVC team has already mentioned that Razor will eventually become the default MVC view engine, which will guarantee continued growth and development of this tool along those lines. And the Razor engine and support tools actually inherit many of the features that MVC pioneered, so there’s some synergy flowing both ways between Razor and MVC. As an existing ASP.NET developer who’s already familiar with Visual Studio and ASP.NET development, the WebMatrix IDE doesn’t give you anything that you want. The tools provided are minimal and provide nothing that you can’t get in Visual Studio today, except the minimal Razor syntax highlighting, so there’s little need to take a step back. With Visual Studio integration coming later there’s little reason to look at WebMatrix for tooling. It’s good to see that Microsoft is giving some thought about the ease of use of .NET as a platform For so many years, we’ve been piling on more and more new features without trying to take a step back and see how complicated the development/configuration/deployment process has become. Sometimes it’s good to take a step - or several steps - back and take another look and realize just how far we’ve come. WebMatrix is one of those reminders and one that likely will result in some positive changes on the platform as a whole. © Rick Strahl, West Wind Technologies, 2005-2010Posted in ASP.NET   IIS7  

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  • Sniffing out SQL Code Smells: Inconsistent use of Symbolic names and Datatypes

    - by Phil Factor
    It is an awkward feeling. You’ve just delivered a database application that seems to be working fine in production, and you just run a few checks on it. You discover that there is a potential bug that, out of sheer good chance, hasn’t kicked in to produce an error; but it lurks, like a smoking bomb. Worse, maybe you find that the bug has started its evil work of corrupting the data, but in ways that nobody has, so far detected. You investigate, and find the damage. You are somehow going to have to repair it. Yes, it still very occasionally happens to me. It is not a nice feeling, and I do anything I can to prevent it happening. That’s why I’m interested in SQL code smells. SQL Code Smells aren’t necessarily bad practices, but just show you where to focus your attention when checking an application. Sometimes with databases the bugs can be subtle. SQL is rather like HTML: the language does its best to try to carry out your wishes, rather than to be picky about your bugs. Most of the time, this is a great benefit, but not always. One particular place where this can be detrimental is where you have implicit conversion between different data types. Most of the time it is completely harmless but we’re  concerned about the occasional time it isn’t. Let’s give an example: String truncation. Let’s give another even more frightening one, rounding errors on assignment to a number of different precision. Each requires a blog-post to explain in detail and I’m not now going to try. Just remember that it is not always a good idea to assign data to variables, parameters or even columns when they aren’t the same datatype, especially if you are relying on implicit conversion to work its magic.For details of the problem and the consequences, see here:  SR0014: Data loss might occur when casting from {Type1} to {Type2} . For any experienced Database Developer, this is a more frightening read than a Vampire Story. This is why one of the SQL Code Smells that makes me edgy, in my own or other peoples’ code, is to see parameters, variables and columns that have the same names and different datatypes. Whereas quite a lot of this is perfectly normal and natural, you need to check in case one of two things have gone wrong. Either sloppy naming, or mixed datatypes. Sure it is hard to remember whether you decided that the length of a log entry was 80 or 100 characters long, or the precision of a number. That is why a little check like this I’m going to show you is excellent for tidying up your code before you check it back into source Control! 1/ Checking Parameters only If you were just going to check parameters, you might just do this. It simply groups all the parameters, either input or output, of all the routines (e.g. stored procedures or functions) by their name and checks to see, in the HAVING clause, whether their data types are all the same. If not, it lists all the examples and their origin (the routine) Even this little check can occasionally be scarily revealing. ;WITH userParameter AS  ( SELECT   c.NAME AS ParameterName,  OBJECT_SCHEMA_NAME(c.object_ID) + '.' + OBJECT_NAME(c.object_ID) AS ObjectName,  t.name + ' '     + CASE     --we may have to put in the length            WHEN t.name IN ('char', 'varchar', 'nchar', 'nvarchar')             THEN '('               + CASE WHEN c.max_length = -1 THEN 'MAX'                ELSE CONVERT(VARCHAR(4),                    CASE WHEN t.name IN ('nchar', 'nvarchar')                      THEN c.max_length / 2 ELSE c.max_length                    END)                END + ')'         WHEN t.name IN ('decimal', 'numeric')             THEN '(' + CONVERT(VARCHAR(4), c.precision)                   + ',' + CONVERT(VARCHAR(4), c.Scale) + ')'         ELSE ''      END  --we've done with putting in the length      + CASE WHEN XML_collection_ID <> 0         THEN --deal with object schema names             '(' + CASE WHEN is_XML_Document = 1                    THEN 'DOCUMENT '                    ELSE 'CONTENT '                   END              + COALESCE(               (SELECT QUOTENAME(ss.name) + '.' + QUOTENAME(sc.name)                FROM sys.xml_schema_collections sc                INNER JOIN Sys.Schemas ss ON sc.schema_ID = ss.schema_ID                WHERE sc.xml_collection_ID = c.XML_collection_ID),'NULL') + ')'          ELSE ''         END        AS [DataType]  FROM sys.parameters c  INNER JOIN sys.types t ON c.user_Type_ID = t.user_Type_ID  WHERE OBJECT_SCHEMA_NAME(c.object_ID) <> 'sys'   AND parameter_id>0)SELECT CONVERT(CHAR(80),objectName+'.'+ParameterName),DataType FROM UserParameterWHERE ParameterName IN   (SELECT ParameterName FROM UserParameter    GROUP BY ParameterName    HAVING MIN(Datatype)<>MAX(DataType))ORDER BY ParameterName   so, in a very small example here, we have a @ClosingDelimiter variable that is only CHAR(1) when, by the looks of it, it should be up to ten characters long, or even worse, a function that should be a char(1) and seems to let in a string of ten characters. Worth investigating. Then we have a @Comment variable that can't decide whether it is a VARCHAR(2000) or a VARCHAR(MAX) 2/ Columns and Parameters Actually, once we’ve cleared up the mess we’ve made of our parameter-naming in the database we’re inspecting, we’re going to be more interested in listing both columns and parameters. We can do this by modifying the routine to list columns as well as parameters. Because of the slight complexity of creating the string version of the datatypes, we will create a fake table of both columns and parameters so that they can both be processed the same way. After all, we want the datatypes to match Unfortunately, parameters do not expose all the attributes we are interested in, such as whether they are nullable (oh yes, subtle bugs happen if this isn’t consistent for a datatype). We’ll have to leave them out for this check. Voila! A slight modification of the first routine ;WITH userObject AS  ( SELECT   Name AS DataName,--the actual name of the parameter or column ('@' removed)  --and the qualified object name of the routine  OBJECT_SCHEMA_NAME(ObjectID) + '.' + OBJECT_NAME(ObjectID) AS ObjectName,  --now the harder bit: the definition of the datatype.  TypeName + ' '     + CASE     --we may have to put in the length. e.g. CHAR (10)           WHEN TypeName IN ('char', 'varchar', 'nchar', 'nvarchar')             THEN '('               + CASE WHEN MaxLength = -1 THEN 'MAX'                ELSE CONVERT(VARCHAR(4),                    CASE WHEN TypeName IN ('nchar', 'nvarchar')                      THEN MaxLength / 2 ELSE MaxLength                    END)                END + ')'         WHEN TypeName IN ('decimal', 'numeric')--a BCD number!             THEN '(' + CONVERT(VARCHAR(4), Precision)                   + ',' + CONVERT(VARCHAR(4), Scale) + ')'         ELSE ''      END  --we've done with putting in the length      + CASE WHEN XML_collection_ID <> 0 --tush tush. XML         THEN --deal with object schema names             '(' + CASE WHEN is_XML_Document = 1                    THEN 'DOCUMENT '                    ELSE 'CONTENT '                   END              + COALESCE(               (SELECT TOP 1 QUOTENAME(ss.name) + '.' + QUOTENAME(sc.Name)                FROM sys.xml_schema_collections sc                INNER JOIN Sys.Schemas ss ON sc.schema_ID = ss.schema_ID                WHERE sc.xml_collection_ID = XML_collection_ID),'NULL') + ')'          ELSE ''         END        AS [DataType],       DataObjectType  FROM   (Select t.name AS TypeName, REPLACE(c.name,'@','') AS Name,          c.max_length AS MaxLength, c.precision AS [Precision],           c.scale AS [Scale], c.[Object_id] AS ObjectID, XML_collection_ID,          is_XML_Document,'P' AS DataobjectType  FROM sys.parameters c  INNER JOIN sys.types t ON c.user_Type_ID = t.user_Type_ID  AND parameter_id>0  UNION all  Select t.name AS TypeName, c.name AS Name, c.max_length AS MaxLength,          c.precision AS [Precision], c.scale AS [Scale],          c.[Object_id] AS ObjectID, XML_collection_ID,is_XML_Document,          'C' AS DataobjectType            FROM sys.columns c  INNER JOIN sys.types t ON c.user_Type_ID = t.user_Type_ID   WHERE OBJECT_SCHEMA_NAME(c.object_ID) <> 'sys'  )f)SELECT CONVERT(CHAR(80),objectName+'.'   + CASE WHEN DataobjectType ='P' THEN '@' ELSE '' END + DataName),DataType FROM UserObjectWHERE DataName IN   (SELECT DataName FROM UserObject   GROUP BY DataName    HAVING MIN(Datatype)<>MAX(DataType))ORDER BY DataName     Hmm. I can tell you I found quite a few minor issues with the various tabases I tested this on, and found some potential bugs that really leap out at you from the results. Here is the start of the result for AdventureWorks. Yes, AccountNumber is, for some reason, a Varchar(10) in the Customer table. Hmm. odd. Why is a city fifty characters long in that view?  The idea of the description of a colour being 256 characters long seems over-ambitious. Go down the list and you'll spot other mistakes. There are no bugs, but just mess. We started out with a listing to examine parameters, then we mixed parameters and columns. Our last listing is for a slightly more in-depth look at table columns. You’ll notice that we’ve delibarately removed the indication of whether a column is persisted, or is an identity column because that gives us false positives for our code smells. If you just want to browse your metadata for other reasons (and it can quite help in some circumstances) then uncomment them! ;WITH userColumns AS  ( SELECT   c.NAME AS columnName,  OBJECT_SCHEMA_NAME(c.object_ID) + '.' + OBJECT_NAME(c.object_ID) AS ObjectName,  REPLACE(t.name + ' '   + CASE WHEN is_computed = 1 THEN ' AS ' + --do DDL for a computed column          (SELECT definition FROM sys.computed_columns cc           WHERE cc.object_id = c.object_id AND cc.column_ID = c.column_ID)     --we may have to put in the length            WHEN t.Name IN ('char', 'varchar', 'nchar', 'nvarchar')             THEN '('               + CASE WHEN c.Max_Length = -1 THEN 'MAX'                ELSE CONVERT(VARCHAR(4),                    CASE WHEN t.Name IN ('nchar', 'nvarchar')                      THEN c.Max_Length / 2 ELSE c.Max_Length                    END)                END + ')'       WHEN t.name IN ('decimal', 'numeric')       THEN '(' + CONVERT(VARCHAR(4), c.precision) + ',' + CONVERT(VARCHAR(4), c.Scale) + ')'       ELSE ''      END + CASE WHEN c.is_rowguidcol = 1          THEN ' ROWGUIDCOL'          ELSE ''         END + CASE WHEN XML_collection_ID <> 0            THEN --deal with object schema names             '(' + CASE WHEN is_XML_Document = 1                THEN 'DOCUMENT '                ELSE 'CONTENT '               END + COALESCE((SELECT                QUOTENAME(ss.name) + '.' + QUOTENAME(sc.name)                FROM                sys.xml_schema_collections sc                INNER JOIN Sys.Schemas ss ON sc.schema_ID = ss.schema_ID                WHERE                sc.xml_collection_ID = c.XML_collection_ID),                'NULL') + ')'            ELSE ''           END + CASE WHEN is_identity = 1             THEN CASE WHEN OBJECTPROPERTY(object_id,                'IsUserTable') = 1 AND COLUMNPROPERTY(object_id,                c.name,                'IsIDNotForRepl') = 0 AND OBJECTPROPERTY(object_id,                'IsMSShipped') = 0                THEN ''                ELSE ' NOT FOR REPLICATION '               END             ELSE ''            END + CASE WHEN c.is_nullable = 0               THEN ' NOT NULL'               ELSE ' NULL'              END + CASE                WHEN c.default_object_id <> 0                THEN ' DEFAULT ' + object_Definition(c.default_object_id)                ELSE ''               END + CASE                WHEN c.collation_name IS NULL                THEN ''                WHEN c.collation_name <> (SELECT                collation_name                FROM                sys.databases                WHERE                name = DB_NAME()) COLLATE Latin1_General_CI_AS                THEN COALESCE(' COLLATE ' + c.collation_name,                '')                ELSE ''                END,'  ',' ') AS [DataType]FROM sys.columns c  INNER JOIN sys.types t ON c.user_Type_ID = t.user_Type_ID  WHERE OBJECT_SCHEMA_NAME(c.object_ID) <> 'sys')SELECT CONVERT(CHAR(80),objectName+'.'+columnName),DataType FROM UserColumnsWHERE columnName IN (SELECT columnName FROM UserColumns  GROUP BY columnName  HAVING MIN(Datatype)<>MAX(DataType))ORDER BY columnName If you take a look down the results against Adventureworks, you'll see once again that there are things to investigate, mostly, in the illustration, discrepancies between null and non-null datatypes So I here you ask, what about temporary variables within routines? If ever there was a source of elusive bugs, you'll find it there. Sadly, these temporary variables are not stored in the metadata so we'll have to find a more subtle way of flushing these out, and that will, I'm afraid, have to wait!

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  • Error when reloading supervisord: unix:///tmp/supervisor.sock no such file

    - by Yarin
    I'm running supervisord on my CentOS 6 box like so, /usr/bin/supervisord -c /etc/supervisord.conf and when I launch supervisorctl all process status are fine, but if I try to reload using supervisorctl I get unix:///tmp/supervisor.sock no such file I'm using the same config file I've used successfully on other boxes, and im running everything as root. I can't undesrtand what the problem is... Config file: ; Sample supervisor config file. [unix_http_server] file=/tmp/supervisor.sock ; (the path to the socket file) ;chmod=0700 ; socket file mode (default 0700) ;chown=nobody:nogroup ; socket file uid:gid owner ;username=user ; (default is no username (open server)) ;password=123 ; (default is no password (open server)) ;[inet_http_server] ; inet (TCP) server disabled by default ;port=127.0.0.1:9001 ; (ip_address:port specifier, *:port for all iface) ;username=user ; (default is no username (open server)) ;password=123 ; (default is no password (open server)) [supervisord] logfile=/tmp/supervisord.log ; (main log file;default $CWD/supervisord.log) logfile_maxbytes=50MB ; (max main logfile bytes b4 rotation;default 50MB) logfile_backups=10 ; (num of main logfile rotation backups;default 10) loglevel=info ; (log level;default info; others: debug,warn,trace) pidfile=/tmp/supervisord.pid ; (supervisord pidfile;default supervisord.pid) nodaemon=false ; (start in foreground if true;default false) minfds=1024 ; (min. avail startup file descriptors;default 1024) minprocs=200 ; (min. avail process descriptors;default 200) ;umask=022 ; (process file creation umask;default 022) ;user=chrism ; (default is current user, required if root) ;identifier=supervisor ; (supervisord identifier, default is 'supervisor') ;directory=/tmp ; (default is not to cd during start) ;nocleanup=true ; (don't clean up tempfiles at start;default false) ;childlogdir=/tmp ; ('AUTO' child log dir, default $TEMP) ;environment=KEY=value ; (key value pairs to add to environment) ;strip_ansi=false ; (strip ansi escape codes in logs; def. false) ; the below section must remain in the config file for RPC ; (supervisorctl/web interface) to work, additional interfaces may be ; added by defining them in separate rpcinterface: sections [rpcinterface:supervisor] supervisor.rpcinterface_factory = supervisor.rpcinterface:make_main_rpcinterface [supervisorctl] serverurl=unix:///tmp/supervisor.sock ; use a unix:// URL for a unix socket ;serverurl=http://127.0.0.1:9001 ; use an http:// url to specify an inet socket ;username=chris ; should be same as http_username if set ;password=123 ; should be same as http_password if set ;prompt=mysupervisor ; cmd line prompt (default "supervisor") ;history_file=~/.sc_history ; use readline history if available ; The below sample program section shows all possible program subsection values, ; create one or more 'real' program: sections to be able to control them under ; supervisor. ;[program:foo] ;command=/bin/cat [program:embed_scheduler] command=/opt/web-apps/mywebsite/custom_process.py process_name=%(program_name)s_%(process_num)d numprocs=3 ;[program:theprogramname] ;command=/bin/cat ; the program (relative uses PATH, can take args) ;process_name=%(program_name)s ; process_name expr (default %(program_name)s) ;numprocs=1 ; number of processes copies to start (def 1) ;directory=/tmp ; directory to cwd to before exec (def no cwd) ;umask=022 ; umask for process (default None) ;priority=999 ; the relative start priority (default 999) ;autostart=true ; start at supervisord start (default: true) ;autorestart=unexpected ; whether/when to restart (default: unexpected) ;startsecs=1 ; number of secs prog must stay running (def. 1) ;startretries=3 ; max # of serial start failures (default 3) ;exitcodes=0,2 ; 'expected' exit codes for process (default 0,2) ;stopsignal=QUIT ; signal used to kill process (default TERM) ;stopwaitsecs=10 ; max num secs to wait b4 SIGKILL (default 10) ;killasgroup=false ; SIGKILL the UNIX process group (def false) ;user=chrism ; setuid to this UNIX account to run the program ;redirect_stderr=true ; redirect proc stderr to stdout (default false) ;stdout_logfile=/a/path ; stdout log path, NONE for none; default AUTO ;stdout_logfile_maxbytes=1MB ; max # logfile bytes b4 rotation (default 50MB) ;stdout_logfile_backups=10 ; # of stdout logfile backups (default 10) ;stdout_capture_maxbytes=1MB ; number of bytes in 'capturemode' (default 0) ;stdout_events_enabled=false ; emit events on stdout writes (default false) ;stderr_logfile=/a/path ; stderr log path, NONE for none; default AUTO ;stderr_logfile_maxbytes=1MB ; max # logfile bytes b4 rotation (default 50MB) ;stderr_logfile_backups=10 ; # of stderr logfile backups (default 10) ;stderr_capture_maxbytes=1MB ; number of bytes in 'capturemode' (default 0) ;stderr_events_enabled=false ; emit events on stderr writes (default false) ;environment=A=1,B=2 ; process environment additions (def no adds) ;serverurl=AUTO ; override serverurl computation (childutils) ; The below sample eventlistener section shows all possible ; eventlistener subsection values, create one or more 'real' ; eventlistener: sections to be able to handle event notifications ; sent by supervisor. ;[eventlistener:theeventlistenername] ;command=/bin/eventlistener ; the program (relative uses PATH, can take args) ;process_name=%(program_name)s ; process_name expr (default %(program_name)s) ;numprocs=1 ; number of processes copies to start (def 1) ;events=EVENT ; event notif. types to subscribe to (req'd) ;buffer_size=10 ; event buffer queue size (default 10) ;directory=/tmp ; directory to cwd to before exec (def no cwd) ;umask=022 ; umask for process (default None) ;priority=-1 ; the relative start priority (default -1) ;autostart=true ; start at supervisord start (default: true) ;autorestart=unexpected ; whether/when to restart (default: unexpected) ;startsecs=1 ; number of secs prog must stay running (def. 1) ;startretries=3 ; max # of serial start failures (default 3) ;exitcodes=0,2 ; 'expected' exit codes for process (default 0,2) ;stopsignal=QUIT ; signal used to kill process (default TERM) ;stopwaitsecs=10 ; max num secs to wait b4 SIGKILL (default 10) ;killasgroup=false ; SIGKILL the UNIX process group (def false) ;user=chrism ; setuid to this UNIX account to run the program ;redirect_stderr=true ; redirect proc stderr to stdout (default false) ;stdout_logfile=/a/path ; stdout log path, NONE for none; default AUTO ;stdout_logfile_maxbytes=1MB ; max # logfile bytes b4 rotation (default 50MB) ;stdout_logfile_backups=10 ; # of stdout logfile backups (default 10) ;stdout_events_enabled=false ; emit events on stdout writes (default false) ;stderr_logfile=/a/path ; stderr log path, NONE for none; default AUTO ;stderr_logfile_maxbytes=1MB ; max # logfile bytes b4 rotation (default 50MB) ;stderr_logfile_backups ; # of stderr logfile backups (default 10) ;stderr_events_enabled=false ; emit events on stderr writes (default false) ;environment=A=1,B=2 ; process environment additions ;serverurl=AUTO ; override serverurl computation (childutils) ; The below sample group section shows all possible group values, ; create one or more 'real' group: sections to create "heterogeneous" ; process groups. ;[group:thegroupname] ;programs=progname1,progname2 ; each refers to 'x' in [program:x] definitions ;priority=999 ; the relative start priority (default 999) ; The [include] section can just contain the "files" setting. This ; setting can list multiple files (separated by whitespace or ; newlines). It can also contain wildcards. The filenames are ; interpreted as relative to this file. Included files *cannot* ; include files themselves. ;[include] ;files = relative/directory/*.ini

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