Search Results

Search found 10206 results on 409 pages for 'upk release'.

Page 405/409 | < Previous Page | 401 402 403 404 405 406 407 408 409  | Next Page >

  • JMS Step 6 - How to Set Up an AQ JMS (Advanced Queueing JMS) for SOA Purposes

    - by John-Brown.Evans
    JMS Step 6 - How to Set Up an AQ JMS (Advanced Queueing JMS) for SOA Purposes .jblist{list-style-type:disc;margin:0;padding:0;padding-left:0pt;margin-left:36pt} ol{margin:0;padding:0} .c17_6{vertical-align:top;width:468pt;border-style:solid;border-color:#000000;border-width:1pt;padding:5pt 5pt 5pt 5pt} .c5_6{vertical-align:top;border-style:solid;border-color:#000000;border-width:1pt;padding:0pt 5pt 0pt 5pt} .c6_6{vertical-align:top;width:156pt;border-style:solid;border-color:#000000;border-width:1pt;padding:5pt 5pt 5pt 5pt} .c15_6{background-color:#ffffff} .c10_6{color:#1155cc;text-decoration:underline} .c1_6{text-align:center;direction:ltr} .c0_6{line-height:1.0;direction:ltr} .c16_6{color:#666666;font-size:12pt} .c18_6{color:inherit;text-decoration:inherit} .c8_6{background-color:#f3f3f3} .c2_6{direction:ltr} .c14_6{font-size:8pt} .c11_6{font-size:10pt} .c7_6{font-weight:bold} .c12_6{height:0pt} .c3_6{height:11pt} .c13_6{border-collapse:collapse} .c4_6{font-family:"Courier New"} .c9_6{font-style:italic} .title{padding-top:24pt;line-height:1.15;text-align:left;color:#000000;font-size:36pt;font-family:"Arial";font-weight:bold;padding-bottom:6pt} .subtitle{padding-top:18pt;line-height:1.15;text-align:left;color:#666666;font-style:italic;font-size:24pt;font-family:"Georgia";padding-bottom:4pt} li{color:#000000;font-size:10pt;font-family:"Arial"} p{color:#000000;font-size:10pt;margin:0;font-family:"Arial"} h1{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:24pt;font-family:"Arial";font-weight:normal} h2{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:18pt;font-family:"Arial";font-weight:normal} h3{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:14pt;font-family:"Arial";font-weight:normal} h4{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:12pt;font-family:"Arial";font-weight:normal} h5{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:11pt;font-family:"Arial";font-weight:normal} h6{padding-top:0pt;line-height:1.15;text-align:left;color:#888;font-size:10pt;font-family:"Arial";font-weight:normal} This post continues the series of JMS articles which demonstrate how to use JMS queues in a SOA context. The previous posts were: JMS Step 1 - How to Create a Simple JMS Queue in Weblogic Server 11g JMS Step 2 - Using the QueueSend.java Sample Program to Send a Message to a JMS Queue JMS Step 3 - Using the QueueReceive.java Sample Program to Read a Message from a JMS Queue JMS Step 4 - How to Create an 11g BPEL Process Which Writes a Message Based on an XML Schema to a JMS Queue JMS Step 5 - How to Create an 11g BPEL Process Which Reads a Message Based on an XML Schema from a JMS Queue This example leads you through the creation of an Oracle database Advanced Queue and the related WebLogic server objects in order to use AQ JMS in connection with a SOA composite. If you have not already done so, I recommend you look at the previous posts in this series, as they include steps which this example builds upon. The following examples will demonstrate how to write and read from the queue from a SOA process. 1. Recap and Prerequisites In the previous examples, we created a JMS Queue, a Connection Factory and a Connection Pool in the WebLogic Server Console. Then we wrote and deployed BPEL composites, which enqueued and dequeued a simple XML payload. AQ JMS allows you to interoperate with database Advanced Queueing via JMS in WebLogic server and therefore take advantage of database features, while maintaining compliance with the JMS architecture. AQ JMS uses the WebLogic JMS Foreign Server framework. A full description of this functionality can be found in the following Oracle documentation Oracle® Fusion Middleware Configuring and Managing JMS for Oracle WebLogic Server 11g Release 1 (10.3.6) Part Number E13738-06 7. Interoperating with Oracle AQ JMS http://docs.oracle.com/cd/E23943_01/web.1111/e13738/aq_jms.htm#CJACBCEJ For easier reference, this sample will use the same names for the objects as in the above document, except for the name of the database user, as it is possible that this user already exists in your database. We will create the following objects Database Objects Name Type AQJMSUSER Database User MyQueueTable Advanced Queue (AQ) Table UserQueue Advanced Queue WebLogic Server Objects Object Name Type JNDI Name aqjmsuserDataSource Data Source jdbc/aqjmsuserDataSource AqJmsModule JMS System Module AqJmsForeignServer JMS Foreign Server AqJmsForeignServerConnectionFactory JMS Foreign Server Connection Factory AqJmsForeignServerConnectionFactory AqJmsForeignDestination AQ JMS Foreign Destination queue/USERQUEUE eis/aqjms/UserQueue Connection Pool eis/aqjms/UserQueue 2. Create a Database User and Advanced Queue The following steps can be executed in the database client of your choice, e.g. JDeveloper or SQL Developer. The examples below use SQL*Plus. Log in to the database as a DBA user, for example SYSTEM or SYS. Create the AQJMSUSER user and grant privileges to enable the user to create AQ objects. Create Database User and Grant AQ Privileges sqlplus system/password as SYSDBA GRANT connect, resource TO aqjmsuser IDENTIFIED BY aqjmsuser; GRANT aq_user_role TO aqjmsuser; GRANT execute ON sys.dbms_aqadm TO aqjmsuser; GRANT execute ON sys.dbms_aq TO aqjmsuser; GRANT execute ON sys.dbms_aqin TO aqjmsuser; GRANT execute ON sys.dbms_aqjms TO aqjmsuser; Create the Queue Table and Advanced Queue and Start the AQ The following commands are executed as the aqjmsuser database user. Create the Queue Table connect aqjmsuser/aqjmsuser; BEGIN dbms_aqadm.create_queue_table ( queue_table = 'myQueueTable', queue_payload_type = 'sys.aq$_jms_text_message', multiple_consumers = false ); END; / Create the AQ BEGIN dbms_aqadm.create_queue ( queue_name = 'userQueue', queue_table = 'myQueueTable' ); END; / Start the AQ BEGIN dbms_aqadm.start_queue ( queue_name = 'userQueue'); END; / The above commands can be executed in a single PL/SQL block, but are shown as separate blocks in this example for ease of reference. You can verify the queue by executing the SQL command SELECT object_name, object_type FROM user_objects; which should display the following objects: OBJECT_NAME OBJECT_TYPE ------------------------------ ------------------- SYS_C0056513 INDEX SYS_LOB0000170822C00041$$ LOB SYS_LOB0000170822C00040$$ LOB SYS_LOB0000170822C00037$$ LOB AQ$_MYQUEUETABLE_T INDEX AQ$_MYQUEUETABLE_I INDEX AQ$_MYQUEUETABLE_E QUEUE AQ$_MYQUEUETABLE_F VIEW AQ$MYQUEUETABLE VIEW MYQUEUETABLE TABLE USERQUEUE QUEUE Similarly, you can view the objects in JDeveloper via a Database Connection to the AQJMSUSER. 3. Configure WebLogic Server and Add JMS Objects All these steps are executed from the WebLogic Server Administration Console. Log in as the webLogic user. Configure a WebLogic Data Source The data source is required for the database connection to the AQ created above. Navigate to domain > Services > Data Sources and press New then Generic Data Source. Use the values:Name: aqjmsuserDataSource JNDI Name: jdbc/aqjmsuserDataSource Database type: Oracle Database Driver: *Oracle’ Driver (Thin XA) for Instance connections; Versions:9.0.1 and later Connection Properties: Enter the connection information to the database containing the AQ created above and enter aqjmsuser for the User Name and Password. Press Test Configuration to verify the connection details and press Next. Target the data source to the soa server. The data source will be displayed in the list. It is a good idea to test the data source at this stage. Click on aqjmsuserDataSource, select Monitoring > Testing > soa_server1 and press Test Data Source. The result is displayed at the top of the page. Configure a JMS System Module The JMS system module is required to host the JMS foreign server for AQ resources. Navigate to Services > Messaging > JMS Modules and select New. Use the values: Name: AqJmsModule (Leave Descriptor File Name and Location in Domain empty.) Target: soa_server1 Click Finish. The other resources will be created in separate steps. The module will be displayed in the list.   Configure a JMS Foreign Server A foreign server is required in order to reference a 3rd-party JMS provider, in this case the database AQ, within a local WebLogic server JNDI tree. Navigate to Services > Messaging > JMS Modules and select (click on) AqJmsModule to configure it. Under Summary of Resources, select New then Foreign Server. Name: AqJmsForeignServer Targets: The foreign server is targeted automatically to soa_server1, based on the JMS module’s target. Press Finish to create the foreign server. The foreign server resource will be listed in the Summary of Resources for the AqJmsModule, but needs additional configuration steps. Click on AqJmsForeignServer and select Configuration > General to complete the configuration: JNDI Initial Context Factory: oracle.jms.AQjmsInitialContextFactory JNDI Connection URL: <empty> JNDI Properties Credential:<empty> Confirm JNDI Properties Credential: <empty> JNDI Properties: datasource=jdbc/aqjmsuserDataSource This is an important property. It is the JNDI name of the data source created above, which points to the AQ schema in the database and must be entered as a name=value pair, as in this example, e.g. datasource=jdbc/aqjmsuserDataSource, including the “datasource=” property name. Default Targeting Enabled: Leave this value checked. Press Save to save the configuration. At this point it is a good idea to verify that the data source was written correctly to the config file. In a terminal window, navigate to $MIDDLEWARE_HOME/user_projects/domains/soa_domain/config/jms  and open the file aqjmsmodule-jms.xml . The foreign server configuration should contain the datasource name-value pair, as follows:   <foreign-server name="AqJmsForeignServer">         <default-targeting-enabled>true</default-targeting-enabled>         <initial-context-factory>oracle.jms.AQjmsInitialContextFactory</initial-context-factory>         <jndi-property>           <key> datasource </key>           <value> jdbc/aqjmsuserDataSource </value>         </jndi-property>   </foreign-server> </weblogic-jms> Configure a JMS Foreign Server Connection Factory When creating the foreign server connection factory, you enter local and remote JNDI names. The name of the connection factory itself and the local JNDI name are arbitrary, but the remote JNDI name must match a specific format, depending on the type of queue or topic to be accessed in the database. This is very important and if the incorrect value is used, the connection to the queue will not be established and the error messages you get will not immediately reflect the cause of the error. The formats required (Remote JNDI names for AQ JMS Connection Factories) are described in the section Configure AQ Destinations  of the Oracle® Fusion Middleware Configuring and Managing JMS for Oracle WebLogic Server document mentioned earlier. In this example, the remote JNDI name used is   XAQueueConnectionFactory  because it matches the AQ and data source created earlier, i.e. thin with AQ. Navigate to JMS Modules > AqJmsModule > AqJmsForeignServer > Connection Factories then New.Name: AqJmsForeignServerConnectionFactory Local JNDI Name: AqJmsForeignServerConnectionFactory Note: this local JNDI name is the JNDI name which your client application, e.g. a later BPEL process, will use to access this connection factory. Remote JNDI Name: XAQueueConnectionFactory Press OK to save the configuration. Configure an AQ JMS Foreign Server Destination A foreign server destination maps the JNDI name on the foreign JNDI provider to the respective local JNDI name, allowing the foreign JNDI name to be accessed via the local server. As with the foreign server connection factory, the local JNDI name is arbitrary (but must be unique), but the remote JNDI name must conform to a specific format defined in the section Configure AQ Destinations  of the Oracle® Fusion Middleware Configuring and Managing JMS for Oracle WebLogic Server document mentioned earlier. In our example, the remote JNDI name is Queues/USERQUEUE , because it references a queue (as opposed to a topic) with the name USERQUEUE. We will name the local JNDI name queue/USERQUEUE, which is a little confusing (note the missing “s” in “queue), but conforms better to the JNDI nomenclature in our SOA server and also allows us to differentiate between the local and remote names for demonstration purposes. Navigate to JMS Modules > AqJmsModule > AqJmsForeignServer > Destinations and select New.Name: AqJmsForeignDestination Local JNDI Name: queue/USERQUEUE Remote JNDI Name:Queues/USERQUEUE After saving the foreign destination configuration, this completes the JMS part of the configuration. We still need to configure the JMS adapter in order to be able to access the queue from a BPEL processt. 4. Create a JMS Adapter Connection Pool in Weblogic Server Create the Connection Pool Access to the AQ JMS queue from a BPEL or other SOA process in our example is done via a JMS adapter. To enable this, the JmsAdapter in WebLogic server needs to be configured to have a connection pool which points to the local connection factory JNDI name which was created earlier. Navigate to Deployments > Next and select (click on) the JmsAdapter. Select Configuration > Outbound Connection Pools and New. Check the radio button for oracle.tip.adapter.jms.IJmsConnectionFactory and press Next. JNDI Name: eis/aqjms/UserQueue Press Finish Expand oracle.tip.adapter.jms.IJmsConnectionFactory and click on eis/aqjms/UserQueue to configure it. The ConnectionFactoryLocation must point to the foreign server’s local connection factory name created earlier. In our example, this is AqJmsForeignServerConnectionFactory . As a reminder, this connection factory is located under JMS Modules > AqJmsModule > AqJmsForeignServer > Connection Factories and the value needed here is under Local JNDI Name. Enter AqJmsForeignServerConnectionFactory  into the Property Value field for ConnectionFactoryLocation. You must then press Return/Enter then Save for the value to be accepted. If your WebLogic server is running in Development mode, you should see the message that the changes have been activated and the deployment plan successfully updated. If not, then you will manually need to activate the changes in the WebLogic server console.Although the changes have been activated, the JmsAdapter needs to be redeployed in order for the changes to become effective. This should be confirmed by the message Remember to update your deployment to reflect the new plan when you are finished with your changes. Redeploy the JmsAdapter Navigate back to the Deployments screen, either by selecting it in the left-hand navigation tree or by selecting the “Summary of Deployments” link in the breadcrumbs list at the top of the screen. Then select the checkbox next to JmsAdapter and press the Update button. On the Update Application Assistant page, select “Redeploy this application using the following deployment files” and press Finish. After a few seconds you should get the message that the selected deployments were updated. The JMS adapter configuration is complete and it can now be used to access the AQ JMS queue. You can verify that the JNDI name was created correctly, by navigating to Environment > Servers > soa_server1 and View JNDI Tree. Then scroll down in the JNDI Tree Structure to eis and select aqjms. This concludes the sample. In the following post, I will show you how to create a BPEL process which sends a message to this advanced queue via JMS. Best regards John-Brown Evans Oracle Technology Proactive Support Delivery

    Read the article

  • The Execute SQL Task

    In this article we are going to take you through the Execute SQL Task in SQL Server Integration Services for SQL Server 2005 (although it appies just as well to SQL Server 2008).  We will be covering all the essentials that you will need to know to effectively use this task and make it as flexible as possible. The things we will be looking at are as follows: A tour of the Task. The properties of the Task. After looking at these introductory topics we will then get into some examples. The examples will show different types of usage for the task: Returning a single value from a SQL query with two input parameters. Returning a rowset from a SQL query. Executing a stored procedure and retrieveing a rowset, a return value, an output parameter value and passing in an input parameter. Passing in the SQL Statement from a variable. Passing in the SQL Statement from a file. Tour Of The Task Before we can start to use the Execute SQL Task in our packages we are going to need to locate it in the toolbox. Let's do that now. Whilst in the Control Flow section of the package expand your toolbox and locate the Execute SQL Task. Below is how we found ours. Now drag the task onto the designer. As you can see from the following image we have a validation error appear telling us that no connection manager has been assigned to the task. This can be easily remedied by creating a connection manager. There are certain types of connection manager that are compatable with this task so we cannot just create any connection manager and these are detailed in a few graphics time. Double click on the task itself to take a look at the custom user interface provided to us for this task. The task will open on the general tab as shown below. Take a bit of time to have a look around here as throughout this article we will be revisting this page many times. Whilst on the general tab, drop down the combobox next to the ConnectionType property. In here you will see the types of connection manager which this task will accept. As with SQL Server 2000 DTS, SSIS allows you to output values from this task in a number of formats. Have a look at the combobox next to the Resultset property. The major difference here is the ability to output into XML. If you drop down the combobox next to the SQLSourceType property you will see the ways in which you can pass a SQL Statement into the task itself. We will have examples of each of these later on but certainly when we saw these for the first time we were very excited. Next to the SQLStatement property if you click in the empty box next to it you will see ellipses appear. Click on them and you will see the very basic query editor that becomes available to you. Alternatively after you have specified a connection manager for the task you can click on the Build Query button to bring up a completely different query editor. This is slightly inconsistent. Once you've finished looking around the general tab, move on to the next tab which is the parameter mapping tab. We shall, again, be visiting this tab throughout the article but to give you an initial heads up this is where you define the input, output and return values from your task. Note this is not where you specify the resultset. If however you now move on to the ResultSet tab this is where you define what variable will receive the output from your SQL Statement in whatever form that is. Property Expressions are one of the most amazing things to happen in SSIS and they will not be covered here as they deserve a whole article to themselves. Watch out for this as their usefulness will astound you. For a more detailed discussion of what should be the parameter markers in the SQL Statements on the General tab and how to map them to variables on the Parameter Mapping tab see Working with Parameters and Return Codes in the Execute SQL Task. Task Properties There are two places where you can specify the properties for your task. One is in the task UI itself and the other is in the property pane which will appear if you right click on your task and select Properties from the context menu. We will be doing plenty of property setting in the UI later so let's take a moment to have a look at the property pane. Below is a graphic showing our properties pane. Now we shall take you through all the properties and tell you exactly what they mean. A lot of these properties you will see across all tasks as well as the package because of everything's base structure The Container. BypassPrepare Should the statement be prepared before sending to the connection manager destination (True/False) Connection This is simply the name of the connection manager that the task will use. We can get this from the connection manager tray at the bottom of the package. DelayValidation Really interesting property and it tells the task to not validate until it actually executes. A usage for this may be that you are operating on table yet to be created but at runtime you know the table will be there. Description Very simply the description of your Task. Disable Should the task be enabled or not? You can also set this through a context menu by right clicking on the task itself. DisableEventHandlers As a result of events that happen in the task, should the event handlers for the container fire? ExecValueVariable The variable assigned here will get or set the execution value of the task. Expressions Expressions as we mentioned earlier are a really powerful tool in SSIS and this graphic below shows us a small peek of what you can do. We select a property on the left and assign an expression to the value of that property on the right causing the value to be dynamically changed at runtime. One of the most obvious uses of this is that the property value can be built dynamically from within the package allowing you a great deal of flexibility FailPackageOnFailure If this task fails does the package? FailParentOnFailure If this task fails does the parent container? A task can he hosted inside another container i.e. the For Each Loop Container and this would then be the parent. ForcedExecutionValue This property allows you to hard code an execution value for the task. ForcedExecutionValueType What is the datatype of the ForcedExecutionValue? ForceExecutionResult Force the task to return a certain execution result. This could then be used by the workflow constraints. Possible values are None, Success, Failure and Completion. ForceExecutionValue Should we force the execution result? IsolationLevel This is the transaction isolation level of the task. IsStoredProcedure Certain optimisations are made by the task if it knows that the query is a Stored Procedure invocation. The docs say this will always be false unless the connection is an ADO connection. LocaleID Gets or sets the LocaleID of the container. LoggingMode Should we log for this container and what settings should we use? The value choices are UseParentSetting, Enabled and Disabled. MaximumErrorCount How many times can the task fail before we call it a day? Name Very simply the name of the task. ResultSetType How do you want the results of your query returned? The choices are ResultSetType_None, ResultSetType_SingleRow, ResultSetType_Rowset and ResultSetType_XML. SqlStatementSource Your Query/SQL Statement. SqlStatementSourceType The method of specifying the query. Your choices here are DirectInput, FileConnection and Variables TimeOut How long should the task wait to receive results? TransactionOption How should the task handle being asked to join a transaction? Usage Examples As we move through the examples we will only cover in them what we think you must know and what we think you should see. This means that some of the more elementary steps like setting up variables will be covered in the early examples but skipped and simply referred to in later ones. All these examples used the AventureWorks database that comes with SQL Server 2005. Returning a Single Value, Passing in Two Input Parameters So the first thing we are going to do is add some variables to our package. The graphic below shows us those variables having been defined. Here the CountOfEmployees variable will be used as the output from the query and EndDate and StartDate will be used as input parameters. As you can see all these variables have been scoped to the package. Scoping allows us to have domains for variables. Each container has a scope and remember a package is a container as well. Variable values of the parent container can be seen in child containers but cannot be passed back up to the parent from a child. Our following graphic has had a number of changes made. The first of those changes is that we have created and assigned an OLEDB connection manager to this Task ExecuteSQL Task Connection. The next thing is we have made sure that the SQLSourceType property is set to Direct Input as we will be writing in our statement ourselves. We have also specified that only a single row will be returned from this query. The expressions we typed in was: SELECT COUNT(*) AS CountOfEmployees FROM HumanResources.Employee WHERE (HireDate BETWEEN ? AND ?) Moving on now to the Parameter Mapping tab this is where we are going to tell the task about our input paramaters. We Add them to the window specifying their direction and datatype. A quick word here about the structure of the variable name. As you can see SSIS has preceeded the variable with the word user. This is a default namespace for variables but you can create your own. When defining your variables if you look at the variables window title bar you will see some icons. If you hover over the last one on the right you will see it says "Choose Variable Columns". If you click the button you will see a list of checkbox options and one of them is namespace. after checking this you will see now where you can define your own namespace. The next tab, result set, is where we need to get back the value(s) returned from our statement and assign to a variable which in our case is CountOfEmployees so we can use it later perhaps. Because we are only returning a single value then if you remember from earlier we are allowed to assign a name to the resultset but it must be the name of the column (or alias) from the query. A really cool feature of Business Intelligence Studio being hosted by Visual Studio is that we get breakpoint support for free. In our package we set a Breakpoint so we can break the package and have a look in a watch window at the variable values as they appear to our task and what the variable value of our resultset is after the task has done the assignment. Here's that window now. As you can see the count of employess that matched the data range was 2. Returning a Rowset In this example we are going to return a resultset back to a variable after the task has executed not just a single row single value. There are no input parameters required so the variables window is nice and straight forward. One variable of type object. Here is the statement that will form the soure for our Resultset. select p.ProductNumber, p.name, pc.Name as ProductCategoryNameFROM Production.ProductCategory pcJOIN Production.ProductSubCategory pscON pc.ProductCategoryID = psc.ProductCategoryIDJOIN Production.Product pON psc.ProductSubCategoryID = p.ProductSubCategoryID We need to make sure that we have selected Full result set as the ResultSet as shown below on the task's General tab. Because there are no input parameters we can skip the parameter mapping tab and move straight to the Result Set tab. Here we need to Add our variable defined earlier and map it to the result name of 0 (remember we covered this earlier) Once we run the task we can again set a breakpoint and have a look at the values coming back from the task. In the following graphic you can see the result set returned to us as a COM object. We can do some pretty interesting things with this COM object and in later articles that is exactly what we shall be doing. Return Values, Input/Output Parameters and Returning a Rowset from a Stored Procedure This example is pretty much going to give us a taste of everything. We have already covered in the previous example how to specify the ResultSet to be a Full result set so we will not cover it again here. For this example we are going to need 4 variables. One for the return value, one for the input parameter, one for the output parameter and one for the result set. Here is the statement we want to execute. Note how much cleaner it is than if you wanted to do it using the current version of DTS. In the Parameter Mapping tab we are going to Add our variables and specify their direction and datatypes. In the Result Set tab we can now map our final variable to the rowset returned from the stored procedure. It really is as simple as that and we were amazed at how much easier it is than in DTS 2000. Passing in the SQL Statement from a Variable SSIS as we have mentioned is hugely more flexible than its predecessor and one of the things you will notice when moving around the tasks and the adapters is that a lot of them accept a variable as an input for something they need. The ExecuteSQL task is no different. It will allow us to pass in a string variable as the SQL Statement. This variable value could have been set earlier on from inside the package or it could have been populated from outside using a configuration. The ResultSet property is set to single row and we'll show you why in a second when we look at the variables. Note also the SQLSourceType property. Here's the General Tab again. Looking at the variable we have in this package you can see we have only two. One for the return value from the statement and one which is obviously for the statement itself. Again we need to map the Result name to our variable and this can be a named Result Name (The column name or alias returned by the query) and not 0. The expected result into our variable should be the amount of rows in the Person.Contact table and if we look in the watch window we see that it is.   Passing in the SQL Statement from a File The final example we are going to show is a really interesting one. We are going to pass in the SQL statement to the task by using a file connection manager. The file itself contains the statement to run. The first thing we are going to need to do is create our file connection mananger to point to our file. Click in the connections tray at the bottom of the designer, right click and choose "New File Connection" As you can see in the graphic below we have chosen to use an existing file and have passed in the name as well. Have a look around at the other "Usage Type" values available whilst you are here. Having set that up we can now see in the connection manager tray our file connection manager sitting alongside our OLE-DB connection we have been using for the rest of these examples. Now we can go back to the familiar General Tab to set up how the task will accept our file connection as the source. All the other properties in this task are set up exactly as we have been doing for other examples depending on the options chosen so we will not cover them again here.   We hope you will agree that the Execute SQL Task has changed considerably in this release from its DTS predecessor. It has a lot of options available but once you have configured it a few times you get to learn what needs to go where. We hope you have found this article useful.

    Read the article

  • Metro: Introduction to CSS 3 Grid Layout

    - by Stephen.Walther
    The purpose of this blog post is to provide you with a quick introduction to the new W3C CSS 3 Grid Layout standard. You can use CSS Grid Layout in Metro style applications written with JavaScript to lay out the content of an HTML page. CSS Grid Layout provides you with all of the benefits of using HTML tables for layout without requiring you to actually use any HTML table elements. Doing Page Layouts without Tables Back in the 1990’s, if you wanted to create a fancy website, then you would use HTML tables for layout. For example, if you wanted to create a standard three-column page layout then you would create an HTML table with three columns like this: <table height="100%"> <tr> <td valign="top" width="300px" bgcolor="red"> Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column </td> <td valign="top" bgcolor="green"> Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column </td> <td valign="top" width="300px" bgcolor="blue"> Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column </td> </tr> </table> When the table above gets rendered out to a browser, you end up with the following three-column layout: The width of the left and right columns is fixed – the width of the middle column expands or contracts depending on the width of the browser. Sometime around the year 2005, everyone decided that using tables for layout was a bad idea. Instead of using tables for layout — it was collectively decided by the spirit of the Web — you should use Cascading Style Sheets instead. Why is using HTML tables for layout bad? Using tables for layout breaks the semantics of the TABLE element. A TABLE element should be used only for displaying tabular information such as train schedules or moon phases. Using tables for layout is bad for accessibility (The Web Content Accessibility Guidelines 1.0 is explicit about this) and using tables for layout is bad for separating content from layout (see http://CSSZenGarden.com). Post 2005, anyone who used HTML tables for layout were encouraged to hold their heads down in shame. That’s all well and good, but the problem with using CSS for layout is that it can be more difficult to work with CSS than HTML tables. For example, to achieve a standard three-column layout, you either need to use absolute positioning or floats. Here’s a three-column layout with floats: <style type="text/css"> #container { min-width: 800px; } #leftColumn { float: left; width: 300px; height: 100%; background-color:red; } #middleColumn { background-color:green; height: 100%; } #rightColumn { float: right; width: 300px; height: 100%; background-color:blue; } </style> <div id="container"> <div id="rightColumn"> Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column </div> <div id="leftColumn"> Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column </div> <div id="middleColumn"> Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column </div> </div> The page above contains four DIV elements: a container DIV which contains a leftColumn, middleColumn, and rightColumn DIV. The leftColumn DIV element is floated to the left and the rightColumn DIV element is floated to the right. Notice that the rightColumn DIV appears in the page before the middleColumn DIV – this unintuitive ordering is necessary to get the floats to work correctly (see http://stackoverflow.com/questions/533607/css-three-column-layout-problem). The page above (almost) works with the most recent versions of most browsers. For example, you get the correct three-column layout in both Firefox and Chrome: And the layout mostly works with Internet Explorer 9 except for the fact that for some strange reason the min-width doesn’t work so when you shrink the width of your browser, you can get the following unwanted layout: Notice how the middle column (the green column) bleeds to the left and right. People have solved these issues with more complicated CSS. For example, see: http://matthewjamestaylor.com/blog/holy-grail-no-quirks-mode.htm But, at this point, no one could argue that using CSS is easier or more intuitive than tables. It takes work to get a layout with CSS and we know that we could achieve the same layout more easily using HTML tables. Using CSS Grid Layout CSS Grid Layout is a new W3C standard which provides you with all of the benefits of using HTML tables for layout without the disadvantage of using an HTML TABLE element. In other words, CSS Grid Layout enables you to perform table layouts using pure Cascading Style Sheets. The CSS Grid Layout standard is still in a “Working Draft” state (it is not finalized) and it is located here: http://www.w3.org/TR/css3-grid-layout/ The CSS Grid Layout standard is only supported by Internet Explorer 10 and there are no signs that any browser other than Internet Explorer will support this standard in the near future. This means that it is only practical to take advantage of CSS Grid Layout when building Metro style applications with JavaScript. Here’s how you can create a standard three-column layout using a CSS Grid Layout: <!DOCTYPE html> <html> <head> <style type="text/css"> html, body, #container { height: 100%; padding: 0px; margin: 0px; } #container { display: -ms-grid; -ms-grid-columns: 300px auto 300px; -ms-grid-rows: 100%; } #leftColumn { -ms-grid-column: 1; background-color:red; } #middleColumn { -ms-grid-column: 2; background-color:green; } #rightColumn { -ms-grid-column: 3; background-color:blue; } </style> </head> <body> <div id="container"> <div id="leftColumn"> Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column </div> <div id="middleColumn"> Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column </div> <div id="rightColumn"> Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column </div> </div> </body> </html> When the page above is rendered in Internet Explorer 10, you get a standard three-column layout: The page above contains four DIV elements: a container DIV which contains a leftColumn DIV, middleColumn DIV, and rightColumn DIV. The container DIV is set to Grid display mode with the following CSS rule: #container { display: -ms-grid; -ms-grid-columns: 300px auto 300px; -ms-grid-rows: 100%; } The display property is set to the value “-ms-grid”. This property causes the container DIV to lay out its child elements in a grid. (Notice that you use “-ms-grid” instead of “grid”. The “-ms-“ prefix is used because the CSS Grid Layout standard is still preliminary. This implementation only works with IE10 and it might change before the final release.) The grid columns and rows are defined with the “-ms-grid-columns” and “-ms-grid-rows” properties. The style rule above creates a grid with three columns and one row. The left and right columns are fixed sized at 300 pixels. The middle column sizes automatically depending on the remaining space available. The leftColumn, middleColumn, and rightColumn DIVs are positioned within the container grid element with the following CSS rules: #leftColumn { -ms-grid-column: 1; background-color:red; } #middleColumn { -ms-grid-column: 2; background-color:green; } #rightColumn { -ms-grid-column: 3; background-color:blue; } The “-ms-grid-column” property is used to specify the column associated with the element selected by the style sheet selector. The leftColumn DIV is positioned in the first grid column, the middleColumn DIV is positioned in the second grid column, and the rightColumn DIV is positioned in the third grid column. I find using CSS Grid Layout to be just as intuitive as using an HTML table for layout. You define your columns and rows and then you position different elements within these columns and rows. Very straightforward. Creating Multiple Columns and Rows In the previous section, we created a super simple three-column layout. This layout contained only a single row. In this section, let’s create a slightly more complicated layout which contains more than one row: The following page contains a header row, a content row, and a footer row. The content row contains three columns: <!DOCTYPE html> <html> <head> <style type="text/css"> html, body, #container { height: 100%; padding: 0px; margin: 0px; } #container { display: -ms-grid; -ms-grid-columns: 300px auto 300px; -ms-grid-rows: 100px 1fr 100px; } #header { -ms-grid-column: 1; -ms-grid-column-span: 3; -ms-grid-row: 1; background-color: yellow; } #leftColumn { -ms-grid-column: 1; -ms-grid-row: 2; background-color:red; } #middleColumn { -ms-grid-column: 2; -ms-grid-row: 2; background-color:green; } #rightColumn { -ms-grid-column: 3; -ms-grid-row: 2; background-color:blue; } #footer { -ms-grid-column: 1; -ms-grid-column-span: 3; -ms-grid-row: 3; background-color: orange; } </style> </head> <body> <div id="container"> <div id="header"> Header, Header, Header </div> <div id="leftColumn"> Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column </div> <div id="middleColumn"> Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column </div> <div id="rightColumn"> Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column </div> <div id="footer"> Footer, Footer, Footer </div> </div> </body> </html> In the page above, the grid layout is created with the following rule which creates a grid with three rows and three columns: #container { display: -ms-grid; -ms-grid-columns: 300px auto 300px; -ms-grid-rows: 100px 1fr 100px; } The header is created with the following rule: #header { -ms-grid-column: 1; -ms-grid-column-span: 3; -ms-grid-row: 1; background-color: yellow; } The header is positioned in column 1 and row 1. Furthermore, notice that the “-ms-grid-column-span” property is used to span the header across three columns. CSS Grid Layout and Fractional Units When you use CSS Grid Layout, you can take advantage of fractional units. Fractional units provide you with an easy way of dividing up remaining space in a page. Imagine, for example, that you want to create a three-column page layout. You want the size of the first column to be fixed at 200 pixels and you want to divide the remaining space among the remaining three columns. The width of the second column is equal to the combined width of the third and fourth columns. The following CSS rule creates four columns with the desired widths: #container { display: -ms-grid; -ms-grid-columns: 200px 2fr 1fr 1fr; -ms-grid-rows: 1fr; } The fr unit represents a fraction. The grid above contains four columns. The second column is two times the size (2fr) of the third (1fr) and fourth (1fr) columns. When you use the fractional unit, the remaining space is divided up using fractional amounts. Notice that the single row is set to a height of 1fr. The single grid row gobbles up the entire vertical space. Here’s the entire HTML page: <!DOCTYPE html> <html> <head> <style type="text/css"> html, body, #container { height: 100%; padding: 0px; margin: 0px; } #container { display: -ms-grid; -ms-grid-columns: 200px 2fr 1fr 1fr; -ms-grid-rows: 1fr; } #firstColumn { -ms-grid-column: 1; background-color:red; } #secondColumn { -ms-grid-column: 2; background-color:green; } #thirdColumn { -ms-grid-column: 3; background-color:blue; } #fourthColumn { -ms-grid-column: 4; background-color:orange; } </style> </head> <body> <div id="container"> <div id="firstColumn"> First Column, First Column, First Column </div> <div id="secondColumn"> Second Column, Second Column, Second Column </div> <div id="thirdColumn"> Third Column, Third Column, Third Column </div> <div id="fourthColumn"> Fourth Column, Fourth Column, Fourth Column </div> </div> </body> </html>   Summary There is more in the CSS 3 Grid Layout standard than discussed in this blog post. My goal was to describe the basics. If you want to learn more than you can read through the entire standard at http://www.w3.org/TR/css3-grid-layout/ In this blog post, I described some of the difficulties that you might encounter when attempting to replace HTML tables with Cascading Style Sheets when laying out a web page. I explained how you can take advantage of the CSS 3 Grid Layout standard to avoid these problems when building Metro style applications using JavaScript. CSS 3 Grid Layout provides you with all of the benefits of using HTML tables for laying out a page without requiring you to use HTML table elements.

    Read the article

  • Entity Framework Code-First, OData & Windows Phone Client

    - by Jon Galloway
    Entity Framework Code-First is the coolest thing since sliced bread, Windows  Phone is the hottest thing since Tickle-Me-Elmo and OData is just too great to ignore. As part of the Full Stack project, we wanted to put them together, which turns out to be pretty easy… once you know how.   EF Code-First CTP5 is available now and there should be very few breaking changes in the release edition, which is due early in 2011.  Note: EF Code-First evolved rapidly and many of the existing documents and blog posts which were written with earlier versions, may now be obsolete or at least misleading.   Code-First? With traditional Entity Framework you start with a database and from that you generate “entities” – classes that bridge between the relational database and your object oriented program. With Code-First (Magic-Unicorn) (see Hanselman’s write up and this later write up by Scott Guthrie) the Entity Framework looks at classes you created and says “if I had created these classes, the database would have to have looked like this…” and creates the database for you! By deriving your entity collections from DbSet and exposing them via a class that derives from DbContext, you "turn on" database backing for your POCO with a minimum of code and no hidden designer or configuration files. POCO == Plain Old CLR Objects Your entity objects can be used throughout your applications - in web applications, console applications, Silverlight and Windows Phone applications, etc. In our case, we'll want to read and update data from a Windows Phone client application, so we'll expose the entities through a DataService and hook the Windows Phone client application to that data via proxies.  Piece of Pie.  Easy as cake. The Demo Architecture To see this at work, we’ll create an ASP.NET/MVC application which will act as the host for our Data Service.  We’ll create an incredibly simple data layer using EF Code-First on top of SQLCE4 and we’ll expose the data in a WCF Data Service using the oData protocol.  Our Windows Phone 7 client will instantiate  the data context via a URI and load the data asynchronously. Setting up the Server project with MVC 3, EF Code First, and SQL CE 4 Create a new application of type ASP.NET MVC 3 and name it DeadSimpleServer.  We need to add the latest SQLCE4 and Entity Framework Code First CTP's to our project. Fortunately, NuGet makes that really easy. Open the Package Manager Console (View / Other Windows / Package Manager Console) and type in "Install-Package EFCodeFirst.SqlServerCompact" at the PM> command prompt. Since NuGet handles dependencies for you, you'll see that it installs everything you need to use Entity Framework Code First in your project. PM> install-package EFCodeFirst.SqlServerCompact 'SQLCE (= 4.0.8435.1)' not installed. Attempting to retrieve dependency from source... Done 'EFCodeFirst (= 0.8)' not installed. Attempting to retrieve dependency from source... Done 'WebActivator (= 1.0.0.0)' not installed. Attempting to retrieve dependency from source... Done You are downloading SQLCE from Microsoft, the license agreement to which is available at http://173.203.67.148/licenses/SQLCE/EULA_ENU.rtf. Check the package for additional dependencies, which may come with their own license agreement(s). Your use of the package and dependencies constitutes your acceptance of their license agreements. If you do not accept the license agreement(s), then delete the relevant components from your device. Successfully installed 'SQLCE 4.0.8435.1' You are downloading EFCodeFirst from Microsoft, the license agreement to which is available at http://go.microsoft.com/fwlink/?LinkID=206497. Check the package for additional dependencies, which may come with their own license agreement(s). Your use of the package and dependencies constitutes your acceptance of their license agreements. If you do not accept the license agreement(s), then delete the relevant components from your device. Successfully installed 'EFCodeFirst 0.8' Successfully installed 'WebActivator 1.0.0.0' You are downloading EFCodeFirst.SqlServerCompact from Microsoft, the license agreement to which is available at http://173.203.67.148/licenses/SQLCE/EULA_ENU.rtf. Check the package for additional dependencies, which may come with their own license agreement(s). Your use of the package and dependencies constitutes your acceptance of their license agreements. If you do not accept the license agreement(s), then delete the relevant components from your device. Successfully installed 'EFCodeFirst.SqlServerCompact 0.8' Successfully added 'SQLCE 4.0.8435.1' to EfCodeFirst-CTP5 Successfully added 'EFCodeFirst 0.8' to EfCodeFirst-CTP5 Successfully added 'WebActivator 1.0.0.0' to EfCodeFirst-CTP5 Successfully added 'EFCodeFirst.SqlServerCompact 0.8' to EfCodeFirst-CTP5 Note: We're using SQLCE 4 with Entity Framework here because they work really well together from a development scenario, but you can of course use Entity Framework Code First with other databases supported by Entity framework. Creating The Model using EF Code First Now we can create our model class. Right-click the Models folder and select Add/Class. Name the Class Person.cs and add the following code: using System.Data.Entity; namespace DeadSimpleServer.Models { public class Person { public int ID { get; set; } public string Name { get; set; } } public class PersonContext : DbContext { public DbSet<Person> People { get; set; } } } Notice that the entity class Person has no special interfaces or base class. There's nothing special needed to make it work - it's just a POCO. The context we'll use to access the entities in the application is called PersonContext, but you could name it anything you wanted. The important thing is that it inherits DbContext and contains one or more DbSet which holds our entity collections. Adding Seed Data We need some testing data to expose from our service. The simplest way to get that into our database is to modify the CreateCeDatabaseIfNotExists class in AppStart_SQLCEEntityFramework.cs by adding some seed data to the Seed method: protected virtual void Seed( TContext context ) { var personContext = context as PersonContext; personContext.People.Add( new Person { ID = 1, Name = "George Washington" } ); personContext.People.Add( new Person { ID = 2, Name = "John Adams" } ); personContext.People.Add( new Person { ID = 3, Name = "Thomas Jefferson" } ); personContext.SaveChanges(); } The CreateCeDatabaseIfNotExists class name is pretty self-explanatory - when our DbContext is accessed and the database isn't found, a new one will be created and populated with the data in the Seed method. There's one more step to make that work - we need to uncomment a line in the Start method at the top of of the AppStart_SQLCEEntityFramework class and set the context name, as shown here, public static class AppStart_SQLCEEntityFramework { public static void Start() { DbDatabase.DefaultConnectionFactory = new SqlCeConnectionFactory("System.Data.SqlServerCe.4.0"); // Sets the default database initialization code for working with Sql Server Compact databases // Uncomment this line and replace CONTEXT_NAME with the name of your DbContext if you are // using your DbContext to create and manage your database DbDatabase.SetInitializer(new CreateCeDatabaseIfNotExists<PersonContext>()); } } Now our database and entity framework are set up, so we can expose data via WCF Data Services. Note: This is a bare-bones implementation with no administration screens. If you'd like to see how those are added, check out The Full Stack screencast series. Creating the oData Service using WCF Data Services Add a new WCF Data Service to the project (right-click the project / Add New Item / Web / WCF Data Service). We’ll be exposing all the data as read/write.  Remember to reconfigure to control and minimize access as appropriate for your own application. Open the code behind for your service. In our case, the service was called PersonTestDataService.svc so the code behind class file is PersonTestDataService.svc.cs. using System.Data.Services; using System.Data.Services.Common; using System.ServiceModel; using DeadSimpleServer.Models; namespace DeadSimpleServer { [ServiceBehavior( IncludeExceptionDetailInFaults = true )] public class PersonTestDataService : DataService<PersonContext> { // This method is called only once to initialize service-wide policies. public static void InitializeService( DataServiceConfiguration config ) { config.SetEntitySetAccessRule( "*", EntitySetRights.All ); config.DataServiceBehavior.MaxProtocolVersion = DataServiceProtocolVersion.V2; config.UseVerboseErrors = true; } } } We're enabling a few additional settings to make it easier to debug if you run into trouble. The ServiceBehavior attribute is set to include exception details in faults, and we're using verbose errors. You can remove both of these when your service is working, as your public production service shouldn't be revealing exception information. You can view the output of the service by running the application and browsing to http://localhost:[portnumber]/PersonTestDataService.svc/: <service xml:base="http://localhost:49786/PersonTestDataService.svc/" xmlns:atom="http://www.w3.org/2005/Atom" xmlns:app="http://www.w3.org/2007/app" xmlns="http://www.w3.org/2007/app"> <workspace> <atom:title>Default</atom:title> <collection href="People"> <atom:title>People</atom:title> </collection> </workspace> </service> This indicates that the service exposes one collection, which is accessible by browsing to http://localhost:[portnumber]/PersonTestDataService.svc/People <?xml version="1.0" encoding="iso-8859-1" standalone="yes"?> <feed xml:base=http://localhost:49786/PersonTestDataService.svc/ xmlns:d="http://schemas.microsoft.com/ado/2007/08/dataservices" xmlns:m="http://schemas.microsoft.com/ado/2007/08/dataservices/metadata" xmlns="http://www.w3.org/2005/Atom"> <title type="text">People</title> <id>http://localhost:49786/PersonTestDataService.svc/People</id> <updated>2010-12-29T01:01:50Z</updated> <link rel="self" title="People" href="People" /> <entry> <id>http://localhost:49786/PersonTestDataService.svc/People(1)</id> <title type="text"></title> <updated>2010-12-29T01:01:50Z</updated> <author> <name /> </author> <link rel="edit" title="Person" href="People(1)" /> <category term="DeadSimpleServer.Models.Person" scheme="http://schemas.microsoft.com/ado/2007/08/dataservices/scheme" /> <content type="application/xml"> <m:properties> <d:ID m:type="Edm.Int32">1</d:ID> <d:Name>George Washington</d:Name> </m:properties> </content> </entry> <entry> ... </entry> </feed> Let's recap what we've done so far. But enough with services and XML - let's get this into our Windows Phone client application. Creating the DataServiceContext for the Client Use the latest DataSvcUtil.exe from http://odata.codeplex.com. As of today, that's in this download: http://odata.codeplex.com/releases/view/54698 You need to run it with a few options: /uri - This will point to the service URI. In this case, it's http://localhost:59342/PersonTestDataService.svc  Pick up the port number from your running server (e.g., the server formerly known as Cassini). /out - This is the DataServiceContext class that will be generated. You can name it whatever you'd like. /Version - should be set to 2.0 /DataServiceCollection - Include this flag to generate collections derived from the DataServiceCollection base, which brings in all the ObservableCollection goodness that handles your INotifyPropertyChanged events for you. Here's the console session from when we ran it: <ListBox x:Name="MainListBox" Margin="0,0,-12,0" ItemsSource="{Binding}" SelectionChanged="MainListBox_SelectionChanged"> Next, to keep things simple, change the Binding on the two TextBlocks within the DataTemplate to Name and ID, <ListBox x:Name="MainListBox" Margin="0,0,-12,0" ItemsSource="{Binding}" SelectionChanged="MainListBox_SelectionChanged"> <ListBox.ItemTemplate> <DataTemplate> <StackPanel Margin="0,0,0,17" Width="432"> <TextBlock Text="{Binding Name}" TextWrapping="Wrap" Style="{StaticResource PhoneTextExtraLargeStyle}" /> <TextBlock Text="{Binding ID}" TextWrapping="Wrap" Margin="12,-6,12,0" Style="{StaticResource PhoneTextSubtleStyle}" /> </StackPanel> </DataTemplate> </ListBox.ItemTemplate> </ListBox> Getting The Context In the code-behind you’ll first declare a member variable to hold the context from the Entity Framework. This is named using convention over configuration. The db type is Person and the context is of type PersonContext, You initialize it by providing the URI, in this case using the URL obtained from the Cassini web server, PersonContext context = new PersonContext( new Uri( "http://localhost:49786/PersonTestDataService.svc/" ) ); Create a second member variable of type DataServiceCollection<Person> but do not initialize it, DataServiceCollection<Person> people; In the constructor you’ll initialize the DataServiceCollection using the PersonContext, public MainPage() { InitializeComponent(); people = new DataServiceCollection<Person>( context ); Finally, you’ll load the people collection using the LoadAsync method, passing in the fully specified URI for the People collection in the web service, people.LoadAsync( new Uri( "http://localhost:49786/PersonTestDataService.svc/People" ) ); Note that this method runs asynchronously and when it is finished the people  collection is already populated. Thus, since we didn’t need or want to override any of the behavior we don’t implement the LoadCompleted. You can use the LoadCompleted event if you need to do any other UI updates, but you don't need to. The final code is as shown below: using System; using System.Data.Services.Client; using System.Windows; using System.Windows.Controls; using DeadSimpleServer.Models; using Microsoft.Phone.Controls; namespace WindowsPhoneODataTest { public partial class MainPage : PhoneApplicationPage { PersonContext context = new PersonContext( new Uri( "http://localhost:49786/PersonTestDataService.svc/" ) ); DataServiceCollection<Person> people; // Constructor public MainPage() { InitializeComponent(); // Set the data context of the listbox control to the sample data // DataContext = App.ViewModel; people = new DataServiceCollection<Person>( context ); people.LoadAsync( new Uri( "http://localhost:49786/PersonTestDataService.svc/People" ) ); DataContext = people; this.Loaded += new RoutedEventHandler( MainPage_Loaded ); } // Handle selection changed on ListBox private void MainListBox_SelectionChanged( object sender, SelectionChangedEventArgs e ) { // If selected index is -1 (no selection) do nothing if ( MainListBox.SelectedIndex == -1 ) return; // Navigate to the new page NavigationService.Navigate( new Uri( "/DetailsPage.xaml?selectedItem=" + MainListBox.SelectedIndex, UriKind.Relative ) ); // Reset selected index to -1 (no selection) MainListBox.SelectedIndex = -1; } // Load data for the ViewModel Items private void MainPage_Loaded( object sender, RoutedEventArgs e ) { if ( !App.ViewModel.IsDataLoaded ) { App.ViewModel.LoadData(); } } } } With people populated we can set it as the DataContext and run the application; you’ll find that the Name and ID are displayed in the list on the Mainpage. Here's how the pieces in the client fit together: Complete source code available here

    Read the article

  • An Introduction to Meteor

    - by Stephen.Walther
    The goal of this blog post is to give you a brief introduction to Meteor which is a framework for building Single Page Apps. In this blog entry, I provide a walkthrough of building a simple Movie database app. What is special about Meteor? Meteor has two jaw-dropping features: Live HTML – If you make any changes to the HTML, CSS, JavaScript, or data on the server then every client shows the changes automatically without a browser refresh. For example, if you change the background color of a page to yellow then every open browser will show the new yellow background color without a refresh. Or, if you add a new movie to a collection of movies, then every open browser will display the new movie automatically. With Live HTML, users no longer need a refresh button. Changes to an application happen everywhere automatically without any effort. The Meteor framework handles all of the messy details of keeping all of the clients in sync with the server for you. Latency Compensation – When you modify data on the client, these modifications appear as if they happened on the server without any delay. For example, if you create a new movie then the movie appears instantly. However, that is all an illusion. In the background, Meteor updates the database with the new movie. If, for whatever reason, the movie cannot be added to the database then Meteor removes the movie from the client automatically. Latency compensation is extremely important for creating a responsive web application. You want the user to be able to make instant modifications in the browser and the framework to handle the details of updating the database without slowing down the user. Installing Meteor Meteor is licensed under the open-source MIT license and you can start building production apps with the framework right now. Be warned that Meteor is still in the “early preview” stage. It has not reached a 1.0 release. According to the Meteor FAQ, Meteor will reach version 1.0 in “More than a month, less than a year.” Don’t be scared away by that. You should be aware that, unlike most open source projects, Meteor has financial backing. The Meteor project received an $11.2 million round of financing from Andreessen Horowitz. So, it would be a good bet that this project will reach the 1.0 mark. And, if it doesn’t, the framework as it exists right now is still very powerful. Meteor runs on top of Node.js. You write Meteor apps by writing JavaScript which runs both on the client and on the server. You can build Meteor apps on Windows, Mac, or Linux (Although the support for Windows is still officially unofficial). If you want to install Meteor on Windows then download the MSI from the following URL: http://win.meteor.com/ If you want to install Meteor on Mac/Linux then run the following CURL command from your terminal: curl https://install.meteor.com | /bin/sh Meteor will install all of its dependencies automatically including Node.js. However, I recommend that you install Node.js before installing Meteor by installing Node.js from the following address: http://nodejs.org/ If you let Meteor install Node.js then Meteor won’t install NPM which is the standard package manager for Node.js. If you install Node.js and then you install Meteor then you get NPM automatically. Creating a New Meteor App To get a sense of how Meteor works, I am going to walk through the steps required to create a simple Movie database app. Our app will display a list of movies and contain a form for creating a new movie. The first thing that we need to do is create our new Meteor app. Open a command prompt/terminal window and execute the following command: Meteor create MovieApp After you execute this command, you should see something like the following: Follow the instructions: execute cd MovieApp to change to your MovieApp directory, and run the meteor command. Executing the meteor command starts Meteor on port 3000. Open up your favorite web browser and navigate to http://localhost:3000 and you should see the default Meteor Hello World page: Open up your favorite development environment to see what the Meteor app looks like. Open the MovieApp folder which we just created. Here’s what the MovieApp looks like in Visual Studio 2012: Notice that our MovieApp contains three files named MovieApp.css, MovieApp.html, and MovieApp.js. In other words, it contains a Cascading Style Sheet file, an HTML file, and a JavaScript file. Just for fun, let’s see how the Live HTML feature works. Open up multiple browsers and point each browser at http://localhost:3000. Now, open the MovieApp.html page and modify the text “Hello World!” to “Hello Cruel World!” and save the change. The text in all of the browsers should update automatically without a browser refresh. Pretty amazing, right? Controlling Where JavaScript Executes You write a Meteor app using JavaScript. Some of the JavaScript executes on the client (the browser) and some of the JavaScript executes on the server and some of the JavaScript executes in both places. For a super simple app, you can use the Meteor.isServer and Meteor.isClient properties to control where your JavaScript code executes. For example, the following JavaScript contains a section of code which executes on the server and a section of code which executes in the browser: if (Meteor.isClient) { console.log("Hello Browser!"); } if (Meteor.isServer) { console.log("Hello Server!"); } console.log("Hello Browser and Server!"); When you run the app, the message “Hello Browser!” is written to the browser JavaScript console. The message “Hello Server!” is written to the command/terminal window where you ran Meteor. Finally, the message “Hello Browser and Server!” is execute on both the browser and server and the message appears in both places. For simple apps, using Meteor.isClient and Meteor.isServer to control where JavaScript executes is fine. For more complex apps, you should create separate folders for your server and client code. Here are the folders which you can use in a Meteor app: · client – This folder contains any JavaScript which executes only on the client. · server – This folder contains any JavaScript which executes only on the server. · common – This folder contains any JavaScript code which executes on both the client and server. · lib – This folder contains any JavaScript files which you want to execute before any other JavaScript files. · public – This folder contains static application assets such as images. For the Movie App, we need the client, server, and common folders. Delete the existing MovieApp.js, MovieApp.html, and MovieApp.css files. We will create new files in the right locations later in this walkthrough. Combining HTML, CSS, and JavaScript Files Meteor combines all of your JavaScript files, and all of your Cascading Style Sheet files, and all of your HTML files automatically. If you want to create one humongous JavaScript file which contains all of the code for your app then that is your business. However, if you want to build a more maintainable application, then you should break your JavaScript files into many separate JavaScript files and let Meteor combine them for you. Meteor also combines all of your HTML files into a single file. HTML files are allowed to have the following top-level elements: <head> — All <head> files are combined into a single <head> and served with the initial page load. <body> — All <body> files are combined into a single <body> and served with the initial page load. <template> — All <template> files are compiled into JavaScript templates. Because you are creating a single page app, a Meteor app typically will contain a single HTML file for the <head> and <body> content. However, a Meteor app typically will contain several template files. In other words, all of the interesting stuff happens within the <template> files. Displaying a List of Movies Let me start building the Movie App by displaying a list of movies. In order to display a list of movies, we need to create the following four files: · client\movies.html – Contains the HTML for the <head> and <body> of the page for the Movie app. · client\moviesTemplate.html – Contains the HTML template for displaying the list of movies. · client\movies.js – Contains the JavaScript for supplying data to the moviesTemplate. · server\movies.js – Contains the JavaScript for seeding the database with movies. After you create these files, your folder structure should looks like this: Here’s what the client\movies.html file looks like: <head> <title>My Movie App</title> </head> <body> <h1>Movies</h1> {{> moviesTemplate }} </body>   Notice that it contains <head> and <body> top-level elements. The <body> element includes the moviesTemplate with the syntax {{> moviesTemplate }}. The moviesTemplate is defined in the client/moviesTemplate.html file: <template name="moviesTemplate"> <ul> {{#each movies}} <li> {{title}} </li> {{/each}} </ul> </template> By default, Meteor uses the Handlebars templating library. In the moviesTemplate above, Handlebars is used to loop through each of the movies using {{#each}}…{{/each}} and display the title for each movie using {{title}}. The client\movies.js JavaScript file is used to bind the moviesTemplate to the Movies collection on the client. Here’s what this JavaScript file looks like: // Declare client Movies collection Movies = new Meteor.Collection("movies"); // Bind moviesTemplate to Movies collection Template.moviesTemplate.movies = function () { return Movies.find(); }; The Movies collection is a client-side proxy for the server-side Movies database collection. Whenever you want to interact with the collection of Movies stored in the database, you use the Movies collection instead of communicating back to the server. The moviesTemplate is bound to the Movies collection by assigning a function to the Template.moviesTemplate.movies property. The function simply returns all of the movies from the Movies collection. The final file which we need is the server-side server\movies.js file: // Declare server Movies collection Movies = new Meteor.Collection("movies"); // Seed the movie database with a few movies Meteor.startup(function () { if (Movies.find().count() == 0) { Movies.insert({ title: "Star Wars", director: "Lucas" }); Movies.insert({ title: "Memento", director: "Nolan" }); Movies.insert({ title: "King Kong", director: "Jackson" }); } }); The server\movies.js file does two things. First, it declares the server-side Meteor Movies collection. When you declare a server-side Meteor collection, a collection is created in the MongoDB database associated with your Meteor app automatically (Meteor uses MongoDB as its database automatically). Second, the server\movies.js file seeds the Movies collection (MongoDB collection) with three movies. Seeding the database gives us some movies to look at when we open the Movies app in a browser. Creating New Movies Let me modify the Movies Database App so that we can add new movies to the database of movies. First, I need to create a new template file – named client\movieForm.html – which contains an HTML form for creating a new movie: <template name="movieForm"> <fieldset> <legend>Add New Movie</legend> <form> <div> <label> Title: <input id="title" /> </label> </div> <div> <label> Director: <input id="director" /> </label> </div> <div> <input type="submit" value="Add Movie" /> </div> </form> </fieldset> </template> In order for the new form to show up, I need to modify the client\movies.html file to include the movieForm.html template. Notice that I added {{> movieForm }} to the client\movies.html file: <head> <title>My Movie App</title> </head> <body> <h1>Movies</h1> {{> moviesTemplate }} {{> movieForm }} </body> After I make these modifications, our Movie app will display the form: The next step is to handle the submit event for the movie form. Below, I’ve modified the client\movies.js file so that it contains a handler for the submit event raised when you submit the form contained in the movieForm.html template: // Declare client Movies collection Movies = new Meteor.Collection("movies"); // Bind moviesTemplate to Movies collection Template.moviesTemplate.movies = function () { return Movies.find(); }; // Handle movieForm events Template.movieForm.events = { 'submit': function (e, tmpl) { // Don't postback e.preventDefault(); // create the new movie var newMovie = { title: tmpl.find("#title").value, director: tmpl.find("#director").value }; // add the movie to the db Movies.insert(newMovie); } }; The Template.movieForm.events property contains an event map which maps event names to handlers. In this case, I am mapping the form submit event to an anonymous function which handles the event. In the event handler, I am first preventing a postback by calling e.preventDefault(). This is a single page app, no postbacks are allowed! Next, I am grabbing the new movie from the HTML form. I’m taking advantage of the template find() method to retrieve the form field values. Finally, I am calling Movies.insert() to insert the new movie into the Movies collection. Here, I am explicitly inserting the new movie into the client-side Movies collection. Meteor inserts the new movie into the server-side Movies collection behind the scenes. When Meteor inserts the movie into the server-side collection, the new movie is added to the MongoDB database associated with the Movies app automatically. If server-side insertion fails for whatever reasons – for example, your internet connection is lost – then Meteor will remove the movie from the client-side Movies collection automatically. In other words, Meteor takes care of keeping the client Movies collection and the server Movies collection in sync. If you open multiple browsers, and add movies, then you should notice that all of the movies appear on all of the open browser automatically. You don’t need to refresh individual browsers to update the client-side Movies collection. Meteor keeps everything synchronized between the browsers and server for you. Removing the Insecure Module To make it easier to develop and debug a new Meteor app, by default, you can modify the database directly from the client. For example, you can delete all of the data in the database by opening up your browser console window and executing multiple Movies.remove() commands. Obviously, enabling anyone to modify your database from the browser is not a good idea in a production application. Before you make a Meteor app public, you should first run the meteor remove insecure command from a command/terminal window: Running meteor remove insecure removes the insecure package from the Movie app. Unfortunately, it also breaks our Movie app. We’ll get an “Access denied” error in our browser console whenever we try to insert a new movie. No worries. I’ll fix this issue in the next section. Creating Meteor Methods By taking advantage of Meteor Methods, you can create methods which can be invoked on both the client and the server. By taking advantage of Meteor Methods you can: 1. Perform form validation on both the client and the server. For example, even if an evil hacker bypasses your client code, you can still prevent the hacker from submitting an invalid value for a form field by enforcing validation on the server. 2. Simulate database operations on the client but actually perform the operations on the server. Let me show you how we can modify our Movie app so it uses Meteor Methods to insert a new movie. First, we need to create a new file named common\methods.js which contains the definition of our Meteor Methods: Meteor.methods({ addMovie: function (newMovie) { // Perform form validation if (newMovie.title == "") { throw new Meteor.Error(413, "Missing title!"); } if (newMovie.director == "") { throw new Meteor.Error(413, "Missing director!"); } // Insert movie (simulate on client, do it on server) return Movies.insert(newMovie); } }); The addMovie() method is called from both the client and the server. This method does two things. First, it performs some basic validation. If you don’t enter a title or you don’t enter a director then an error is thrown. Second, the addMovie() method inserts the new movie into the Movies collection. When called on the client, inserting the new movie into the Movies collection just updates the collection. When called on the server, inserting the new movie into the Movies collection causes the database (MongoDB) to be updated with the new movie. You must add the common\methods.js file to the common folder so it will get executed on both the client and the server. Our folder structure now looks like this: We actually call the addMovie() method within our client code in the client\movies.js file. Here’s what the updated file looks like: // Declare client Movies collection Movies = new Meteor.Collection("movies"); // Bind moviesTemplate to Movies collection Template.moviesTemplate.movies = function () { return Movies.find(); }; // Handle movieForm events Template.movieForm.events = { 'submit': function (e, tmpl) { // Don't postback e.preventDefault(); // create the new movie var newMovie = { title: tmpl.find("#title").value, director: tmpl.find("#director").value }; // add the movie to the db Meteor.call( "addMovie", newMovie, function (err, result) { if (err) { alert("Could not add movie " + err.reason); } } ); } }; The addMovie() method is called – on both the client and the server – by calling the Meteor.call() method. This method accepts the following parameters: · The string name of the method to call. · The data to pass to the method (You can actually pass multiple params for the data if you like). · A callback function to invoke after the method completes. In the JavaScript code above, the addMovie() method is called with the new movie retrieved from the HTML form. The callback checks for an error. If there is an error then the error reason is displayed in an alert (please don’t use alerts for validation errors in a production app because they are ugly!). Summary The goal of this blog post was to provide you with a brief walk through of a simple Meteor app. I showed you how you can create a simple Movie Database app which enables you to display a list of movies and create new movies. I also explained why it is important to remove the Meteor insecure package from a production app. I showed you how to use Meteor Methods to insert data into the database instead of doing it directly from the client. I’m very impressed with the Meteor framework. The support for Live HTML and Latency Compensation are required features for many real world Single Page Apps but implementing these features by hand is not easy. Meteor makes it easy.

    Read the article

  • Toorcon14

    - by danx
    Toorcon 2012 Information Security Conference San Diego, CA, http://www.toorcon.org/ Dan Anderson, October 2012 It's almost Halloween, and we all know what that means—yes, of course, it's time for another Toorcon Conference! Toorcon is an annual conference for people interested in computer security. This includes the whole range of hackers, computer hobbyists, professionals, security consultants, press, law enforcement, prosecutors, FBI, etc. We're at Toorcon 14—see earlier blogs for some of the previous Toorcon's I've attended (back to 2003). This year's "con" was held at the Westin on Broadway in downtown San Diego, California. The following are not necessarily my views—I'm just the messenger—although I could have misquoted or misparaphrased the speakers. Also, I only reviewed some of the talks, below, which I attended and interested me. MalAndroid—the Crux of Android Infections, Aditya K. Sood Programming Weird Machines with ELF Metadata, Rebecca "bx" Shapiro Privacy at the Handset: New FCC Rules?, Valkyrie Hacking Measured Boot and UEFI, Dan Griffin You Can't Buy Security: Building the Open Source InfoSec Program, Boris Sverdlik What Journalists Want: The Investigative Reporters' Perspective on Hacking, Dave Maas & Jason Leopold Accessibility and Security, Anna Shubina Stop Patching, for Stronger PCI Compliance, Adam Brand McAfee Secure & Trustmarks — a Hacker's Best Friend, Jay James & Shane MacDougall MalAndroid—the Crux of Android Infections Aditya K. Sood, IOActive, Michigan State PhD candidate Aditya talked about Android smartphone malware. There's a lot of old Android software out there—over 50% Gingerbread (2.3.x)—and most have unpatched vulnerabilities. Of 9 Android vulnerabilities, 8 have known exploits (such as the old Gingerbread Global Object Table exploit). Android protection includes sandboxing, security scanner, app permissions, and screened Android app market. The Android permission checker has fine-grain resource control, policy enforcement. Android static analysis also includes a static analysis app checker (bouncer), and a vulnerablity checker. What security problems does Android have? User-centric security, which depends on the user to grant permission and make smart decisions. But users don't care or think about malware (the're not aware, not paranoid). All they want is functionality, extensibility, mobility Android had no "proper" encryption before Android 3.0 No built-in protection against social engineering and web tricks Alternative Android app markets are unsafe. Simply visiting some markets can infect Android Aditya classified Android Malware types as: Type A—Apps. These interact with the Android app framework. For example, a fake Netflix app. Or Android Gold Dream (game), which uploads user files stealthy manner to a remote location. Type K—Kernel. Exploits underlying Linux libraries or kernel Type H—Hybrid. These use multiple layers (app framework, libraries, kernel). These are most commonly used by Android botnets, which are popular with Chinese botnet authors What are the threats from Android malware? These incude leak info (contacts), banking fraud, corporate network attacks, malware advertising, malware "Hackivism" (the promotion of social causes. For example, promiting specific leaders of the Tunisian or Iranian revolutions. Android malware is frequently "masquerated". That is, repackaged inside a legit app with malware. To avoid detection, the hidden malware is not unwrapped until runtime. The malware payload can be hidden in, for example, PNG files. Less common are Android bootkits—there's not many around. What they do is hijack the Android init framework—alteering system programs and daemons, then deletes itself. For example, the DKF Bootkit (China). Android App Problems: no code signing! all self-signed native code execution permission sandbox — all or none alternate market places no robust Android malware detection at network level delayed patch process Programming Weird Machines with ELF Metadata Rebecca "bx" Shapiro, Dartmouth College, NH https://github.com/bx/elf-bf-tools @bxsays on twitter Definitions. "ELF" is an executable file format used in linking and loading executables (on UNIX/Linux-class machines). "Weird machine" uses undocumented computation sources (I think of them as unintended virtual machines). Some examples of "weird machines" are those that: return to weird location, does SQL injection, corrupts the heap. Bx then talked about using ELF metadata as (an uintended) "weird machine". Some ELF background: A compiler takes source code and generates a ELF object file (hello.o). A static linker makes an ELF executable from the object file. A runtime linker and loader takes ELF executable and loads and relocates it in memory. The ELF file has symbols to relocate functions and variables. ELF has two relocation tables—one at link time and another one at loading time: .rela.dyn (link time) and .dynsym (dynamic table). GOT: Global Offset Table of addresses for dynamically-linked functions. PLT: Procedure Linkage Tables—works with GOT. The memory layout of a process (not the ELF file) is, in order: program (+ heap), dynamic libraries, libc, ld.so, stack (which includes the dynamic table loaded into memory) For ELF, the "weird machine" is found and exploited in the loader. ELF can be crafted for executing viruses, by tricking runtime into executing interpreted "code" in the ELF symbol table. One can inject parasitic "code" without modifying the actual ELF code portions. Think of the ELF symbol table as an "assembly language" interpreter. It has these elements: instructions: Add, move, jump if not 0 (jnz) Think of symbol table entries as "registers" symbol table value is "contents" immediate values are constants direct values are addresses (e.g., 0xdeadbeef) move instruction: is a relocation table entry add instruction: relocation table "addend" entry jnz instruction: takes multiple relocation table entries The ELF weird machine exploits the loader by relocating relocation table entries. The loader will go on forever until told to stop. It stores state on stack at "end" and uses IFUNC table entries (containing function pointer address). The ELF weird machine, called "Brainfu*k" (BF) has: 8 instructions: pointer inc, dec, inc indirect, dec indirect, jump forward, jump backward, print. Three registers - 3 registers Bx showed example BF source code that implemented a Turing machine printing "hello, world". More interesting was the next demo, where bx modified ping. Ping runs suid as root, but quickly drops privilege. BF modified the loader to disable the library function call dropping privilege, so it remained as root. Then BF modified the ping -t argument to execute the -t filename as root. It's best to show what this modified ping does with an example: $ whoami bx $ ping localhost -t backdoor.sh # executes backdoor $ whoami root $ The modified code increased from 285948 bytes to 290209 bytes. A BF tool compiles "executable" by modifying the symbol table in an existing ELF executable. The tool modifies .dynsym and .rela.dyn table, but not code or data. Privacy at the Handset: New FCC Rules? "Valkyrie" (Christie Dudley, Santa Clara Law JD candidate) Valkyrie talked about mobile handset privacy. Some background: Senator Franken (also a comedian) became alarmed about CarrierIQ, where the carriers track their customers. Franken asked the FCC to find out what obligations carriers think they have to protect privacy. The carriers' response was that they are doing just fine with self-regulation—no worries! Carriers need to collect data, such as missed calls, to maintain network quality. But carriers also sell data for marketing. Verizon sells customer data and enables this with a narrow privacy policy (only 1 month to opt out, with difficulties). The data sold is not individually identifiable and is aggregated. But Verizon recommends, as an aggregation workaround to "recollate" data to other databases to identify customers indirectly. The FCC has regulated telephone privacy since 1934 and mobile network privacy since 2007. Also, the carriers say mobile phone privacy is a FTC responsibility (not FCC). FTC is trying to improve mobile app privacy, but FTC has no authority over carrier / customer relationships. As a side note, Apple iPhones are unique as carriers have extra control over iPhones they don't have with other smartphones. As a result iPhones may be more regulated. Who are the consumer advocates? Everyone knows EFF, but EPIC (Electrnic Privacy Info Center), although more obsecure, is more relevant. What to do? Carriers must be accountable. Opt-in and opt-out at any time. Carriers need incentive to grant users control for those who want it, by holding them liable and responsible for breeches on their clock. Location information should be added current CPNI privacy protection, and require "Pen/trap" judicial order to obtain (and would still be a lower standard than 4th Amendment). Politics are on a pro-privacy swing now, with many senators and the Whitehouse. There will probably be new regulation soon, and enforcement will be a problem, but consumers will still have some benefit. Hacking Measured Boot and UEFI Dan Griffin, JWSecure, Inc., Seattle, @JWSdan Dan talked about hacking measured UEFI boot. First some terms: UEFI is a boot technology that is replacing BIOS (has whitelisting and blacklisting). UEFI protects devices against rootkits. TPM - hardware security device to store hashs and hardware-protected keys "secure boot" can control at firmware level what boot images can boot "measured boot" OS feature that tracks hashes (from BIOS, boot loader, krnel, early drivers). "remote attestation" allows remote validation and control based on policy on a remote attestation server. Microsoft pushing TPM (Windows 8 required), but Google is not. Intel TianoCore is the only open source for UEFI. Dan has Measured Boot Tool at http://mbt.codeplex.com/ with a demo where you can also view TPM data. TPM support already on enterprise-class machines. UEFI Weaknesses. UEFI toolkits are evolving rapidly, but UEFI has weaknesses: assume user is an ally trust TPM implicitly, and attached to computer hibernate file is unprotected (disk encryption protects against this) protection migrating from hardware to firmware delays in patching and whitelist updates will UEFI really be adopted by the mainstream (smartphone hardware support, bank support, apathetic consumer support) You Can't Buy Security: Building the Open Source InfoSec Program Boris Sverdlik, ISDPodcast.com co-host Boris talked about problems typical with current security audits. "IT Security" is an oxymoron—IT exists to enable buiness, uptime, utilization, reporting, but don't care about security—IT has conflict of interest. There's no Magic Bullet ("blinky box"), no one-size-fits-all solution (e.g., Intrusion Detection Systems (IDSs)). Regulations don't make you secure. The cloud is not secure (because of shared data and admin access). Defense and pen testing is not sexy. Auditors are not solution (security not a checklist)—what's needed is experience and adaptability—need soft skills. Step 1: First thing is to Google and learn the company end-to-end before you start. Get to know the management team (not IT team), meet as many people as you can. Don't use arbitrary values such as CISSP scores. Quantitive risk assessment is a myth (e.g. AV*EF-SLE). Learn different Business Units, legal/regulatory obligations, learn the business and where the money is made, verify company is protected from script kiddies (easy), learn sensitive information (IP, internal use only), and start with low-hanging fruit (customer service reps and social engineering). Step 2: Policies. Keep policies short and relevant. Generic SANS "security" boilerplate policies don't make sense and are not followed. Focus on acceptable use, data usage, communications, physical security. Step 3: Implementation: keep it simple stupid. Open source, although useful, is not free (implementation cost). Access controls with authentication & authorization for local and remote access. MS Windows has it, otherwise use OpenLDAP, OpenIAM, etc. Application security Everyone tries to reinvent the wheel—use existing static analysis tools. Review high-risk apps and major revisions. Don't run different risk level apps on same system. Assume host/client compromised and use app-level security control. Network security VLAN != segregated because there's too many workarounds. Use explicit firwall rules, active and passive network monitoring (snort is free), disallow end user access to production environment, have a proxy instead of direct Internet access. Also, SSL certificates are not good two-factor auth and SSL does not mean "safe." Operational Controls Have change, patch, asset, & vulnerability management (OSSI is free). For change management, always review code before pushing to production For logging, have centralized security logging for business-critical systems, separate security logging from administrative/IT logging, and lock down log (as it has everything). Monitor with OSSIM (open source). Use intrusion detection, but not just to fulfill a checkbox: build rules from a whitelist perspective (snort). OSSEC has 95% of what you need. Vulnerability management is a QA function when done right: OpenVas and Seccubus are free. Security awareness The reality is users will always click everything. Build real awareness, not compliance driven checkbox, and have it integrated into the culture. Pen test by crowd sourcing—test with logging COSSP http://www.cossp.org/ - Comprehensive Open Source Security Project What Journalists Want: The Investigative Reporters' Perspective on Hacking Dave Maas, San Diego CityBeat Jason Leopold, Truthout.org The difference between hackers and investigative journalists: For hackers, the motivation varies, but method is same, technological specialties. For investigative journalists, it's about one thing—The Story, and they need broad info-gathering skills. J-School in 60 Seconds: Generic formula: Person or issue of pubic interest, new info, or angle. Generic criteria: proximity, prominence, timeliness, human interest, oddity, or consequence. Media awareness of hackers and trends: journalists becoming extremely aware of hackers with congressional debates (privacy, data breaches), demand for data-mining Journalists, use of coding and web development for Journalists, and Journalists busted for hacking (Murdock). Info gathering by investigative journalists include Public records laws. Federal Freedom of Information Act (FOIA) is good, but slow. California Public Records Act is a lot stronger. FOIA takes forever because of foot-dragging—it helps to be specific. Often need to sue (especially FBI). CPRA is faster, and requests can be vague. Dumps and leaks (a la Wikileaks) Journalists want: leads, protecting ourselves, our sources, and adapting tools for news gathering (Google hacking). Anonomity is important to whistleblowers. They want no digital footprint left behind (e.g., email, web log). They don't trust encryption, want to feel safe and secure. Whistleblower laws are very weak—there's no upside for whistleblowers—they have to be very passionate to do it. Accessibility and Security or: How I Learned to Stop Worrying and Love the Halting Problem Anna Shubina, Dartmouth College Anna talked about how accessibility and security are related. Accessibility of digital content (not real world accessibility). mostly refers to blind users and screenreaders, for our purpose. Accessibility is about parsing documents, as are many security issues. "Rich" executable content causes accessibility to fail, and often causes security to fail. For example MS Word has executable format—it's not a document exchange format—more dangerous than PDF or HTML. Accessibility is often the first and maybe only sanity check with parsing. They have no choice because someone may want to read what you write. Google, for example, is very particular about web browser you use and are bad at supporting other browsers. Uses JavaScript instead of links, often requiring mouseover to display content. PDF is a security nightmare. Executible format, embedded flash, JavaScript, etc. 15 million lines of code. Google Chrome doesn't handle PDF correctly, causing several security bugs. PDF has an accessibility checker and PDF tagging, to help with accessibility. But no PDF checker checks for incorrect tags, untagged content, or validates lists or tables. None check executable content at all. The "Halting Problem" is: can one decide whether a program will ever stop? The answer, in general, is no (Rice's theorem). The same holds true for accessibility checkers. Language-theoretic Security says complicated data formats are hard to parse and cannot be solved due to the Halting Problem. W3C Web Accessibility Guidelines: "Perceivable, Operable, Understandable, Robust" Not much help though, except for "Robust", but here's some gems: * all information should be parsable (paraphrasing) * if not parsable, cannot be converted to alternate formats * maximize compatibility in new document formats Executible webpages are bad for security and accessibility. They say it's for a better web experience. But is it necessary to stuff web pages with JavaScript for a better experience? A good example is The Drudge Report—it has hand-written HTML with no JavaScript, yet drives a lot of web traffic due to good content. A bad example is Google News—hidden scrollbars, guessing user input. Solutions: Accessibility and security problems come from same source Expose "better user experience" myth Keep your corner of Internet parsable Remember "Halting Problem"—recognize false solutions (checking and verifying tools) Stop Patching, for Stronger PCI Compliance Adam Brand, protiviti @adamrbrand, http://www.picfun.com/ Adam talked about PCI compliance for retail sales. Take an example: for PCI compliance, 50% of Brian's time (a IT guy), 960 hours/year was spent patching POSs in 850 restaurants. Often applying some patches make no sense (like fixing a browser vulnerability on a server). "Scanner worship" is overuse of vulnerability scanners—it gives a warm and fuzzy and it's simple (red or green results—fix reds). Scanners give a false sense of security. In reality, breeches from missing patches are uncommon—more common problems are: default passwords, cleartext authentication, misconfiguration (firewall ports open). Patching Myths: Myth 1: install within 30 days of patch release (but PCI §6.1 allows a "risk-based approach" instead). Myth 2: vendor decides what's critical (also PCI §6.1). But §6.2 requires user ranking of vulnerabilities instead. Myth 3: scan and rescan until it passes. But PCI §11.2.1b says this applies only to high-risk vulnerabilities. Adam says good recommendations come from NIST 800-40. Instead use sane patching and focus on what's really important. From NIST 800-40: Proactive: Use a proactive vulnerability management process: use change control, configuration management, monitor file integrity. Monitor: start with NVD and other vulnerability alerts, not scanner results. Evaluate: public-facing system? workstation? internal server? (risk rank) Decide:on action and timeline Test: pre-test patches (stability, functionality, rollback) for change control Install: notify, change control, tickets McAfee Secure & Trustmarks — a Hacker's Best Friend Jay James, Shane MacDougall, Tactical Intelligence Inc., Canada "McAfee Secure Trustmark" is a website seal marketed by McAfee. A website gets this badge if they pass their remote scanning. The problem is a removal of trustmarks act as flags that you're vulnerable. Easy to view status change by viewing McAfee list on website or on Google. "Secure TrustGuard" is similar to McAfee. Jay and Shane wrote Perl scripts to gather sites from McAfee and search engines. If their certification image changes to a 1x1 pixel image, then they are longer certified. Their scripts take deltas of scans to see what changed daily. The bottom line is change in TrustGuard status is a flag for hackers to attack your site. Entire idea of seals is silly—you're raising a flag saying if you're vulnerable.

    Read the article

  • Error in installing ZTE AC2738 on ubuntu 3.0.0-12-generic

    - by Netro
    I am getting this error ,struct usb_serial_driver has no member named shutdown. I am installing on 64bit ubuntu 3.0.0-12-generic ... Beginning Verify CD ... ... Verify CD Succeed! ... Beginning Copy Install Package Files ... ... will take a long time, waiting 5 seconds, please ... Copy Install Package Files Succeed! ... 'ztemtApp' previous version not found. and install now Beginning install ... ... Current linux release version is 'Ubuntu' ... Checking 'App' process ... Checking old installation ... Installing ... Current Path is : . : /tmp/ztemt_datacard/Linux 1. Checking Previous Version ... 2. Copying Data Bin ... ... will take a few seconds, please waiting ... /tmp/ztemt_datacard/Linux 3. Auto Load Usb Driver Module ... Rather than invoking init scripts through /etc/init.d, use the service(8) utility, e.g. service acpid restart Since the script you are attempting to invoke has been converted to an Upstart job, you may also use the stop(8) and then start(8) utilities, e.g. stop acpid ; start acpid. The restart(8) utility is also available. acpid stop/waiting acpid start/running, process 11802 4. Changing pppd Options ... 5. Changing File Permission ... 6. Deleting Qt lib When Local QT Vertion > V4.4.0 ... ... Package 'libqtgui4' exist ... QT_VERSION = 4 7. Deleting process id file: EVDOApp.pid ... 8. Making USB Serial Driver Module : ztemt.ko ... ... will take a few seconds, please waiting ... make -C /lib/modules/3.0.0-12-generic/build M=/usr/local/bin/ztemtApp/zteusbserial/below2.6.27 modules make[1]: Entering directory `/usr/src/linux-headers-3.0.0-12-generic' CC [M] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.o /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘destroy_serial’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:159:14: error: ‘struct usb_serial_driver’ has no member named ‘shutdown’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:165:18: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_open’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:241:36: error: ‘struct usb_serial_port’ has no member named ‘mutex’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:246:8: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:251:6: error: ‘struct usb_serial_port’ has no member named ‘tty’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:253:10: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:265:3: warning: passing argument 1 of ‘serial->type->open’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:265:3: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:265:3: warning: passing argument 2 of ‘serial->type->open’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:265:3: note: expected ‘struct usb_serial_port *’ but argument is of type ‘struct file *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:270:20: error: ‘struct usb_serial_port’ has no member named ‘mutex’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:276:6: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:278:6: error: ‘struct usb_serial_port’ has no member named ‘tty’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:279:20: error: ‘struct usb_serial_port’ has no member named ‘mutex’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_close’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:355:18: error: ‘struct usb_serial_port’ has no member named ‘mutex’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:357:10: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:358:21: error: ‘struct usb_serial_port’ has no member named ‘mutex’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:371:8: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:372:10: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:375:3: error: too many arguments to function ‘port->serial->type->close’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:377:11: error: ‘struct usb_serial_port’ has no member named ‘tty’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:378:12: error: ‘struct usb_serial_port’ has no member named ‘tty’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:379:9: error: ‘struct usb_serial_port’ has no member named ‘tty’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:380:8: error: ‘struct usb_serial_port’ has no member named ‘tty’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:386:20: error: ‘struct usb_serial_port’ has no member named ‘mutex’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_write’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:407:11: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:413:2: warning: passing argument 1 of ‘port->serial->type->write’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:413:2: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:413:2: warning: passing argument 2 of ‘port->serial->type->write’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:413:2: note: expected ‘struct usb_serial_port *’ but argument is of type ‘const unsigned char *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:413:2: warning: passing argument 3 of ‘port->serial->type->write’ makes pointer from integer without a cast [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:413:2: note: expected ‘const unsigned char *’ but argument is of type ‘int’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:413:2: error: too few arguments to function ‘port->serial->type->write’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_write_room’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:429:11: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:435:2: warning: passing argument 1 of ‘port->serial->type->write_room’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:435:2: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_chars_in_buffer’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:451:11: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:457:2: warning: passing argument 1 of ‘port->serial->type->chars_in_buffer’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:457:2: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_throttle’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:472:11: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:479:3: warning: passing argument 1 of ‘port->serial->type->throttle’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:479:3: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_unthrottle’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:491:11: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:498:3: warning: passing argument 1 of ‘port->serial->type->unthrottle’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:498:3: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_ioctl’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:511:11: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:518:3: warning: passing argument 1 of ‘port->serial->type->ioctl’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:518:3: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:518:3: warning: passing argument 2 of ‘port->serial->type->ioctl’ makes integer from pointer without a cast [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:518:3: note: expected ‘unsigned int’ but argument is of type ‘struct file *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:518:3: error: too many arguments to function ‘port->serial->type->ioctl’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_set_termios’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:535:11: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:542:3: warning: passing argument 1 of ‘port->serial->type->set_termios’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:542:3: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:542:3: warning: passing argument 2 of ‘port->serial->type->set_termios’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:542:3: note: expected ‘struct usb_serial_port *’ but argument is of type ‘struct termios *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:542:3: error: too few arguments to function ‘port->serial->type->set_termios’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_break’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:554:11: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:561:3: warning: passing argument 1 of ‘port->serial->type->break_ctl’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:561:3: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_tiocmget’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:629:11: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:635:3: warning: passing argument 1 of ‘port->serial->type->tiocmget’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:635:3: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:635:3: error: too many arguments to function ‘port->serial->type->tiocmget’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘serial_tiocmset’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:651:11: error: ‘struct usb_serial_port’ has no member named ‘open_count’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:657:3: warning: passing argument 1 of ‘port->serial->type->tiocmset’ from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:657:3: note: expected ‘struct tty_struct *’ but argument is of type ‘struct usb_serial_port *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:657:3: warning: passing argument 2 of ‘port->serial->type->tiocmset’ makes integer from pointer without a cast [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:657:3: note: expected ‘unsigned int’ but argument is of type ‘struct file *’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:657:3: error: too many arguments to function ‘port->serial->type->tiocmset’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘usb_serial_port_work’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:697:12: error: ‘struct usb_serial_port’ has no member named ‘tty’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘port_release’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:709:2: error: ‘struct device’ has no member named ‘bus_id’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘usb_serial_probe’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:858:2: error: implicit declaration of function ‘lock_kernel’ [-Werror=implicit-function-declaration] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:861:3: error: implicit declaration of function ‘unlock_kernel’ [-Werror=implicit-function-declaration] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1034:3: error: ‘struct usb_serial_port’ has no member named ‘mutex’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1182:23: error: ‘struct device’ has no member named ‘bus_id’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1182:51: error: ‘struct device’ has no member named ‘bus_id’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1183:3: error: ‘struct device’ has no member named ‘bus_id’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘usb_serial_disconnect’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1257:13: error: ‘struct usb_serial_port’ has no member named ‘tty’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1258:21: error: ‘struct usb_serial_port’ has no member named ‘tty’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: At top level: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1280:2: warning: initialization from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1280:2: warning: (near initialization for ‘serial_ops.ioctl’) [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1281:2: warning: initialization from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1281:2: warning: (near initialization for ‘serial_ops.set_termios’) [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1284:2: warning: initialization from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1284:2: warning: (near initialization for ‘serial_ops.break_ctl’) [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1286:2: error: unknown field ‘read_proc’ specified in initializer /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1286:2: warning: initialization from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1286:2: warning: (near initialization for ‘serial_ops.ioctl’) [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1287:2: warning: initialization from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1287:2: warning: (near initialization for ‘serial_ops.tiocmget’) [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1288:2: warning: initialization from incompatible pointer type [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1288:2: warning: (near initialization for ‘serial_ops.tiocmset’) [enabled by default] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘usb_serial_init’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1352:2: error: implicit declaration of function ‘info’ [-Werror=implicit-function-declaration] /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c: In function ‘fixup_generic’: /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1406:2: error: ‘struct usb_serial_driver’ has no member named ‘shutdown’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1406:2: error: ‘struct usb_serial_driver’ has no member named ‘shutdown’ /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1406:1: error: ‘usb_serial_generic_shutdown’ undeclared (first use in this function) /usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.c:1406:1: note: each undeclared identifier is reported only once for each function it appears in cc1: some warnings being treated as errors make[2]: *** [/usr/local/bin/ztemtApp/zteusbserial/below2.6.27/usb-serial.o] Error 1 make[1]: *** [_module_/usr/local/bin/ztemtApp/zteusbserial/below2.6.27] Error 2 make[1]: Leaving directory `/usr/src/linux-headers-3.0.0-12-generic' make: *** [modules] Error 2 Install finished! Any suggestion? Update: from installation document , In some special cases, the setup package can’t automatically compile the driver, so you need change the configurations yourself and manually compile the driver. Method: enter the directory: /usr/local/bin/ztemtApp/zteusbserial, and find the corresponding kernel version of current system. I can see 2.6.27 2.6.28 2.6.29 2.6.30 2.6.31 2.6.32 2.6.33 2.6.34 2.6.35 2.6.36 2.6.37 2.6.38 2.6.39 below2.6.2 in the directory. Which version shall I pick up for installation? Readme file says, we have to do this to insert ztemt.ko in kernel.

    Read the article

  • Problems configuring nameserver in plesk

    - by Saif Bechan
    Hello, i have some troubles with setting up a nameserver in PLESK for months now. I have tried all possible scenario's but i can not get this to work. I am really in need for some help, and if you can i will really appreciate it. Basically what i want is to just set up a nameserver in PLESK. I have a primary IP, and my host gave me a secondary nameserver i can use. My host is leaseweb in the netherlands. I have made some screenshots of the important parts in my opinion, maybe you guys can see some errors in them. To use the secondary nameserver provided by leaseweb i had to enable ACL on that account, i did so and made a screenshot of that too. The DNS recursion is set to localnets. These settings have not changed for months, so the dns should be fully updated everywhere. The check i run is the following: https://www.sidn.nl/over-nl/aanvraag...-server-check/ Domeinnaam (inclusief .nl): rdshosting.nl Eerste Nameserver: ns1.rdshosting.nl Eerste IP: 62.212.66.33 Tweede Nameserver: ns7.leaseweb.net Tweede ip: 62.212.76.50 If i run the dns check of the netherlands it gives me the following errors: primary name server "ns1.rdshosting.nl." Error: specified name server is not listed as NS record. All public name servers for a domain must also be listed as NS records in the zone of the domain. This domain was specified explicitly as a name server, but not found in the zone description of the primary name server. TE.6a rdshosting.nl. 86400 IN SOA ns1.rdspartners.nl. saif2k.hotmail.com. (2010031102 12H 1H 7D 3H) Error: the MNAME in SOA says "ns1.rdspartners.nl." is the primary name server. The MNAME field in the SOA record (first parameter) lists a different primary name server from the one specified for this check. RFC1035 section 3.3.13 rdshosting.nl. 86400 IN NS ns1.rdspartners.nl. Warning: hidden name server "ns1.rdspartners.nl." never used for first contact. The zone contains an NS record for a host which is not in the list of specified name servers. Hence, this name server will not be used to initiate contact to the domain. It may be used in sequential lookups, so it may still be useful. secondary name server "ns1.rdspartners.nl." [BROKEN] [HIDDEN] Failure: name server at 77.232.85.129 cannot be reached: (unknown error) The name server could not be contacted, which may be due to temporary technical problems or global DNS configuration mistakes. The internal error is shown, but not always clear about the cause. secondary name server "ns7.leaseweb.net." Info: name server looks correctly configured. I have the content of the file etc/named.conf also: // $Id: named.conf,v 1.1.1.1 2001/10/15 07:44:36 kap Exp $ // // Refer to the named(8) man page for details. If you are ever going // to setup a primary server, make sure you've understood the hairy // details of how DNS is working. Even with simple mistakes, you can // break connectivity for affected parties, or cause huge amount of // useless Internet traffic. options { allow-recursion { localnets; }; directory "/var"; auth-nxdomain no; pid-file "/var/run/named/named.pid"; // In addition to the "forwarders" clause, you can force your name // server to never initiate queries of its own, but always ask its // forwarders only, by enabling the following line: // // forward only; // If you've got a DNS server around at your upstream provider, enter // its IP address here, and enable the line below. This will make you // benefit from its cache, thus reduce overall DNS traffic in the Internet. /* forwarders { 127.0.0.1; }; */ /* * If there is a firewall between you and nameservers you want * to talk to, you might need to uncomment the query-source * directive below. Previous versions of BIND always asked * questions using port 53, but BIND 8.1 uses an unprivileged * port by default. */ // query-source address * port 53; /* * If running in a sandbox, you may have to specify a different * location for the dumpfile. */ // dump-file "s/named_dump.db"; }; //Use with the following in named.conf, adjusting the allow list as needed: key "rndc-key" { algorithm hmac-md5; secret "CeMgS23y0oWE20nyv0x40Q=="; }; controls { inet 127.0.0.1 port 953 allow { 127.0.0.1; } keys { "rndc-key"; }; }; // Note: the following will be supported in a future release. /* host { any; } { topology { 127.0.0.0/8; }; }; */ // Setting up secondaries is way easier and the rough picture for this // is explained below. // // If you enable a local name server, don't forget to enter 127.0.0.1 // into your /etc/resolv.conf so this server will be queried first. // Also, make sure to enable it in /etc/rc.conf. zone "." { type hint; file "named.root"; }; zone "0.0.127.IN-ADDR.ARPA" { type master; file "localhost.rev"; }; // NB: Do not use the IP addresses below, they are faked, and only // serve demonstration/documentation purposes! // // Example secondary config entries. It can be convenient to become // a secondary at least for the zone where your own domain is in. Ask // your network administrator for the IP address of the responsible // primary. // // Never forget to include the reverse lookup (IN-ADDR.ARPA) zone! // (This is the first bytes of the respective IP address, in reverse // order, with ".IN-ADDR.ARPA" appended.) // // Before starting to setup a primary zone, better make sure you fully // understand how DNS and BIND works, however. There are sometimes // unobvious pitfalls. Setting up a secondary is comparably simpler. // // NB: Don't blindly enable the examples below. :-) Use actual names // and addresses instead. // // NOTE!!! FreeBSD runs bind in a sandbox (see named_flags in rc.conf). // The directory containing the secondary zones must be write accessible // to bind. The following sequence is suggested: // // mkdir /etc/namedb/s // chown bind.bind /etc/namedb/s // chmod 750 /etc/namedb/s zone "rdshosting.nl" { type master; file "rdshosting.nl"; allow-transfer { 77.232.85.129; 62.212.76.50; common-allow-transfer; }; }; zone "66.212.62.in-addr.arpa" { type master; file "66.212.62.in-addr.arpa"; allow-transfer { common-allow-transfer; }; }; acl common-allow-transfer { 62.212.76.50; }; As i mentioned i made some screenshots of some parts: First the dns settings in plesk: http://www.freeimagehosting.net/uploads/2480faed5e.jpg Second the acl settings in plesk: http://www.freeimagehosting.net/uploads/777f5e69b0.jpg Third my settings at leaseweb: http://www.freeimagehosting.net/uploads/de7122b19c.jpg And last the secondary nameserver settings from leaseweb: http://www.freeimagehosting.net/uploads/fd1da38a8f.jpg If someone has anysuggestion at all on this this will be highly appriciated. Thank you for your time! PS. I am dutch so dutch answers are welcome aswell

    Read the article

  • Toorcon 15 (2013)

    - by danx
    The Toorcon gang (senior staff): h1kari (founder), nfiltr8, and Geo Introduction to Toorcon 15 (2013) A Tale of One Software Bypass of MS Windows 8 Secure Boot Breaching SSL, One Byte at a Time Running at 99%: Surviving an Application DoS Security Response in the Age of Mass Customized Attacks x86 Rewriting: Defeating RoP and other Shinanighans Clowntown Express: interesting bugs and running a bug bounty program Active Fingerprinting of Encrypted VPNs Making Attacks Go Backwards Mask Your Checksums—The Gorry Details Adventures with weird machines thirty years after "Reflections on Trusting Trust" Introduction to Toorcon 15 (2013) Toorcon 15 is the 15th annual security conference held in San Diego. I've attended about a third of them and blogged about previous conferences I attended here starting in 2003. As always, I've only summarized the talks I attended and interested me enough to write about them. Be aware that I may have misrepresented the speaker's remarks and that they are not my remarks or opinion, or those of my employer, so don't quote me or them. Those seeking further details may contact the speakers directly or use The Google. For some talks, I have a URL for further information. A Tale of One Software Bypass of MS Windows 8 Secure Boot Andrew Furtak and Oleksandr Bazhaniuk Yuri Bulygin, Oleksandr ("Alex") Bazhaniuk, and (not present) Andrew Furtak Yuri and Alex talked about UEFI and Bootkits and bypassing MS Windows 8 Secure Boot, with vendor recommendations. They previously gave this talk at the BlackHat 2013 conference. MS Windows 8 Secure Boot Overview UEFI (Unified Extensible Firmware Interface) is interface between hardware and OS. UEFI is processor and architecture independent. Malware can replace bootloader (bootx64.efi, bootmgfw.efi). Once replaced can modify kernel. Trivial to replace bootloader. Today many legacy bootkits—UEFI replaces them most of them. MS Windows 8 Secure Boot verifies everything you load, either through signatures or hashes. UEFI firmware relies on secure update (with signed update). You would think Secure Boot would rely on ROM (such as used for phones0, but you can't do that for PCs—PCs use writable memory with signatures DXE core verifies the UEFI boat loader(s) OS Loader (winload.efi, winresume.efi) verifies the OS kernel A chain of trust is established with a root key (Platform Key, PK), which is a cert belonging to the platform vendor. Key Exchange Keys (KEKs) verify an "authorized" database (db), and "forbidden" database (dbx). X.509 certs with SHA-1/SHA-256 hashes. Keys are stored in non-volatile (NV) flash-based NVRAM. Boot Services (BS) allow adding/deleting keys (can't be accessed once OS starts—which uses Run-Time (RT)). Root cert uses RSA-2048 public keys and PKCS#7 format signatures. SecureBoot — enable disable image signature checks SetupMode — update keys, self-signed keys, and secure boot variables CustomMode — allows updating keys Secure Boot policy settings are: always execute, never execute, allow execute on security violation, defer execute on security violation, deny execute on security violation, query user on security violation Attacking MS Windows 8 Secure Boot Secure Boot does NOT protect from physical access. Can disable from console. Each BIOS vendor implements Secure Boot differently. There are several platform and BIOS vendors. It becomes a "zoo" of implementations—which can be taken advantage of. Secure Boot is secure only when all vendors implement it correctly. Allow only UEFI firmware signed updates protect UEFI firmware from direct modification in flash memory protect FW update components program SPI controller securely protect secure boot policy settings in nvram protect runtime api disable compatibility support module which allows unsigned legacy Can corrupt the Platform Key (PK) EFI root certificate variable in SPI flash. If PK is not found, FW enters setup mode wich secure boot turned off. Can also exploit TPM in a similar manner. One is not supposed to be able to directly modify the PK in SPI flash from the OS though. But they found a bug that they can exploit from User Mode (undisclosed) and demoed the exploit. It loaded and ran their own bootkit. The exploit requires a reboot. Multiple vendors are vulnerable. They will disclose this exploit to vendors in the future. Recommendations: allow only signed updates protect UEFI fw in ROM protect EFI variable store in ROM Breaching SSL, One Byte at a Time Yoel Gluck and Angelo Prado Angelo Prado and Yoel Gluck, Salesforce.com CRIME is software that performs a "compression oracle attack." This is possible because the SSL protocol doesn't hide length, and because SSL compresses the header. CRIME requests with every possible character and measures the ciphertext length. Look for the plaintext which compresses the most and looks for the cookie one byte-at-a-time. SSL Compression uses LZ77 to reduce redundancy. Huffman coding replaces common byte sequences with shorter codes. US CERT thinks the SSL compression problem is fixed, but it isn't. They convinced CERT that it wasn't fixed and they issued a CVE. BREACH, breachattrack.com BREACH exploits the SSL response body (Accept-Encoding response, Content-Encoding). It takes advantage of the fact that the response is not compressed. BREACH uses gzip and needs fairly "stable" pages that are static for ~30 seconds. It needs attacker-supplied content (say from a web form or added to a URL parameter). BREACH listens to a session's requests and responses, then inserts extra requests and responses. Eventually, BREACH guesses a session's secret key. Can use compression to guess contents one byte at-a-time. For example, "Supersecret SupersecreX" (a wrong guess) compresses 10 bytes, and "Supersecret Supersecret" (a correct guess) compresses 11 bytes, so it can find each character by guessing every character. To start the guess, BREACH needs at least three known initial characters in the response sequence. Compression length then "leaks" information. Some roadblocks include no winners (all guesses wrong) or too many winners (multiple possibilities that compress the same). The solutions include: lookahead (guess 2 or 3 characters at-a-time instead of 1 character). Expensive rollback to last known conflict check compression ratio can brute-force first 3 "bootstrap" characters, if needed (expensive) block ciphers hide exact plain text length. Solution is to align response in advance to block size Mitigations length: use variable padding secrets: dynamic CSRF tokens per request secret: change over time separate secret to input-less servlets Future work eiter understand DEFLATE/GZIP HTTPS extensions Running at 99%: Surviving an Application DoS Ryan Huber Ryan Huber, Risk I/O Ryan first discussed various ways to do a denial of service (DoS) attack against web services. One usual method is to find a slow web page and do several wgets. Or download large files. Apache is not well suited at handling a large number of connections, but one can put something in front of it Can use Apache alternatives, such as nginx How to identify malicious hosts short, sudden web requests user-agent is obvious (curl, python) same url requested repeatedly no web page referer (not normal) hidden links. hide a link and see if a bot gets it restricted access if not your geo IP (unless the website is global) missing common headers in request regular timing first seen IP at beginning of attack count requests per hosts (usually a very large number) Use of captcha can mitigate attacks, but you'll lose a lot of genuine users. Bouncer, goo.gl/c2vyEc and www.github.com/rawdigits/Bouncer Bouncer is software written by Ryan in netflow. Bouncer has a small, unobtrusive footprint and detects DoS attempts. It closes blacklisted sockets immediately (not nice about it, no proper close connection). Aggregator collects requests and controls your web proxies. Need NTP on the front end web servers for clean data for use by bouncer. Bouncer is also useful for a popularity storm ("Slashdotting") and scraper storms. Future features: gzip collection data, documentation, consumer library, multitask, logging destroyed connections. Takeaways: DoS mitigation is easier with a complete picture Bouncer designed to make it easier to detect and defend DoS—not a complete cure Security Response in the Age of Mass Customized Attacks Peleus Uhley and Karthik Raman Peleus Uhley and Karthik Raman, Adobe ASSET, blogs.adobe.com/asset/ Peleus and Karthik talked about response to mass-customized exploits. Attackers behave much like a business. "Mass customization" refers to concept discussed in the book Future Perfect by Stan Davis of Harvard Business School. Mass customization is differentiating a product for an individual customer, but at a mass production price. For example, the same individual with a debit card receives basically the same customized ATM experience around the world. Or designing your own PC from commodity parts. Exploit kits are another example of mass customization. The kits support multiple browsers and plugins, allows new modules. Exploit kits are cheap and customizable. Organized gangs use exploit kits. A group at Berkeley looked at 77,000 malicious websites (Grier et al., "Manufacturing Compromise: The Emergence of Exploit-as-a-Service", 2012). They found 10,000 distinct binaries among them, but derived from only a dozen or so exploit kits. Characteristics of Mass Malware: potent, resilient, relatively low cost Technical characteristics: multiple OS, multipe payloads, multiple scenarios, multiple languages, obfuscation Response time for 0-day exploits has gone down from ~40 days 5 years ago to about ~10 days now. So the drive with malware is towards mass customized exploits, to avoid detection There's plenty of evicence that exploit development has Project Manager bureaucracy. They infer from the malware edicts to: support all versions of reader support all versions of windows support all versions of flash support all browsers write large complex, difficult to main code (8750 lines of JavaScript for example Exploits have "loose coupling" of multipe versions of software (adobe), OS, and browser. This allows specific attacks against specific versions of multiple pieces of software. Also allows exploits of more obscure software/OS/browsers and obscure versions. Gave examples of exploits that exploited 2, 3, 6, or 14 separate bugs. However, these complete exploits are more likely to be buggy or fragile in themselves and easier to defeat. Future research includes normalizing malware and Javascript. Conclusion: The coming trend is that mass-malware with mass zero-day attacks will result in mass customization of attacks. x86 Rewriting: Defeating RoP and other Shinanighans Richard Wartell Richard Wartell The attack vector we are addressing here is: First some malware causes a buffer overflow. The malware has no program access, but input access and buffer overflow code onto stack Later the stack became non-executable. The workaround malware used was to write a bogus return address to the stack jumping to malware Later came ASLR (Address Space Layout Randomization) to randomize memory layout and make addresses non-deterministic. The workaround malware used was to jump t existing code segments in the program that can be used in bad ways "RoP" is Return-oriented Programming attacks. RoP attacks use your own code and write return address on stack to (existing) expoitable code found in program ("gadgets"). Pinkie Pie was paid $60K last year for a RoP attack. One solution is using anti-RoP compilers that compile source code with NO return instructions. ASLR does not randomize address space, just "gadgets". IPR/ILR ("Instruction Location Randomization") randomizes each instruction with a virtual machine. Richard's goal was to randomize a binary with no source code access. He created "STIR" (Self-Transofrming Instruction Relocation). STIR disassembles binary and operates on "basic blocks" of code. The STIR disassembler is conservative in what to disassemble. Each basic block is moved to a random location in memory. Next, STIR writes new code sections with copies of "basic blocks" of code in randomized locations. The old code is copied and rewritten with jumps to new code. the original code sections in the file is marked non-executible. STIR has better entropy than ASLR in location of code. Makes brute force attacks much harder. STIR runs on MS Windows (PEM) and Linux (ELF). It eliminated 99.96% or more "gadgets" (i.e., moved the address). Overhead usually 5-10% on MS Windows, about 1.5-4% on Linux (but some code actually runs faster!). The unique thing about STIR is it requires no source access and the modified binary fully works! Current work is to rewrite code to enforce security policies. For example, don't create a *.{exe,msi,bat} file. Or don't connect to the network after reading from the disk. Clowntown Express: interesting bugs and running a bug bounty program Collin Greene Collin Greene, Facebook Collin talked about Facebook's bug bounty program. Background at FB: FB has good security frameworks, such as security teams, external audits, and cc'ing on diffs. But there's lots of "deep, dark, forgotten" parts of legacy FB code. Collin gave several examples of bountied bugs. Some bounty submissions were on software purchased from a third-party (but bounty claimers don't know and don't care). We use security questions, as does everyone else, but they are basically insecure (often easily discoverable). Collin didn't expect many bugs from the bounty program, but they ended getting 20+ good bugs in first 24 hours and good submissions continue to come in. Bug bounties bring people in with different perspectives, and are paid only for success. Bug bounty is a better use of a fixed amount of time and money versus just code review or static code analysis. The Bounty program started July 2011 and paid out $1.5 million to date. 14% of the submissions have been high priority problems that needed to be fixed immediately. The best bugs come from a small % of submitters (as with everything else)—the top paid submitters are paid 6 figures a year. Spammers like to backstab competitors. The youngest sumitter was 13. Some submitters have been hired. Bug bounties also allows to see bugs that were missed by tools or reviews, allowing improvement in the process. Bug bounties might not work for traditional software companies where the product has release cycle or is not on Internet. Active Fingerprinting of Encrypted VPNs Anna Shubina Anna Shubina, Dartmouth Institute for Security, Technology, and Society (I missed the start of her talk because another track went overtime. But I have the DVD of the talk, so I'll expand later) IPsec leaves fingerprints. Using netcat, one can easily visually distinguish various crypto chaining modes just from packet timing on a chart (example, DES-CBC versus AES-CBC) One can tell a lot about VPNs just from ping roundtrips (such as what router is used) Delayed packets are not informative about a network, especially if far away from the network More needed to explore about how TCP works in real life with respect to timing Making Attacks Go Backwards Fuzzynop FuzzyNop, Mandiant This talk is not about threat attribution (finding who), product solutions, politics, or sales pitches. But who are making these malware threats? It's not a single person or group—they have diverse skill levels. There's a lot of fat-fingered fumblers out there. Always look for low-hanging fruit first: "hiding" malware in the temp, recycle, or root directories creation of unnamed scheduled tasks obvious names of files and syscalls ("ClearEventLog") uncleared event logs. Clearing event log in itself, and time of clearing, is a red flag and good first clue to look for on a suspect system Reverse engineering is hard. Disassembler use takes practice and skill. A popular tool is IDA Pro, but it takes multiple interactive iterations to get a clean disassembly. Key loggers are used a lot in targeted attacks. They are typically custom code or built in a backdoor. A big tip-off is that non-printable characters need to be printed out (such as "[Ctrl]" "[RightShift]") or time stamp printf strings. Look for these in files. Presence is not proof they are used. Absence is not proof they are not used. Java exploits. Can parse jar file with idxparser.py and decomile Java file. Java typially used to target tech companies. Backdoors are the main persistence mechanism (provided externally) for malware. Also malware typically needs command and control. Application of Artificial Intelligence in Ad-Hoc Static Code Analysis John Ashaman John Ashaman, Security Innovation Initially John tried to analyze open source files with open source static analysis tools, but these showed thousands of false positives. Also tried using grep, but tis fails to find anything even mildly complex. So next John decided to write his own tool. His approach was to first generate a call graph then analyze the graph. However, the problem is that making a call graph is really hard. For example, one problem is "evil" coding techniques, such as passing function pointer. First the tool generated an Abstract Syntax Tree (AST) with the nodes created from method declarations and edges created from method use. Then the tool generated a control flow graph with the goal to find a path through the AST (a maze) from source to sink. The algorithm is to look at adjacent nodes to see if any are "scary" (a vulnerability), using heuristics for search order. The tool, called "Scat" (Static Code Analysis Tool), currently looks for C# vulnerabilities and some simple PHP. Later, he plans to add more PHP, then JSP and Java. For more information see his posts in Security Innovation blog and NRefactory on GitHub. Mask Your Checksums—The Gorry Details Eric (XlogicX) Davisson Eric (XlogicX) Davisson Sometimes in emailing or posting TCP/IP packets to analyze problems, you may want to mask the IP address. But to do this correctly, you need to mask the checksum too, or you'll leak information about the IP. Problem reports found in stackoverflow.com, sans.org, and pastebin.org are usually not masked, but a few companies do care. If only the IP is masked, the IP may be guessed from checksum (that is, it leaks data). Other parts of packet may leak more data about the IP. TCP and IP checksums both refer to the same data, so can get more bits of information out of using both checksums than just using one checksum. Also, one can usually determine the OS from the TTL field and ports in a packet header. If we get hundreds of possible results (16x each masked nibble that is unknown), one can do other things to narrow the results, such as look at packet contents for domain or geo information. With hundreds of results, can import as CSV format into a spreadsheet. Can corelate with geo data and see where each possibility is located. Eric then demoed a real email report with a masked IP packet attached. Was able to find the exact IP address, given the geo and university of the sender. Point is if you're going to mask a packet, do it right. Eric wouldn't usually bother, but do it correctly if at all, to not create a false impression of security. Adventures with weird machines thirty years after "Reflections on Trusting Trust" Sergey Bratus Sergey Bratus, Dartmouth College (and Julian Bangert and Rebecca Shapiro, not present) "Reflections on Trusting Trust" refers to Ken Thompson's classic 1984 paper. "You can't trust code that you did not totally create yourself." There's invisible links in the chain-of-trust, such as "well-installed microcode bugs" or in the compiler, and other planted bugs. Thompson showed how a compiler can introduce and propagate bugs in unmodified source. But suppose if there's no bugs and you trust the author, can you trust the code? Hell No! There's too many factors—it's Babylonian in nature. Why not? Well, Input is not well-defined/recognized (code's assumptions about "checked" input will be violated (bug/vunerabiliy). For example, HTML is recursive, but Regex checking is not recursive. Input well-formed but so complex there's no telling what it does For example, ELF file parsing is complex and has multiple ways of parsing. Input is seen differently by different pieces of program or toolchain Any Input is a program input executes on input handlers (drives state changes & transitions) only a well-defined execution model can be trusted (regex/DFA, PDA, CFG) Input handler either is a "recognizer" for the inputs as a well-defined language (see langsec.org) or it's a "virtual machine" for inputs to drive into pwn-age ELF ABI (UNIX/Linux executible file format) case study. Problems can arise from these steps (without planting bugs): compiler linker loader ld.so/rtld relocator DWARF (debugger info) exceptions The problem is you can't really automatically analyze code (it's the "halting problem" and undecidable). Only solution is to freeze code and sign it. But you can't freeze everything! Can't freeze ASLR or loading—must have tables and metadata. Any sufficiently complex input data is the same as VM byte code Example, ELF relocation entries + dynamic symbols == a Turing Complete Machine (TM). @bxsays created a Turing machine in Linux from relocation data (not code) in an ELF file. For more information, see Rebecca "bx" Shapiro's presentation from last year's Toorcon, "Programming Weird Machines with ELF Metadata" @bxsays did same thing with Mach-O bytecode Or a DWARF exception handling data .eh_frame + glibc == Turning Machine X86 MMU (IDT, GDT, TSS): used address translation to create a Turning Machine. Page handler reads and writes (on page fault) memory. Uses a page table, which can be used as Turning Machine byte code. Example on Github using this TM that will fly a glider across the screen Next Sergey talked about "Parser Differentials". That having one input format, but two parsers, will create confusion and opportunity for exploitation. For example, CSRs are parsed during creation by cert requestor and again by another parser at the CA. Another example is ELF—several parsers in OS tool chain, which are all different. Can have two different Program Headers (PHDRs) because ld.so parses multiple PHDRs. The second PHDR can completely transform the executable. This is described in paper in the first issue of International Journal of PoC. Conclusions trusting computers not only about bugs! Bugs are part of a problem, but no by far all of it complex data formats means bugs no "chain of trust" in Babylon! (that is, with parser differentials) we need to squeeze complexity out of data until data stops being "code equivalent" Further information See and langsec.org. USENIX WOOT 2013 (Workshop on Offensive Technologies) for "weird machines" papers and videos.

    Read the article

  • Elfsign Object Signing on Solaris

    - by danx
    Elfsign Object Signing on Solaris Don't let this happen to you—use elfsign! Solaris elfsign(1) is a command that signs and verifies ELF format executables. That includes not just executable programs (such as ls or cp), but other ELF format files including libraries (such as libnvpair.so) and kernel modules (such as autofs). Elfsign has been available since Solaris 10 and ELF format files distributed with Solaris, since Solaris 10, are signed by either Sun Microsystems or its successor, Oracle Corporation. When an ELF file is signed, elfsign adds a new section the ELF file, .SUNW_signature, that contains a RSA public key signature and other information about the signer. That is, the algorithm used, algorithm OID, signer CN/OU, and time stamp. The signature section can later be verified by elfsign or other software by matching the signature in the file agains the ELF file contents (excluding the signature). ELF executable files may also be signed by a 3rd-party or by the customer. This is useful for verifying the origin and authenticity of executable files installed on a system. The 3rd-party or customer public key certificate should be installed in /etc/certs/ to allow verification by elfsign. For currently-released versions of Solaris, only cryptographic framework plugin libraries are verified by Solaris. However, all ELF files may be verified by the elfsign command at any time. Elfsign Algorithms Elfsign signatures are created by taking a digest of the ELF section contents, then signing the digest with RSA. To verify, one takes a digest of ELF file and compares with the expected digest that's computed from the signature and RSA public key. Originally elfsign took a MD5 digest of a SHA-1 digest of the ELF file sections, then signed the resulting digest with RSA. In Solaris 11.1 then Solaris 11.1 SRU 7 (5/2013), the elfsign crypto algorithms available have been expanded to keep up with evolving cryptography. The following table shows the available elfsign algorithms: Elfsign Algorithm Solaris Release Comments elfsign sign -F rsa_md5_sha1   S10, S11.0, S11.1 Default for S10. Not recommended* elfsign sign -F rsa_sha1 S11.1 Default for S11.1. Not recommended elfsign sign -F rsa_sha256 S11.1 patch SRU7+   Recommended ___ *Most or all CAs do not accept MD5 CSRs and do not issue MD5 certs due to MD5 hash collision problems. RSA Key Length. I recommend using RSA-2048 key length with elfsign is RSA-2048 as the best balance between a long expected "life time", interoperability, and performance. RSA-2048 keys have an expected lifetime through 2030 (and probably beyond). For details, see Recommendation for Key Management: Part 1: General, NIST Publication SP 800-57 part 1 (rev. 3, 7/2012, PDF), tables 2 and 4 (pp. 64, 67). Step 1: create or obtain a key and cert The first step in using elfsign is to obtain a key and cert from a public Certificate Authority (CA), or create your own self-signed key and cert. I'll briefly explain both methods. Obtaining a Certificate from a CA To obtain a cert from a CA, such as Verisign, Thawte, or Go Daddy (to name a few random examples), you create a private key and a Certificate Signing Request (CSR) file and send it to the CA, following the instructions of the CA on their website. They send back a signed public key certificate. The public key cert, along with the private key you created is used by elfsign to sign an ELF file. The public key cert is distributed with the software and is used by elfsign to verify elfsign signatures in ELF files. You need to request a RSA "Class 3 public key certificate", which is used for servers and software signing. Elfsign uses RSA and we recommend RSA-2048 keys. The private key and CSR can be generated with openssl(1) or pktool(1) on Solaris. Here's a simple example that uses pktool to generate a private RSA_2048 key and a CSR for sending to a CA: $ pktool gencsr keystore=file format=pem outcsr=MYCSR.p10 \ subject="CN=canineswworks.com,OU=Canine SW object signing" \ outkey=MYPRIVATEKEY.key $ openssl rsa -noout -text -in MYPRIVATEKEY.key Private-Key: (2048 bit) modulus: 00:d2:ef:42:f2:0b:8c:96:9f:45:32:fc:fe:54:94: . . . [omitted for brevity] . . . c9:c7 publicExponent: 65537 (0x10001) privateExponent: 26:14:fc:49:26:bc:a3:14:ee:31:5e:6b:ac:69:83: . . . [omitted for brevity] . . . 81 prime1: 00:f6:b7:52:73:bc:26:57:26:c8:11:eb:6c:dc:cb: . . . [omitted for brevity] . . . bc:91:d0:40:d6:9d:ac:b5:69 prime2: 00:da:df:3f:56:b2:18:46:e1:89:5b:6c:f1:1a:41: . . . [omitted for brevity] . . . f3:b7:48:de:c3:d9:ce:af:af exponent1: 00:b9:a2:00:11:02:ed:9a:3f:9c:e4:16:ce:c7:67: . . . [omitted for brevity] . . . 55:50:25:70:d3:ca:b9:ab:99 exponent2: 00:c8:fc:f5:57:11:98:85:8e:9a:ea:1f:f2:8f:df: . . . [omitted for brevity] . . . 23:57:0e:4d:b2:a0:12:d2:f5 coefficient: 2f:60:21:cd:dc:52:76:67:1a:d8:75:3e:7f:b0:64: . . . [omitted for brevity] . . . 06:94:56:d8:9d:5c:8e:9b $ openssl req -noout -text -in MYCSR.p10 Certificate Request: Data: Version: 2 (0x2) Subject: OU=Canine SW object signing, CN=canineswworks.com Subject Public Key Info: Public Key Algorithm: rsaEncryption Public-Key: (2048 bit) Modulus: 00:d2:ef:42:f2:0b:8c:96:9f:45:32:fc:fe:54:94: . . . [omitted for brevity] . . . c9:c7 Exponent: 65537 (0x10001) Attributes: Signature Algorithm: sha1WithRSAEncryption b3:e8:30:5b:88:37:68:1c:26:6b:45:af:5e:de:ea:60:87:ea: . . . [omitted for brevity] . . . 06:f9:ed:b4 Secure storage of RSA private key. The private key needs to be protected if the key signing is used for production (as opposed to just testing). That is, protect the key to protect against unauthorized signatures by others. One method is to use a PIN-protected PKCS#11 keystore. The private key you generate should be stored in a secure manner, such as in a PKCS#11 keystore using pktool(1). Otherwise others can sign your signature. Other secure key storage mechanisms include a SCA-6000 crypto card, a USB thumb drive stored in a locked area, a dedicated server with restricted access, Oracle Key Manager (OKM), or some combination of these. I also recommend secure backup of the private key. Here's an example of generating a private key protected in the PKCS#11 keystore, and a CSR. $ pktool setpin # use if PIN not set yet Enter token passphrase: changeme Create new passphrase: Re-enter new passphrase: Passphrase changed. $ pktool gencsr keystore=pkcs11 label=MYPRIVATEKEY \ format=pem outcsr=MYCSR.p10 \ subject="CN=canineswworks.com,OU=Canine SW object signing" $ pktool list keystore=pkcs11 Enter PIN for Sun Software PKCS#11 softtoken: Found 1 asymmetric public keys. Key #1 - RSA public key: MYPRIVATEKEY Here's another example that uses openssl instead of pktool to generate a private key and CSR: $ openssl genrsa -out cert.key 2048 $ openssl req -new -key cert.key -out MYCSR.p10 Self-Signed Cert You can use openssl or pktool to create a private key and a self-signed public key certificate. A self-signed cert is useful for development, testing, and internal use. The private key created should be stored in a secure manner, as mentioned above. The following example creates a private key, MYSELFSIGNED.key, and a public key cert, MYSELFSIGNED.pem, using pktool and displays the contents with the openssl command. $ pktool gencert keystore=file format=pem serial=0xD06F00D lifetime=20-year \ keytype=rsa hash=sha256 outcert=MYSELFSIGNED.pem outkey=MYSELFSIGNED.key \ subject="O=Canine Software Works, OU=Self-signed CA, CN=canineswworks.com" $ pktool list keystore=file objtype=cert infile=MYSELFSIGNED.pem Found 1 certificates. 1. (X.509 certificate) Filename: MYSELFSIGNED.pem ID: c8:24:59:08:2b:ae:6e:5c:bc:26:bd:ef:0a:9c:54:de:dd:0f:60:46 Subject: O=Canine Software Works, OU=Self-signed CA, CN=canineswworks.com Issuer: O=Canine Software Works, OU=Self-signed CA, CN=canineswworks.com Not Before: Oct 17 23:18:00 2013 GMT Not After: Oct 12 23:18:00 2033 GMT Serial: 0xD06F00D0 Signature Algorithm: sha256WithRSAEncryption $ openssl x509 -noout -text -in MYSELFSIGNED.pem Certificate: Data: Version: 3 (0x2) Serial Number: 3496935632 (0xd06f00d0) Signature Algorithm: sha256WithRSAEncryption Issuer: O=Canine Software Works, OU=Self-signed CA, CN=canineswworks.com Validity Not Before: Oct 17 23:18:00 2013 GMT Not After : Oct 12 23:18:00 2033 GMT Subject: O=Canine Software Works, OU=Self-signed CA, CN=canineswworks.com Subject Public Key Info: Public Key Algorithm: rsaEncryption Public-Key: (2048 bit) Modulus: 00:bb:e8:11:21:d9:4b:88:53:8b:6c:5a:7a:38:8b: . . . [omitted for brevity] . . . bf:77 Exponent: 65537 (0x10001) Signature Algorithm: sha256WithRSAEncryption 9e:39:fe:c8:44:5c:87:2c:8f:f4:24:f6:0c:9a:2f:64:84:d1: . . . [omitted for brevity] . . . 5f:78:8e:e8 $ openssl rsa -noout -text -in MYSELFSIGNED.key Private-Key: (2048 bit) modulus: 00:bb:e8:11:21:d9:4b:88:53:8b:6c:5a:7a:38:8b: . . . [omitted for brevity] . . . bf:77 publicExponent: 65537 (0x10001) privateExponent: 0a:06:0f:23:e7:1b:88:62:2c:85:d3:2d:c1:e6:6e: . . . [omitted for brevity] . . . 9c:e1:e0:0a:52:77:29:4a:75:aa:02:d8:af:53:24: c1 prime1: 00:ea:12:02:bb:5a:0f:5a:d8:a9:95:b2:ba:30:15: . . . [omitted for brevity] . . . 5b:ca:9c:7c:19:48:77:1e:5d prime2: 00:cd:82:da:84:71:1d:18:52:cb:c6:4d:74:14:be: . . . [omitted for brevity] . . . 5f:db:d5:5e:47:89:a7:ef:e3 exponent1: 32:37:62:f6:a6:bf:9c:91:d6:f0:12:c3:f7:04:e9: . . . [omitted for brevity] . . . 97:3e:33:31:89:66:64:d1 exponent2: 00:88:a2:e8:90:47:f8:75:34:8f:41:50:3b:ce:93: . . . [omitted for brevity] . . . ff:74:d4:be:f3:47:45:bd:cb coefficient: 4d:7c:09:4c:34:73:c4:26:f0:58:f5:e1:45:3c:af: . . . [omitted for brevity] . . . af:01:5f:af:ad:6a:09:bf Step 2: Sign the ELF File object By now you should have your private key, and obtained, by hook or crook, a cert (either from a CA or use one you created (a self-signed cert). The next step is to sign one or more objects with your private key and cert. Here's a simple example that creates an object file, signs, verifies, and lists the contents of the ELF signature. $ echo '#include <stdio.h>\nint main(){printf("Hello\\n");}'>hello.c $ make hello cc -o hello hello.c $ elfsign verify -v -c MYSELFSIGNED.pem -e hello elfsign: no signature found in hello. $ elfsign sign -F rsa_sha256 -v -k MYSELFSIGNED.key -c MYSELFSIGNED.pem -e hello elfsign: hello signed successfully. format: rsa_sha256. signer: O=Canine Software Works, OU=Self-signed CA, CN=canineswworks.com. signed on: October 17, 2013 04:22:49 PM PDT. $ elfsign list -f format -e hello rsa_sha256 $ elfsign list -f signer -e hello O=Canine Software Works, OU=Self-signed CA, CN=canineswworks.com $ elfsign list -f time -e hello October 17, 2013 04:22:49 PM PDT $ elfsign verify -v -c MYSELFSIGNED.key -e hello elfsign: verification of hello failed. format: rsa_sha256. signer: O=Canine Software Works, OU=Self-signed CA, CN=canineswworks.com. signed on: October 17, 2013 04:22:49 PM PDT. Signing using the pkcs11 keystore To sign the ELF file using a private key in the secure pkcs11 keystore, replace "-K MYSELFSIGNED.key" in the "elfsign sign" command line with "-T MYPRIVATEKEY", where MYPRIVATKEY is the pkcs11 token label. Step 3: Install the cert and test on another system Just signing the object isn't enough. You need to copy or install the cert and the signed ELF file(s) on another system to test that the signature is OK. Your public key cert should be installed in /etc/certs. Use elfsign verify to verify the signature. Elfsign verify checks each cert in /etc/certs until it finds one that matches the elfsign signature in the file. If one isn't found, the verification fails. Here's an example: $ su Password: # rm /etc/certs/MYSELFSIGNED.key # cp MYSELFSIGNED.pem /etc/certs # exit $ elfsign verify -v hello elfsign: verification of hello passed. format: rsa_sha256. signer: O=Canine Software Works, OU=Self-signed CA, CN=canineswworks.com. signed on: October 17, 2013 04:24:20 PM PDT. After testing, package your cert along with your ELF object to allow elfsign verification after your cert and object are installed or copied. Under the Hood: elfsign verification Here's the steps taken to verify a ELF file signed with elfsign. The steps to sign the file are similar except the private key exponent is used instead of the public key exponent and the .SUNW_signature section is written to the ELF file instead of being read from the file. Generate a digest (SHA-256) of the ELF file sections. This digest uses all ELF sections loaded in memory, but excludes the ELF header, the .SUNW_signature section, and the symbol table Extract the RSA signature (RSA-2048) from the .SUNW_signature section Extract the RSA public key modulus and public key exponent (65537) from the public key cert Calculate the expected digest as follows:     signaturepublicKeyExponent % publicKeyModulus Strip the PKCS#1 padding (most significant bytes) from the above. The padding is 0x00, 0x01, 0xff, 0xff, . . ., 0xff, 0x00. If the actual digest == expected digest, the ELF file is verified (OK). Further Information elfsign(1), pktool(1), and openssl(1) man pages. "Signed Solaris 10 Binaries?" blog by Darren Moffat (2005) shows how to use elfsign. "Simple CLI based CA on Solaris" blog by Darren Moffat (2008) shows how to set up a simple CA for use with self-signed certificates. "How to Create a Certificate by Using the pktool gencert Command" System Administration Guide: Security Services (available at docs.oracle.com)

    Read the article

  • Node.js Adventure - Host Node.js on Windows Azure Worker Role

    - by Shaun
    In my previous post I demonstrated about how to develop and deploy a Node.js application on Windows Azure Web Site (a.k.a. WAWS). WAWS is a new feature in Windows Azure platform. Since it’s low-cost, and it provides IIS and IISNode components so that we can host our Node.js application though Git, FTP and WebMatrix without any configuration and component installation. But sometimes we need to use the Windows Azure Cloud Service (a.k.a. WACS) and host our Node.js on worker role. Below are some benefits of using worker role. - WAWS leverages IIS and IISNode to host Node.js application, which runs in x86 WOW mode. It reduces the performance comparing with x64 in some cases. - WACS worker role does not need IIS, hence there’s no restriction of IIS, such as 8000 concurrent requests limitation. - WACS provides more flexibility and controls to the developers. For example, we can RDP to the virtual machines of our worker role instances. - WACS provides the service configuration features which can be changed when the role is running. - WACS provides more scaling capability than WAWS. In WAWS we can have at most 3 reserved instances per web site while in WACS we can have up to 20 instances in a subscription. - Since when using WACS worker role we starts the node by ourselves in a process, we can control the input, output and error stream. We can also control the version of Node.js.   Run Node.js in Worker Role Node.js can be started by just having its execution file. This means in Windows Azure, we can have a worker role with the “node.exe” and the Node.js source files, then start it in Run method of the worker role entry class. Let’s create a new windows azure project in Visual Studio and add a new worker role. Since we need our worker role execute the “node.exe” with our application code we need to add the “node.exe” into our project. Right click on the worker role project and add an existing item. By default the Node.js will be installed in the “Program Files\nodejs” folder so we can navigate there and add the “node.exe”. Then we need to create the entry code of Node.js. In WAWS the entry file must be named “server.js”, which is because it’s hosted by IIS and IISNode and IISNode only accept “server.js”. But here as we control everything we can choose any files as the entry code. For example, I created a new JavaScript file named “index.js” in project root. Since we created a C# Windows Azure project we cannot create a JavaScript file from the context menu “Add new item”. We have to create a text file, and then rename it to JavaScript extension. After we added these two files we should set their “Copy to Output Directory” property to “Copy Always”, or “Copy if Newer”. Otherwise they will not be involved in the package when deployed. Let’s paste a very simple Node.js code in the “index.js” as below. As you can see I created a web server listening at port 12345. 1: var http = require("http"); 2: var port = 12345; 3:  4: http.createServer(function (req, res) { 5: res.writeHead(200, { "Content-Type": "text/plain" }); 6: res.end("Hello World\n"); 7: }).listen(port); 8:  9: console.log("Server running at port %d", port); Then we need to start “node.exe” with this file when our worker role was started. This can be done in its Run method. I found the Node.js and entry JavaScript file name, and then create a new process to run it. Our worker role will wait for the process to be exited. If everything is OK once our web server was opened the process will be there listening for incoming requests, and should not be terminated. The code in worker role would be like this. 1: public override void Run() 2: { 3: // This is a sample worker implementation. Replace with your logic. 4: Trace.WriteLine("NodejsHost entry point called", "Information"); 5:  6: // retrieve the node.exe and entry node.js source code file name. 7: var node = Environment.ExpandEnvironmentVariables(@"%RoleRoot%\approot\node.exe"); 8: var js = "index.js"; 9:  10: // prepare the process starting of node.exe 11: var info = new ProcessStartInfo(node, js) 12: { 13: CreateNoWindow = false, 14: ErrorDialog = true, 15: WindowStyle = ProcessWindowStyle.Normal, 16: UseShellExecute = false, 17: WorkingDirectory = Environment.ExpandEnvironmentVariables(@"%RoleRoot%\approot") 18: }; 19: Trace.WriteLine(string.Format("{0} {1}", node, js), "Information"); 20:  21: // start the node.exe with entry code and wait for exit 22: var process = Process.Start(info); 23: process.WaitForExit(); 24: } Then we can run it locally. In the computer emulator UI the worker role started and it executed the Node.js, then Node.js windows appeared. Open the browser to verify the website hosted by our worker role. Next let’s deploy it to azure. But we need some additional steps. First, we need to create an input endpoint. By default there’s no endpoint defined in a worker role. So we will open the role property window in Visual Studio, create a new input TCP endpoint to the port we want our website to use. In this case I will use 80. Even though we created a web server we should add a TCP endpoint of the worker role, since Node.js always listen on TCP instead of HTTP. And then changed the “index.js”, let our web server listen on 80. 1: var http = require("http"); 2: var port = 80; 3:  4: http.createServer(function (req, res) { 5: res.writeHead(200, { "Content-Type": "text/plain" }); 6: res.end("Hello World\n"); 7: }).listen(port); 8:  9: console.log("Server running at port %d", port); Then publish it to Windows Azure. And then in browser we can see our Node.js website was running on WACS worker role. We may encounter an error if we tried to run our Node.js website on 80 port at local emulator. This is because the compute emulator registered 80 and map the 80 endpoint to 81. But our Node.js cannot detect this operation. So when it tried to listen on 80 it will failed since 80 have been used.   Use NPM Modules When we are using WAWS to host Node.js, we can simply install modules we need, and then just publish or upload all files to WAWS. But if we are using WACS worker role, we have to do some extra steps to make the modules work. Assuming that we plan to use “express” in our application. Firstly of all we should download and install this module through NPM command. But after the install finished, they are just in the disk but not included in the worker role project. If we deploy the worker role right now the module will not be packaged and uploaded to azure. Hence we need to add them to the project. On solution explorer window click the “Show all files” button, select the “node_modules” folder and in the context menu select “Include In Project”. But that not enough. We also need to make all files in this module to “Copy always” or “Copy if newer”, so that they can be uploaded to azure with the “node.exe” and “index.js”. This is painful step since there might be many files in a module. So I created a small tool which can update a C# project file, make its all items as “Copy always”. The code is very simple. 1: static void Main(string[] args) 2: { 3: if (args.Length < 1) 4: { 5: Console.WriteLine("Usage: copyallalways [project file]"); 6: return; 7: } 8:  9: var proj = args[0]; 10: File.Copy(proj, string.Format("{0}.bak", proj)); 11:  12: var xml = new XmlDocument(); 13: xml.Load(proj); 14: var nsManager = new XmlNamespaceManager(xml.NameTable); 15: nsManager.AddNamespace("pf", "http://schemas.microsoft.com/developer/msbuild/2003"); 16:  17: // add the output setting to copy always 18: var contentNodes = xml.SelectNodes("//pf:Project/pf:ItemGroup/pf:Content", nsManager); 19: UpdateNodes(contentNodes, xml, nsManager); 20: var noneNodes = xml.SelectNodes("//pf:Project/pf:ItemGroup/pf:None", nsManager); 21: UpdateNodes(noneNodes, xml, nsManager); 22: xml.Save(proj); 23:  24: // remove the namespace attributes 25: var content = xml.InnerXml.Replace("<CopyToOutputDirectory xmlns=\"\">", "<CopyToOutputDirectory>"); 26: xml.LoadXml(content); 27: xml.Save(proj); 28: } 29:  30: static void UpdateNodes(XmlNodeList nodes, XmlDocument xml, XmlNamespaceManager nsManager) 31: { 32: foreach (XmlNode node in nodes) 33: { 34: var copyToOutputDirectoryNode = node.SelectSingleNode("pf:CopyToOutputDirectory", nsManager); 35: if (copyToOutputDirectoryNode == null) 36: { 37: var n = xml.CreateNode(XmlNodeType.Element, "CopyToOutputDirectory", null); 38: n.InnerText = "Always"; 39: node.AppendChild(n); 40: } 41: else 42: { 43: if (string.Compare(copyToOutputDirectoryNode.InnerText, "Always", true) != 0) 44: { 45: copyToOutputDirectoryNode.InnerText = "Always"; 46: } 47: } 48: } 49: } Please be careful when use this tool. I created only for demo so do not use it directly in a production environment. Unload the worker role project, execute this tool with the worker role project file name as the command line argument, it will set all items as “Copy always”. Then reload this worker role project. Now let’s change the “index.js” to use express. 1: var express = require("express"); 2: var app = express(); 3:  4: var port = 80; 5:  6: app.configure(function () { 7: }); 8:  9: app.get("/", function (req, res) { 10: res.send("Hello Node.js!"); 11: }); 12:  13: app.get("/User/:id", function (req, res) { 14: var id = req.params.id; 15: res.json({ 16: "id": id, 17: "name": "user " + id, 18: "company": "IGT" 19: }); 20: }); 21:  22: app.listen(port); Finally let’s publish it and have a look in browser.   Use Windows Azure SQL Database We can use Windows Azure SQL Database (a.k.a. WACD) from Node.js as well on worker role hosting. Since we can control the version of Node.js, here we can use x64 version of “node-sqlserver” now. This is better than if we host Node.js on WAWS since it only support x86. Just install the “node-sqlserver” module from NPM, copy the “sqlserver.node” from “Build\Release” folder to “Lib” folder. Include them in worker role project and run my tool to make them to “Copy always”. Finally update the “index.js” to use WASD. 1: var express = require("express"); 2: var sql = require("node-sqlserver"); 3:  4: var connectionString = "Driver={SQL Server Native Client 10.0};Server=tcp:{SERVER NAME}.database.windows.net,1433;Database={DATABASE NAME};Uid={LOGIN}@{SERVER NAME};Pwd={PASSWORD};Encrypt=yes;Connection Timeout=30;"; 5: var port = 80; 6:  7: var app = express(); 8:  9: app.configure(function () { 10: app.use(express.bodyParser()); 11: }); 12:  13: app.get("/", function (req, res) { 14: sql.open(connectionString, function (err, conn) { 15: if (err) { 16: console.log(err); 17: res.send(500, "Cannot open connection."); 18: } 19: else { 20: conn.queryRaw("SELECT * FROM [Resource]", function (err, results) { 21: if (err) { 22: console.log(err); 23: res.send(500, "Cannot retrieve records."); 24: } 25: else { 26: res.json(results); 27: } 28: }); 29: } 30: }); 31: }); 32:  33: app.get("/text/:key/:culture", function (req, res) { 34: sql.open(connectionString, function (err, conn) { 35: if (err) { 36: console.log(err); 37: res.send(500, "Cannot open connection."); 38: } 39: else { 40: var key = req.params.key; 41: var culture = req.params.culture; 42: var command = "SELECT * FROM [Resource] WHERE [Key] = '" + key + "' AND [Culture] = '" + culture + "'"; 43: conn.queryRaw(command, function (err, results) { 44: if (err) { 45: console.log(err); 46: res.send(500, "Cannot retrieve records."); 47: } 48: else { 49: res.json(results); 50: } 51: }); 52: } 53: }); 54: }); 55:  56: app.get("/sproc/:key/:culture", function (req, res) { 57: sql.open(connectionString, function (err, conn) { 58: if (err) { 59: console.log(err); 60: res.send(500, "Cannot open connection."); 61: } 62: else { 63: var key = req.params.key; 64: var culture = req.params.culture; 65: var command = "EXEC GetItem '" + key + "', '" + culture + "'"; 66: conn.queryRaw(command, function (err, results) { 67: if (err) { 68: console.log(err); 69: res.send(500, "Cannot retrieve records."); 70: } 71: else { 72: res.json(results); 73: } 74: }); 75: } 76: }); 77: }); 78:  79: app.post("/new", function (req, res) { 80: var key = req.body.key; 81: var culture = req.body.culture; 82: var val = req.body.val; 83:  84: sql.open(connectionString, function (err, conn) { 85: if (err) { 86: console.log(err); 87: res.send(500, "Cannot open connection."); 88: } 89: else { 90: var command = "INSERT INTO [Resource] VALUES ('" + key + "', '" + culture + "', N'" + val + "')"; 91: conn.queryRaw(command, function (err, results) { 92: if (err) { 93: console.log(err); 94: res.send(500, "Cannot retrieve records."); 95: } 96: else { 97: res.send(200, "Inserted Successful"); 98: } 99: }); 100: } 101: }); 102: }); 103:  104: app.listen(port); Publish to azure and now we can see our Node.js is working with WASD through x64 version “node-sqlserver”.   Summary In this post I demonstrated how to host our Node.js in Windows Azure Cloud Service worker role. By using worker role we can control the version of Node.js, as well as the entry code. And it’s possible to do some pre jobs before the Node.js application started. It also removed the IIS and IISNode limitation. I personally recommended to use worker role as our Node.js hosting. But there are some problem if you use the approach I mentioned here. The first one is, we need to set all JavaScript files and module files as “Copy always” or “Copy if newer” manually. The second one is, in this way we cannot retrieve the cloud service configuration information. For example, we defined the endpoint in worker role property but we also specified the listening port in Node.js hardcoded. It should be changed that our Node.js can retrieve the endpoint. But I can tell you it won’t be working here. In the next post I will describe another way to execute the “node.exe” and Node.js application, so that we can get the cloud service configuration in Node.js. I will also demonstrate how to use Windows Azure Storage from Node.js by using the Windows Azure Node.js SDK.   Hope this helps, Shaun All documents and related graphics, codes are provided "AS IS" without warranty of any kind. Copyright © Shaun Ziyan Xu. This work is licensed under the Creative Commons License.

    Read the article

  • Metro, Authentication, and the ASP.NET Web API

    - by Stephen.Walther
    Imagine that you want to create a Metro style app written with JavaScript and you want to communicate with a remote web service. For example, you are creating a movie app which retrieves a list of movies from a movies service. In this situation, how do you authenticate your Metro app and the Metro user so not just anyone can call the movies service? How can you identify the user making the request so you can return user specific data from the service? The Windows Live SDK supports a feature named Single Sign-On. When a user logs into a Windows 8 machine using their Live ID, you can authenticate the user’s identity automatically. Even better, when the Metro app performs a call to a remote web service, you can pass an authentication token to the remote service and prevent unauthorized access to the service. The documentation for Single Sign-On is located here: http://msdn.microsoft.com/en-us/library/live/hh826544.aspx In this blog entry, I describe the steps that you need to follow to use Single Sign-On with a (very) simple movie app. We build a Metro app which communicates with a web service created using the ASP.NET Web API. Creating the Visual Studio Solution Let’s start by creating a Visual Studio solution which contains two projects: a Windows Metro style Blank App project and an ASP.NET MVC 4 Web Application project. Name the Metro app MovieApp and the ASP.NET MVC application MovieApp.Services. When you create the ASP.NET MVC application, select the Web API template: After you create the two projects, your Visual Studio Solution Explorer window should look like this: Configuring the Live SDK You need to get your hands on the Live SDK and register your Metro app. You can download the latest version of the SDK (version 5.2) from the following address: http://www.microsoft.com/en-us/download/details.aspx?id=29938 After you download the Live SDK, you need to visit the following website to register your Metro app: https://manage.dev.live.com/build Don’t let the title of the website — Windows Push Notifications & Live Connect – confuse you, this is the right place. Follow the instructions at the website to register your Metro app. Don’t forget to follow the instructions in Step 3 for updating the information in your Metro app’s manifest. After you register, your client secret is displayed. Record this client secret because you will need it later (we use it with the web service): You need to configure one more thing. You must enter your Redirect Domain by visiting the following website: https://manage.dev.live.com/Applications/Index Click on your application name, click Edit Settings, click the API Settings tab, and enter a value for the Redirect Domain field. You can enter any domain that you please just as long as the domain has not already been taken: For the Redirect Domain, I entered http://superexpertmovieapp.com. Create the Metro MovieApp Next, we need to create the MovieApp. The MovieApp will: 1. Use Single Sign-On to log the current user into Live 2. Call the MoviesService web service 3. Display the results in a ListView control Because we use the Live SDK in the MovieApp, we need to add a reference to it. Right-click your References folder in the Solution Explorer window and add the reference: Here’s the HTML page for the Metro App: <!DOCTYPE html> <html> <head> <meta charset="utf-8" /> <title>MovieApp</title> <!-- WinJS references --> <link href="//Microsoft.WinJS.1.0.RC/css/ui-dark.css" rel="stylesheet" /> <script src="//Microsoft.WinJS.1.0.RC/js/base.js"></script> <script src="//Microsoft.WinJS.1.0.RC/js/ui.js"></script> <!-- Live SDK --> <script type="text/javascript" src="/LiveSDKHTML/js/wl.js"></script> <!-- WebServices references --> <link href="/css/default.css" rel="stylesheet" /> <script src="/js/default.js"></script> </head> <body> <div id="tmplMovie" data-win-control="WinJS.Binding.Template"> <div class="movieItem"> <span data-win-bind="innerText:title"></span> <br /><span data-win-bind="innerText:director"></span> </div> </div> <div id="lvMovies" data-win-control="WinJS.UI.ListView" data-win-options="{ itemTemplate: select('#tmplMovie') }"> </div> </body> </html> The HTML page above contains a Template and ListView control. These controls are used to display the movies when the movies are returned from the movies service. Notice that the page includes a reference to the Live script that we registered earlier: <!-- Live SDK --> <script type="text/javascript" src="/LiveSDKHTML/js/wl.js"></script> The JavaScript code looks like this: (function () { "use strict"; var REDIRECT_DOMAIN = "http://superexpertmovieapp.com"; var WEBSERVICE_URL = "http://localhost:49743/api/movies"; function init() { WinJS.UI.processAll().done(function () { // Get element and control references var lvMovies = document.getElementById("lvMovies").winControl; // Login to Windows Live var scopes = ["wl.signin"]; WL.init({ scope: scopes, redirect_uri: REDIRECT_DOMAIN }); WL.login().then( function(response) { // Get the authentication token var authenticationToken = response.session.authentication_token; // Call the web service var options = { url: WEBSERVICE_URL, headers: { authenticationToken: authenticationToken } }; WinJS.xhr(options).done( function (xhr) { var movies = JSON.parse(xhr.response); var listMovies = new WinJS.Binding.List(movies); lvMovies.itemDataSource = listMovies.dataSource; }, function (xhr) { console.log(xhr.statusText); } ); }, function(response) { throw WinJS.ErrorFromName("Failed to login!"); } ); }); } document.addEventListener("DOMContentLoaded", init); })(); There are two constants which you need to set to get the code above to work: REDIRECT_DOMAIN and WEBSERVICE_URL. The REDIRECT_DOMAIN is the domain that you entered when registering your app with Live. The WEBSERVICE_URL is the path to your web service. You can get the correct value for WEBSERVICE_URL by opening the Project Properties for the MovieApp.Services project, clicking the Web tab, and getting the correct URL. The port number is randomly generated. In my code, I used the URL  “http://localhost:49743/api/movies”. Assuming that the user is logged into Windows 8 with a Live account, when the user runs the MovieApp, the user is logged into Live automatically. The user is logged in with the following code: // Login to Windows Live var scopes = ["wl.signin"]; WL.init({ scope: scopes, redirect_uri: REDIRECT_DOMAIN }); WL.login().then(function(response) { // Do something }); The scopes setting determines what the user has permission to do. For example, access the user’s SkyDrive or access the user’s calendar or contacts. The available scopes are listed here: http://msdn.microsoft.com/en-us/library/live/hh243646.aspx In our case, we only need the wl.signin scope which enables Single Sign-On. After the user signs in, you can retrieve the user’s Live authentication token. The authentication token is passed to the movies service to authenticate the user. Creating the Movies Service The Movies Service is implemented as an API controller in an ASP.NET MVC 4 Web API project. Here’s what the MoviesController looks like: using System.Collections.Generic; using System.Linq; using System.Net; using System.Net.Http; using System.Web.Http; using JWTSample; using MovieApp.Services.Models; namespace MovieApp.Services.Controllers { public class MoviesController : ApiController { const string CLIENT_SECRET = "NtxjF2wu7JeY1unvVN-lb0hoeWOMUFoR"; // GET api/values public HttpResponseMessage Get() { // Authenticate // Get authenticationToken var authenticationToken = Request.Headers.GetValues("authenticationToken").FirstOrDefault(); if (authenticationToken == null) { return new HttpResponseMessage(HttpStatusCode.Unauthorized); } // Validate token var d = new Dictionary<int, string>(); d.Add(0, CLIENT_SECRET); try { var myJWT = new JsonWebToken(authenticationToken, d); } catch { return new HttpResponseMessage(HttpStatusCode.Unauthorized); } // Return results return Request.CreateResponse( HttpStatusCode.OK, new List<Movie> { new Movie {Title="Star Wars", Director="Lucas"}, new Movie {Title="King Kong", Director="Jackson"}, new Movie {Title="Memento", Director="Nolan"} } ); } } } Because the Metro app performs an HTTP GET request, the MovieController Get() action is invoked. This action returns a set of three movies when, and only when, the authentication token is validated. The Movie class looks like this: using Newtonsoft.Json; namespace MovieApp.Services.Models { public class Movie { [JsonProperty(PropertyName="title")] public string Title { get; set; } [JsonProperty(PropertyName="director")] public string Director { get; set; } } } Notice that the Movie class uses the JsonProperty attribute to change Title to title and Director to director to make JavaScript developers happy. The Get() method validates the authentication token before returning the movies to the Metro app. To get authentication to work, you need to provide the client secret which you created at the Live management site. If you forgot to write down the secret, you can get it again here: https://manage.dev.live.com/Applications/Index The client secret is assigned to a constant at the top of the MoviesController class. The MoviesController class uses a helper class named JsonWebToken to validate the authentication token. This class was created by the Windows Live team. You can get the source code for the JsonWebToken class from the following GitHub repository: https://github.com/liveservices/LiveSDK/blob/master/Samples/Asp.net/AuthenticationTokenSample/JsonWebToken.cs You need to add an additional reference to your MVC project to use the JsonWebToken class: System.Runtime.Serialization. You can use the JsonWebToken class to get a unique and validated user ID like this: var user = myJWT.Claims.UserId; If you need to store user specific information then you can use the UserId property to uniquely identify the user making the web service call. Running the MovieApp When you first run the Metro MovieApp, you get a screen which asks whether the app should have permission to use Single Sign-On. This screen never appears again after you give permission once. Actually, when I first ran the app, I get the following error: According to the error, the app is blocked because “We detected some suspicious activity with your Online Id account. To help protect you, we’ve temporarily blocked your account.” This appears to be a bug in the current preview release of the Live SDK and there is more information about this bug here: http://social.msdn.microsoft.com/Forums/en-US/messengerconnect/thread/866c495f-2127-429d-ab07-842ef84f16ae/ If you click continue, and continue running the app, the error message does not appear again.  Summary The goal of this blog entry was to describe how you can validate Metro apps and Metro users when performing a call to a remote web service. First, I explained how you can create a Metro app which takes advantage of Single Sign-On to authenticate the current user against Live automatically. You learned how to register your Metro app with Live and how to include an authentication token in an Ajax call. Next, I explained how you can validate the authentication token – retrieved from the request header – in a web service. I discussed how you can use the JsonWebToken class to validate the authentication token and retrieve the unique user ID.

    Read the article

  • On automating a split-mirror ASM backup with EMC TimeFinder ...

    - by [email protected]
    Normal 0 21 false false false MicrosoftInternetExplorer4 st1\:*{behavior:url(#ieooui) } /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-parent:""; mso-padding-alt:0cm 5.4pt 0cm 5.4pt; mso-para-margin:0cm; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:10.0pt; font-family:"Times New Roman"; mso-ansi-language:#0400; mso-fareast-language:#0400; mso-bidi-language:#0400;} Hi clerks,   Offloading the backup operation to another host using disk cloning could really improve the performance on highly busy databases ( 24x7, zero downtime and all this stuff ...) There are well know white papers on this subject, ASM included, but today Im showing you a nice way to automate the procedure using shell scripting with EMC TimeFinder technologies:   Assumptions: *********** ASM diskgroups name:   +data_${db_name} : asm data diskgroup +fra_${db_name} :  asm fra  diskgroup   EMC Time Finder sync groups name:   rac_${DB_NAME}_data_tf : data group rac_${DB_NAME}_fra_tf:   fra group     There are two scripts, one located on the production box ( bck_database.sh ) and the other one on the backup server node ( bck_database_mirror.sh ) The second one is remotly executed from the production host There are a bunch of variables along the code with selfexplanatory names I guess, anyway let me know if you want some help     #!/bin/ksh ### ###  Copyright (c) 1988, 2010, Oracle Corporation.  All Rights Reserved. ### ###    NAME ###     bck_database.sh ### ###    DESCRIPTION ###     Database backup on third mirror ### ###    RETURNS ### ###    NOTES ### ###    MODIFIED                                 (DD/MM/YY) ###    Oracle            28/01/10             - Creacion ###   V_DATE=`/bin/date +%Y%m%d_%H%M%S` V_FICH_LOG=`dirname $0`/trace_dir_location/`basename $0`.${V_DATE}.log exec 4>&1 tee ${V_FICH_LOG} >&4 |& exec 1>&p 2>&1     ADMIN_DIR=`dirname $0` . ${ADMIN_DIR}/setenv_instance.sh -- This script should set the instance vars like Oracle Home, Sid, db_name ... if [ $? -ne 0 ] then   echo "Error when setting the environment."   exit 1 fi   echo "${V_DATE} ####################################################" echo "Executing database backup: ${DB_NAME}" echo "####################################################################"   V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Sync asm data diskgroups ..." echo "####################################################################" sudo symmir -g rac_${DB_NAME}_data_tf establish -noprompt if [ $? -ne 0 ] then   echo "Error when sync asm data diskgroups"   exit 2 fi V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Verifying asm data disks ..." echo "####################################################################" sudo symmir -g rac_${DB_NAME}_data_tf -i 30 verify if [ $? -ne 0 ] then   echo "Error when verifying asm data diskgroups"   exit 3 fi     V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Sync asm fra diskgroups ..." echo "####################################################################" sudo symmir -g rac_${DB_NAME}_fra_tf establish -noprompt if [ $? -ne 0 ] then   echo "Error when sync asm fra diskgroups"   exit 4 fi V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Verifying asm fra disks ..." echo "####################################################################" sudo symmir -g rac_${DB_NAME}_fra_tf -i 30 verify if [ $? -ne 0 ] then   echo "Error when verifying asm fra diskgroups"   exit 5 fi   V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "ASM sync sucessfully completed!" echo "####################################################################"     V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Updating status ${DB_NAME} to BEGIN BACKUP ..." echo "####################################################################" sqlplus -s /nolog <<-!   whenever sqlerror exit 1   connect / as sysdba   whenever sqlerror exit   alter system archive log current;   alter database ${DB_NAME} begin backup; ! if [ $? -ne 0 ] then   echo "Error when updating database status to BEGIN backup"   exit 6 fi   V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Splitting asm data disks....." echo "####################################################################" sudo symmir -g rac_${DB_NAME}_data_tf split -noprompt if [ $? -ne 0 ] then   echo "Error when splitting asm data disks"   exit 7 fi   V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Updating status ${DB_NAME} to END BACKUP ..." echo "####################################################################" sqlplus -s /nolog <<-!   whenever sqlerror exit 1   connect / as sysdba   whenever sqlerror exit   alter database ${DB_NAME} end backup;   alter system archive log current; ! if [ $? -ne 0 ] then   echo "Error when updating database status to END backup"   exit 8 fi   V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Generating controlfile copies...." echo "####################################################################" rman<<-! connect target / run { allocate channel ch1 type DISK; copy current controlfile to '+FRA_${DB_NAME}/${DB_NAME}/CONTROLFILE/control_mount.ctl'; copy current controlfile to '+FRA_${DB_NAME}/${DB_NAME}/CONTROLFILE/control_backup.ctl'; } ! if [ $? -ne 0 ] then   echo "Error generating controlfile copies"   exit 9 fi V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Resync RMAN catalog ....." echo "####################################################################" rman<<-! connect target / connect catalog ${V_RMAN_USR}/${V_RMAN_PWD}@${V_DB_CATALOG} resync catalog; ! if [ $? -ne 0 ] then   echo "Error when resyncing RMAN catalog"   exit 10 fi   V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Splitting asm fra disks....." echo "####################################################################" sudo symmir -g rac_${DB_NAME}_fra_tf split -noprompt if [ $? -ne 0 ] then   echo "Error when splitting asm fra disks"   exit 11 fi     echo "WARNING!: Calling bck_database_mirror.sh host ${NODE_BCK_SERVER}..." ssh ${NODO_BCK_SERVER} ${ADMIN_DIR_BCK}/bck_database_mirror.sh if [ $? -ne 0 ] then   echo "Error, when remote executing the backup "   exit 12 fi V_DATE=`/bin/date +%Y%m%d_%H%M%S` echo "${V_DATE} ####################################################" echo "Cleaning the archived redo logs already copied to tape ..." echo "####################################################################" rman<<-! connect target / connect catalog ${V_RMAN_USR}/${V_RMAN_PWD}@${V_DB_CATALOG} run { resync catalog; delete noprompt archivelog all backed up 1 times to device type sbt; } ! if [ $? -ne 0 ] then   echo "Error when cleaning the archived redo logs"   exit 13 fi echo "${V_DATE} ####################################################" echo "Backup sucessfully executed!!" echo "####################################################################" exit 0   ------------------------------------------------------------------------------ ------------------------** BACKUP SERVER NODE ** ----------------------------- ------------------------------------------------------------------------------   #!/bin/ksh ### ###  Copyright (c) 1988, 2010, Oracle Corporation.  All Rights Reserved. ### ###    ###    NAME ###     bck_database_mirror.sh ### ###    DESCRIPTION ###      Backup @ backup server ### ###    RETURNS ### ###    NOTES ### ###    MODIFIED                                 (DD/MM/YY) ###      Oracle                    28/01/10     - Creacion         V_DATE=`/bin/date +%Y%m%d_%H%M%S`   echo "${V_DATE} ####################################################"   echo "Starting ASM instance ..."   echo "####################################################################"   ${V_ADMIN_DIR}/start_asm.sh -- This script is supposed to start the ASM instance in the backup server   if [ $? -ne 0 ]   then     echo "Error when tying to start ASM instance."     exit 1   fi       . ${V_ADMIN_DIR}/setenv_asm.sh -- This script is supposed to set the env. variables of the ASM instance   if [ $? -ne 0 ]   then     echo "Error when setting the ASM environment"     exit 1   fi       V_DATE=`/bin/date +%Y%m%d_%H%M%S`   echo "${V_DATE} ####################################################"   echo "The asm diskgroups/disks dettected are the following ..."   echo "####################################################################"     sqlplus /nolog <<-!     whenever sqlerror exit 1     connect / as sysdba     whenever sqlerror exit     SET LINES 200     COL PATH FORMAT A25     SELECT DISK.MOUNT_STATUS, DISK.PATH, DISK.NAME, DISK_GROUP.NAME, DISK_GROUP.TOTAL_MB FROM V\$ASM_DISK DISK, V\$ASM_DISKGROUP DISK_GROUP WHERE DISK.GROUP_NUMBER=DISK_GROUP.GROUP_NUMBER; !       V_ADMIN_DIR=`dirname $0`   . ${V_ADMIN_DIR}/setenv_instance.sh -- This script is supposed to set the env. variables of the database instance   if [ $? -ne 0 ]   then     echo "Error when setting the database instance environment"     exit 1   fi     V_DATE=`/bin/date +%Y%m%d_%H%M%S`   echo "${V_DATE} ####################################################"   echo "Starting ${DB_NAME} in MOUNT mode..."   echo "####################################################################"   ${V_ADMIN_DIR}/start_instance_mount.sh -- This script is supposed to do a startup mount   if [ $? -ne 0 ]   then     echo "Error starting  ${DB_NAME} in MOUNT mode"     exit 1   fi   V_DATE=`/bin/date +%Y%m%d_%H%M%S`   echo "${V_DATE} ####################################################"   echo "Executing RMAN backup..."   echo "####################################################################"   rman<<-!   connect target /   connect catalog ${V_RMAN_USR}/${V_RMAN_PWD}@${V_DB_CATALOG}   run {   allocate channel ch1 type 'SBT_TAPE' parms'ENV=(TDPO_OPTFILE=/opt/tivoli/tsm/client/oracle/bin64/tdpo.opt)'; -- TDPO Media Library   crosscheck archivelog all;   backup tag BCK_CONTROLFILE_ST_${DB_NAME}   format 'ctl_%d_%s__%p_%t'   controlfilecopy '+FRA_${DB_NAME}/${DB_NAME}/CONTROLFILE/control_backup.ctl';   backup tag BCK_DATAFILE_ST_${DB_NAME} full   format 'db_%d_%s_%p_%t'database;   backup tag BCK_ARCHLOG_ST_${DB_NAME} format 'al_%d_%s_%p_%t' archivelog all;   release channel ch1;   } !   if [ $? -ne 0 ]   then     echo "Error executing the RMAN backup"     exit 1   fi     ${V_ADMIN_DIR}/stop_instance_immediate.sh -- This script is supposed to do a shutdown immediate of the database instance   ${ADMIN_DIR}/stop_asm_immediate.sh -- This script is supposed to do a shutdown immediate of the ASM instance   exit 0     fi   Hope it helps someone! --L

    Read the article

  • Install Control Center Agent on Oracle Application Server

    - by qianqian.wu
    Control Center Agent (CCA) The Control Center Agent is the OWB component that runs the Template Mappings in the Oracle Containers for J2EE (OC4J) server; also referred to as the J2EE Runtime. The Control Center Agent provides a Java-based runtime environment that can be installed on Oracle and non-Oracle database hosts. The Control Center Agent provides fundamental infrastructure for the heterogeneous, Code Template-based mapping support and Web services-related features of OWB in this release. In Oracle Warehouse Builder 11gR2 the Control Center Agent, by default will run in the built-in OC4J that is bundled in the Oracle Home. Besides that, you also have ability to install the Control Center Agent in an Oracle Application Server install. In this article, you will find step-by-step instructions how to install the Control Center Agent on an Oracle Application Server instance. The instructions cover the following tasks: Task 1: Install and Configure the Application Server Task 2: Deploy the Control Center Agent to the Application Server Task 3: Optional Configuration Tasks   Task 1: Install and Configure the Application Server Before configuring the Application Server, you need to install it from Oracle Application Server CD-ROM, or by downloading the installation program from Oracle Technology Network (OTN). Once the installation is completed, you are ready to configure the Application Server. The purpose of the configuration task is to make sure the Control Center Agent ear file can be deployed and runs in the Application Server successfully. The essential configuration tasks are outlined below: · Modify the OC4J Startup Script · Set up Control Center Agent Server Side Logging · Set up Audit Table Data Source · Copy ct_permissions.properties File · Set up Security Roles for Control Center Agent · Create JMS Queues · Install JDBC Drivers to OC4J Modify the OC4J Startup Script The OC4J startup script “opmn.xml” is located in Application Server configuration directory, $AS_HOME/opmn/conf. $AS_HOME stands for the root home directory of the application server. Open the file opmn.xml in a text editor, and alter the contents of the file as displayed in the following sample. You need to make sure that: The MaxPerSize is set to 128M. This is to ensure that you allocate enough PermGen space to OC4J to run Control Center Agent. This will prevent java.lang.OutOfMemoryError when running the agent. The Python.path sets the path for the Python library files used by the Control Center Agent: jython_lib.zip and jython_owblib.jar. These two files are in the $OWB_HOME/owb/lib/int directory, where $OWB_HOME is the directory where owb is installed. · The km_security_needed determines whether restrictions will be applied to the kinds of operating system commands allowed to be executed by the OWB Code Template script executed by Control Center Agent. Setting km_security_needed to “true” enforces such restriction while setting it to “false” removes such restrictions. Set up Control Center Agent Server Side Logging Ensure that you are in the Application Server configuration directory, $AS_HOME/j2ee/home/config. Open the file j2ee-logging.xml in a text editor and add the following lines to the log handler section. The jrt-internal-log-handler is the handler used by Control Center Agent runtime logger to create log files. Then add the following entry into the loggers section to create the logger for Control Center Agent runtime auditing. Set up Audit Table Data Source To enable Audit Table logging, a managed data source and connection pool need to be set up before Control Center Agent deployment. Ensure that you are in the Application Server configuration directory, $AS_HOME/j2ee/home/config. Open the file data-sources.xml in a text editor. Define the audit data source shown below in the file, <managed-data-source name="AuditDS" connection-pool-name="OWBSYS Audit   Connection Pool" jndi-name="jdbc/AuditDS"/> <connection-pool name="OWBSYS Audit Connection Pool">   <connection-factory factory-class="oracle.jdbc.pool.OracleDataSource"     user="owbsys_audit" password="owbsys_audit"     url="jdbc:oracle:thin:@//localhost:1521/ORCL"/> </connection-pool> Copy ct_permissions.properties File The ct_permissions.properties can be obtained from $OWB_HOME /owb/jrt/config/ directory. You need to copy the file to $AS_HOME/j2ee/home/config directory.This properties file takes effect when the setting km-security is set to true in Control Center Agent. By default the ALLOWED_CMD is commented out in ct_permissions.properties file. This prevents all system command from being invoked from scripts executed in Control Center Agent (when km-security is set to true). To allow certain system commands to be invoked, ALLOWED_CMD needs to be uncommented out, and the system commands (allowed to be invoked) need to be added to the ALLOWED_CMD. Set up Security Roles for Control Center Agent You can set up the Control Center Agent security roles through Oracle Enterprise Manager. In a web browser, navigate to Enterprise Manager Homepage (e.g. http://hostname:8889/em). 1. Log in using the oc4jadmin credentials. After the Cluster Topology page is loaded, click home (the OC4J instance). This takes you to the home page of the OC4J instance. On the OC4J home screen, click the Administration tab. On the Administration Tasks screen, expand Security. Click the task icon next to Security Providers. 2. On Security Providers page click on the button “Instance Level Security”. On Instance Level Security page, go to “Realms” tab. You will see a row for the default realm “jazn.com” in the results table. It has a “Roles” column and a “Users” column. Click on the number in “Roles” column. In the “Roles” page it will display all the roles available for the realm. Click on “Create” button to create a new role “OWB_J2EE_ EXECUTOR”. 3. On the Add Role screen, enter Name OWB_J2EE_EXECUTOR, and click OK. 4. Follow the same steps as before, and create a new role “OWB_J2EE_OPERATOR”. 5. Assign role “oc4j-administrators” and “OWB_J2EE_EXECUTOR” to the role “OWB_J2EE_OPERATOR” by moving these roles from “Available Roles” and click “OK” to save. 6. Go back to Instance Level Security page and create a new role “OWB_J2EE_ADMINISTRATOR”. 7. Assign roles “OWB_J2EE_ OPERATOR” and “OWB_J2EE_EXECUTOR” to the role “OWB_J2EE_ ADMINISTRATOR” by moving these roles from “Available Roles” and click “OK” to save. 8.Go back to Instance Level Security page. This time, click on the number in “Users” column for the realm “jazn.com”. In the “Users” page, it shows all the users defined for this realm. Locate the user “oc4jadmin” in the results table and click on it. 9. Assign the roles “OWB_J2EE_ADMINISTRATOR” and “oc4j-app-administrators” to this user by moving the role from the “Available Roles” selection box to “Selected Roles” box and click “Apply” to save. 10. Go back to Instance Level Security page and create a new role “OWB_INTERNAL_USERS”, assign no user or role to this role. Simply click “OK” to create this role. Now you have finished creating the security roles required for Control Center Agent. Create JMS Queues You need to create two JMS queues for Control Center Agent: owbQueue and abort_owbQueue. 1. Now go to OC4J home Page. On the OC4J home screen, click the Administration tab. On the Administration Tasks screen, expand Services and then expand Enterprise Messaging Service. Click the task icon next to JMS Destinations. 2. On JMS Destinations page, click “Create New” button to create a new JMS queue. On Add Destination page, choose “Queue” as Destination Type. Put “owbQueue” as Destination Name. Select “In Memory Persistence Only” as the Persistence Type and put “jms/owbQueue” as JNDI Location and click on “OK” to finish. 3. Follow the same instruction as above to create the owb_abortQueue. Now you have finished creating the JMS queues required for Control Center Agent. Install JDBC Drivers to OC4J In order to execute Code Templates using commercial databases other than Oracle, e.g. DB2, SQL Server etc, the corresponding jdbc driver files need to be added to $AS_HOME/j2ee/home/applib directory. 1. To install other JDBC drivers to OC4J, first obtain the .jar file containing the JDBC driver. All the external JDBC drivers .jar files can be found in the directory: $OWB_HOME/owb/lib/ext/. For DB2, the files needed are db2jcc.jar and db2jcc_license_cu.jar. For SQL Server the file is sqljdbc.jar. For sunopsis JDBC drivers, the file needed is snpsxmlo.jar. 2. Copy the required JDBC driver file into the directory $AS_HOME/j2ee/home/applib. Now you have finished the Application Server configuration. To make the configuration to take an effect, you need to restart the Application Server.   Task 2: Deploy the Control Center Agent to the Application Server Now you can deploy the Control Center Agent to the Application Server. In a web browser, navigate to Enterprise Manager Homepage (e.g. http://hostname:8889/em). 1. Log in using the oc4jadmin credentials. After the Cluster Topology page is loaded, click home (the OC4J instance). This takes you to the home page of the OC4J instance. On the OC4J home screen, click the Applications tab. Click Deploy to begin deploying Control Center Agent. 2. On the Deploy: Select Archive screen, under Archive, select Archive is present on local host. Upload the archive to the server where Application Server Control is running. Click Browse and locate the jrt.ear file in the $OWB_HOME/owb/jrt/applications directory. Under Deployment Plan, select Automatically create a new deployment plan. Click Next. 3. Wait for the ear file to be uploaded to Application Server. On the Deploy: Application Attributes screen, enter Application Name jrt, and Context Root jrt. Leave the other attributes at their default values. Click Next. 4. On Deploy: Deployment Settings screen, leave all attributes at their default values, and click Deploy. This will take about 1 minute or so and when the application is deployed successfully, a confirmation message will be displayed. Now the Control Center Agent is started automatically. Go back to OC4J home page and click on Applications tab to make sure the deployed application jrt is showing in the applications list.   Task 3: Optional Configuration Tasks The optional configuration tasks contain: · Secure Control Center Agent Web Service · Setting the PATH Environment Variable Secure Control Center Agent Web Service If you want to use JRTWebService with a secure website, you need to do the following steps, 1. Create a file “secure-web-site.xml” in the $AS_HOME/j2ee/home/config directory. The file can be obtained from $OWB_HOME/owb/jrt/config directory. A sample secure-web-site.xml is shown as below. We need to modify the “protocol” to “https”, and “secure” to “true”, also choose an port as the secure http port. Also we need to add the entry “ssl-config” in the file. Remember to use the absolute path for the key store file. 2. Modify the file “server.xml” that is located at $AS_HOME/j2ee/home/config directory. Then add the <web-site> element in the file for the secure-web-site. 3. Create a key store file “serverkeystore.jks” in the $AS_HOME/j2ee/home/config directory. The file can be obtained from $OWB_HOME/owb/jrt/config directory. After the three files are altered, restart the application server. Now you can access the JRTWebService in SSL way through https://hostname:4443/jrt/webservice. Setting the PATH Environment Variable Sometimes, some system commands such as linux ls, sh etc, can not be executed successfully during the script execution due to they are not found in PATH. To ensure they work normally, you can setup the environment variable PATH. Let’s navigate to the Enterprise Manager Homepage. 1. Go to OC4J home screen and click the Administration tab. Expand Administration Tasks, then expand Properties. Click the task icon next to Server Properties. 2. On the Server Properties screen, scroll down to Environment Variables section. Under Environment Variables, click Add Another Row. Enter PATH in Name, and fill Value with directories that contain the system commands. Click Apply.   After you work through this article, I believe you have developed a deeper understanding of the Control Center Agent installation process, and you can apply this knowledge in other installation plan such as Control Center Agent installation on Standalone OC4J.

    Read the article

  • Introducing Data Annotations Extensions

    - by srkirkland
    Validation of user input is integral to building a modern web application, and ASP.NET MVC offers us a way to enforce business rules on both the client and server using Model Validation.  The recent release of ASP.NET MVC 3 has improved these offerings on the client side by introducing an unobtrusive validation library built on top of jquery.validation.  Out of the box MVC comes with support for Data Annotations (that is, System.ComponentModel.DataAnnotations) and can be extended to support other frameworks.  Data Annotations Validation is becoming more popular and is being baked in to many other Microsoft offerings, including Entity Framework, though with MVC it only contains four validators: Range, Required, StringLength and Regular Expression.  The Data Annotations Extensions project attempts to augment these validators with additional attributes while maintaining the clean integration Data Annotations provides. A Quick Word About Data Annotations Extensions The Data Annotations Extensions project can be found at http://dataannotationsextensions.org/, and currently provides 11 additional validation attributes (ex: Email, EqualTo, Min/Max) on top of Data Annotations’ original 4.  You can find a current list of the validation attributes on the afore mentioned website. The core library provides server-side validation attributes that can be used in any .NET 4.0 project (no MVC dependency). There is also an easily pluggable client-side validation library which can be used in ASP.NET MVC 3 projects using unobtrusive jquery validation (only MVC3 included javascript files are required). On to the Preview Let’s say you had the following “Customer” domain model (or view model, depending on your project structure) in an MVC 3 project: public class Customer { public string Email { get; set; } public int Age { get; set; } public string ProfilePictureLocation { get; set; } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } When it comes time to create/edit this Customer, you will probably have a CustomerController and a simple form that just uses one of the Html.EditorFor() methods that the ASP.NET MVC tooling generates for you (or you can write yourself).  It should look something like this: With no validation, the customer can enter nonsense for an email address, and then can even report their age as a negative number!  With the built-in Data Annotations validation, I could do a bit better by adding a Range to the age, adding a RegularExpression for email (yuck!), and adding some required attributes.  However, I’d still be able to report my age as 10.75 years old, and my profile picture could still be any string.  Let’s use Data Annotations along with this project, Data Annotations Extensions, and see what we can get: public class Customer { [Email] [Required] public string Email { get; set; }   [Integer] [Min(1, ErrorMessage="Unless you are benjamin button you are lying.")] [Required] public int Age { get; set; }   [FileExtensions("png|jpg|jpeg|gif")] public string ProfilePictureLocation { get; set; } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Now let’s try to put in some invalid values and see what happens: That is very nice validation, all done on the client side (will also be validated on the server).  Also, the Customer class validation attributes are very easy to read and understand. Another bonus: Since Data Annotations Extensions can integrate with MVC 3’s unobtrusive validation, no additional scripts are required! Now that we’ve seen our target, let’s take a look at how to get there within a new MVC 3 project. Adding Data Annotations Extensions To Your Project First we will File->New Project and create an ASP.NET MVC 3 project.  I am going to use Razor for these examples, but any view engine can be used in practice.  Now go into the NuGet Extension Manager (right click on references and select add Library Package Reference) and search for “DataAnnotationsExtensions.”  You should see the following two packages: The first package is for server-side validation scenarios, but since we are using MVC 3 and would like comprehensive sever and client validation support, click on the DataAnnotationsExtensions.MVC3 project and then click Install.  This will install the Data Annotations Extensions server and client validation DLLs along with David Ebbo’s web activator (which enables the validation attributes to be registered with MVC 3). Now that Data Annotations Extensions is installed you have all you need to start doing advanced model validation.  If you are already using Data Annotations in your project, just making use of the additional validation attributes will provide client and server validation automatically.  However, assuming you are starting with a blank project I’ll walk you through setting up a controller and model to test with. Creating Your Model In the Models folder, create a new User.cs file with a User class that you can use as a model.  To start with, I’ll use the following class: public class User { public string Email { get; set; } public string Password { get; set; } public string PasswordConfirm { get; set; } public string HomePage { get; set; } public int Age { get; set; } } Next, create a simple controller with at least a Create method, and then a matching Create view (note, you can do all of this via the MVC built-in tooling).  Your files will look something like this: UserController.cs: public class UserController : Controller { public ActionResult Create() { return View(new User()); }   [HttpPost] public ActionResult Create(User user) { if (!ModelState.IsValid) { return View(user); }   return Content("User valid!"); } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Create.cshtml: @model NuGetValidationTester.Models.User   @{ ViewBag.Title = "Create"; }   <h2>Create</h2>   <script src="@Url.Content("~/Scripts/jquery.validate.min.js")" type="text/javascript"></script> <script src="@Url.Content("~/Scripts/jquery.validate.unobtrusive.min.js")" type="text/javascript"></script>   @using (Html.BeginForm()) { @Html.ValidationSummary(true) <fieldset> <legend>User</legend> @Html.EditorForModel() <p> <input type="submit" value="Create" /> </p> </fieldset> } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } In the Create.cshtml view, note that we are referencing jquery validation and jquery unobtrusive (jquery is referenced in the layout page).  These MVC 3 included scripts are the only ones you need to enjoy both the basic Data Annotations validation as well as the validation additions available in Data Annotations Extensions.  These references are added by default when you use the MVC 3 “Add View” dialog on a modification template type. Now when we go to /User/Create we should see a form for editing a User Since we haven’t yet added any validation attributes, this form is valid as shown (including no password, email and an age of 0).  With the built-in Data Annotations attributes we can make some of the fields required, and we could use a range validator of maybe 1 to 110 on Age (of course we don’t want to leave out supercentenarians) but let’s go further and validate our input comprehensively using Data Annotations Extensions.  The new and improved User.cs model class. { [Required] [Email] public string Email { get; set; }   [Required] public string Password { get; set; }   [Required] [EqualTo("Password")] public string PasswordConfirm { get; set; }   [Url] public string HomePage { get; set; }   [Integer] [Min(1)] public int Age { get; set; } } .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Now let’s re-run our form and try to use some invalid values: All of the validation errors you see above occurred on the client, without ever even hitting submit.  The validation is also checked on the server, which is a good practice since client validation is easily bypassed. That’s all you need to do to start a new project and include Data Annotations Extensions, and of course you can integrate it into an existing project just as easily. Nitpickers Corner ASP.NET MVC 3 futures defines four new data annotations attributes which this project has as well: CreditCard, Email, Url and EqualTo.  Unfortunately referencing MVC 3 futures necessitates taking an dependency on MVC 3 in your model layer, which may be unadvisable in a multi-tiered project.  Data Annotations Extensions keeps the server and client side libraries separate so using the project’s validation attributes don’t require you to take any additional dependencies in your model layer which still allowing for the rich client validation experience if you are using MVC 3. Custom Error Message and Globalization: Since the Data Annotations Extensions are build on top of Data Annotations, you have the ability to define your own static error messages and even to use resource files for very customizable error messages. Available Validators: Please see the project site at http://dataannotationsextensions.org/ for an up-to-date list of the new validators included in this project.  As of this post, the following validators are available: CreditCard Date Digits Email EqualTo FileExtensions Integer Max Min Numeric Url Conclusion Hopefully I’ve illustrated how easy it is to add server and client validation to your MVC 3 projects, and how to easily you can extend the available validation options to meet real world needs. The Data Annotations Extensions project is fully open source under the BSD license.  Any feedback would be greatly appreciated.  More information than you require, along with links to the source code, is available at http://dataannotationsextensions.org/. Enjoy!

    Read the article

  • Basic Spatial Data with SQL Server and Entity Framework 5.0

    - by Rick Strahl
    In my most recent project we needed to do a bit of geo-spatial referencing. While spatial features have been in SQL Server for a while using those features inside of .NET applications hasn't been as straight forward as could be, because .NET natively doesn't support spatial types. There are workarounds for this with a few custom project like SharpMap or a hack using the Sql Server specific Geo types found in the Microsoft.SqlTypes assembly that ships with SQL server. While these approaches work for manipulating spatial data from .NET code, they didn't work with database access if you're using Entity Framework. Other ORM vendors have been rolling their own versions of spatial integration. In Entity Framework 5.0 running on .NET 4.5 the Microsoft ORM finally adds support for spatial types as well. In this post I'll describe basic geography features that deal with single location and distance calculations which is probably the most common usage scenario. SQL Server Transact-SQL Syntax for Spatial Data Before we look at how things work with Entity framework, lets take a look at how SQL Server allows you to use spatial data to get an understanding of the underlying semantics. The following SQL examples should work with SQL 2008 and forward. Let's start by creating a test table that includes a Geography field and also a pair of Long/Lat fields that demonstrate how you can work with the geography functions even if you don't have geography/geometry fields in the database. Here's the CREATE command:CREATE TABLE [dbo].[Geo]( [id] [int] IDENTITY(1,1) NOT NULL, [Location] [geography] NULL, [Long] [float] NOT NULL, [Lat] [float] NOT NULL ) Now using plain SQL you can insert data into the table using geography::STGeoFromText SQL CLR function:insert into Geo( Location , long, lat ) values ( geography::STGeomFromText ('POINT(-121.527200 45.712113)', 4326), -121.527200, 45.712113 ) insert into Geo( Location , long, lat ) values ( geography::STGeomFromText ('POINT(-121.517265 45.714240)', 4326), -121.517265, 45.714240 ) insert into Geo( Location , long, lat ) values ( geography::STGeomFromText ('POINT(-121.511536 45.714825)', 4326), -121.511536, 45.714825) The STGeomFromText function accepts a string that points to a geometric item (a point here but can also be a line or path or polygon and many others). You also need to provide an SRID (Spatial Reference System Identifier) which is an integer value that determines the rules for how geography/geometry values are calculated and returned. For mapping/distance functionality you typically want to use 4326 as this is the format used by most mapping software and geo-location libraries like Google and Bing. The spatial data in the Location field is stored in binary format which looks something like this: Once the location data is in the database you can query the data and do simple distance computations very easily. For example to calculate the distance of each of the values in the database to another spatial point is very easy to calculate. Distance calculations compare two points in space using a direct line calculation. For our example I'll compare a new point to all the points in the database. Using the Location field the SQL looks like this:-- create a source point DECLARE @s geography SET @s = geography:: STGeomFromText('POINT(-121.527200 45.712113)' , 4326); --- return the ids select ID, Location as Geo , Location .ToString() as Point , @s.STDistance( Location) as distance from Geo order by distance The code defines a new point which is the base point to compare each of the values to. You can also compare values from the database directly, but typically you'll want to match a location to another location and determine the difference for which you can use the geography::STDistance function. This query produces the following output: The STDistance function returns the straight line distance between the passed in point and the point in the database field. The result for SRID 4326 is always in meters. Notice that the first value passed was the same point so the difference is 0. The other two points are two points here in town in Hood River a little ways away - 808 and 1256 meters respectively. Notice also that you can order the result by the resulting distance, which effectively gives you results that are ordered radially out from closer to further away. This is great for searches of points of interest near a central location (YOU typically!). These geolocation functions are also available to you if you don't use the Geography/Geometry types, but plain float values. It's a little more work, as each point has to be created in the query using the string syntax, but the following code doesn't use a geography field but produces the same result as the previous query.--- using float fields select ID, geography::STGeomFromText ('POINT(' + STR (long, 15,7 ) + ' ' + Str(lat ,15, 7) + ')' , 4326), geography::STGeomFromText ('POINT(' + STR (long, 15,7 ) + ' ' + Str(lat ,15, 7) + ')' , 4326). ToString(), @s.STDistance( geography::STGeomFromText ('POINT(' + STR(long ,15, 7) + ' ' + Str(lat ,15, 7) + ')' , 4326)) as distance from geo order by distance Spatial Data in the Entity Framework Prior to Entity Framework 5.0 on .NET 4.5 consuming of the data above required using stored procedures or raw SQL commands to access the spatial data. In Entity Framework 5 however, Microsoft introduced the new DbGeometry and DbGeography types. These immutable location types provide a bunch of functionality for manipulating spatial points using geometry functions which in turn can be used to do common spatial queries like I described in the SQL syntax above. The DbGeography/DbGeometry types are immutable, meaning that you can't write to them once they've been created. They are a bit odd in that you need to use factory methods in order to instantiate them - they have no constructor() and you can't assign to properties like Latitude and Longitude. Creating a Model with Spatial Data Let's start by creating a simple Entity Framework model that includes a Location property of type DbGeography: public class GeoLocationContext : DbContext { public DbSet<GeoLocation> Locations { get; set; } } public class GeoLocation { public int Id { get; set; } public DbGeography Location { get; set; } public string Address { get; set; } } That's all there's to it. When you run this now against SQL Server, you get a Geography field for the Location property, which looks the same as the Location field in the SQL examples earlier. Adding Spatial Data to the Database Next let's add some data to the table that includes some latitude and longitude data. An easy way to find lat/long locations is to use Google Maps to pinpoint your location, then right click and click on What's Here. Click on the green marker to get the GPS coordinates. To add the actual geolocation data create an instance of the GeoLocation type and use the DbGeography.PointFromText() factory method to create a new point to assign to the Location property:[TestMethod] public void AddLocationsToDataBase() { var context = new GeoLocationContext(); // remove all context.Locations.ToList().ForEach( loc => context.Locations.Remove(loc)); context.SaveChanges(); var location = new GeoLocation() { // Create a point using native DbGeography Factory method Location = DbGeography.PointFromText( string.Format("POINT({0} {1})", -121.527200,45.712113) ,4326), Address = "301 15th Street, Hood River" }; context.Locations.Add(location); location = new GeoLocation() { Location = CreatePoint(45.714240, -121.517265), Address = "The Hatchery, Bingen" }; context.Locations.Add(location); location = new GeoLocation() { // Create a point using a helper function (lat/long) Location = CreatePoint(45.708457, -121.514432), Address = "Kaze Sushi, Hood River" }; context.Locations.Add(location); location = new GeoLocation() { Location = CreatePoint(45.722780, -120.209227), Address = "Arlington, OR" }; context.Locations.Add(location); context.SaveChanges(); } As promised, a DbGeography object has to be created with one of the static factory methods provided on the type as the Location.Longitude and Location.Latitude properties are read only. Here I'm using PointFromText() which uses a "Well Known Text" format to specify spatial data. In the first example I'm specifying to create a Point from a longitude and latitude value, using an SRID of 4326 (just like earlier in the SQL examples). You'll probably want to create a helper method to make the creation of Points easier to avoid that string format and instead just pass in a couple of double values. Here's my helper called CreatePoint that's used for all but the first point creation in the sample above:public static DbGeography CreatePoint(double latitude, double longitude) { var text = string.Format(CultureInfo.InvariantCulture.NumberFormat, "POINT({0} {1})", longitude, latitude); // 4326 is most common coordinate system used by GPS/Maps return DbGeography.PointFromText(text, 4326); } Using the helper the syntax becomes a bit cleaner, requiring only a latitude and longitude respectively. Note that my method intentionally swaps the parameters around because Latitude and Longitude is the common format I've seen with mapping libraries (especially Google Mapping/Geolocation APIs with their LatLng type). When the context is changed the data is written into the database using the SQL Geography type which looks the same as in the earlier SQL examples shown. Querying Once you have some location data in the database it's now super easy to query the data and find out the distance between locations. A common query is to ask for a number of locations that are near a fixed point - typically your current location and order it by distance. Using LINQ to Entities a query like this is easy to construct:[TestMethod] public void QueryLocationsTest() { var sourcePoint = CreatePoint(45.712113, -121.527200); var context = new GeoLocationContext(); // find any locations within 5 kilometers ordered by distance var matches = context.Locations .Where(loc => loc.Location.Distance(sourcePoint) < 5000) .OrderBy( loc=> loc.Location.Distance(sourcePoint) ) .Select( loc=> new { Address = loc.Address, Distance = loc.Location.Distance(sourcePoint) }); Assert.IsTrue(matches.Count() > 0); foreach (var location in matches) { Console.WriteLine("{0} ({1:n0} meters)", location.Address, location.Distance); } } This example produces: 301 15th Street, Hood River (0 meters)The Hatchery, Bingen (809 meters)Kaze Sushi, Hood River (1,074 meters)   The first point in the database is the same as my source point I'm comparing against so the distance is 0. The other two are within the 5 mile radius, while the Arlington location which is 65 miles or so out is not returned. The result is ordered by distance from closest to furthest away. In the code, I first create a source point that is the basis for comparison. The LINQ query then selects all locations that are within 5km of the source point using the Location.Distance() function, which takes a source point as a parameter. You can either use a pre-defined value as I'm doing here, or compare against another database DbGeography property (say when you have to points in the same database for things like routes). What's nice about this query syntax is that it's very clean and easy to read and understand. You can calculate the distance and also easily order by the distance to provide a result that shows locations from closest to furthest away which is a common scenario for any application that places a user in the context of several locations. It's now super easy to accomplish this. Meters vs. Miles As with the SQL Server functions, the Distance() method returns data in meters, so if you need to work with miles or feet you need to do some conversion. Here are a couple of helpers that might be useful (can be found in GeoUtils.cs of the sample project):/// <summary> /// Convert meters to miles /// </summary> /// <param name="meters"></param> /// <returns></returns> public static double MetersToMiles(double? meters) { if (meters == null) return 0F; return meters.Value * 0.000621371192; } /// <summary> /// Convert miles to meters /// </summary> /// <param name="miles"></param> /// <returns></returns> public static double MilesToMeters(double? miles) { if (miles == null) return 0; return miles.Value * 1609.344; } Using these two helpers you can query on miles like this:[TestMethod] public void QueryLocationsMilesTest() { var sourcePoint = CreatePoint(45.712113, -121.527200); var context = new GeoLocationContext(); // find any locations within 5 miles ordered by distance var fiveMiles = GeoUtils.MilesToMeters(5); var matches = context.Locations .Where(loc => loc.Location.Distance(sourcePoint) <= fiveMiles) .OrderBy(loc => loc.Location.Distance(sourcePoint)) .Select(loc => new { Address = loc.Address, Distance = loc.Location.Distance(sourcePoint) }); Assert.IsTrue(matches.Count() > 0); foreach (var location in matches) { Console.WriteLine("{0} ({1:n1} miles)", location.Address, GeoUtils.MetersToMiles(location.Distance)); } } which produces: 301 15th Street, Hood River (0.0 miles)The Hatchery, Bingen (0.5 miles)Kaze Sushi, Hood River (0.7 miles) Nice 'n simple. .NET 4.5 Only Note that DbGeography and DbGeometry are exclusive to Entity Framework 5.0 (not 4.4 which ships in the same NuGet package or installer) and requires .NET 4.5. That's because the new DbGeometry and DbGeography (and related) types are defined in the 4.5 version of System.Data.Entity which is a CLR assembly and is only updated by major versions of .NET. Why this decision was made to add these types to System.Data.Entity rather than to the frequently updated EntityFramework assembly that would have possibly made this work in .NET 4.0 is beyond me, especially given that there are no native .NET framework spatial types to begin with. I find it also odd that there is no native CLR spatial type. The DbGeography and DbGeometry types are specific to Entity Framework and live on those assemblies. They will also work for general purpose, non-database spatial data manipulation, but then you are forced into having a dependency on System.Data.Entity, which seems a bit silly. There's also a System.Spatial assembly that's apparently part of WCF Data Services which in turn don't work with Entity framework. Another example of multiple teams at Microsoft not communicating and implementing the same functionality (differently) in several different places. Perplexed as a I may be, for EF specific code the Entity framework specific types are easy to use and work well. Working with pre-.NET 4.5 Entity Framework and Spatial Data If you can't go to .NET 4.5 just yet you can also still use spatial features in Entity Framework, but it's a lot more work as you can't use the DbContext directly to manipulate the location data. You can still run raw SQL statements to write data into the database and retrieve results using the same TSQL syntax I showed earlier using Context.Database.ExecuteSqlCommand(). Here's code that you can use to add location data into the database:[TestMethod] public void RawSqlEfAddTest() { string sqlFormat = @"insert into GeoLocations( Location, Address) values ( geography::STGeomFromText('POINT({0} {1})', 4326),@p0 )"; var sql = string.Format(sqlFormat,-121.527200, 45.712113); Console.WriteLine(sql); var context = new GeoLocationContext(); Assert.IsTrue(context.Database.ExecuteSqlCommand(sql,"301 N. 15th Street") > 0); } Here I'm using the STGeomFromText() function to add the location data. Note that I'm using string.Format here, which usually would be a bad practice but is required here. I was unable to use ExecuteSqlCommand() and its named parameter syntax as the longitude and latitude parameters are embedded into a string. Rest assured it's required as the following does not work:string sqlFormat = @"insert into GeoLocations( Location, Address) values ( geography::STGeomFromText('POINT(@p0 @p1)', 4326),@p2 )";context.Database.ExecuteSqlCommand(sql, -121.527200, 45.712113, "301 N. 15th Street") Explicitly assigning the point value with string.format works however. There are a number of ways to query location data. You can't get the location data directly, but you can retrieve the point string (which can then be parsed to get Latitude and Longitude) and you can return calculated values like distance. Here's an example of how to retrieve some geo data into a resultset using EF's and SqlQuery method:[TestMethod] public void RawSqlEfQueryTest() { var sqlFormat = @" DECLARE @s geography SET @s = geography:: STGeomFromText('POINT({0} {1})' , 4326); SELECT Address, Location.ToString() as GeoString, @s.STDistance( Location) as Distance FROM GeoLocations ORDER BY Distance"; var sql = string.Format(sqlFormat, -121.527200, 45.712113); var context = new GeoLocationContext(); var locations = context.Database.SqlQuery<ResultData>(sql); Assert.IsTrue(locations.Count() > 0); foreach (var location in locations) { Console.WriteLine(location.Address + " " + location.GeoString + " " + location.Distance); } } public class ResultData { public string GeoString { get; set; } public double Distance { get; set; } public string Address { get; set; } } Hopefully you don't have to resort to this approach as it's fairly limited. Using the new DbGeography/DbGeometry types makes this sort of thing so much easier. When I had to use code like this before I typically ended up retrieving data pks only and then running another query with just the PKs to retrieve the actual underlying DbContext entities. This was very inefficient and tedious but it did work. Summary For the current project I'm working on we actually made the switch to .NET 4.5 purely for the spatial features in EF 5.0. This app heavily relies on spatial queries and it was worth taking a chance with pre-release code to get this ease of integration as opposed to manually falling back to stored procedures or raw SQL string queries to return spatial specific queries. Using native Entity Framework code makes life a lot easier than the alternatives. It might be a late addition to Entity Framework, but it sure makes location calculations and storage easy. Where do you want to go today? ;-) Resources Download Sample Project© Rick Strahl, West Wind Technologies, 2005-2012Posted in ADO.NET  Sql Server  .NET   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

    Read the article

  • lxc containers hangs after upgrade to 13.10

    - by doug123
    I have 3 lxc containers. They were all working fine on 12.10 and I upgraded the containers with do-release-upgrade on the containers to 13.04 and 13.10 and that worked great. Then I upgraded the host to 13.04 and then 13.10 and now the 3 containers hang with this: >lxc-start -n as1 -l DEBUG -o $(tty) lxc-start 1383145786.513 INFO lxc_start_ui - using rcfile /var/lib/lxc/as1/config lxc-start 1383145786.513 WARN lxc_log - lxc_log_init called with log already initialized lxc-start 1383145786.513 INFO lxc_apparmor - aa_enabled set to 1 lxc-start 1383145786.514 DEBUG lxc_conf - allocated pty '/dev/pts/2' (5/6) lxc-start 1383145786.514 DEBUG lxc_conf - allocated pty '/dev/pts/13' (7/8) lxc-start 1383145786.514 DEBUG lxc_conf - allocated pty '/dev/pts/14' (9/10) lxc-start 1383145786.514 DEBUG lxc_conf - allocated pty '/dev/pts/15' (11/12) lxc-start 1383145786.514 DEBUG lxc_conf - allocated pty '/dev/pts/17' (13/14) lxc-start 1383145786.514 DEBUG lxc_conf - allocated pty '/dev/pts/18' (15/16) lxc-start 1383145786.514 DEBUG lxc_conf - allocated pty '/dev/pts/19' (17/18) lxc-start 1383145786.514 DEBUG lxc_conf - allocated pty '/dev/pts/20' (19/20) lxc-start 1383145786.514 INFO lxc_conf - tty's configured lxc-start 1383145786.514 DEBUG lxc_start - sigchild handler set lxc-start 1383145786.514 DEBUG lxc_console - opening /dev/tty for console peer lxc-start 1383145786.514 DEBUG lxc_console - using '/dev/tty' as console lxc-start 1383145786.514 DEBUG lxc_console - 6242 got SIGWINCH fd 25 lxc-start 1383145786.514 DEBUG lxc_console - set winsz dstfd:22 cols:177 rows:53 lxc-start 1383145786.514 INFO lxc_start - 'as1' is initialized lxc-start 1383145786.522 DEBUG lxc_start - Not dropping cap_sys_boot or watching utmp lxc-start 1383145786.524 DEBUG lxc_conf - mac address of host interface 'vethB4L35W' changed to private fe:7c:96:a0:ae:29 lxc-start 1383145786.525 DEBUG lxc_conf - instanciated veth 'vethB4L35W/vethVC61K2', index is '26' lxc-start 1383145786.529 DEBUG lxc_cgroup - cgroup 'memory.limit_in_bytes' set to '20G' lxc-start 1383145786.529 DEBUG lxc_cgroup - cgroup 'cpuset.cpus' set to '12-23' lxc-start 1383145786.529 INFO lxc_cgroup - cgroup has been setup lxc-start 1383145786.555 DEBUG lxc_conf - move 'eth0' to '6249' lxc-start 1383145786.555 INFO lxc_conf - 'as1' hostname has been setup lxc-start 1383145786.575 DEBUG lxc_conf - 'eth0' has been setup lxc-start 1383145786.575 INFO lxc_conf - network has been setup lxc-start 1383145786.575 INFO lxc_conf - looking at .44 42 252:0 / / rw,relatime - ext4 /dev/mapper/limitorderbook1-root rw,errors=remount-ro,data=ordered . lxc-start 1383145786.575 INFO lxc_conf - now p is . /. lxc-start 1383145786.575 INFO lxc_conf - looking at .52 44 0:5 / /dev rw,relatime - devtmpfs udev rw,size=32961632k,nr_inodes=8240408,mode=755 . lxc-start 1383145786.575 INFO lxc_conf - now p is . /dev. lxc-start 1383145786.575 INFO lxc_conf - looking at .61 52 0:11 / /dev/pts rw,nosuid,noexec,relatime - devpts devpts rw,mode=600,ptmxmode=000 . lxc-start 1383145786.575 INFO lxc_conf - now p is . /dev/pts. lxc-start 1383145786.575 INFO lxc_conf - looking at .68 44 0:15 / /run rw,nosuid,noexec,relatime - tmpfs tmpfs rw,size=6594456k,mode=755 . lxc-start 1383145786.575 INFO lxc_conf - now p is . /run. lxc-start 1383145786.575 INFO lxc_conf - looking at .69 68 0:18 / /run/lock rw,nosuid,nodev,noexec,relatime - tmpfs none rw,size=5120k . lxc-start 1383145786.575 INFO lxc_conf - now p is . /run/lock. lxc-start 1383145786.575 INFO lxc_conf - looking at .72 68 0:19 / /run/shm rw,nosuid,nodev,relatime - tmpfs none rw . lxc-start 1383145786.575 INFO lxc_conf - now p is . /run/shm. lxc-start 1383145786.575 INFO lxc_conf - looking at .73 68 0:21 / /run/user rw,nosuid,nodev,noexec,relatime - tmpfs none rw,size=102400k,mode=755 . lxc-start 1383145786.575 INFO lxc_conf - now p is . /run/user. lxc-start 1383145786.575 INFO lxc_conf - looking at .76 44 0:14 / /sys rw,nosuid,nodev,noexec,relatime - sysfs sysfs rw . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys. lxc-start 1383145786.575 INFO lxc_conf - looking at .77 76 0:16 / /sys/fs/cgroup rw,relatime - tmpfs none rw,size=4k,mode=755 . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup. lxc-start 1383145786.575 INFO lxc_conf - looking at .78 77 0:20 / /sys/fs/cgroup/cpuset rw,relatime - cgroup cgroup rw,cpuset,clone_children . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup/cpuset. lxc-start 1383145786.575 INFO lxc_conf - looking at .79 77 0:23 / /sys/fs/cgroup/cpu rw,relatime - cgroup cgroup rw,cpu . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup/cpu. lxc-start 1383145786.575 INFO lxc_conf - looking at .80 77 0:24 / /sys/fs/cgroup/cpuacct rw,relatime - cgroup cgroup rw,cpuacct . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup/cpuacct. lxc-start 1383145786.575 INFO lxc_conf - looking at .81 77 0:25 / /sys/fs/cgroup/memory rw,relatime - cgroup cgroup rw,memory . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup/memory. lxc-start 1383145786.575 INFO lxc_conf - looking at .82 77 0:26 / /sys/fs/cgroup/devices rw,relatime - cgroup cgroup rw,devices . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup/devices. lxc-start 1383145786.575 INFO lxc_conf - looking at .83 77 0:27 / /sys/fs/cgroup/freezer rw,relatime - cgroup cgroup rw,freezer . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup/freezer. lxc-start 1383145786.575 INFO lxc_conf - looking at .84 77 0:28 / /sys/fs/cgroup/blkio rw,relatime - cgroup cgroup rw,blkio . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup/blkio. lxc-start 1383145786.575 INFO lxc_conf - looking at .85 77 0:29 / /sys/fs/cgroup/perf_event rw,relatime - cgroup cgroup rw,perf_event . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup/perf_event. lxc-start 1383145786.575 INFO lxc_conf - looking at .94 77 0:30 / /sys/fs/cgroup/hugetlb rw,relatime - cgroup cgroup rw,hugetlb . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup/hugetlb. lxc-start 1383145786.575 INFO lxc_conf - looking at .95 77 0:31 / /sys/fs/cgroup/systemd rw,nosuid,nodev,noexec,relatime - cgroup systemd rw,name=systemd . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/cgroup/systemd. lxc-start 1383145786.575 INFO lxc_conf - looking at .96 76 0:17 / /sys/fs/fuse/connections rw,relatime - fusectl none rw . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/fuse/connections. lxc-start 1383145786.575 INFO lxc_conf - looking at .98 76 0:6 / /sys/kernel/debug rw,relatime - debugfs none rw . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/kernel/debug. lxc-start 1383145786.575 INFO lxc_conf - looking at .101 76 0:10 / /sys/kernel/security rw,relatime - securityfs none rw . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/kernel/security. lxc-start 1383145786.575 INFO lxc_conf - looking at .102 76 0:22 / /sys/fs/pstore rw,relatime - pstore none rw . lxc-start 1383145786.575 INFO lxc_conf - now p is . /sys/fs/pstore. lxc-start 1383145786.575 INFO lxc_conf - looking at .103 44 0:3 / /proc rw,nosuid,nodev,noexec,relatime - proc proc rw . lxc-start 1383145786.575 INFO lxc_conf - now p is . /proc. lxc-start 1383145786.575 INFO lxc_conf - looking at .104 44 9:2 / /data rw,relatime - ext4 /dev/md2 rw,errors=remount-ro,data=ordered . lxc-start 1383145786.575 INFO lxc_conf - now p is . /data. lxc-start 1383145786.575 INFO lxc_conf - looking at .105 44 8:1 / /boot rw,relatime - ext2 /dev/sda1 rw,errors=continue . lxc-start 1383145786.575 INFO lxc_conf - now p is . /boot. lxc-start 1383145786.576 DEBUG lxc_conf - mounted '/data/srv/lxc/as1' on '/usr/lib/x86_64-linux-gnu/lxc' lxc-start 1383145786.576 DEBUG lxc_conf - mounted 'none' on '/usr/lib/x86_64-linux-gnu/lxc//dev/pts', type 'devpts' lxc-start 1383145786.576 DEBUG lxc_conf - mounted 'none' on '/usr/lib/x86_64-linux-gnu/lxc//proc', type 'proc' lxc-start 1383145786.576 DEBUG lxc_conf - mounted 'none' on '/usr/lib/x86_64-linux-gnu/lxc//sys', type 'sysfs' lxc-start 1383145786.576 DEBUG lxc_conf - mounted 'none' on '/usr/lib/x86_64-linux-gnu/lxc//run', type 'tmpfs' lxc-start 1383145786.576 INFO lxc_conf - mount points have been setup lxc-start 1383145786.577 INFO lxc_conf - console has been setup lxc-start 1383145786.577 INFO lxc_conf - 8 tty(s) has been setup lxc-start 1383145786.577 INFO lxc_conf - rootfs path is ./data/srv/lxc/as1., mount is ./usr/lib/x86_64-linux-gnu/lxc. lxc-start 1383145786.577 INFO lxc_apparmor - I am 1, /proc/self points to 1 lxc-start 1383145786.577 DEBUG lxc_conf - created '/usr/lib/x86_64-linux-gnu/lxc/lxc_putold' directory lxc-start 1383145786.577 DEBUG lxc_conf - mountpoint for old rootfs is '/usr/lib/x86_64-linux-gnu/lxc/lxc_putold' lxc-start 1383145786.577 DEBUG lxc_conf - pivot_root syscall to '/usr/lib/x86_64-linux-gnu/lxc' successful lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/dev/pts' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/run/lock' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/run/shm' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/run/user' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup/cpuset' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup/cpu' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup/cpuacct' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup/memory' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup/devices' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup/freezer' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup/blkio' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup/perf_event' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup/hugetlb' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup/systemd' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/fuse/connections' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/kernel/debug' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/kernel/security' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/pstore' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/proc' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/data' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/boot' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/dev' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/run' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys/fs/cgroup' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold/sys' lxc-start 1383145786.577 DEBUG lxc_conf - umounted '/lxc_putold' lxc-start 1383145786.577 INFO lxc_conf - created new pts instance lxc-start 1383145786.578 DEBUG lxc_conf - drop capability 'sys_boot' (22) lxc-start 1383145786.578 DEBUG lxc_conf - capabilities have been setup lxc-start 1383145786.578 NOTICE lxc_conf - 'as1' is setup. lxc-start 1383145786.578 DEBUG lxc_cgroup - cgroup 'memory.limit_in_bytes' set to '20G' lxc-start 1383145786.578 DEBUG lxc_cgroup - cgroup 'cpuset.cpus' set to '12-23' lxc-start 1383145786.578 INFO lxc_cgroup - cgroup has been setup lxc-start 1383145786.578 INFO lxc_apparmor - setting up apparmor lxc-start 1383145786.578 INFO lxc_apparmor - changed apparmor profile to lxc-container-default lxc-start 1383145786.578 NOTICE lxc_start - exec'ing '/sbin/init' lxc-start 1383145786.578 INFO lxc_conf - looking at .15 20 0:14 / /sys rw,nosuid,nodev,noexec,relatime - sysfs sysfs rw . lxc-start 1383145786.578 INFO lxc_conf - now p is . /sys. lxc-start 1383145786.578 INFO lxc_conf - looking at .16 20 0:3 / /proc rw,nosuid,nodev,noexec,relatime - proc proc rw . lxc-start 1383145786.578 INFO lxc_conf - now p is . /proc. lxc-start 1383145786.578 INFO lxc_conf - looking at .17 20 0:5 / /dev rw,relatime - devtmpfs udev rw,size=32961632k,nr_inodes=8240408,mode=755 . lxc-start 1383145786.578 INFO lxc_conf - now p is . /dev. lxc-start 1383145786.578 INFO lxc_conf - looking at .18 17 0:11 / /dev/pts rw,nosuid,noexec,relatime - devpts devpts rw,mode=600,ptmxmode=000 . lxc-start 1383145786.578 INFO lxc_conf - now p is . /dev/pts. lxc-start 1383145786.578 INFO lxc_conf - looking at .19 20 0:15 / /run rw,nosuid,noexec,relatime - tmpfs tmpfs rw,size=6594456k,mode=755 . lxc-start 1383145786.578 INFO lxc_conf - now p is . /run. lxc-start 1383145786.578 INFO lxc_conf - looking at .20 1 252:0 / / rw,relatime - ext4 /dev/mapper/limitorderbook1-root rw,errors=remount-ro,data=ordered . lxc-start 1383145786.578 INFO lxc_conf - now p is . /. lxc-start 1383145786.578 INFO lxc_conf - looking at .22 15 0:16 / /sys/fs/cgroup rw,relatime - tmpfs none rw,size=4k,mode=755 . lxc-start 1383145786.578 INFO lxc_conf - now p is . /sys/fs/cgroup. lxc-start 1383145786.578 INFO lxc_conf - looking at .23 15 0:17 / /sys/fs/fuse/connections rw,relatime - fusectl none rw . lxc-start 1383145786.578 INFO lxc_conf - now p is . /sys/fs/fuse/connections. lxc-start 1383145786.578 INFO lxc_conf - looking at .24 15 0:6 / /sys/kernel/debug rw,relatime - debugfs none rw . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/kernel/debug. lxc-start 1383145786.579 INFO lxc_conf - looking at .25 15 0:10 / /sys/kernel/security rw,relatime - securityfs none rw . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/kernel/security. lxc-start 1383145786.579 INFO lxc_conf - looking at .26 19 0:18 / /run/lock rw,nosuid,nodev,noexec,relatime - tmpfs none rw,size=5120k . lxc-start 1383145786.579 INFO lxc_conf - now p is . /run/lock. lxc-start 1383145786.579 INFO lxc_conf - looking at .27 19 0:19 / /run/shm rw,nosuid,nodev,relatime - tmpfs none rw . lxc-start 1383145786.579 INFO lxc_conf - now p is . /run/shm. lxc-start 1383145786.579 INFO lxc_conf - looking at .28 22 0:20 / /sys/fs/cgroup/cpuset rw,relatime - cgroup cgroup rw,cpuset,clone_children . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/cgroup/cpuset. lxc-start 1383145786.579 INFO lxc_conf - looking at .29 19 0:21 / /run/user rw,nosuid,nodev,noexec,relatime - tmpfs none rw,size=102400k,mode=755 . lxc-start 1383145786.579 INFO lxc_conf - now p is . /run/user. lxc-start 1383145786.579 INFO lxc_conf - looking at .30 15 0:22 / /sys/fs/pstore rw,relatime - pstore none rw . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/pstore. lxc-start 1383145786.579 INFO lxc_conf - looking at .31 22 0:23 / /sys/fs/cgroup/cpu rw,relatime - cgroup cgroup rw,cpu . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/cgroup/cpu. lxc-start 1383145786.579 INFO lxc_conf - looking at .32 22 0:24 / /sys/fs/cgroup/cpuacct rw,relatime - cgroup cgroup rw,cpuacct . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/cgroup/cpuacct. lxc-start 1383145786.579 INFO lxc_conf - looking at .33 22 0:25 / /sys/fs/cgroup/memory rw,relatime - cgroup cgroup rw,memory . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/cgroup/memory. lxc-start 1383145786.579 INFO lxc_conf - looking at .34 22 0:26 / /sys/fs/cgroup/devices rw,relatime - cgroup cgroup rw,devices . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/cgroup/devices. lxc-start 1383145786.579 INFO lxc_conf - looking at .35 22 0:27 / /sys/fs/cgroup/freezer rw,relatime - cgroup cgroup rw,freezer . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/cgroup/freezer. lxc-start 1383145786.579 INFO lxc_conf - looking at .36 22 0:28 / /sys/fs/cgroup/blkio rw,relatime - cgroup cgroup rw,blkio . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/cgroup/blkio. lxc-start 1383145786.579 INFO lxc_conf - looking at .37 22 0:29 / /sys/fs/cgroup/perf_event rw,relatime - cgroup cgroup rw,perf_event . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/cgroup/perf_event. lxc-start 1383145786.579 INFO lxc_conf - looking at .38 22 0:30 / /sys/fs/cgroup/hugetlb rw,relatime - cgroup cgroup rw,hugetlb . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/cgroup/hugetlb. lxc-start 1383145786.579 INFO lxc_conf - looking at .39 20 9:2 / /data rw,relatime - ext4 /dev/md2 rw,errors=remount-ro,data=ordered . lxc-start 1383145786.579 INFO lxc_conf - now p is . /data. lxc-start 1383145786.579 INFO lxc_conf - looking at .40 20 8:1 / /boot rw,relatime - ext2 /dev/sda1 rw,errors=continue . lxc-start 1383145786.579 INFO lxc_conf - now p is . /boot. lxc-start 1383145786.579 INFO lxc_conf - looking at .41 22 0:31 / /sys/fs/cgroup/systemd rw,nosuid,nodev,noexec,relatime - cgroup systemd rw,name=systemd . lxc-start 1383145786.579 INFO lxc_conf - now p is . /sys/fs/cgroup/systemd. lxc-start 1383145786.579 NOTICE lxc_start - '/sbin/init' started with pid '6249' lxc-start 1383145786.579 WARN lxc_start - invalid pid for SIGCHLD <4>init: ureadahead main process (7) terminated with status 5 <4>init: console-font main process (94) terminated with status 1 And it will just sit there like that for hours at least. The container becomes pingable but I can't ssh and if I try lxc-console -n as1 I get a blank screen. If I do lxc-stop -n as1 or ^C in the window where it has hung I get: ^CTERM environment variable not set. <4>init: plymouth-upstart-bridge main process (192) terminated with status 1 <4>init: hwclock-save main process (187) terminated with status 70 * Asking all remaining processes to terminate... ...done. * All processes ended within 1 seconds... ...done. * Deactivating swap... ...fail! mount: cannot mount block device /dev/md2 read-only * Will now restart But after 20 minutes it hasn't restarted. Any ideas why these containers are hanging?

    Read the article

  • Windows Azure Evolution &ndash; Caching (Preview)

    - by Shaun
    Caching is a popular topic when we are building a high performance and high scalable system not only on top of the cloud platform but the on-premise environment as well. On March 2011 the Windows Azure AppFabric Caching had been production launched. It provides an in-memory, distributed caching service over the cloud. And now, in this June 2012 update, the cache team announce a grand new caching solution on Windows Azure, which is called Windows Azure Caching (Preview). And the original Windows Azure AppFabric Caching was renamed to Windows Azure Shared Caching.   What’s Caching (Preview) If you had been using the Shared Caching you should know that it is constructed by a bunch of cache servers. And when you want to use you should firstly create a cache account from the developer portal and specify the size you want to use, which means how much memory you can use to store your data that wanted to be cached. Then you can add, get and remove them through your code through the cache URL. The Shared Caching is a multi-tenancy system which host all cached items across all users. So you don’t know which server your data was located. This caching mode works well and can take most of the cases. But it has some problems. The first one is the performance. Since the Shared Caching is a multi-tenancy system, which means all cache operations should go through the Shared Caching gateway and then routed to the server which have the data your are looking for. Even though there are some caches in the Shared Caching system it also takes time from your cloud services to the cache service. Secondary, the Shared Caching service works as a block box to the developer. The only thing we know is my cache endpoint, and that’s all. Someone may satisfied since they don’t want to care about anything underlying. But if you need to know more and want more control that’s impossible in the Shared Caching. The last problem would be the price and cost-efficiency. You pay the bill based on how much cache you requested per month. But when we host a web role or worker role, it seldom consumes all of the memory and CPU in the virtual machine (service instance). If using Shared Caching we have to pay for the cache service while waste of some of our memory and CPU locally. Since the issues above Microsoft offered a new caching mode over to us, which is the Caching (Preview). Instead of having a separated cache service, the Caching (Preview) leverage the memory and CPU in our cloud services (web role and worker role) as the cache clusters. Hence the Caching (Preview) runs on the virtual machines which hosted or near our cloud applications. Without any gateway and routing, since it located in the same data center and same racks, it provides really high performance than the Shared Caching. The Caching (Preview) works side-by-side to our application, initialized and worked as a Windows Service running in the virtual machines invoked by the startup tasks from our roles, we could get more information and control to them. And since the Caching (Preview) utilizes the memory and CPU from our existing cloud services, so it’s free. What we need to pay is the original computing price. And the resource on each machines could be used more efficiently.   Enable Caching (Preview) It’s very simple to enable the Caching (Preview) in a cloud service. Let’s create a new windows azure cloud project from Visual Studio and added an ASP.NET Web Role. Then open the role setting and select the Caching page. This is where we enable and configure the Caching (Preview) on a role. To enable the Caching (Preview) just open the “Enable Caching (Preview Release)” check box. And then we need to specify which mode of the caching clusters we want to use. There are two kinds of caching mode, co-located and dedicate. The co-located mode means we use the memory in the instances we run our cloud services (web role or worker role). By using this mode we must specify how many percentage of the memory will be used as the cache. The default value is 30%. So make sure it will not affect the role business execution. The dedicate mode will use all memory in the virtual machine as the cache. In fact it will reserve some for operation system, azure hosting etc.. But it will try to use as much as the available memory to be the cache. As you can see, the Caching (Preview) was defined based on roles, which means all instances of this role will apply the same setting and play as a whole cache pool, and you can consume it by specifying the name of the role, which I will demonstrate later. And in a windows azure project we can have more than one role have the Caching (Preview) enabled. Then we will have more caches. For example, let’s say I have a web role and worker role. The web role I specified 30% co-located caching and the worker role I specified dedicated caching. If I have 3 instances of my web role and 2 instances of my worker role, then I will have two caches. As the figure above, cache 1 was contributed by three web role instances while cache 2 was contributed by 2 worker role instances. Then we can add items into cache 1 and retrieve it from web role code and worker role code. But the items stored in cache 1 cannot be retrieved from cache 2 since they are isolated. Back to our Visual Studio we specify 30% of co-located cache and use the local storage emulator to store the cache cluster runtime status. Then at the bottom we can specify the named caches. Now we just use the default one. Now we had enabled the Caching (Preview) in our web role settings. Next, let’s have a look on how to consume our cache.   Consume Caching (Preview) The Caching (Preview) can only be consumed by the roles in the same cloud services. As I mentioned earlier, a cache contributed by web role can be connected from a worker role if they are in the same cloud service. But you cannot consume a Caching (Preview) from other cloud services. This is different from the Shared Caching. The Shared Caching is opened to all services if it has the connection URL and authentication token. To consume the Caching (Preview) we need to add some references into our project as well as some configuration in the Web.config. NuGet makes our life easy. Right click on our web role project and select “Manage NuGet packages”, and then search the package named “WindowsAzure.Caching”. In the package list install the “Windows Azure Caching Preview”. It will download all necessary references from the NuGet repository and update our Web.config as well. Open the Web.config of our web role and find the “dataCacheClients” node. Under this node we can specify the cache clients we are going to use. For each cache client it will use the role name to identity and find the cache. Since we only have this web role with the Caching (Preview) enabled so I pasted the current role name in the configuration. Then, in the default page I will add some code to show how to use the cache. I will have a textbox on the page where user can input his or her name, then press a button to generate the email address for him/her. And in backend code I will check if this name had been added in cache. If yes I will return the email back immediately. Otherwise, I will sleep the tread for 2 seconds to simulate the latency, then add it into cache and return back to the page. 1: protected void btnGenerate_Click(object sender, EventArgs e) 2: { 3: // check if name is specified 4: var name = txtName.Text; 5: if (string.IsNullOrWhiteSpace(name)) 6: { 7: lblResult.Text = "Error. Please specify name."; 8: return; 9: } 10:  11: bool cached; 12: var sw = new Stopwatch(); 13: sw.Start(); 14:  15: // create the cache factory and cache 16: var factory = new DataCacheFactory(); 17: var cache = factory.GetDefaultCache(); 18:  19: // check if the name specified is in cache 20: var email = cache.Get(name) as string; 21: if (email != null) 22: { 23: cached = true; 24: sw.Stop(); 25: } 26: else 27: { 28: cached = false; 29: // simulate the letancy 30: Thread.Sleep(2000); 31: email = string.Format("{0}@igt.com", name); 32: // add to cache 33: cache.Add(name, email); 34: } 35:  36: sw.Stop(); 37: lblResult.Text = string.Format( 38: "Cached = {0}. Duration: {1}s. {2} => {3}", 39: cached, sw.Elapsed.TotalSeconds.ToString("0.00"), name, email); 40: } The Caching (Preview) can be used on the local emulator so we just F5. The first time I entered my name it will take about 2 seconds to get the email back to me since it was not in the cache. But if we re-enter my name it will be back at once from the cache. Since the Caching (Preview) is distributed across all instances of the role, so we can scaling-out it by scaling-out our web role. Just use 2 instances and tweak some code to show the current instance ID in the page, and have another try. Then we can see the cache can be retrieved even though it was added by another instance.   Consume Caching (Preview) Across Roles As I mentioned, the Caching (Preview) can be consumed by all other roles within the same cloud service. For example, let’s add another web role in our cloud solution and add the same code in its default page. In the Web.config we add the cache client to one enabled in the last role, by specifying its role name here. Then we start the solution locally and go to web role 1, specify the name and let it generate the email to us. Since there’s no cache for this name so it will take about 2 seconds but will save the email into cache. And then we go to web role 2 and specify the same name. Then you can see it retrieve the email saved by the web role 1 and returned back very quickly. Finally then we can upload our application to Windows Azure and test again. Make sure you had changed the cache cluster status storage account to the real azure account.   More Awesome Features As a in-memory distributed caching solution, the Caching (Preview) has some fancy features I would like to highlight here. The first one is the high availability support. This is the first time I have heard that a distributed cache support high availability. In the distributed cache world if a cache cluster was failed, the data it stored will be lost. This behavior was introduced by Memcached and is followed by almost all distributed cache productions. But Caching (Preview) provides high availability, which means you can specify if the named cache will be backup automatically. If yes then the data belongs to this named cache will be replicated on another role instance of this role. Then if one of the instance was failed the data can be retrieved from its backup instance. To enable the backup just open the Caching page in Visual Studio. In the named cache you want to enable backup, change the Backup Copies value from 0 to 1. The value of Backup Copies only for 0 and 1. “0” means no backup and no high availability while “1” means enabled high availability with backup the data into another instance. But by using the high availability feature there are something we need to make sure. Firstly the high availability does NOT means the data in cache will never be lost for any kind of failure. For example, if we have a role with cache enabled that has 10 instances, and 9 of them was failed, then most of the cached data will be lost since the primary and backup instance may failed together. But normally is will not be happened since MS guarantees that it will use the instance in the different fault domain for backup cache. Another one is that, enabling the backup means you store two copies of your data. For example if you think 100MB memory is OK for cache, but you need at least 200MB if you enabled backup. Besides the high availability, the Caching (Preview) support more features introduced in Windows Server AppFabric Caching than the Windows Azure Shared Caching. It supports local cache with notification. It also support absolute and slide window expiration types as well. And the Caching (Preview) also support the Memcached protocol as well. This means if you have an application based on Memcached, you can use Caching (Preview) without any code changes. What you need to do is to change the configuration of how you connect to the cache. Similar as the Windows Azure Shared Caching, MS also offers the out-of-box ASP.NET session provider and output cache provide on top of the Caching (Preview).   Summary Caching is very important component when we building a cloud-based application. In the June 2012 update MS provides a new cache solution named Caching (Preview). Different from the existing Windows Azure Shared Caching, Caching (Preview) runs the cache cluster within the role instances we have deployed to the cloud. It gives more control, more performance and more cost-effect. So now we have two caching solutions in Windows Azure, the Shared Caching and Caching (Preview). If you need a central cache service which can be used by many cloud services and web sites, then you have to use the Shared Caching. But if you only need a fast, near distributed cache, then you’d better use Caching (Preview).   Hope this helps, Shaun All documents and related graphics, codes are provided "AS IS" without warranty of any kind. Copyright © Shaun Ziyan Xu. This work is licensed under the Creative Commons License.

    Read the article

  • 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

    Read the article

  • Can't make AWUS036H work in Ubuntu 12.10

    - by sfrj
    I am using 64 bit Ubuntu 12.10. This is my kernel version: Linux 3.5.0-19-generic #30-Ubuntu SMP Tue Nov 13 17:48:01 UTC 2012 x86_64 x86_64 x86_64 GNU/Linux My wireless card is an AWUSU36H The first thing I do to install the driver is copy the driver from the CD to the Downloads folder. cd /media/me/AWUS036H/Drivers/RTL8187L/Unix (Linux)/Linux driver for kernel 2.6.X$ cp rtl8187_linux_26.1025.0328.2007.tar.gz ~/Downloads/ Then I extract the tar tar xvfz rtl8187_linux_26.1025.0328.2007.tar.gz I navigate into the extracted folder, and I try to follow the instructions in the Readme.txt cd rtl8187_linux_26.1025.0328.2007 This are the contents of the folder: drv.tar.gz makedrv stack.tar.gz wlan0rmv ieee80211 ReadMe.txt wlan0dhcp wlan0up ifcfg-wlan0 rtl8187 wlan0down wpa_supplicant-0.4.9 This is what the Readme.txt says: Release Date: 2006-02-09, ver 1.2^M RTL8187 Linux driver version 1.2^M ^M --This driver supports RealTek RTL8187 Wireless LAN driver for ^M Fedora Core 2/3/4/5, Debian 3.1, Mandrake 10.2/Mandriva 2006, ^M SUSE 9.3/10.1/10.2, Gentoo 3.1, etc.^M - Support Client mode for either infrastructure or adhoc mode^M - Support WEP and WPAPSK connection^M ^M < Component >^M The driver is composed of several parts:^M 1. Module source code^M stack.tar.gz^M drv.tar.gz^M ^M 2. Script ot build the modules^M makedrv^M ^M 3. Script to load/unload modules^M wlan0up^M wlan0down ^M ^M 4. Script and configuration for DHCP^M "ReadMe.txt" [readonly] 140 lines, 4590 characters So what I do know is extract both of the compressed files: sudo tar xvfz drv.tar.gz sudo tar xvfz stack.tar.gz This 2 commands will add some data to the folders ieee80211 and rtl8187 At this point I get lost, and I don't know what to do. If I go in each of this 2 folders and I run the sudo make command then I get errors like this one: sudo makemake -C /lib/modules/3.5.0-19-generic/build M=/home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187 modules make[1]: Entering directory `/usr/src/linux-headers-3.5.0-19-generic' CC [M] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187/r8187_core.o In file included from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187/r8187_core.c:64:0: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187/r8187.h:29:26: fatal error: linux/config.h: No such file or directory compilation terminated. make[2]: *** [/home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187/r8187_core.o] Error 1 make[1]: *** [_module_/home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187] Error 2 make[1]: Leaving directory `/usr/src/linux-headers-3.5.0-19-generic' make: *** [modules] Error 2 If I try to run any of the script ./makedrv that the instructions describe, then I also get an error: ~/Downloads/rtl8187_linux_26.1025.0328.2007$ sudo ./makedrv [sudo] password for me: ieee80211/ ieee80211/license ieee80211/ieee80211_crypt.c ieee80211/ieee80211_tx.c ieee80211/ieee80211_softmac.c ieee80211/ieee80211_softmac_wx.c ieee80211/ieee80211_module.c ieee80211/ieee80211_crypt_ccmp.c ieee80211/ieee80211_rx.c ieee80211/tags ieee80211/ieee80211_crypt_tkip.c ieee80211/Makefile ieee80211/readme ieee80211/.tmp_versions/ ieee80211/.tmp_versions/ieee80211-rtl.mod ieee80211/.tmp_versions/ieee80211_crypt_wep-rtl.mod ieee80211/.tmp_versions/ieee80211_crypt_tkip-rtl.mod ieee80211/.tmp_versions/ieee80211_crypt-rtl.mod ieee80211/.tmp_versions/ieee80211_crypt_ccmp-rtl.mod ieee80211/ieee80211_crypt_wep.c ieee80211/ieee80211.h ieee80211/ieee80211_wx.c ieee80211/ieee80211_crypt.h rtl8187/ rtl8187/license rtl8187/r8180_rtl8225z2.c rtl8187/r8180_rtl8225.h rtl8187/r8187_led.c rtl8187/r8180_93cx6.h rtl8187/r8180_wx.h rtl8187/r8180_hw.h rtl8187/copying rtl8187/r8187_led.h rtl8187/r8180_pm.h rtl8187/tags rtl8187/r8187.h rtl8187/Makefile rtl8187/r8180_rtl8225.c rtl8187/readme rtl8187/install rtl8187/.tmp_versions/ rtl8187/.tmp_versions/r8187.mod rtl8187/changes rtl8187/r8180_wx.c rtl8187/r8180_pm.c rtl8187/r8187_core.c rtl8187/r8180_93cx6.c rtl8187/authors rtl8187/ieee80211.h rtl8187/ieee80211_crypt.h rm -f *.mod.c *.mod *.o .*.cmd *.ko *~ rm -rf /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/tmp make -C /lib/modules/3.5.0-19-generic/build M=/home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211 modules make[1]: Entering directory `/usr/src/linux-headers-3.5.0-19-generic' CC [M] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.o In file included from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:17:0: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211.h:1019:24: error: field ‘ps_task’ has incomplete type /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c: In function ‘ieee80211_softmac_scan_wq’: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:421:2: warning: passing argument 2 of ‘queue_delayed_work’ from incompatible pointer type [enabled by default] In file included from include/linux/srcu.h:32:0, from include/linux/notifier.h:15, from /usr/src/linux-headers-3.5.0-19-generic/arch/x86/include/asm/uprobes.h:26, from include/linux/uprobes.h:35, from include/linux/mm_types.h:15, from include/linux/kmemcheck.h:4, from include/linux/skbuff.h:18, from include/linux/if_ether.h:134, from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211.h:26, from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:17: include/linux/workqueue.h:371:12: note: expected ‘struct delayed_work *’ but argument is of type ‘struct work_struct *’ /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c: In function ‘ieee80211_softmac_stop_scan’: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:495:3: warning: passing argument 1 of ‘cancel_delayed_work’ from incompatible pointer type [enabled by default] In file included from include/linux/srcu.h:32:0, from include/linux/notifier.h:15, from /usr/src/linux-headers-3.5.0-19-generic/arch/x86/include/asm/uprobes.h:26, from include/linux/uprobes.h:35, from include/linux/mm_types.h:15, from include/linux/kmemcheck.h:4, from include/linux/skbuff.h:18, from include/linux/if_ether.h:134, from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211.h:26, from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:17: include/linux/workqueue.h:410:20: note: expected ‘struct delayed_work *’ but argument is of type ‘struct work_struct *’ /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c: In function ‘ieee80211_associate_abort’: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:915:2: warning: passing argument 2 of ‘queue_delayed_work’ from incompatible pointer type [enabled by default] In file included from include/linux/srcu.h:32:0, from include/linux/notifier.h:15, from /usr/src/linux-headers-3.5.0-19-generic/arch/x86/include/asm/uprobes.h:26, from include/linux/uprobes.h:35, from include/linux/mm_types.h:15, from include/linux/kmemcheck.h:4, from include/linux/skbuff.h:18, from include/linux/if_ether.h:134, from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211.h:26, from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:17: include/linux/workqueue.h:371:12: note: expected ‘struct delayed_work *’ but argument is of type ‘struct work_struct *’ /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c: In function ‘ieee80211_rx_frame_softmac’: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:1527:3: error: implicit declaration of function ‘tasklet_schedule’ [-Werror=implicit-function-declaration] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c: In function ‘ieee80211_stop_protocol_rtl’: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2120:2: warning: passing argument 1 of ‘cancel_delayed_work’ from incompatible pointer type [enabled by default] In file included from include/linux/srcu.h:32:0, from include/linux/notifier.h:15, from /usr/src/linux-headers-3.5.0-19-generic/arch/x86/include/asm/uprobes.h:26, from include/linux/uprobes.h:35, from include/linux/mm_types.h:15, from include/linux/kmemcheck.h:4, from include/linux/skbuff.h:18, from include/linux/if_ether.h:134, from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211.h:26, from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:17: include/linux/workqueue.h:410:20: note: expected ‘struct delayed_work *’ but argument is of type ‘struct work_struct *’ /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c: In function ‘ieee80211_softmac_init’: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2229:78: error: macro "INIT_WORK" passed 3 arguments, but takes just 2 /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2229:2: error: ‘INIT_WORK’ undeclared (first use in this function) /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2229:2: note: each undeclared identifier is reported only once for each function it appears in /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2230:88: error: macro "INIT_WORK" passed 3 arguments, but takes just 2 /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2231:94: error: macro "INIT_WORK" passed 3 arguments, but takes just 2 /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2232:96: error: macro "INIT_WORK" passed 3 arguments, but takes just 2 /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2233:82: error: macro "INIT_WORK" passed 3 arguments, but takes just 2 /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2234:82: error: macro "INIT_WORK" passed 3 arguments, but takes just 2 /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2244:2: error: implicit declaration of function ‘tasklet_init’ [-Werror=implicit-function-declaration] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c: In function ‘ieee80211_softmac_free’: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2255:2: warning: passing argument 1 of ‘cancel_delayed_work’ from incompatible pointer type [enabled by default] In file included from include/linux/srcu.h:32:0, from include/linux/notifier.h:15, from /usr/src/linux-headers-3.5.0-19-generic/arch/x86/include/asm/uprobes.h:26, from include/linux/uprobes.h:35, from include/linux/mm_types.h:15, from include/linux/kmemcheck.h:4, from include/linux/skbuff.h:18, from include/linux/if_ether.h:134, from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211.h:26, from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:17: include/linux/workqueue.h:410:20: note: expected ‘struct delayed_work *’ but argument is of type ‘struct work_struct *’ /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c: In function ‘ieee80211_wpa_set_encryption’: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2489:3: error: implicit declaration of function ‘request_module’ [-Werror=implicit-function-declaration] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2518:3: error: implicit declaration of function ‘try_module_get’ [-Werror=implicit-function-declaration] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c: At top level: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2663:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2663:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2663:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2664:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2664:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2664:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2665:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2665:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2665:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2666:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2666:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2666:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2667:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2667:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2667:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2668:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2668:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2668:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2669:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2669:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2669:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2670:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2670:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2670:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2671:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2671:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2671:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2672:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2672:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2672:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2673:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2673:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2673:1: warning: parameter names (without types) in function declaration [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2674:1: warning: data definition has no type or storage class [enabled by default] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2674:1: warning: type defaults to ‘int’ in declaration of ‘EXPORT_SYMBOL’ [-Wimplicit-int] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:2674:1: warning: parameter names (without types) in function declaration [enabled by default] In file included from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.c:17:0: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211.h:1212:37: warning: ‘netdev_priv’ is static but used in inline function ‘ieee80211_priv’ which is not static [enabled by default] cc1: some warnings being treated as errors make[2]: *** [/home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211/ieee80211_softmac.o] Error 1 make[1]: *** [_module_/home/me/Downloads/rtl8187_linux_26.1025.0328.2007/ieee80211] Error 2 make[1]: Leaving directory `/usr/src/linux-headers-3.5.0-19-generic' make: *** [modules] Error 2 rm -f *.mod.c *.mod *.o .*.cmd *.ko *~ rm -rf /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187/tmp make -C /lib/modules/3.5.0-19-generic/build M=/home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187 modules make[1]: Entering directory `/usr/src/linux-headers-3.5.0-19-generic' CC [M] /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187/r8187_core.o In file included from /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187/r8187_core.c:64:0: /home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187/r8187.h:29:26: fatal error: linux/config.h: No such file or directory compilation terminated. make[2]: *** [/home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187/r8187_core.o] Error 1 make[1]: *** [_module_/home/me/Downloads/rtl8187_linux_26.1025.0328.2007/rtl8187] Error 2 make[1]: Leaving directory `/usr/src/linux-headers-3.5.0-19-generic' make: *** [modules] Error 2 Can somebody give me a hand finding out what I need to do to make my wifi card work? Update This is the output of the lsusb command lsusb Bus 003 Device 002: ID 147e:1000 Upek Biometric Touchchip/Touchstrip Fingerprint Sensor Bus 001 Device 001: ID 1d6b:0002 Linux Foundation 2.0 root hub Bus 002 Device 001: ID 1d6b:0002 Linux Foundation 2.0 root hub Bus 003 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub Bus 004 Device 001: ID 1d6b:0001 Linux Foundation 1.1 root hub

    Read the article

  • Metro: Introduction to CSS 3 Grid Layout

    - by Stephen.Walther
    The purpose of this blog post is to provide you with a quick introduction to the new W3C CSS 3 Grid Layout standard. You can use CSS Grid Layout in Metro style applications written with JavaScript to lay out the content of an HTML page. CSS Grid Layout provides you with all of the benefits of using HTML tables for layout without requiring you to actually use any HTML table elements. Doing Page Layouts without Tables Back in the 1990’s, if you wanted to create a fancy website, then you would use HTML tables for layout. For example, if you wanted to create a standard three-column page layout then you would create an HTML table with three columns like this: <table height="100%"> <tr> <td valign="top" width="300px" bgcolor="red"> Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column </td> <td valign="top" bgcolor="green"> Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column </td> <td valign="top" width="300px" bgcolor="blue"> Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column </td> </tr> </table> When the table above gets rendered out to a browser, you end up with the following three-column layout: The width of the left and right columns is fixed – the width of the middle column expands or contracts depending on the width of the browser. Sometime around the year 2005, everyone decided that using tables for layout was a bad idea. Instead of using tables for layout — it was collectively decided by the spirit of the Web — you should use Cascading Style Sheets instead. Why is using HTML tables for layout bad? Using tables for layout breaks the semantics of the TABLE element. A TABLE element should be used only for displaying tabular information such as train schedules or moon phases. Using tables for layout is bad for accessibility (The Web Content Accessibility Guidelines 1.0 is explicit about this) and using tables for layout is bad for separating content from layout (see http://CSSZenGarden.com). Post 2005, anyone who used HTML tables for layout were encouraged to hold their heads down in shame. That’s all well and good, but the problem with using CSS for layout is that it can be more difficult to work with CSS than HTML tables. For example, to achieve a standard three-column layout, you either need to use absolute positioning or floats. Here’s a three-column layout with floats: <style type="text/css"> #container { min-width: 800px; } #leftColumn { float: left; width: 300px; height: 100%; background-color:red; } #middleColumn { background-color:green; height: 100%; } #rightColumn { float: right; width: 300px; height: 100%; background-color:blue; } </style> <div id="container"> <div id="rightColumn"> Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column </div> <div id="leftColumn"> Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column </div> <div id="middleColumn"> Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column </div> </div> The page above contains four DIV elements: a container DIV which contains a leftColumn, middleColumn, and rightColumn DIV. The leftColumn DIV element is floated to the left and the rightColumn DIV element is floated to the right. Notice that the rightColumn DIV appears in the page before the middleColumn DIV – this unintuitive ordering is necessary to get the floats to work correctly (see http://stackoverflow.com/questions/533607/css-three-column-layout-problem). The page above (almost) works with the most recent versions of most browsers. For example, you get the correct three-column layout in both Firefox and Chrome: And the layout mostly works with Internet Explorer 9 except for the fact that for some strange reason the min-width doesn’t work so when you shrink the width of your browser, you can get the following unwanted layout: Notice how the middle column (the green column) bleeds to the left and right. People have solved these issues with more complicated CSS. For example, see: http://matthewjamestaylor.com/blog/holy-grail-no-quirks-mode.htm But, at this point, no one could argue that using CSS is easier or more intuitive than tables. It takes work to get a layout with CSS and we know that we could achieve the same layout more easily using HTML tables. Using CSS Grid Layout CSS Grid Layout is a new W3C standard which provides you with all of the benefits of using HTML tables for layout without the disadvantage of using an HTML TABLE element. In other words, CSS Grid Layout enables you to perform table layouts using pure Cascading Style Sheets. The CSS Grid Layout standard is still in a “Working Draft” state (it is not finalized) and it is located here: http://www.w3.org/TR/css3-grid-layout/ The CSS Grid Layout standard is only supported by Internet Explorer 10 and there are no signs that any browser other than Internet Explorer will support this standard in the near future. This means that it is only practical to take advantage of CSS Grid Layout when building Metro style applications with JavaScript. Here’s how you can create a standard three-column layout using a CSS Grid Layout: <!DOCTYPE html> <html> <head> <style type="text/css"> html, body, #container { height: 100%; padding: 0px; margin: 0px; } #container { display: -ms-grid; -ms-grid-columns: 300px auto 300px; -ms-grid-rows: 100%; } #leftColumn { -ms-grid-column: 1; background-color:red; } #middleColumn { -ms-grid-column: 2; background-color:green; } #rightColumn { -ms-grid-column: 3; background-color:blue; } </style> </head> <body> <div id="container"> <div id="leftColumn"> Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column </div> <div id="middleColumn"> Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column </div> <div id="rightColumn"> Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column </div> </div> </body> </html> When the page above is rendered in Internet Explorer 10, you get a standard three-column layout: The page above contains four DIV elements: a container DIV which contains a leftColumn DIV, middleColumn DIV, and rightColumn DIV. The container DIV is set to Grid display mode with the following CSS rule: #container { display: -ms-grid; -ms-grid-columns: 300px auto 300px; -ms-grid-rows: 100%; } The display property is set to the value “-ms-grid”. This property causes the container DIV to lay out its child elements in a grid. (Notice that you use “-ms-grid” instead of “grid”. The “-ms-“ prefix is used because the CSS Grid Layout standard is still preliminary. This implementation only works with IE10 and it might change before the final release.) The grid columns and rows are defined with the “-ms-grid-columns” and “-ms-grid-rows” properties. The style rule above creates a grid with three columns and one row. The left and right columns are fixed sized at 300 pixels. The middle column sizes automatically depending on the remaining space available. The leftColumn, middleColumn, and rightColumn DIVs are positioned within the container grid element with the following CSS rules: #leftColumn { -ms-grid-column: 1; background-color:red; } #middleColumn { -ms-grid-column: 2; background-color:green; } #rightColumn { -ms-grid-column: 3; background-color:blue; } The “-ms-grid-column” property is used to specify the column associated with the element selected by the style sheet selector. The leftColumn DIV is positioned in the first grid column, the middleColumn DIV is positioned in the second grid column, and the rightColumn DIV is positioned in the third grid column. I find using CSS Grid Layout to be just as intuitive as using an HTML table for layout. You define your columns and rows and then you position different elements within these columns and rows. Very straightforward. Creating Multiple Columns and Rows In the previous section, we created a super simple three-column layout. This layout contained only a single row. In this section, let’s create a slightly more complicated layout which contains more than one row: The following page contains a header row, a content row, and a footer row. The content row contains three columns: <!DOCTYPE html> <html> <head> <style type="text/css"> html, body, #container { height: 100%; padding: 0px; margin: 0px; } #container { display: -ms-grid; -ms-grid-columns: 300px auto 300px; -ms-grid-rows: 100px 1fr 100px; } #header { -ms-grid-column: 1; -ms-grid-column-span: 3; -ms-grid-row: 1; background-color: yellow; } #leftColumn { -ms-grid-column: 1; -ms-grid-row: 2; background-color:red; } #middleColumn { -ms-grid-column: 2; -ms-grid-row: 2; background-color:green; } #rightColumn { -ms-grid-column: 3; -ms-grid-row: 2; background-color:blue; } #footer { -ms-grid-column: 1; -ms-grid-column-span: 3; -ms-grid-row: 3; background-color: orange; } </style> </head> <body> <div id="container"> <div id="header"> Header, Header, Header </div> <div id="leftColumn"> Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column, Left Column </div> <div id="middleColumn"> Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column, Middle Column </div> <div id="rightColumn"> Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column, Right Column </div> <div id="footer"> Footer, Footer, Footer </div> </div> </body> </html> In the page above, the grid layout is created with the following rule which creates a grid with three rows and three columns: #container { display: -ms-grid; -ms-grid-columns: 300px auto 300px; -ms-grid-rows: 100px 1fr 100px; } The header is created with the following rule: #header { -ms-grid-column: 1; -ms-grid-column-span: 3; -ms-grid-row: 1; background-color: yellow; } The header is positioned in column 1 and row 1. Furthermore, notice that the “-ms-grid-column-span” property is used to span the header across three columns. CSS Grid Layout and Fractional Units When you use CSS Grid Layout, you can take advantage of fractional units. Fractional units provide you with an easy way of dividing up remaining space in a page. Imagine, for example, that you want to create a three-column page layout. You want the size of the first column to be fixed at 200 pixels and you want to divide the remaining space among the remaining three columns. The width of the second column is equal to the combined width of the third and fourth columns. The following CSS rule creates four columns with the desired widths: #container { display: -ms-grid; -ms-grid-columns: 200px 2fr 1fr 1fr; -ms-grid-rows: 1fr; } The fr unit represents a fraction. The grid above contains four columns. The second column is two times the size (2fr) of the third (1fr) and fourth (1fr) columns. When you use the fractional unit, the remaining space is divided up using fractional amounts. Notice that the single row is set to a height of 1fr. The single grid row gobbles up the entire vertical space. Here’s the entire HTML page: <!DOCTYPE html> <html> <head> <style type="text/css"> html, body, #container { height: 100%; padding: 0px; margin: 0px; } #container { display: -ms-grid; -ms-grid-columns: 200px 2fr 1fr 1fr; -ms-grid-rows: 1fr; } #firstColumn { -ms-grid-column: 1; background-color:red; } #secondColumn { -ms-grid-column: 2; background-color:green; } #thirdColumn { -ms-grid-column: 3; background-color:blue; } #fourthColumn { -ms-grid-column: 4; background-color:orange; } </style> </head> <body> <div id="container"> <div id="firstColumn"> First Column, First Column, First Column </div> <div id="secondColumn"> Second Column, Second Column, Second Column </div> <div id="thirdColumn"> Third Column, Third Column, Third Column </div> <div id="fourthColumn"> Fourth Column, Fourth Column, Fourth Column </div> </div> </body> </html>   Summary There is more in the CSS 3 Grid Layout standard than discussed in this blog post. My goal was to describe the basics. If you want to learn more than you can read through the entire standard at http://www.w3.org/TR/css3-grid-layout/ In this blog post, I described some of the difficulties that you might encounter when attempting to replace HTML tables with Cascading Style Sheets when laying out a web page. I explained how you can take advantage of the CSS 3 Grid Layout standard to avoid these problems when building Metro style applications using JavaScript. CSS 3 Grid Layout provides you with all of the benefits of using HTML tables for laying out a page without requiring you to use HTML table elements.

    Read the article

  • Transformation of Product Management in Telecommunications for Rapid Launch of Next Generation Products

    - by raul.goycoolea
    @font-face { font-family: "Arial"; }@font-face { font-family: "Courier New"; }@font-face { font-family: "Wingdings"; }@font-face { font-family: "Cambria"; }p.MsoNormal, li.MsoNormal, div.MsoNormal { margin: 0cm 0cm 0.0001pt; font-size: 12pt; font-family: "Times New Roman"; }a:link, span.MsoHyperlink { color: blue; text-decoration: underline; }a:visited, span.MsoHyperlinkFollowed { color: purple; text-decoration: underline; }p.MsoListParagraph, li.MsoListParagraph, div.MsoListParagraph { margin: 0cm 0cm 0.0001pt 36pt; font-size: 12pt; font-family: "Times New Roman"; }p.MsoListParagraphCxSpFirst, li.MsoListParagraphCxSpFirst, div.MsoListParagraphCxSpFirst { margin: 0cm 0cm 0.0001pt 36pt; font-size: 12pt; font-family: "Times New Roman"; }p.MsoListParagraphCxSpMiddle, li.MsoListParagraphCxSpMiddle, div.MsoListParagraphCxSpMiddle { margin: 0cm 0cm 0.0001pt 36pt; font-size: 12pt; font-family: "Times New Roman"; }p.MsoListParagraphCxSpLast, li.MsoListParagraphCxSpLast, div.MsoListParagraphCxSpLast { margin: 0cm 0cm 0.0001pt 36pt; font-size: 12pt; font-family: "Times New Roman"; }div.Section1 { page: Section1; }ol { margin-bottom: 0cm; }ul { margin-bottom: 0cm; } The Telecom industry continues to evolve through disruptive products, uncertain markets, shorter product lifecycles and convergence of technologies. Today’s market has moved from network centric to consumer centric and focuses primarily on the customer experience. It has resulted in several product management challenges such as an increased complexity and volume of offerings, creating product variants, accelerating time-to-market, ability to provide multiple product views for varied stakeholders, leveraging OSS intelligence to BSS layer, product co-creation and increasing audit and security concerns for service providers. The document discusses how enterprise product management enabled by PLM-based product catalogue solutions helps to launch next generation products rapidly in the context of the Telecommunication Industry.   1.0.       Introduction   Figure 1: Business Scenario   Modern business demands the launch of complex products in a very short timeframe and effecting changes in the price plan faster without IT intervention. One of the key transformation initiatives companies are focusing on is in the area of product management transformation and operational efficiency improvement. As part of these initiatives, companies are investing in best- in-class COTs-based Product Management solutions developed on industry-wide standards.   The new COTs packages are planned to integrate with existing or new B/OSS systems to provide a strategic end-to-end agile solution for reduced time-to-market and order journey time. In addition, system rationalization is being undertaken to phase out legacy systems and migrate to strategic systems.   2.0.       An Overview of Product Management in Telecom   Product data in telecom is multi- dimensional and difficult to manage. It increased significantly due to the complexity of the product, product offerings on the converged network, increased volume of offerings, bundled offering structures and ever increasing regulatory requirements.   In addition, the shrinking product lifecycle in telecom makes it difficult to manage the dynamic product data. Mergers and acquisitions coupled with organic growth pose major challenges in product portfolio management. It is a roadblock in the journey towards becoming an agile organization.       Figure 2: Complexity in Product Management   Network Technology’ is the new dimension in telecom product management where the same products are realized through different networks i.e., Soiled network to Converged network. Consequently, the product solution is different.     Figure 3: Current Scenario - Pain Points in Product Management   The major business implications arising out of the current scenario are slow time-to-market and an inefficient process that affects innovation.   3.0. Transformation of Next Generation Product Management   Companies must focus on their Product Management Transformation Journey in the areas of:   ·       Management of single truth of product information across the organization/geographies which is currently managed in heterogeneous systems   ·       Management of the Intellectual Property (IP) on the product concept and partnership in the design of discrete components to integrate into the system   ·       Leveraging structured and unstructured product data within the extended enterprise to extract consumer insights and drive innovation   ·       Management of effective operational separation to comply with regulatory bodies   ·       Reuse of existing designs and add relevant features such as value-added services to enable effective product bundling     Figure 4: Next generation needs   PLM-based Enterprise Product Catalogue solutions efficiently address the above requirements and act as an enabler towards product management transformation and rapid product launch.   4.0. PLM-based Enterprise Product Management     Figure 5: PLM-based Enterprise Product Mastering   Enterprise Product Management (EPM) enables the business to manage complex product attributes of data in complex environments. Product Mastering helps create a 'single view' of the product by creating a business-driven, IT-supported environment where a global 'single truth record' is created, managed and reused.   4.1 The Business Case for Telco PLM-based solutions for Enterprise Product Management   ·       Telco PLM-based Product Mastering solutions provide a centralized authoring environment for product definition and control of all product data and rules   ·       PLM packages are designed to support multiple perspectives of product data (ordering perspective, billing perspective, provisioning perspective)   ·       Maintains relationships/links between different elements of the entire product definition   ·       Telco PLM packages are specialized in next generation lifecycle management requirements of products such as revision and state management, test and release management, role management and impact analysis)   ·       Takes into consideration all aspects of OSS product requirements compared to CRM product catalogue solutions where the product data managed is mostly order oriented and transactional     ·       New breed of Telco PLM packages are designed with 'open' standards such as SID and eTOM. They are interoperable, support integration frameworks such as subscription and notification.   ·       Telco PLM packages have developed good collaboration frameworks to integrate suppliers and partners into the product development value chain   4.2 Various Architectures/Approaches for Product Mastering using Telco PLM systems   4. 2.a Single Central Product Management (Mastering) Approach   Figure 6: Single Central Product Management (Master) Approach       This approach is implemented across verticals such as aerospace and automotive. It focuses on a physically centralized product master to which other sources are dependent on. The product definition data (Product bundles, service bundles, price plans, offers and discounts, product configuration rules and market campaigns) is created and maintained physically in a centralized environment. In addition, the product definition/authoring environment is centralized. The existing legacy product definition data available in CRM product catalogue, billing catalogue and the legacy product catalogue is migrated to the centralized PLM-based Enterprise Product Management solution.   Architectural changes must be made in the existing business landscape of applications to create and revise data because the applications have to refer to the central repository for approvals and validation of product configurations. It is achieved by modifying how the applications write data or how the applications can be adapted to use the rules to be managed and published.   Complete product configuration validation will be done in enterprise / central product catalogue and final configuration will be sent to the B/OSS system through the SOA compliant product distribution architecture. The approach/architecture enables greater control in terms of product data management and product data governance.   4.2.b Federated Product Management (Mastering) Architecture     Figure 7: Federated Product Management (Mastering) Architecture   In the federated product mastering approach, the basic unique product definition data (product id, description product hierarchy, basic price plans and simple product design rules) will be centrally created and will be maintained. And, the advanced product definition (Product bundling, promotions, offers & discount plans) will be created in respective down stream OSS systems. The advanced product definition (Product bundling, promotions, offers and discount plans) will be created in respective downstream OSS systems.   For example, basic product definitions such as attributes, product hierarchy and basic price plans will be created and maintained in Enterprise/Central product reference catalogue and distributed to downstream OSS systems. Respective downstream OSS systems build product bundles, promotions, advanced price plans over the basic product definition and master the advanced product definition. Central reference database accesses the respective other source product master data and assembles a point-in-time consolidated view of the product. The approach is typically adapted in some merger and acquisition scenarios where there is a low probability of a central physical authority managing the data. In addition, the migration effort in this case is minimal and there are no big architectural changes to the organization application landscape. However, this approach will not result in better product data management and data governance.   5.0 Customer Scenario – Before EPC deployment   A leading global telecommunications service provider wanted to launch a quad play and triple play service offering in the shortest possible lead time. The service provider was offering Broadband and VoIP services to customers. The company wanted to reuse a majority of the Broadband services and price plans and bundle them with new wireless and IPTV services for quad play and triple play. The challenges in launching the new service offerings were:       Figure 8: Triple Play Plan   ·       Broadband product data was stored in multiple product catalogues (CRM catalogue, Billing catalogue, spread sheets)   ·       Product managers spent a lot of time performing tasks involving duplication or re-keying of data. Manual effort caused errors, cost and time over-runs.   ·       No effective product and price data governance mechanism. Price change issues arising from the lack of data consistency across systems resulted in leakage of customer value and revenue.   ·       Product data had re-usability issues and was not in a structured format. It resulted in uncontrolled product portfolio creation and product management issues.   ·       Lack of enterprise product model resulted into product distribution challenges and thus delays in product launch.   ·       Designers are constrained by existing legacy product management solutions to model product/service requirements and product configuration rules such as upgrading, downgrading and cross selling.    5.1 Customer Scenario - After EPC deployment     Figure 9: SOA-based end-to-end EPC Solution   The company deployed PLM-based Enterprise Product Catalogue solutions to launch quad play service after evaluating various product catalogues. The broadband product offering, service and price data were migrated to the new system, and the product and price plan hierarchy for new offerings were created using the entities defined in the Enterprise Product Model. Supplier product catalogue data such as routers and set up boxes were loaded onto the new solution through SOA-based web service. Price plans and configuration rules were built in the new system. The validated final product configurations were extracted from the product catalogue in a SID format and were distributed to the downstream B/OSS systems through exposed SOA-based web services. The transformations required for the B/OSS system were handled using the transformation layer as part of the solution.   6.0 How PLM enabled Product Management Transformation         Figure 10: Product Management Transformation     PLM-based Product Catalogue Solution helped the customer reduce the product launch cycle time by 30% and enable transformation of Product Management for next generation services.   7.0 Conclusion   On the one hand, the telecom industry is undergoing changes due to disruptions, uncertain product markets and increased complexity of products. On the other hand, the ARPU is decreasing year-on-year. Communications Service Providers are embarking on convergence, bundled service offerings, flexibility to cross-sell and up-sell, introduce new value-added services, leverage Web 2.0 concepts and network capabilities. Consequently, large scale IT transformation initiatives to improve their ARPU supporting network and business transformations are a business imperative. Product Management has become a focus area. Companies are investing in best-in- class COTS solutions to reduce time-to-market, ensure rapid service delivery and improve operational efficiency. An efficient PLM-based enterprise product mastering solution plays a key role in achieving zero touch automation and rapid product launch.   References:   1.     Preston G.Smith, Donald G.Reineristsem, Van Nostrand Reinhold “Developing Products in Half the time”.   2.     John G. Innes, "Achieving Successful Product Change", Pitman Publishing.   3.     D T Pham and R M Setchi (16th Jan, 2001) "Authoring environment for documentation development" University of Wales Cardiff, U.K., Proceedings on Institution of Mechanical Engineers, Vol. 215, Part B.   4.     Oracle Product Hub for Communications:   http://www.oracle.com/us/products/applications/master-data-management/product-hub-082059.html  

    Read the article

  • Oracle BI Server Modeling, Part 1- Designing a Query Factory

    - by bob.ertl(at)oracle.com
      Welcome to Oracle BI Development's BI Foundation blog, focused on helping you get the most value from your Oracle Business Intelligence Enterprise Edition (BI EE) platform deployments.  In my first series of posts, I plan to show developers the concepts and best practices for modeling in the Common Enterprise Information Model (CEIM), the semantic layer of Oracle BI EE.  In this segment, I will lay the groundwork for the modeling concepts.  First, I will cover the big picture of how the BI Server fits into the system, and how the CEIM controls the query processing. Oracle BI EE Query Cycle The purpose of the Oracle BI Server is to bridge the gap between the presentation services and the data sources.  There are typically a variety of data sources in a variety of technologies: relational, normalized transaction systems; relational star-schema data warehouses and marts; multidimensional analytic cubes and financial applications; flat files, Excel files, XML files, and so on. Business datasets can reside in a single type of source, or, most of the time, are spread across various types of sources. Presentation services users are generally business people who need to be able to query that set of sources without any knowledge of technologies, schemas, or how sources are organized in their company. They think of business analysis in terms of measures with specific calculations, hierarchical dimensions for breaking those measures down, and detailed reports of the business transactions themselves.  Most of them create queries without knowing it, by picking a dashboard page and some filters.  Others create their own analysis by selecting metrics and dimensional attributes, and possibly creating additional calculations. The BI Server bridges that gap from simple business terms to technical physical queries by exposing just the business focused measures and dimensional attributes that business people can use in their analyses and dashboards.   After they make their selections and start the analysis, the BI Server plans the best way to query the data sources, writes the optimized sequence of physical queries to those sources, post-processes the results, and presents them to the client as a single result set suitable for tables, pivots and charts. The CEIM is a model that controls the processing of the BI Server.  It provides the subject areas that presentation services exposes for business users to select simplified metrics and dimensional attributes for their analysis.  It models the mappings to the physical data access, the calculations and logical transformations, and the data access security rules.  The CEIM consists of metadata stored in the repository, authored by developers using the Administration Tool client.     Presentation services and other query clients create their queries in BI EE's SQL-92 language, called Logical SQL or LSQL.  The API simply uses ODBC or JDBC to pass the query to the BI Server.  Presentation services writes the LSQL query in terms of the simplified objects presented to the users.  The BI Server creates a query plan, and rewrites the LSQL into fully-detailed SQL or other languages suitable for querying the physical sources.  For example, the LSQL on the left below was rewritten into the physical SQL for an Oracle 11g database on the right. Logical SQL   Physical SQL SELECT "D0 Time"."T02 Per Name Month" saw_0, "D4 Product"."P01  Product" saw_1, "F2 Units"."2-01  Billed Qty  (Sum All)" saw_2 FROM "Sample Sales" ORDER BY saw_0, saw_1       WITH SAWITH0 AS ( select T986.Per_Name_Month as c1, T879.Prod_Dsc as c2,      sum(T835.Units) as c3, T879.Prod_Key as c4 from      Product T879 /* A05 Product */ ,      Time_Mth T986 /* A08 Time Mth */ ,      FactsRev T835 /* A11 Revenue (Billed Time Join) */ where ( T835.Prod_Key = T879.Prod_Key and T835.Bill_Mth = T986.Row_Wid) group by T879.Prod_Dsc, T879.Prod_Key, T986.Per_Name_Month ) select SAWITH0.c1 as c1, SAWITH0.c2 as c2, SAWITH0.c3 as c3 from SAWITH0 order by c1, c2   Probably everybody reading this blog can write SQL or MDX.  However, the trick in designing the CEIM is that you are modeling a query-generation factory.  Rather than hand-crafting individual queries, you model behavior and relationships, thus configuring the BI Server machinery to manufacture millions of different queries in response to random user requests.  This mass production requires a different mindset and approach than when you are designing individual SQL statements in tools such as Oracle SQL Developer, Oracle Hyperion Interactive Reporting (formerly Brio), or Oracle BI Publisher.   The Structure of the Common Enterprise Information Model (CEIM) The CEIM has a unique structure specifically for modeling the relationships and behaviors that fill the gap from logical user requests to physical data source queries and back to the result.  The model divides the functionality into three specialized layers, called Presentation, Business Model and Mapping, and Physical, as shown below. Presentation services clients can generally only see the presentation layer, and the objects in the presentation layer are normally the only ones used in the LSQL request.  When a request comes into the BI Server from presentation services or another client, the relationships and objects in the model allow the BI Server to select the appropriate data sources, create a query plan, and generate the physical queries.  That's the left to right flow in the diagram below.  When the results come back from the data source queries, the right to left relationships in the model show how to transform the results and perform any final calculations and functions that could not be pushed down to the databases.   Business Model Think of the business model as the heart of the CEIM you are designing.  This is where you define the analytic behavior seen by the users, and the superset library of metric and dimension objects available to the user community as a whole.  It also provides the baseline business-friendly names and user-readable dictionary.  For these reasons, it is often called the "logical" model--it is a virtual database schema that persists no data, but can be queried as if it is a database. The business model always has a dimensional shape (more on this in future posts), and its simple shape and terminology hides the complexity of the source data models. Besides hiding complexity and normalizing terminology, this layer adds most of the analytic value, as well.  This is where you define the rich, dimensional behavior of the metrics and complex business calculations, as well as the conformed dimensions and hierarchies.  It contributes to the ease of use for business users, since the dimensional metric definitions apply in any context of filters and drill-downs, and the conformed dimensions enable dashboard-wide filters and guided analysis links that bring context along from one page to the next.  The conformed dimensions also provide a key to hiding the complexity of many sources, including federation of different databases, behind the simple business model. Note that the expression language in this layer is LSQL, so that any expression can be rewritten into any data source's query language at run time.  This is important for federation, where a given logical object can map to several different physical objects in different databases.  It is also important to portability of the CEIM to different database brands, which is a key requirement for Oracle's BI Applications products. Your requirements process with your user community will mostly affect the business model.  This is where you will define most of the things they specifically ask for, such as metric definitions.  For this reason, many of the best-practice methodologies of our consulting partners start with the high-level definition of this layer. Physical Model The physical model connects the business model that meets your users' requirements to the reality of the data sources you have available. In the query factory analogy, think of the physical layer as the bill of materials for generating physical queries.  Every schema, table, column, join, cube, hierarchy, etc., that will appear in any physical query manufactured at run time must be modeled here at design time. Each physical data source will have its own physical model, or "database" object in the CEIM.  The shape of each physical model matches the shape of its physical source.  In other words, if the source is normalized relational, the physical model will mimic that normalized shape.  If it is a hypercube, the physical model will have a hypercube shape.  If it is a flat file, it will have a denormalized tabular shape. To aid in query optimization, the physical layer also tracks the specifics of the database brand and release.  This allows the BI Server to make the most of each physical source's distinct capabilities, writing queries in its syntax, and using its specific functions. This allows the BI Server to push processing work as deep as possible into the physical source, which minimizes data movement and takes full advantage of the database's own optimizer.  For most data sources, native APIs are used to further optimize performance and functionality. The value of having a distinct separation between the logical (business) and physical models is encapsulation of the physical characteristics.  This encapsulation is another enabler of packaged BI applications and federation.  It is also key to hiding the complex shapes and relationships in the physical sources from the end users.  Consider a routine drill-down in the business model: physically, it can require a drill-through where the first query is MDX to a multidimensional cube, followed by the drill-down query in SQL to a normalized relational database.  The only difference from the user's point of view is that the 2nd query added a more detailed dimension level column - everything else was the same. Mappings Within the Business Model and Mapping Layer, the mappings provide the binding from each logical column and join in the dimensional business model, to each of the objects that can provide its data in the physical layer.  When there is more than one option for a physical source, rules in the mappings are applied to the query context to determine which of the data sources should be hit, and how to combine their results if more than one is used.  These rules specify aggregate navigation, vertical partitioning (fragmentation), and horizontal partitioning, any of which can be federated across multiple, heterogeneous sources.  These mappings are usually the most sophisticated part of the CEIM. Presentation You might think of the presentation layer as a set of very simple relational-like views into the business model.  Over ODBC/JDBC, they present a relational catalog consisting of databases, tables and columns.  For business users, presentation services interprets these as subject areas, folders and columns, respectively.  (Note that in 10g, subject areas were called presentation catalogs in the CEIM.  In this blog, I will stick to 11g terminology.)  Generally speaking, presentation services and other clients can query only these objects (there are exceptions for certain clients such as BI Publisher and Essbase Studio). The purpose of the presentation layer is to specialize the business model for different categories of users.  Based on a user's role, they will be restricted to specific subject areas, tables and columns for security.  The breakdown of the model into multiple subject areas organizes the content for users, and subjects superfluous to a particular business role can be hidden from that set of users.  Customized names and descriptions can be used to override the business model names for a specific audience.  Variables in the object names can be used for localization. For these reasons, you are better off thinking of the tables in the presentation layer as folders than as strict relational tables.  The real semantics of tables and how they function is in the business model, and any grouping of columns can be included in any table in the presentation layer.  In 11g, an LSQL query can also span multiple presentation subject areas, as long as they map to the same business model. Other Model Objects There are some objects that apply to multiple layers.  These include security-related objects, such as application roles, users, data filters, and query limits (governors).  There are also variables you can use in parameters and expressions, and initialization blocks for loading their initial values on a static or user session basis.  Finally, there are Multi-User Development (MUD) projects for developers to check out units of work, and objects for the marketing feature used by our packaged customer relationship management (CRM) software.   The Query Factory At this point, you should have a grasp on the query factory concept.  When developing the CEIM model, you are configuring the BI Server to automatically manufacture millions of queries in response to random user requests. You do this by defining the analytic behavior in the business model, mapping that to the physical data sources, and exposing it through the presentation layer's role-based subject areas. While configuring mass production requires a different mindset than when you hand-craft individual SQL or MDX statements, it builds on the modeling and query concepts you already understand. The following posts in this series will walk through the CEIM modeling concepts and best practices in detail.  We will initially review dimensional concepts so you can understand the business model, and then present a pattern-based approach to learning the mappings from a variety of physical schema shapes and deployments to the dimensional model.  Along the way, we will also present the dimensional calculation template, and learn how to configure the many additivity patterns.

    Read the article

  • Microsoft Declares the Future of ASP.NET is Web API

    - by sbwalker
    Sitting on a plane on my way home from Tech Ed 2012 in Orlando, I thought it would be a good time to jot down some key takeaways from this year’s conference. Some of these items I have known since the Microsoft MVP Summit which occurred in Redmond in late February ( but due to NDA restrictions I could not share them with the developer community at large ) and some of them are a result of insightful conversations with a wide variety of industry insiders and Microsoft employees at the conference. First, let’s travel back in time 4 years to the Microsoft MVP Summit in 2008. Microsoft was facing some heat from market newcomer Ruby on Rails and responded with a new web development framework of its own, ASP.NET MVC. At the Summit they estimated that MVC would only be applicable for ~10% of all new web development projects. Based on that prediction I questioned why they were investing such considerable resources for such a relative edge case, but my guess is that they felt it was an important edge case at the time as some of the more vocal .NET evangelists as well as some very high profile start-ups ( ie. Twitter ) had publicly announced their intent to use Rails. Microsoft made a lot of noise about MVC. In fact, they focused so much of their messaging and marketing hype around MVC that it appeared that WebForms was essentially dead. Yes, it may have been true that Microsoft continued to invest in WebForms, but from an outside perspective it really appeared that MVC was the only framework getting any real attention. As a result, MVC started to gain market share. An inside source at Microsoft told me that MVC usage has grown at a rate of about 5% per year and now sits at ~30%. Essentially by focusing so much marketing effort on MVC, Microsoft actually created a larger market demand for it.  This is because in the Microsoft ecosystem there is somewhat of a bandwagon mentality amongst developers. If Microsoft spends a lot of time talking about a specific technology, developers get the perception that it must be really important. So rather than choosing the right tool for the job, they often choose the tool with the most marketing hype and then try to sell it to the customer. In 2010, I blogged about the fact that MVC did not make any business sense for the DotNetNuke platform. This was because our ecosystem relied on third party extensions which were dependent on the WebForms model. If we migrated the core to MVC it would mean that all of the third party extensions would no longer be compatible, which would be an irresponsible business decision for us to make at the expense of our users and customers. However, this did not stop the debate from continuing to occur in our ecosystem. Clearly some developers had drunk Microsoft’s Kool-Aid about MVC and were of the mindset, to paraphrase an old Scottish saying, “If its not MVC, it’s crap”. Now, this is a rather ignorant position to take as most of the benefits of MVC can be achieved in WebForms with solid architecture and responsible coding practices. Clean separation of concerns, unit testing, and direct control over page output are all possible in the WebForms model – it just requires diligence and discipline. So over the past few years some horror stories have begun to bubble to the surface of software development projects focused on ground-up rewrites of web applications for the sole purpose of migrating from WebForms to MVC. These large scale rewrites were typically initiated by engineering teams with only a single argument driving the business decision, that Microsoft was promoting MVC as “the future”. These ill-fated rewrites offered no benefit to end users or customers and in fact resulted in a less stable, less scalable and more complicated systems – basically taking one step forward and two full steps back. A case in point is the announcement earlier this week that a popular open source .NET CMS provider has decided to pull the plug on their new MVC product which has been under active development for more than 18 months and revert back to WebForms. The availability of multiple server-side development models has deeply fragmented the Microsoft developer community. Some folks like to compare it to the age-old VB vs. C# language debate. However, the VB vs. C# language debate was ultimately more of a religious war because at least the two dominant programming languages were compatible with one another and could be used interchangeably. The issue with WebForms vs. MVC is much more challenging. This is because the messaging from Microsoft has positioned the two solutions as being incompatible with one another and as a result web developers feel like they are forced to choose one path or another. Yes, it is true that it has always been technically possible to use WebForms and MVC in the same project, but the tooling support has always made this feel “dirty”. The fragmentation has also made it difficult to attract newcomers as the perceived barrier to entry for learning ASP.NET has become higher. As a result many new software developers entering the market are gravitating to environments where the development model seems more simple and intuitive ( ie. PHP or Ruby ). At the same time that the Web Platform team was busy promoting ASP.NET MVC, the Microsoft Office team has been promoting Sharepoint as a platform for building internal enterprise web applications. Sharepoint has great penetration in the enterprise and over time has been enhanced with improved extensibility capabilities for software developers. But, like many other mature enterprise ASP.NET web applications, it is built on the WebForms development model. Similar to DotNetNuke, Sharepoint leverages a rich third party ecosystem for both generic web controls and more specialized WebParts – both of which rely on WebForms. So basically this resulted in a situation where the Web Platform group had headed off in one direction and the Office team had gone in another direction, and the end customer was stuck in the middle trying to figure out what to do with their existing investments in Microsoft technology. It really emphasized the perception that the left hand was not speaking to the right hand, as strategically speaking there did not seem to be any high level plan from Microsoft to ensure consistency and continuity across the different product lines. With the introduction of ASP.NET MVC, it also made some of the third party control vendors scratch their heads, and wonder what the heck Microsoft was thinking. The original value proposition of ASP.NET over Classic ASP was the ability for web developers to emulate the highly productive desktop development model by using abstract components for creating rich, interactive web interfaces. Web control vendors like Telerik, Infragistics, DevExpress, and ComponentArt had all built sizable businesses offering powerful user interface components to WebForms developers. And even after MVC was introduced these vendors continued to improve their products, offering greater productivity and a superior user experience via AJAX to what was possible in MVC. And since many developers were comfortable and satisfied with these third party solutions, the demand remained strong and the third party web control market continued to prosper despite the availability of MVC. While all of this was going on in the Microsoft ecosystem, there has also been a fundamental shift in the general software development industry. Driven by the explosion of Internet-enabled devices, the focus has now centered on service-oriented architecture (SOA). Service-oriented architecture is all about defining a public API for your product that any client can consume; whether it’s a native application running on a smart phone or tablet, a web browser taking advantage of HTML5 and Javascript, or a rich desktop application running on a PC. REST-based services which utilize the less verbose characteristics of JSON as a transport mechanism, have become the preferred approach over older, more bloated SOAP-based techniques. SOA also has the benefit of producing a cross-platform API, as every major technology stack is able to interact with standard REST-based web services. And for web applications, more and more developers are turning to robust Javascript libraries like JQuery and Knockout for browser-based client-side development techniques for calling web services and rendering content to end users. In fact, traditional server-side page rendering has largely fallen out of favor, resulting in decreased demand for server-side frameworks like Ruby on Rails, WebForms, and (gasp) MVC. In response to these new industry trends, Microsoft did what it always does – it immediately poured some resources into developing a solution which will ensure they remain relevant and competitive in the web space. This work culminated in a new framework which was branded as Web API. It is convention-based and designed to embrace native HTTP standards without copious layers of abstraction. This framework is designed to be the ultimate replacement for both the REST aspects of WCF and ASP.NET MVC Web Services. And since it was developed out of band with a dependency only on ASP.NET 4.0, it means that it can be used immediately in a variety of production scenarios. So at Tech Ed 2012 it was made abundantly clear in numerous sessions that Microsoft views Web API as the “Future of ASP.NET”. In fact, one Microsoft PM even went as far as to say that if we look 3-4 years into the future, that all ASP.NET web applications will be developed using the Web API approach. This is a fairly bold prediction and clearly telegraphs where Microsoft plans to allocate its resources going forward. Currently Web API is being delivered as part of the MVC4 package, but this is only temporary for the sake of convenience. It also sounds like there are still internal discussions going on in terms of how to brand the various aspects of ASP.NET going forward – perhaps the moniker of “ASP.NET Web Stack” coined a couple years ago by Scott Hanselman and utilized as part of the open source release of ASP.NET bits on Codeplex a few months back will eventually stick. Web API is being positioned as the unification of ASP.NET – the glue that is able to pull this fragmented mess back together again. The  “One ASP.NET” strategy will promote the use of all frameworks - WebForms, MVC, and Web API, even within the same web project. Basically the message is utilize the appropriate aspects of each framework to solve your business problems. Instead of navigating developers to a fork in the road, the plan is to educate them that “hybrid” applications are a great strategy for delivering solutions to customers. In addition, the service-oriented approach coupled with client-side development promoted by Web API can effectively be used in both WebForms and MVC applications. So this means it is also relevant to application platforms like DotNetNuke and Sharepoint, which means that it starts to create a unified development strategy across all ASP.NET product lines once again. And so what about MVC? There have actually been rumors floated that MVC has reached a stage of maturity where, similar to WebForms, it will be treated more as a maintenance product line going forward ( MVC4 may in fact be the last significant iteration of this framework ). This may sound alarming to some folks who have recently adopted MVC but it really shouldn’t, as both WebForms and MVC will continue to play a vital role in delivering solutions to customers. They will just not be the primary area where Microsoft is spending the majority of its R&D resources. That distinction will obviously go to Web API. And when the question comes up of why not enhance MVC to make it work with Web API, you must take a step back and look at this from the higher level to see that it really makes no sense. MVC is a server-side page compositing framework; whereas, Web API promotes client-side page compositing with a heavy focus on web services. In order to make MVC work well with Web API, would require a complete rewrite of MVC and at the end of the day, there would be no upgrade path for existing MVC applications. So it really does not make much business sense. So what does this have to do with DotNetNuke? Well, around 8-12 months ago we recognized the software industry trends towards web services and client-side development. We decided to utilize a “hybrid” model which would provide compatibility for existing modules while at the same time provide a bridge for developers who wanted to utilize more modern web techniques. Customers who like the productivity and familiarity of WebForms can continue to build custom modules using the traditional approach. However, in DotNetNuke 6.2 we also introduced a new Service Framework which is actually built on top of MVC2 ( we chose to leverage MVC because it had the most intuitive, light-weight REST implementation in the .NET stack ). The Services Framework allowed us to build some rich interactive features in DotNetNuke 6.2, including the Messaging and Notification Center and Activity Feed. But based on where we know Microsoft is heading, it makes sense for the next major version of DotNetNuke ( which is expected to be released in Q4 2012 ) to migrate from MVC2 to Web API. This will likely result in some breaking changes in the Services Framework but we feel it is the best approach for ensuring the platform remains highly modern and relevant. The fact that our development strategy is perfectly aligned with the “One ASP.NET” strategy from Microsoft means that our customers and developer community can be confident in their current and future investments in the DotNetNuke platform.

    Read the article

  • Frequent Disconnects ubuntu desktop 12.10 x64 intel 82579V e1000e

    - by user112055
    I'm having frequent disconnects with my new install of Ubuntu 12.10. I tried updating the kernel driver to the latest intel release to no avail. My expertise is spent. It happens anywhere between 1 min and 10 min. Any ideas? syslog: Dec 1 13:51:39 andromeda kernel: [ 972.188809] audit_printk_skb: 6 callbacks suppressed Dec 1 13:51:39 andromeda kernel: [ 972.188813] type=1701 audit(1354398699.418:199): auid=4294967295 uid=1000 gid=1000 ses=4294967295 pid=6039 comm="chrome" reason="seccomp" sig=0 syscall=4 compat=0 ip=0x7f26777d9205 code=0x50000 Dec 1 13:51:39 andromeda kernel: [ 972.188817] type=1701 audit(1354398699.418:200): auid=4294967295 uid=1000 gid=1000 ses=4294967295 pid=6039 comm="chrome" reason="seccomp" sig=0 syscall=4 compat=0 ip=0x7f26777d9205 code=0x50000 Dec 1 13:51:39 andromeda kernel: [ 972.188820] type=1701 audit(1354398699.418:201): auid=4294967295 uid=1000 gid=1000 ses=4294967295 pid=6039 comm="chrome" reason="seccomp" sig=0 syscall=4 compat=0 ip=0x7f26777d9205 code=0x50000 Dec 1 13:51:39 andromeda kernel: [ 972.188823] type=1701 audit(1354398699.418:202): auid=4294967295 uid=1000 gid=1000 ses=4294967295 pid=6039 comm="chrome" reason="seccomp" sig=0 syscall=4 compat=0 ip=0x7f26777d9205 code=0x50000 Dec 1 13:51:39 andromeda kernel: [ 972.188825] type=1701 audit(1354398699.418:203): auid=4294967295 uid=1000 gid=1000 ses=4294967295 pid=6039 comm="chrome" reason="seccomp" sig=0 syscall=4 compat=0 ip=0x7f26777d9205 code=0x50000 Dec 1 13:51:39 andromeda kernel: [ 972.331419] type=1701 audit(1354398699.558:204): auid=4294967295 uid=1000 gid=1000 ses=4294967295 pid=6039 comm="chrome" reason="seccomp" sig=0 syscall=2 compat=0 ip=0x7f26777d96b0 code=0x50000 Dec 1 13:53:12 andromeda NetworkManager[1115]: <info> (eth0): carrier now OFF (device state 100, deferring action for 4 seconds) Dec 1 13:53:12 andromeda kernel: [ 1064.894387] e1000e: e1000e: eth0 NIC Link is Down Dec 1 13:53:16 andromeda NetworkManager[1115]: <info> (eth0): device state change: activated -> unavailable (reason 'carrier-changed') [100 20 40] Dec 1 13:53:16 andromeda NetworkManager[1115]: <info> (eth0): deactivating device (reason 'carrier-changed') [40] Dec 1 13:53:16 andromeda NetworkManager[1115]: <info> (eth0): canceled DHCP transaction, DHCP client pid 5946 Dec 1 13:53:16 andromeda avahi-daemon[890]: Withdrawing address record for fe80::ea40:f2ff:fee2:4d86 on eth0. Dec 1 13:53:16 andromeda avahi-daemon[890]: Leaving mDNS multicast group on interface eth0.IPv6 with address fe80::ea40:f2ff:fee2:4d86. Dec 1 13:53:16 andromeda avahi-daemon[890]: Interface eth0.IPv6 no longer relevant for mDNS. Dec 1 13:53:16 andromeda kernel: [ 1069.025288] IPv6: ADDRCONF(NETDEV_UP): eth0: link is not ready Dec 1 13:53:16 andromeda avahi-daemon[890]: Withdrawing address record for 192.168.11.17 on eth0. Dec 1 13:53:16 andromeda avahi-daemon[890]: Leaving mDNS multicast group on interface eth0.IPv4 with address 192.168.11.17. Dec 1 13:53:16 andromeda avahi-daemon[890]: Interface eth0.IPv4 no longer relevant for mDNS. Dec 1 13:53:16 andromeda NetworkManager[1115]: <warn> DNS: plugin dnsmasq update failed Dec 1 13:53:16 andromeda NetworkManager[1115]: <info> ((null)): removing resolv.conf from /sbin/resolvconf Dec 1 13:53:16 andromeda dnsmasq[1907]: setting upstream servers from DBus Dec 1 13:53:16 andromeda dbus[800]: [system] Activating service name='org.freedesktop.nm_dispatcher' (using servicehelper) Dec 1 13:53:16 andromeda dbus[800]: [system] Successfully activated service 'org.freedesktop.nm_dispatcher' Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> (eth0): carrier now ON (device state 20) Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> (eth0): device state change: unavailable -> disconnected (reason 'carrier-changed') [20 30 40] Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Auto-activating connection '82579V'. Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) starting connection '82579V' Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> (eth0): device state change: disconnected -> prepare (reason 'none') [30 40 0] Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 1 of 5 (Device Prepare) scheduled... Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 1 of 5 (Device Prepare) started... Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 2 of 5 (Device Configure) scheduled... Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 1 of 5 (Device Prepare) complete. Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 2 of 5 (Device Configure) starting... Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> (eth0): device state change: prepare -> config (reason 'none') [40 50 0] Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 2 of 5 (Device Configure) successful. Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 3 of 5 (IP Configure Start) scheduled. Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 2 of 5 (Device Configure) complete. Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 3 of 5 (IP Configure Start) started... Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> (eth0): device state change: config -> ip-config (reason 'none') [50 70 0] Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Beginning DHCPv4 transaction (timeout in 45 seconds) Dec 1 13:53:32 andromeda kernel: [ 1084.938042] e1000e: e1000e: eth0 NIC Link is Up 100 Mbps Full Duplex, Flow Control: Rx/Tx Dec 1 13:53:32 andromeda kernel: [ 1084.938049] e1000e 0000:00:19.0: eth0: 10/100 speed: disabling TSO Dec 1 13:53:32 andromeda kernel: [ 1084.938815] IPv6: ADDRCONF(NETDEV_CHANGE): eth0: link becomes ready Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> dhclient started with pid 6080 Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 3 of 5 (IP Configure Start) complete. Dec 1 13:53:32 andromeda dhclient: Internet Systems Consortium DHCP Client 4.2.4 Dec 1 13:53:32 andromeda dhclient: Copyright 2004-2012 Internet Systems Consortium. Dec 1 13:53:32 andromeda dhclient: All rights reserved. Dec 1 13:53:32 andromeda dhclient: For info, please visit https://www.isc.org/software/dhcp/ Dec 1 13:53:32 andromeda dhclient: Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> (eth0): DHCPv4 state changed nbi -> preinit Dec 1 13:53:32 andromeda dhclient: Listening on LPF/eth0/e8:40:f2:e2:4d:86 Dec 1 13:53:32 andromeda dhclient: Sending on LPF/eth0/e8:40:f2:e2:4d:86 Dec 1 13:53:32 andromeda dhclient: Sending on Socket/fallback Dec 1 13:53:32 andromeda dhclient: DHCPREQUEST of 192.168.11.17 on eth0 to 255.255.255.255 port 67 Dec 1 13:53:32 andromeda dhclient: DHCPACK of 192.168.11.17 from 192.168.11.1 Dec 1 13:53:32 andromeda dhclient: bound to 192.168.11.17 -- renewal in 33576 seconds. Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> (eth0): DHCPv4 state changed preinit -> reboot Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> address 192.168.11.17 Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> prefix 24 (255.255.255.0) Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> gateway 192.168.11.1 Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> hostname 'andromeda' Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> nameserver '192.168.11.1' Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> domain name 'hsd1.ca.comcast.net' Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 5 of 5 (IPv4 Configure Commit) scheduled... Dec 1 13:53:32 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 5 of 5 (IPv4 Commit) started... Dec 1 13:53:32 andromeda avahi-daemon[890]: Joining mDNS multicast group on interface eth0.IPv4 with address 192.168.11.17. Dec 1 13:53:32 andromeda avahi-daemon[890]: New relevant interface eth0.IPv4 for mDNS. Dec 1 13:53:32 andromeda avahi-daemon[890]: Registering new address record for 192.168.11.17 on eth0.IPv4. Dec 1 13:53:33 andromeda NetworkManager[1115]: <info> (eth0): device state change: ip-config -> activated (reason 'none') [70 100 0] Dec 1 13:53:33 andromeda NetworkManager[1115]: <info> ((null)): writing resolv.conf to /sbin/resolvconf Dec 1 13:53:33 andromeda dnsmasq[1907]: setting upstream servers from DBus Dec 1 13:53:33 andromeda dnsmasq[1907]: using nameserver 192.168.11.1#53 Dec 1 13:53:33 andromeda NetworkManager[1115]: <info> Policy set '82579V' (eth0) as default for IPv4 routing and DNS. Dec 1 13:53:33 andromeda NetworkManager[1115]: <info> Activation (eth0) successful, device activated. Dec 1 13:53:33 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 5 of 5 (IPv4 Commit) complete. Dec 1 13:53:33 andromeda dbus[800]: [system] Activating service name='org.freedesktop.nm_dispatcher' (using servicehelper) Dec 1 13:53:33 andromeda dbus[800]: [system] Successfully activated service 'org.freedesktop.nm_dispatcher' Dec 1 13:53:33 andromeda avahi-daemon[890]: Joining mDNS multicast group on interface eth0.IPv6 with address fe80::ea40:f2ff:fee2:4d86. Dec 1 13:53:33 andromeda avahi-daemon[890]: New relevant interface eth0.IPv6 for mDNS. Dec 1 13:53:33 andromeda avahi-daemon[890]: Registering new address record for fe80::ea40:f2ff:fee2:4d86 on eth0.*. Dec 1 13:53:41 andromeda ntpdate[6154]: adjust time server 91.189.94.4 offset 0.000928 sec Dec 1 13:53:50 andromeda NetworkManager[1115]: <info> (eth0): carrier now OFF (device state 100, deferring action for 4 seconds) Dec 1 13:53:50 andromeda kernel: [ 1102.980003] e1000e: e1000e: eth0 NIC Link is Down Dec 1 13:53:54 andromeda NetworkManager[1115]: <info> (eth0): device state change: activated -> unavailable (reason 'carrier-changed') [100 20 40] Dec 1 13:53:54 andromeda NetworkManager[1115]: <info> (eth0): deactivating device (reason 'carrier-changed') [40] Dec 1 13:53:54 andromeda NetworkManager[1115]: <info> (eth0): canceled DHCP transaction, DHCP client pid 6080 Dec 1 13:53:54 andromeda avahi-daemon[890]: Withdrawing address record for fe80::ea40:f2ff:fee2:4d86 on eth0. Dec 1 13:53:54 andromeda avahi-daemon[890]: Leaving mDNS multicast group on interface eth0.IPv6 with address fe80::ea40:f2ff:fee2:4d86. Dec 1 13:53:54 andromeda avahi-daemon[890]: Interface eth0.IPv6 no longer relevant for mDNS. Dec 1 13:53:54 andromeda avahi-daemon[890]: Withdrawing address record for 192.168.11.17 on eth0. Dec 1 13:53:54 andromeda avahi-daemon[890]: Leaving mDNS multicast group on interface eth0.IPv4 with address 192.168.11.17. Dec 1 13:53:54 andromeda kernel: [ 1107.025959] IPv6: ADDRCONF(NETDEV_UP): eth0: link is not ready Dec 1 13:53:54 andromeda NetworkManager[1115]: <warn> DNS: plugin dnsmasq update failed Dec 1 13:53:54 andromeda NetworkManager[1115]: <info> ((null)): removing resolv.conf from /sbin/resolvconf Dec 1 13:53:54 andromeda avahi-daemon[890]: Interface eth0.IPv4 no longer relevant for mDNS. Dec 1 13:53:54 andromeda dnsmasq[1907]: setting upstream servers from DBus Dec 1 13:53:54 andromeda dbus[800]: [system] Activating service name='org.freedesktop.nm_dispatcher' (using servicehelper) Dec 1 13:53:54 andromeda dbus[800]: [system] Successfully activated service 'org.freedesktop.nm_dispatcher' Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> (eth0): carrier now ON (device state 20) Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> (eth0): device state change: unavailable -> disconnected (reason 'carrier-changed') [20 30 40] Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Auto-activating connection '82579V'. Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) starting connection '82579V' Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> (eth0): device state change: disconnected -> prepare (reason 'none') [30 40 0] Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 1 of 5 (Device Prepare) scheduled... Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 1 of 5 (Device Prepare) started... Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 2 of 5 (Device Configure) scheduled... Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 1 of 5 (Device Prepare) complete. Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 2 of 5 (Device Configure) starting... Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> (eth0): device state change: prepare -> config (reason 'none') [40 50 0] Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 2 of 5 (Device Configure) successful. Dec 1 13:54:10 andromeda kernel: [ 1123.167668] e1000e: e1000e: eth0 NIC Link is Up 100 Mbps Full Duplex, Flow Control: Rx/Tx Dec 1 13:54:10 andromeda kernel: [ 1123.167675] e1000e 0000:00:19.0: eth0: 10/100 speed: disabling TSO Dec 1 13:54:10 andromeda kernel: [ 1123.168430] IPv6: ADDRCONF(NETDEV_CHANGE): eth0: link becomes ready Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 3 of 5 (IP Configure Start) scheduled. Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 2 of 5 (Device Configure) complete. Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 3 of 5 (IP Configure Start) started... Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> (eth0): device state change: config -> ip-config (reason 'none') [50 70 0] Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Beginning DHCPv4 transaction (timeout in 45 seconds) Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> dhclient started with pid 6212 Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 3 of 5 (IP Configure Start) complete. Dec 1 13:54:10 andromeda dhclient: Internet Systems Consortium DHCP Client 4.2.4 Dec 1 13:54:10 andromeda dhclient: Copyright 2004-2012 Internet Systems Consortium. Dec 1 13:54:10 andromeda dhclient: All rights reserved. Dec 1 13:54:10 andromeda dhclient: For info, please visit https://www.isc.org/software/dhcp/ Dec 1 13:54:10 andromeda dhclient: Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> (eth0): DHCPv4 state changed nbi -> preinit Dec 1 13:54:10 andromeda dhclient: Listening on LPF/eth0/e8:40:f2:e2:4d:86 Dec 1 13:54:10 andromeda dhclient: Sending on LPF/eth0/e8:40:f2:e2:4d:86 Dec 1 13:54:10 andromeda dhclient: Sending on Socket/fallback Dec 1 13:54:10 andromeda dhclient: DHCPREQUEST of 192.168.11.17 on eth0 to 255.255.255.255 port 67 Dec 1 13:54:10 andromeda dhclient: DHCPACK of 192.168.11.17 from 192.168.11.1 Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> (eth0): DHCPv4 state changed preinit -> reboot Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> address 192.168.11.17 Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> prefix 24 (255.255.255.0) Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> gateway 192.168.11.1 Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> hostname 'andromeda' Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> nameserver '192.168.11.1' Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> domain name 'hsd1.ca.comcast.net' Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 5 of 5 (IPv4 Configure Commit) scheduled... Dec 1 13:54:10 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 5 of 5 (IPv4 Commit) started... Dec 1 13:54:10 andromeda avahi-daemon[890]: Joining mDNS multicast group on interface eth0.IPv4 with address 192.168.11.17. Dec 1 13:54:10 andromeda dhclient: bound to 192.168.11.17 -- renewal in 35416 seconds. Dec 1 13:54:10 andromeda avahi-daemon[890]: New relevant interface eth0.IPv4 for mDNS. Dec 1 13:54:10 andromeda avahi-daemon[890]: Registering new address record for 192.168.11.17 on eth0.IPv4. Dec 1 13:54:11 andromeda NetworkManager[1115]: <info> (eth0): device state change: ip-config -> activated (reason 'none') [70 100 0] Dec 1 13:54:11 andromeda NetworkManager[1115]: <info> ((null)): writing resolv.conf to /sbin/resolvconf Dec 1 13:54:11 andromeda dnsmasq[1907]: setting upstream servers from DBus Dec 1 13:54:11 andromeda dnsmasq[1907]: using nameserver 192.168.11.1#53 Dec 1 13:54:11 andromeda NetworkManager[1115]: <info> Policy set '82579V' (eth0) as default for IPv4 routing and DNS. Dec 1 13:54:11 andromeda NetworkManager[1115]: <info> Activation (eth0) successful, device activated. Dec 1 13:54:11 andromeda NetworkManager[1115]: <info> Activation (eth0) Stage 5 of 5 (IPv4 Commit) complete. Dec 1 13:54:11 andromeda dbus[800]: [system] Activating service name='org.freedesktop.nm_dispatcher' (using servicehelper) Dec 1 13:54:11 andromeda dbus[800]: [system] Successfully activated service 'org.freedesktop.nm_dispatcher' Dec 1 13:54:12 andromeda avahi-daemon[890]: Joining mDNS multicast group on interface eth0.IPv6 with address fe80::ea40:f2ff:fee2:4d86. Dec 1 13:54:12 andromeda avahi-daemon[890]: New relevant interface eth0.IPv6 for mDNS. Dec 1 13:54:12 andromeda avahi-daemon[890]: Registering new address record for fe80::ea40:f2ff:fee2:4d86 on eth0.*. Dec 1 13:54:19 andromeda ntpdate[6286]: adjust time server 91.189.94.4 offset 0.001142 sec $ lspci -v 00:19.0 Ethernet controller: Intel Corporation 82579V Gigabit Network Connection (rev 04) Subsystem: Intel Corporation Device 2031 Flags: bus master, fast devsel, latency 0, IRQ 45 Memory at f7f00000 (32-bit, non-prefetchable) [size=128K] Memory at f7f39000 (32-bit, non-prefetchable) [size=4K] I/O ports at f040 [size=32] Capabilities: [c8] Power Management version 2 Capabilities: [d0] MSI: Enable+ Count=1/1 Maskable- 64bit+ Capabilities: [e0] PCI Advanced Features Kernel driver in use: e1000e Kernel modules: e1000e $ modinfo e1000e filename: /lib/modules/3.5.0-19-generic/kernel/drivers/net/e1000e/e1000e.ko version: 2.1.4-NAPI license: GPL description: Intel(R) PRO/1000 Network Driver author: Intel Corporation, <[email protected]> srcversion: 0809529BE0BBC44883956AF alias: pci:v00008086d0000153Bsv*sd*bc*sc*i* alias: pci:v00008086d0000153Asv*sd*bc*sc*i* alias: pci:v00008086d00001503sv*sd*bc*sc*i* alias: pci:v00008086d00001502sv*sd*bc*sc*i* alias: pci:v00008086d000010F0sv*sd*bc*sc*i* alias: pci:v00008086d000010EFsv*sd*bc*sc*i* alias: pci:v00008086d000010EBsv*sd*bc*sc*i* alias: pci:v00008086d000010EAsv*sd*bc*sc*i* alias: pci:v00008086d00001525sv*sd*bc*sc*i* alias: pci:v00008086d000010DFsv*sd*bc*sc*i* alias: pci:v00008086d000010DEsv*sd*bc*sc*i* alias: pci:v00008086d000010CEsv*sd*bc*sc*i* alias: pci:v00008086d000010CDsv*sd*bc*sc*i* alias: pci:v00008086d000010CCsv*sd*bc*sc*i* alias: pci:v00008086d000010CBsv*sd*bc*sc*i* alias: pci:v00008086d000010F5sv*sd*bc*sc*i* alias: pci:v00008086d000010BFsv*sd*bc*sc*i* alias: pci:v00008086d000010E5sv*sd*bc*sc*i* alias: pci:v00008086d0000294Csv*sd*bc*sc*i* alias: pci:v00008086d000010BDsv*sd*bc*sc*i* alias: pci:v00008086d000010C3sv*sd*bc*sc*i* alias: pci:v00008086d000010C2sv*sd*bc*sc*i* alias: pci:v00008086d000010C0sv*sd*bc*sc*i* alias: pci:v00008086d00001501sv*sd*bc*sc*i* alias: pci:v00008086d00001049sv*sd*bc*sc*i* alias: pci:v00008086d0000104Dsv*sd*bc*sc*i* alias: pci:v00008086d0000104Bsv*sd*bc*sc*i* alias: pci:v00008086d0000104Asv*sd*bc*sc*i* alias: pci:v00008086d000010C4sv*sd*bc*sc*i* alias: pci:v00008086d000010C5sv*sd*bc*sc*i* alias: pci:v00008086d0000104Csv*sd*bc*sc*i* alias: pci:v00008086d000010BBsv*sd*bc*sc*i* alias: pci:v00008086d00001098sv*sd*bc*sc*i* alias: pci:v00008086d000010BAsv*sd*bc*sc*i* alias: pci:v00008086d00001096sv*sd*bc*sc*i* alias: pci:v00008086d0000150Csv*sd*bc*sc*i* alias: pci:v00008086d000010F6sv*sd*bc*sc*i* alias: pci:v00008086d000010D3sv*sd*bc*sc*i* alias: pci:v00008086d0000109Asv*sd*bc*sc*i* alias: pci:v00008086d0000108Csv*sd*bc*sc*i* alias: pci:v00008086d0000108Bsv*sd*bc*sc*i* alias: pci:v00008086d0000107Fsv*sd*bc*sc*i* alias: pci:v00008086d0000107Esv*sd*bc*sc*i* alias: pci:v00008086d0000107Dsv*sd*bc*sc*i* alias: pci:v00008086d000010B9sv*sd*bc*sc*i* alias: pci:v00008086d000010D5sv*sd*bc*sc*i* alias: pci:v00008086d000010DAsv*sd*bc*sc*i* alias: pci:v00008086d000010D9sv*sd*bc*sc*i* alias: pci:v00008086d00001060sv*sd*bc*sc*i* alias: pci:v00008086d000010A5sv*sd*bc*sc*i* alias: pci:v00008086d000010BCsv*sd*bc*sc*i* alias: pci:v00008086d000010A4sv*sd*bc*sc*i* alias: pci:v00008086d0000105Fsv*sd*bc*sc*i* alias: pci:v00008086d0000105Esv*sd*bc*sc*i* depends: vermagic: 3.5.0-19-generic SMP mod_unload modversions parm: copybreak:Maximum size of packet that is copied to a new buffer on receive (uint) parm: TxIntDelay:Transmit Interrupt Delay (array of int) parm: TxAbsIntDelay:Transmit Absolute Interrupt Delay (array of int) parm: RxIntDelay:Receive Interrupt Delay (array of int) parm: RxAbsIntDelay:Receive Absolute Interrupt Delay (array of int) parm: InterruptThrottleRate:Interrupt Throttling Rate (array of int) parm: IntMode:Interrupt Mode (array of int) parm: SmartPowerDownEnable:Enable PHY smart power down (array of int) parm: KumeranLockLoss:Enable Kumeran lock loss workaround (array of int) parm: CrcStripping:Enable CRC Stripping, disable if your BMC needs the CRC (array of int) parm: EEE:Enable/disable on parts that support the feature (array of int) parm: Node:[ROUTING] Node to allocate memory on, default -1 (array of int) parm: debug:Debug level (0=none,...,16=all) (int)

    Read the article

< Previous Page | 401 402 403 404 405 406 407 408 409  | Next Page >