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  • SQL SERVER – Solution – Puzzle – Statistics are not Updated but are Created Once

    - by pinaldave
    Earlier I asked puzzle why statistics are not updated. Read the complete details over here: Statistics are not Updated but are Created Once In the question I have demonstrated even though statistics should have been updated after lots of insert in the table are not updated.(Read the details SQL SERVER – When are Statistics Updated – What triggers Statistics to Update) In this example I have created following situation: Create Table Insert 1000 Records Check the Statistics Now insert 10 times more 10,000 indexes Check the Statistics – it will be NOT updated Auto Update Statistics and Auto Create Statistics for database is TRUE Now I have requested two things in the example 1) Why this is happening? 2) How to fix this issue? I have many answers – here is the how I fixed it which has resolved the issue for me. NOTE: There are multiple answers to this problem and I will do my best to list all. Solution: Create nonclustered Index on column City Here is the working example for the same. Let us understand this script and there is added explanation at the end. -- Execution Plans Difference -- Estimated Execution Plan Vs Actual Execution Plan -- Create Sample Database CREATE DATABASE SampleDB GO USE SampleDB GO -- Create Table CREATE TABLE ExecTable (ID INT, FirstName VARCHAR(100), LastName VARCHAR(100), City VARCHAR(100)) GO CREATE NONCLUSTERED INDEX IX_ExecTable1 ON ExecTable (City); GO -- Insert One Thousand Records -- INSERT 1 INSERT INTO ExecTable (ID,FirstName,LastName,City) SELECT TOP 1000 ROW_NUMBER() OVER (ORDER BY a.name) RowID, 'Bob', CASE WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%2 = 1 THEN 'Smith' ELSE 'Brown' END, CASE WHEN ROW_NUMBER() OVER (ORDER BY a.name)%20 = 1 THEN 'New York' WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%20 = 5 THEN 'San Marino' WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%20 = 3 THEN 'Los Angeles' WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%20 = 7 THEN 'La Cinega' WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%20 = 13 THEN 'San Diego' WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%20 = 17 THEN 'Las Vegas' ELSE 'Houston' END FROM sys.all_objects a CROSS JOIN sys.all_objects b GO -- Display statistics of the table sp_helpstats N'ExecTable', 'ALL' GO -- Select Statement SELECT FirstName, LastName, City FROM ExecTable WHERE City  = 'New York' GO -- Display statistics of the table sp_helpstats N'ExecTable', 'ALL' GO -- Replace your Statistics over here DBCC SHOW_STATISTICS('ExecTable', IX_ExecTable1); GO -------------------------------------------------------------- -- Round 2 -- Insert One Thousand Records -- INSERT 2 INSERT INTO ExecTable (ID,FirstName,LastName,City) SELECT TOP 1000 ROW_NUMBER() OVER (ORDER BY a.name) RowID, 'Bob', CASE WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%2 = 1 THEN 'Smith' ELSE 'Brown' END, CASE WHEN ROW_NUMBER() OVER (ORDER BY a.name)%20 = 1 THEN 'New York' WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%20 = 5 THEN 'San Marino' WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%20 = 3 THEN 'Los Angeles' WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%20 = 7 THEN 'La Cinega' WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%20 = 13 THEN 'San Diego' WHEN  ROW_NUMBER() OVER (ORDER BY a.name)%20 = 17 THEN 'Las Vegas' ELSE 'Houston' END FROM sys.all_objects a CROSS JOIN sys.all_objects b GO -- Select Statement SELECT FirstName, LastName, City FROM ExecTable WHERE City  = 'New York' GO -- Display statistics of the table sp_helpstats N'ExecTable', 'ALL' GO -- Replace your Statistics over here DBCC SHOW_STATISTICS('ExecTable', IX_ExecTable1); GO -- Clean up Database DROP TABLE ExecTable GO When I created non clustered index on the column city, it also created statistics on the same column with same name as index. When we populate the data in the column the index is update – resulting execution plan to be invalided – this leads to the statistics to be updated in next execution of SELECT. This behavior does not happen on Heap or column where index is auto created. If you explicitly update the index, often you can see the statistics are updated as well. You can see this is for sure happening if you follow the tell of John Sansom. John Sansom‘s suggestion: That was fun! Although the column statistics are invalidated by the time the second select statement is executed, the query is not compiled/recompiled but instead the existing query plan is reused. It is the “next” compiled query against the column statistics that will see that they are out of date and will then in turn instantiate the action of updating statistics. You can see this in action by forcing the second statement to recompile. SELECT FirstName, LastName, City FROM ExecTable WHERE City = ‘New York’ option(RECOMPILE) GO Kevin Cross also have another suggestion: I agree with John. It is reusing the Execution Plan. Aside from OPTION(RECOMPILE), clearing the Execution Plan Cache before the subsequent tests will also work. i.e., run this before round 2: ————————————————————– – Clear execution plan cache before next test DBCC FREEPROCCACHE WITH NO_INFOMSGS; ————————————————————– Nice puzzle! Kevin As this was puzzle John and Kevin both got the correct answer, there was no condition for answer to be part of best practices. I know John and he is finest DBA around – his tremendous knowledge has always impressed me. John and Kevin both will agree that clearing cache either using DBCC FREEPROCCACHE and recompiling each query every time is for sure not good advice on production server. It is correct answer but not best practice. By the way, if you have better solution or have better suggestion please advise. I am open to change my answer and publish further improvement to this solution. On very separate note, I like to have clustered index on my Primary Key, which I have not mentioned here as it is out of the scope of this puzzle. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Pinal Dave, PostADay, Readers Contribution, Readers Question, SQL, SQL Authority, SQL Index, SQL Puzzle, SQL Query, SQL Scripts, SQL Server, SQL Tips and Tricks, T SQL, Technology Tagged: Statistics

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  • SQL SERVER – Guest Post by Sandip Pani – SQL Server Statistics Name and Index Creation

    - by pinaldave
    Sometimes something very small or a common error which we observe in daily life teaches us new things. SQL Server Expert Sandip Pani (winner of Joes 2 Pros Contests) has come across similar experience. Sandip has written a guest post on an error he faced in his daily work. Sandip is working for QSI Healthcare as an Associate Technical Specialist and have more than 5 years of total experience. He blogs at SQLcommitted.com and contribute in various forums. His social media hands are LinkedIn, Facebook and Twitter. Once I faced following error when I was working on performance tuning project and attempt to create an Index. Mug 1913, Level 16, State 1, Line 1 The operation failed because an index or statistics with name ‘Ix_Table1_1′ already exists on table ‘Table1′. The immediate reaction to the error was that I might have created that index earlier and when I researched it further I found the same as the index was indeed created two times. This totally makes sense. This can happen due to many reasons for example if the user is careless and executes the same code two times as well, when he attempts to create index without checking if there was index already on the object. However when I paid attention to the details of the error, I realize that error message also talks about statistics along with the index. I got curious if the same would happen if I attempt to create indexes with the same name as statistics already created. There are a few other questions also prompted in my mind. I decided to do a small demonstration of the subject and build following demonstration script. The goal of my experiment is to find out the relation between statistics and the index. Statistics is one of the important input parameter for the optimizer during query optimization process. If the query is nontrivial then only optimizer uses statistics to perform a cost based optimization to select a plan. For accuracy and further learning I suggest to read MSDN. Now let’s find out the relationship between index and statistics. We will do the experiment in two parts. i) Creating Index ii) Creating Statistics We will be using the following T-SQL script for our example. IF (OBJECT_ID('Table1') IS NOT NULL) DROP TABLE Table1 GO CREATE TABLE Table1 (Col1 INT NOT NULL, Col2 VARCHAR(20) NOT NULL) GO We will be using following two queries to check if there are any index or statistics on our sample table Table1. -- Details of Index SELECT OBJECT_NAME(OBJECT_ID) AS TableName, Name AS IndexName, type_desc FROM sys.indexes WHERE OBJECT_NAME(OBJECT_ID) = 'table1' GO -- Details of Statistics SELECT OBJECT_NAME(OBJECT_ID) TableName, Name AS StatisticsName FROM sys.stats WHERE OBJECT_NAME(OBJECT_ID) = 'table1' GO When I ran above two scripts on the table right after it was created it did not give us any result which was expected. Now let us begin our test. 1) Create an index on the table Create following index on the table. CREATE NONCLUSTERED INDEX Ix_Table1_1 ON Table1(Col1) GO Now let us use above two scripts and see their results. We can see that when we created index at the same time it created statistics also with the same name. Before continuing to next set of demo – drop the table using following script and re-create the table using a script provided at the beginning of the table. DROP TABLE table1 GO 2) Create a statistic on the table Create following statistics on the table. CREATE STATISTICS Ix_table1_1 ON Table1 (Col1) GO Now let us use above two scripts and see their results. We can see that when we created statistics Index is not created. The behavior of this experiment is different from the earlier experiment. Clean up the table setup using the following script: DROP TABLE table1 GO Above two experiments teach us very valuable lesson that when we create indexes, SQL Server generates the index and statistics (with the same name as the index name) together. Now due to the reason if we have already had statistics with the same name but not the index, it is quite possible that we will face the error to create the index even though there is no index with the same name. A Quick Check To validate that if we create statistics first and then index after that with the same name, it will throw an error let us run following script in SSMS. Make sure to drop the table and clean up our sample table at the end of the experiment. -- Create sample table CREATE TABLE TestTable (Col1 INT NOT NULL, Col2 VARCHAR(20) NOT NULL) GO -- Create Statistics CREATE STATISTICS IX_TestTable_1 ON TestTable (Col1) GO -- Create Index CREATE NONCLUSTERED INDEX IX_TestTable_1 ON TestTable(Col1) GO -- Check error /*Msg 1913, Level 16, State 1, Line 2 The operation failed because an index or statistics with name 'IX_TestTable_1' already exists on table 'TestTable'. */ -- Clean up DROP TABLE TestTable GO While creating index it will throw the following error as statistics with the same name is already created. In simple words – when we create index the name of the index should be different from any of the existing indexes and statistics. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: PostADay, SQL, SQL Authority, SQL Error Messages, SQL Index, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology Tagged: SQL Statistics

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  • Posting from ASP.NET WebForms page to another URL

    - by hajan
    Few days ago I had a case when I needed to make FORM POST from my ASP.NET WebForms page to an external site URL. More specifically, I was working on implementing Simple Payment System (like Amazon, PayPal, MoneyBookers). The operator asks to make FORM POST request to a given URL in their website, sending parameters together with the post which are computed on my application level (access keys, secret keys, signature, return-URL… etc). So, since we are not allowed nesting another form inside the <form runat=”server”> … </form>, which is required because other controls in my ASPX code work on server-side, I thought to inject the HTML and create FORM with method=”POST”. After making some proof of concept and testing some scenarios, I’ve concluded that I can do this very fast in two ways: Using jQuery to create form on fly with the needed parameters and make submit() Using HttpContext.Current.Response.Write to write the form on server-side (code-behind) and embed JavaScript code that will do the post Both ways seemed fine. 1. Using jQuery to create FORM html code and Submit it. Let’s say we have ‘PAY NOW’ button in our ASPX code: <asp:Button ID="btnPayNow" runat="server" Text="Pay Now" /> Now, if we want to make this button submit a FORM using POST method to another website, the jQuery way should be as follows: <script src="http://ajax.aspnetcdn.com/ajax/jquery/jquery-1.5.1.js" type="text/javascript"></script> <script type="text/javascript">     $(function () {         $("#btnPayNow").click(function (event) {             event.preventDefault();             //construct htmlForm string             var htmlForm = "<form id='myform' method='POST' action='http://www.microsoft.com'>" +                 "<input type='hidden' id='name' value='hajan' />" +             "</form>";             //Submit the form             $(htmlForm).appendTo("body").submit();         });     }); </script> Yes, as you see, the code fires on btnPayNow click. It removes the default button behavior, then creates htmlForm string. After that using jQuery we append the form to the body and submit it. Inside the form, you can see I have set the htttp://www.microsoft.com URL, so after clicking the button you should be automatically redirected to the Microsoft website (just for test, of course for Payment I’m using Operator's URL). 2. Using HttpContext.Current.Response.Write to write the form on server-side (code-behind) and embed JavaScript code that will do the post The C# code behind should be something like this: public void btnPayNow_Click(object sender, EventArgs e) {     string Url = "http://www.microsoft.com";     string formId = "myForm1";     StringBuilder htmlForm = new StringBuilder();     htmlForm.AppendLine("<html>");     htmlForm.AppendLine(String.Format("<body onload='document.forms[\"{0}\"].submit()'>",formId));     htmlForm.AppendLine(String.Format("<form id='{0}' method='POST' action='{1}'>", formId, Url));     htmlForm.AppendLine("<input type='hidden' id='name' value='hajan' />");     htmlForm.AppendLine("</form>");     htmlForm.AppendLine("</body>");     htmlForm.AppendLine("</html>");     HttpContext.Current.Response.Clear();     HttpContext.Current.Response.Write(htmlForm.ToString());     HttpContext.Current.Response.End();             } So, with this code we create htmlForm string using StringBuilder class and then just write the html to the page using HttpContext.Current.Response.Write. The interesting part here is that we submit the form using JavaScript code: document.forms["myForm1"].submit() This code runs on body load event, which means once the body is loaded the form is automatically submitted. Note: In order to test both solutions, create two applications on your web server and post the form from first to the second website, then get the values in the second website using Request.Form[“input-field-id”] I hope this was useful post for you. Regards, Hajan

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  • NoSQL Memcached API for MySQL: Latest Updates

    - by Mat Keep
    With data volumes exploding, it is vital to be able to ingest and query data at high speed. For this reason, MySQL has implemented NoSQL interfaces directly to the InnoDB and MySQL Cluster (NDB) storage engines, which bypass the SQL layer completely. Without SQL parsing and optimization, Key-Value data can be written directly to MySQL tables up to 9x faster, while maintaining ACID guarantees. In addition, users can continue to run complex queries with SQL across the same data set, providing real-time analytics to the business or anonymizing sensitive data before loading to big data platforms such as Hadoop, while still maintaining all of the advantages of their existing relational database infrastructure. This and more is discussed in the latest Guide to MySQL and NoSQL where you can learn more about using the APIs to scale new generations of web, cloud, mobile and social applications on the world's most widely deployed open source database The native Memcached API is part of the MySQL 5.6 Release Candidate, and is already available in the GA release of MySQL Cluster. By using the ubiquitous Memcached API for writing and reading data, developers can preserve their investments in Memcached infrastructure by re-using existing Memcached clients, while also eliminating the need for application changes. Speed, when combined with flexibility, is essential in the world of growing data volumes and variability. Complementing NoSQL access, support for on-line DDL (Data Definition Language) operations in MySQL 5.6 and MySQL Cluster enables DevOps teams to dynamically update their database schema to accommodate rapidly changing requirements, such as the need to capture additional data generated by their applications. These changes can be made without database downtime. Using the Memcached interface, developers do not need to define a schema at all when using MySQL Cluster. Lets look a little more closely at the Memcached implementations for both InnoDB and MySQL Cluster. Memcached Implementation for InnoDB The Memcached API for InnoDB is previewed as part of the MySQL 5.6 Release Candidate. As illustrated in the following figure, Memcached for InnoDB is implemented via a Memcached daemon plug-in to the mysqld process, with the Memcached protocol mapped to the native InnoDB API. Figure 1: Memcached API Implementation for InnoDB With the Memcached daemon running in the same process space, users get very low latency access to their data while also leveraging the scalability enhancements delivered with InnoDB and a simple deployment and management model. Multiple web / application servers can remotely access the Memcached / InnoDB server to get direct access to a shared data set. With simultaneous SQL access, users can maintain all the advanced functionality offered by InnoDB including support for Foreign Keys, XA transactions and complex JOIN operations. Benchmarks demonstrate that the NoSQL Memcached API for InnoDB delivers up to 9x higher performance than the SQL interface when inserting new key/value pairs, with a single low-end commodity server supporting nearly 70,000 Transactions per Second. Figure 2: Over 9x Faster INSERT Operations The delivered performance demonstrates MySQL with the native Memcached NoSQL interface is well suited for high-speed inserts with the added assurance of transactional guarantees. You can check out the latest Memcached / InnoDB developments and benchmarks here You can learn how to configure the Memcached API for InnoDB here Memcached Implementation for MySQL Cluster Memcached API support for MySQL Cluster was introduced with General Availability (GA) of the 7.2 release, and joins an extensive range of NoSQL interfaces that are already available for MySQL Cluster Like Memcached, MySQL Cluster provides a distributed hash table with in-memory performance. MySQL Cluster extends Memcached functionality by adding support for write-intensive workloads, a full relational model with ACID compliance (including persistence), rich query support, auto-sharding and 99.999% availability, with extensive management and monitoring capabilities. All writes are committed directly to MySQL Cluster, eliminating cache invalidation and the overhead of data consistency checking to ensure complete synchronization between the database and cache. Figure 3: Memcached API Implementation with MySQL Cluster Implementation is simple: 1. The application sends reads and writes to the Memcached process (using the standard Memcached API). 2. This invokes the Memcached Driver for NDB (which is part of the same process) 3. The NDB API is called, providing for very quick access to the data held in MySQL Cluster’s data nodes. The solution has been designed to be very flexible, allowing the application architect to find a configuration that best fits their needs. It is possible to co-locate the Memcached API in either the data nodes or application nodes, or alternatively within a dedicated Memcached layer. The benefit of this flexible approach to deployment is that users can configure behavior on a per-key-prefix basis (through tables in MySQL Cluster) and the application doesn’t have to care – it just uses the Memcached API and relies on the software to store data in the right place(s) and to keep everything synchronized. Using Memcached for Schema-less Data By default, every Key / Value is written to the same table with each Key / Value pair stored in a single row – thus allowing schema-less data storage. Alternatively, the developer can define a key-prefix so that each value is linked to a pre-defined column in a specific table. Of course if the application needs to access the same data through SQL then developers can map key prefixes to existing table columns, enabling Memcached access to schema-structured data already stored in MySQL Cluster. Conclusion Download the Guide to MySQL and NoSQL to learn more about NoSQL APIs and how you can use them to scale new generations of web, cloud, mobile and social applications on the world's most widely deployed open source database See how to build a social app with MySQL Cluster and the Memcached API from our on-demand webinar or take a look at the docs Don't hesitate to use the comments section below for any questions you may have 

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  • Handling Trailing Delimiters in HL7 Messages

    - by Thomas Canter
    Applies to: BizTalk Server 2006 with the HL7 1.3 Accelerator Outline of the problem Trailing Delimiters are empty values at the end of an object in a HL7 ER7 formatted message. Examples: Empty Field NTE|P| NTE|P|| Empty component ORC|1|725^ Empty Subcomponent ORC|1|||||27& Empty repeat OBR|1||||||||027~ Trailing delimiters indicate the following object exists and is empty, which is quite different from null, null is an explicit value indicated by a pair of double quotes -> "". The BizTalk HL7 Accelerator by default does not allow trailing delimiters. There are three methods to allow trailing delimiters. NOTE: All Schemas always allow trailing delimiters in the MSH Segment Using party identifiers MSH3.1 – Receive/inbound processing, using this value as a party allows you to configure the system to allow inbound trailing delimiters. MSH5.1 – Send/outbound processing, using this value as a party allows you to configure the system to allow outbound trailing delimiters. Generally, if you allow inbound trailing delimiters, unless you are willing to programmatically remove all trailing delimiters, then you need to configure the send to allow trailing delimiters. Add the appropriate parties to the BizTalk Parties list from these two fields in your message stream. Open the BizTalk HL7 Configuration tool and for each party check the "Allow trailing delimiters (separators)" check box on the Validation tab. Disadvantage – Each MSH3.1 and MSH5.1 value must be represented in the parties list and configured. Advantage – granular control over system behavior for each inbound/outbound system. Using instance properties of a pipeline used in a send port or receive location. Open the BizTalk Server Administration console locate the send port or receive location that contains the BTAHL72XReceivePipeline or BTAHL72XSendPipeline pipeline. Open the properties To the right of the pipeline selected locate the […] ellipses button In the property list, locate the "TrailingDelimiterAllowed" property and set it to True. Advantage – All messages through a particular Send Port or Receive Location will allow trailing delimiters. Disadvantage – Must configure each Send Port or Receive Location. No granular control over which remote parties will send or receive messages with trailing delimiters. Using a custom pipeline that uses a pre-configured BTA HL7 Pipeline component. Use Visual Studio to construct a custom receive and send pipeline using the appropriate assembler or dissasembler. Set the component property to "TrailingDelimitersAllowed" to True Compile and deploy the custom pipeline Use the custom pipeline instead of the standard pipeline for all HL7 message processing Advantage – All messages using the custom pipeline will automatically allow trailing delimiters. Disadvantage – Requires custom coding and development to create and deploy the custom pipeline. No granular control over which remote parties will send or receive messages with trailing delimiters. What does a Trailing Delimiter do to the XML Schema? Allowing trailing delimiters does not have the impact often expected in the actual XML Schema.The Schema reproduces the message with no data loss.Thus, the message when represented in XML must contain the extra fields, in order to reproduce the outbound message.Thus, a trialing delimiter results in an empty XML field.Trailing Delmiters are not stripped from the inbound message. Example:<PID_21>44172</PID_21><PID_21>9257</PID_21> -> the original maximum number of repeats<PID_21></PID_21> -> The empty repeated field Allowing trailing delimiters not remove the trailing delimiters from the message, it simply suppresses the check that will cause the message to fail parse with trailing delimiters. When can you not fix the problem by enabling trailing delimiters Each object in a message must have a location in the target BTAHL7 schema for its content to reside.If you have more objects in the message than are contained at that location, then enabling trailing delimiters will not resolve the problem. The schema must be extended to accommodate the empty message content.Examples: Extra Field NTE|P||||Only 4 fields in NTE Segment, the 4th field exists, but is empty. Extra component PID|1|1523|47^^^^^^^Only 5 components in a CX data type, the 5th component exists, but is empty Extra subcomponent ORC|1|||||27&&Only 2 subcomponents in a CQ data type, the 3rd subcomponent is empty, but exists. Extra Repeat PID|1||||||||||||||||||||4419~5217~Only 2 repeats allowed for the field "Mother's identifier", the repeat is empty, but exists. In each of these cases, you must locate the failing object and extend the type to allow an additional object of that type. FieldAdd a field of ST to the end of the segment with a suitable name in the segments_nnn.xsd Component Create a new Custom CX data type (i.e. CX_XtraComp) in the datatypes_nnn.xsd and add a new component to the custom CX data type. Update the field in the segments_nnn.xsd file to use the custom data type instead of the standard datatype. Subcomponent Create a new Custom CQ data type that accepts an additional TS value at the end of the data type. Create a custom TQ data type that uses the new custom CQ data type as the first subcomponent. Modify the ORC segment to use the new CQ data type at ORC.7 instead of the standard CQ data type. RepeatModify the Field definition for PID.21 in the segments_nnn.xsd to allow more repeats in the field.

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  • Building InstallShield based Installers using Team Build 2010

    - by jehan
    Last few weeks, I have been working on Application Packaging stuff using all the widely used tools like InstallShield, WISE, WiX and Visual Studio Installer. So, I thought it would be good to post about how to Build the Installers developed using these tools with Team Build 2010. This post will focus on how to build the InstallShield generated packages using Team Build 2010. For the release of VS2010, Microsoft has partnered with Flexera who are the makers of InstallShield to create InstallShield Limited Edition, especially for the customers of Visual Studio. First Microsoft planned to release WiX (Windows Installer Xml) with VS2010, but later Microsoft dropped  WiX from VS2010 due to reasons which are best known to them and partnered with InstallShield for Limited Edition. It disappointed lot of people because InstallShield Limited Edition provides only few features of InstallShield and it may not feasable to build complex installer packages using this and it also requires License, where as WiX is an open source with no license costs and it has proved efficient in building most complex packages. Only the last three features are available in InstallShield Limited Edition from the total features offered by InstallShield as shown in below list.                                                                                            Feature Limited Edition for Visual Studio 2010 Standalone Build System Maintain a clean build machine by using only the part of InstallShield that compiles the installations. InstallShield Best Practices Validation Suite Avoid common installation issues. Try and Die Functionality RCreate a fully functional trial version of your product. InstallShield Repackager Create Windows Installer setups from any legacy installation. Multilingual Support Present installation text in up to 35 languages. Microsoft App-V™ Support Deploy your applications as App-V virtual packages that run without conflict. Industry-Standard InstallScript Achieve maximum flexibility in your installations. Dialog Editor Modify the layout of existing end-user dialogs, create new custom dialogs, and more. Patch Creation Build updates and patches for your products. Setup Prerequisite Editor Easily control prerequisite restart behavior and source locations. String Editor View Control the localizable text strings displayed at run time with this spreadsheet-like table. Text File Changes View Configure search-and-replace actions for content in text files to be modified at run time. Virtual Machine Detection Block your installations from running on virtual machines. Unicode Support Improve multi-language installation development. Support for 64-Bit COM Extraction Extract COM data from a 64-bit COM server. Windows Installer Installation Chaining Add MSI packages to your main installation and chain them together. XML Support Save time by quickly testing XML configuration changes to installation projects. Billboard Support for Custom Branding Display Adobe Flash billboards and other graphic files during the install process. SaaS Support (IIS 7 and SSL Technologies) Easily deploy Windows-based Web applications. Project Assistant Jumpstart a project by using a simplified set of views. Support for Digital Signatures Save time by digitally signing all your files at build time. Easily Run Custom Actions Schedule a custom action to run at precisely the right moment in your installation. Installation Prerequisites Check for and install prerequisites before your installation is executed. To create a InstallShield project in Visual Studio and Build it using Team Build 2010, first you have to add the InstallShield Project template  to your Solution file. If you want to use InstallShield Limited edition you can add it from FileàNewà project àother Project Types àSetup and Deploymentà InstallShield LE and if you are using other versions of InstallShield, then you have to add it from  from FileàNewà project àInstallShield Projects. Here, I’m using  InstallShield 2011 Premier edition as I already have it Installed. I have created a simple package for TailSpin Application which has a Feature called Web, few components and a IIS Web Site for  TailSpin application.   Before started working on this, I thought I may need to build the package by calling invoke process activity in build process template or have to create a new custom activity. But, it got build without any changes to build process template. But, it was failing with below error message. C:\Program Files (x86)\MSBuild\InstallShield\2011\InstallShield.targets (68): The "InstallShield.Tasks.InstallShield" task could not be loaded from the assembly C:\Program Files (x86)\MSBuild\InstallShield\2010Limited\InstallShield.Tasks.dll. Could not load file or assembly 'file:///C:\Program Files(x86)\MSBuild\InstallShield\2011\InstallShield.Tasks.dll' or one of its dependencies. An attempt was made to load a program with an incorrect format. Confirm that the <UsingTask> declaration is correct, that the assembly and all its dependencies are available, and that the task contains a public class that implements Microsoft.Build.Framework.ITask. This error is due to 64-bit build machine which I’m using. This issue will be replicable if you are queuing a build on a 64-bit build machine. To avoid this you have to ensure that you configured the build definition for your InstallShield project to load the InstallShield.Tasks.dll file (which is a 32-bit file); otherwise, you will encounter this build error informing you that the InstallShield.Tasks.dll file could not be loaded. To select the 32-bit version of MSBuild, click the Process tab of your build definition in Team Explorer. Then, under the Advanced node, find the MSBuild Platform setting, and select x86. Note that if you are using a 32-bit build machine, you can select either Auto or x86 for the MSBuild Platform setting.  Once I did above changes, the build got successful.

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  • ASP.NET Web API and Simple Value Parameters from POSTed data

    - by Rick Strahl
    In testing out various features of Web API I've found a few oddities in the way that the serialization is handled. These are probably not super common but they may throw you for a loop. Here's what I found. Simple Parameters from Xml or JSON Content Web API makes it very easy to create action methods that accept parameters that are automatically parsed from XML or JSON request bodies. For example, you can send a JavaScript JSON object to the server and Web API happily deserializes it for you. This works just fine:public string ReturnAlbumInfo(Album album) { return album.AlbumName + " (" + album.YearReleased.ToString() + ")"; } However, if you have methods that accept simple parameter types like strings, dates, number etc., those methods don't receive their parameters from XML or JSON body by default and you may end up with failures. Take the following two very simple methods:public string ReturnString(string message) { return message; } public HttpResponseMessage ReturnDateTime(DateTime time) { return Request.CreateResponse<DateTime>(HttpStatusCode.OK, time); } The first one accepts a string and if called with a JSON string from the client like this:var client = new HttpClient(); var result = client.PostAsJsonAsync<string>(http://rasxps/AspNetWebApi/albums/rpc/ReturnString, "Hello World").Result; which results in a trace like this: POST http://rasxps/AspNetWebApi/albums/rpc/ReturnString HTTP/1.1Content-Type: application/json; charset=utf-8Host: rasxpsContent-Length: 13Expect: 100-continueConnection: Keep-Alive "Hello World" produces… wait for it: null. Sending a date in the same fashion:var client = new HttpClient(); var result = client.PostAsJsonAsync<DateTime>(http://rasxps/AspNetWebApi/albums/rpc/ReturnDateTime, new DateTime(2012, 1, 1)).Result; results in this trace: POST http://rasxps/AspNetWebApi/albums/rpc/ReturnDateTime HTTP/1.1Content-Type: application/json; charset=utf-8Host: rasxpsContent-Length: 30Expect: 100-continueConnection: Keep-Alive "\/Date(1325412000000-1000)\/" (yes still the ugly MS AJAX date, yuk! This will supposedly change by RTM with Json.net used for client serialization) produces an error response: The parameters dictionary contains a null entry for parameter 'time' of non-nullable type 'System.DateTime' for method 'System.Net.Http.HttpResponseMessage ReturnDateTime(System.DateTime)' in 'AspNetWebApi.Controllers.AlbumApiController'. An optional parameter must be a reference type, a nullable type, or be declared as an optional parameter. Basically any simple parameters are not parsed properly resulting in null being sent to the method. For the string the call doesn't fail, but for the non-nullable date it produces an error because the method can't handle a null value. This behavior is a bit unexpected to say the least, but there's a simple solution to make this work using an explicit [FromBody] attribute:public string ReturnString([FromBody] string message) andpublic HttpResponseMessage ReturnDateTime([FromBody] DateTime time) which explicitly instructs Web API to read the value from the body. UrlEncoded Form Variable Parsing Another similar issue I ran into is with POST Form Variable binding. Web API can retrieve parameters from the QueryString and Route Values but it doesn't explicitly map parameters from POST values either. Taking our same ReturnString function from earlier and posting a message POST variable like this:var formVars = new Dictionary<string,string>(); formVars.Add("message", "Some Value"); var content = new FormUrlEncodedContent(formVars); var client = new HttpClient(); var result = client.PostAsync(http://rasxps/AspNetWebApi/albums/rpc/ReturnString, content).Result; which produces this trace: POST http://rasxps/AspNetWebApi/albums/rpc/ReturnString HTTP/1.1Content-Type: application/x-www-form-urlencodedHost: rasxpsContent-Length: 18Expect: 100-continue message=Some+Value When calling ReturnString:public string ReturnString(string message) { return message; } unfortunately it does not map the message value to the message parameter. This sort of mapping unfortunately is not available in Web API. Web API does support binding to form variables but only as part of model binding, which binds object properties to the POST variables. Sending the same message as in the previous example you can use the following code to pick up POST variable data:public string ReturnMessageModel(MessageModel model) { return model.Message; } public class MessageModel { public string Message { get; set; }} Note that the model is bound and the message form variable is mapped to the Message property as would other variables to properties if there were more. This works but it's not very dynamic. There's no real easy way to retrieve form variables (or query string values for that matter) in Web API's Request object as far as I can discern. Well only if you consider this easy:public string ReturnString() { var formData = Request.Content.ReadAsAsync<FormDataCollection>().Result; return formData.Get("message"); } Oddly FormDataCollection does not allow for indexers to work so you have to use the .Get() method which is rather odd. If you're running under IIS/Cassini you can always resort to the old and trusty HttpContext access for request data:public string ReturnString() { return HttpContext.Current.Request.Form["message"]; } which works fine and is easier. It's kind of a bummer that HttpRequestMessage doesn't expose some sort of raw Request object that has access to dynamic data - given that it's meant to serve as a generic REST/HTTP API that seems like a crucial missing piece. I don't see any way to read query string values either. To me personally HttpContext works, since I don't see myself using self-hosted code much.© Rick Strahl, West Wind Technologies, 2005-2012Posted in Web Api   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); })();

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  • Workarounds for supporting MVVM in the Silverlight TreeView Control

    - by cibrax
    MVVM (Model-View-ViewModel) is the pattern that you will typically choose for building testable user interfaces either in WPF or Silverlight. This pattern basically relies on the data binding support in those two technologies for mapping an existing model class (the view model) to the different parts of the UI or view. Unfortunately, MVVM was not threated as first citizen for some of controls released out of the box in the Silverlight runtime or the Silverlight toolkit. That means that using data binding for implementing MVVM is not always something trivial and usually requires some customization in the existing controls. In ran into different problems myself trying to fully support data binding in controls like the tree view or the context menu or things like drag & drop.  For that reason, I decided to write this post to show how the tree view control or the tree view items can be customized to support data binding in many of its properties. In first place, you will typically use a tree view for showing hierarchical data so the view model somehow must reflect that hierarchy. An easy way to implement hierarchy in a model is to use a base item element like this one, public abstract class TreeItemModel { public abstract IEnumerable<TreeItemModel> Children; } You can later derive your concrete model classes from that base class. For example, public class CustomerModel { public string FullName { get; set; } public string Address { get; set; } public IEnumerable<OrderModel> Orders { get; set; } }   public class CustomerTreeItemModel : TreeItemModel { public CustomerTreeItemModel(CustomerModel customer) { }   public override IEnumerable<TreeItemModel> Children { get { // Return orders } } } The Children property in the CustomerTreeItem model implementation can return for instance an ObservableCollection<TreeItemModel> with the orders, so the tree view will automatically subscribe to all the changes in the collection. You can bind this model to the tree view control in the UI by using a Hierarchical data template. <e:TreeView x:Name="TreeView" ItemsSource="{Binding Customers}"> <e:TreeView.ItemTemplate> <sdk:HierarchicalDataTemplate ItemsSource="{Binding Children}"> <!-- TEMPLATE --> </sdk:HierarchicalDataTemplate> </e:TreeView.ItemTemplate> </e:TreeView> An interesting behavior with the Children property and the Hierarchical data template is that the Children property is only invoked before the expansion, so you can use lazy load at this point (The tree view control will not expand the whole tree in the first expansion). The problem with using MVVM in this control is that you can not bind properties in model with specific properties of the TreeView item such as IsSelected or IsExpanded. Here is where you need to customize the existing tree view control to support data binding in tree items. public class CustomTreeView : TreeView { public CustomTreeView() { }   protected override DependencyObject GetContainerForItemOverride() { CustomTreeViewItem tvi = new CustomTreeViewItem(); Binding expandedBinding = new Binding("IsExpanded"); expandedBinding.Mode = BindingMode.TwoWay; tvi.SetBinding(CustomTreeViewItem.IsExpandedProperty, expandedBinding); Binding selectedBinding = new Binding("IsSelected"); selectedBinding.Mode = BindingMode.TwoWay; tvi.SetBinding(CustomTreeViewItem.IsSelectedProperty, selectedBinding); return tvi; } }   public class CustomTreeViewItem : TreeViewItem { public CustomTreeViewItem() { }   protected override DependencyObject GetContainerForItemOverride() { CustomTreeViewItem tvi = new CustomTreeViewItem(); Binding expandedBinding = new Binding("IsExpanded"); expandedBinding.Mode = BindingMode.TwoWay; tvi.SetBinding(CustomTreeViewItem.IsExpandedProperty, expandedBinding); Binding selectedBinding = new Binding("IsSelected"); selectedBinding.Mode = BindingMode.TwoWay; tvi.SetBinding(CustomTreeViewItem.IsSelectedProperty, selectedBinding); return tvi; } } You basically need to derive the TreeView and TreeViewItem controls to manually add a binding for the properties you need. In the example above, I am adding a binding for the “IsExpanded” and “IsSelected” properties in the items. The model for the tree items now needs to be extended to support those properties as well, public abstract class TreeItemModel : INotifyPropertyChanged { bool isExpanded = false; bool isSelected = false;   public abstract IEnumerable<TreeItemModel> Children { get; }   public bool IsExpanded { get { return isExpanded; } set { isExpanded = value; if (PropertyChanged != null) PropertyChanged(this, new PropertyChangedEventArgs("IsExpanded")); } }   public bool IsSelected { get { return isSelected; } set { isSelected = value; if (PropertyChanged != null) PropertyChanged(this, new PropertyChangedEventArgs("IsSelected")); } }   public event PropertyChangedEventHandler PropertyChanged; } However, as soon as you use this custom tree view control, you lose all the automatic styles from the built-in toolkit themes because they are tied to the control type (TreeView in this case).  The only ugly workaround I found so far for this problem is to copy the styles from the Toolkit source code and reuse them in the application.

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  • Alien deletes .deb when converting from .rpm

    - by Stann
    I'm trying to convert .rpm to .deb using alien. sudo alien -k libtetra-1.0.0-2.i386.rpm Alien says that: libtetra-1.0.0-2.i386.deb generated But when I check the folder - there is just original .rpm and no .deb. Also - I can see that for a split second there is a .deb file in a folder. so it looks like alien create .deb and deletes it right away. I suspect that it's maybe because I run 64 bit os and package is 32? Can somebody explain why alien deletes .deb automatically? Verbose output: LANG=C rpm -qp --queryformat %{NAME} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{VERSION} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{RELEASE} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{ARCH} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{CHANGELOGTEXT} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{SUMMARY} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{DESCRIPTION} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{PREFIXES} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{POSTIN} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{POSTUN} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{PREUN} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{LICENSE} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qp --queryformat %{PREIN} libtetra-1.0.0-2.i386.rpm LANG=C rpm -qcp libtetra-1.0.0-2.i386.rpm rpm -qpi libtetra-1.0.0-2.i386.rpm LANG=C rpm -qpl libtetra-1.0.0-2.i386.rpm mkdir libtetra-1.0.0 chmod 755 libtetra-1.0.0 rpm2cpio libtetra-1.0.0-2.i386.rpm | lzma -t -q > /dev/null 2>&1 rpm2cpio libtetra-1.0.0-2.i386.rpm | (cd libtetra-1.0.0; cpio --extract --make-directories --no-absolute-filenames --preserve-modification-time) 2>&1 chmod 755 libtetra-1.0.0/./ chmod 755 libtetra-1.0.0/./usr chmod 755 libtetra-1.0.0/./usr/lib chown 0:0 libtetra-1.0.0//usr/lib/libtetra.so.1.0.0 chmod 755 libtetra-1.0.0//usr/lib/libtetra.so.1.0.0 mkdir libtetra-1.0.0/debian date -R date -R chmod 755 libtetra-1.0.0/debian/rules debian/rules binary 2>&1 libtetra_1.0.0-3_i386.deb generated find libtetra-1.0.0 -type d -exec chmod 755 {} ; rm -rf libtetra-1.0.0 Very Verbose output LANG=C rpm -qp --queryformat %{NAME} libtetra-1.0.0-2.i386.rpm libtetra LANG=C rpm -qp --queryformat %{VERSION} libtetra-1.0.0-2.i386.rpm 1.0.0 LANG=C rpm -qp --queryformat %{RELEASE} libtetra-1.0.0-2.i386.rpm 2 LANG=C rpm -qp --queryformat %{ARCH} libtetra-1.0.0-2.i386.rpm i386 LANG=C rpm -qp --queryformat %{CHANGELOGTEXT} libtetra-1.0.0-2.i386.rpm - First RPM Package LANG=C rpm -qp --queryformat %{SUMMARY} libtetra-1.0.0-2.i386.rpm Panasonic KX-MC6000 series Printer Driver for Linux. LANG=C rpm -qp --queryformat %{DESCRIPTION} libtetra-1.0.0-2.i386.rpm This software is Panasonic KX-MC6000 series Printer Driver for Linux. You can print from applications by using CUPS(Common Unix Printing System) which is the printing system for Linux. Other functions for KX-MC6000 series are not supported by this software. LANG=C rpm -qp --queryformat %{PREFIXES} libtetra-1.0.0-2.i386.rpm (none) LANG=C rpm -qp --queryformat %{POSTIN} libtetra-1.0.0-2.i386.rpm (none) LANG=C rpm -qp --queryformat %{POSTUN} libtetra-1.0.0-2.i386.rpm (none) LANG=C rpm -qp --queryformat %{PREUN} libtetra-1.0.0-2.i386.rpm (none) LANG=C rpm -qp --queryformat %{LICENSE} libtetra-1.0.0-2.i386.rpm GPL and LGPL (Version2) LANG=C rpm -qp --queryformat %{PREIN} libtetra-1.0.0-2.i386.rpm (none) LANG=C rpm -qcp libtetra-1.0.0-2.i386.rpm rpm -qpi libtetra-1.0.0-2.i386.rpm Name : libtetra Relocations: (not relocatable) Version : 1.0.0 Vendor: Panasonic Communications Co., Ltd. Release : 2 Build Date: Tue 27 Apr 2010 05:16:40 AM EDT Install Date: (not installed) Build Host: localhost.localdomain Group : System Environment/Daemons Source RPM: libtetra-1.0.0-2.src.rpm Size : 31808 License: GPL and LGPL (Version2) Signature : (none) URL : http://panasonic.net/pcc/support/fax/world.htm Summary : Panasonic KX-MC6000 series Printer Driver for Linux. Description : This software is Panasonic KX-MC6000 series Printer Driver for Linux. You can print from applications by using CUPS(Common Unix Printing System) which is the printing system for Linux. Other functions for KX-MC6000 series are not supported by this software. LANG=C rpm -qpl libtetra-1.0.0-2.i386.rpm /usr/lib/libtetra.so /usr/lib/libtetra.so.1.0.0 mkdir libtetra-1.0.0 chmod 755 libtetra-1.0.0 rpm2cpio libtetra-1.0.0-2.i386.rpm | lzma -t -q > /dev/null 2>&1 rpm2cpio libtetra-1.0.0-2.i386.rpm | (cd libtetra-1.0.0; cpio --extract --make-directories --no-absolute-filenames --preserve-modification-time) 2>&1 63 blocks chmod 755 libtetra-1.0.0/./ chmod 755 libtetra-1.0.0/./usr chmod 755 libtetra-1.0.0/./usr/lib chown 0:0 libtetra-1.0.0//usr/lib/libtetra.so.1.0.0 chmod 755 libtetra-1.0.0//usr/lib/libtetra.so.1.0.0 mkdir libtetra-1.0.0/debian date -R Mon, 07 Feb 2011 11:03:58 -0500 date -R Mon, 07 Feb 2011 11:03:58 -0500 chmod 755 libtetra-1.0.0/debian/rules debian/rules binary 2>&1 dh_testdir dh_testdir dh_testroot dh_clean -k -d dh_clean: No packages to build. dh_installdirs dh_installdocs dh_installchangelogs find . -maxdepth 1 -mindepth 1 -not -name debian -print0 | \ xargs -0 -r -i cp -a {} debian/ dh_compress dh_makeshlibs dh_installdeb dh_shlibdeps dh_gencontrol dh_md5sums dh_builddeb libtetra_1.0.0-2_i386.deb generated find libtetra-1.0.0 -type d -exec chmod 755 {} ; rm -rf libtetra-1.0.0 Resolution Oh well. It looks like it's perhaps a bug? or I don't know. I simply installed 32-bit version of Ubuntu in VirtualBox and converted package there. For some reason I couldn't convert 32-bit package in 64 OS. and that is that. If someone ever finds the reason ffor this behavior - plz. post somewhere in comments. Thanks

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  • What Did You Do? is a Bad Question

    - by Ajarn Mark Caldwell
    Brian Moran (blog | Twitter) did a great presentation today for the PASS Professional Development Virtual Chapter on The Art of Questions.  One of the points that Brian made was that there are good questions and bad (or at least not-as-good) questions.  Good questions tend to open-up the conversation and engender positive reactions (perhaps even trust and respect) between the participants; and bad questions tend to close-down a conversation either through the narrow list of possible responses (e.g. strictly Yes/No) or through the negative reactions they can produce.  And this explains why I so frequently had problems troubleshooting real-time problems with users in the past.  I’ll explain that in more detail below, but before we go on, let me recommend that you watch the recording of Brian’s presentation to learn why the question Why is often problematic in the U.S. and yet we so often resort to it. For a short portion (3 years) of my career, I taught basic computer skills and Office applications in an adult vocational school, and this gave me ample opportunity to do live troubleshooting of user challenges with computers.  And like many people who ended up in computer related jobs, I also have had numerous times where I was called upon by less computer-savvy individuals to help them with some challenge they were having, whether it was part of my job or not.  One of the things that I noticed, especially during my time as a teacher, was that when I was helping somebody, typically the first question I would ask them was, “What did you do?”  This seemed to me like a good way to start my detective work trying to figure out what happened, what went wrong, how to fix it, and how to help the person avoid it again in the future.  I always asked it in a polite tone of voice as I was just trying to gather the facts before diving in deeper.  However; 99.999% of the time, I always got the same answer, “Nothing!”  For a long time this frustrated me because (remember I’m in detective mode at that point) I knew it could not possibly be true.  They HAD to have done SOMETHING…just tell me what were the last actions you took before this problem presented itself.  But no, they always stuck with “Nothing”.  At which point, with frustration growing, and not a little bit of disdain for their lack of helpfulness, I would usually ask them to move aside while I took over their machine and got them out of whatever they had gotten themselves into.  After a while I just grew used to the fact that this was the answer I would usually receive, but I always kept asking because for the .001% of the people who would actually tell me, I could then help them understand what went wrong and how to avoid it in the future. Now, after hearing Brian’s talk, I understand what the problem was.  Even though I meant to just be in an information gathering mode, the words I was using, “What did YOU do?” have such a strong negative connotation that people would instinctively go into defense-mode and stop sharing information that might make them look bad.  Many of them probably were not even consciously aware that they had gone on the defensive, but the self-preservation instinct, especially self-preservation of the ego, is so strong that people would end up there without even realizing it. So, if “What did you do” is a bad question, what would have been better?  Well, one suggestion that Brian makes in his talk is something along the lines of, “Can you tell me what led up to this?” or “what was happening on the computer right before this came up?”  It’s subtle, but the point is to take the focus off of the person and their behavior; instead depersonalizing it and talk about events from more of a 3rd-party observer point of view.  With this approach, people will be more likely to talk about what the computer did and what they did in response to it without feeling the interrogation spotlight is on them.  They are also more likely to mention other events that occurred around the same time that may or may not be related, but which could certainly help you troubleshoot a larger problem if it is not just user actions.  And that is the ultimate goal of your asking the questions.  So yes, it does matter how you ask the question; and there are such things as good questions and bad questions.  Excellent topic Brian!  Thanks for getting the thinking gears churning! (Cross-posted to the Professional Development Virtual Chapter blog.)

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  • AS11 Oracle B2B Sync Support - Series 2

    - by sinkarbabu.kirubanithi
    In the earlier series, we discussed about how to model "Sync Support" in Oracle B2B. And, we haven't discussed how the response can be consumed synchronously by the back-end application or initiator of sync request. In this sequel, we will see how we can extend it to the SOA composite applications to model the end-to-end usecase, this would help the initiator of sync request to receive the response synchronously. Series 2 - is little lengthier for blog standards so be prepared before you continue further :). Let's start our discussion with a high-level scenario where one need to initiate a synchronous request and get response synchronously. There are various approaches available, we will see one simplest approach here. Components Involved: 1. Oracle B2B 2. Oracle JCA JMS Adapter 3. Oracle BPEL 4. All of the above are wrapped up in a single SOA composite application. Oracle B2B: Skipping the "Sync Support" setup part in B2B, as we have already discussed that in the earlier series 1. Here we have provided "Sync Support" samples that can be imported to B2B directly and users can start testing the same in few minutes. Initiator Sample: This requires two JMS queues to be created, one for B2B to receive initial outbound sync request and the other is for B2B to deliver the incoming sync response to the back-end. Please enable "Use JMS Id" option in both internal listening and delivery channels. This would enable JCA JMS Adapter to correlate the initial B2B request and response and in turn it would be returned as synchronous response of BPEL. Internal Listening Channel Image: Internal Delivery Channel Image: To get going without much challenges, just create queues in Weblogic with the JNDI mentioned in the above two screenshots. If you want to use different names, then you may have to change the queue jndi names in sample after importing it into B2B. Here are the Queue related JNDI names used in the sample, 1. Internal Listening Channel Queue details, Name: JNDI Name: jms/b2b/syncreplyqueue 2. Internal Delivery Channel Queue details, Name: JNDI Name: jms/b2b/syncrequestqueue Here is the Initiator Sample Acme.zip Note: You may have to adjust the ip address of GlobalChips endpoint in the Delivery Channel. Responder Sample: Contains B2B meta-data and the Callout. Just import the sample and place the callout binary under "/tmp/callout" directory. If you choose to use a different location for callout, then you may have to change the same in B2B Configuration after importing the sample. Here are the artifacts, 1. Callout Source SampleCallout.java 2. Callout Binary sample-callout.jar 3. Responder Sample GlobalChips.zip Callout Details: Just gives the static response XML that needs to be sent back as response for the inbound sync request. For a sample purpose, we have given static response but in production you may have to invoke a web service or something similar to get the response. IMPORTANT NOTE: For Sync Support use case, responder is not expected to deliver the inbound sync request to backend as the process of delivering and getting the response from backend are expected from the Callout. This default behavior can be overridden by enabling the config property "b2b.SyncAppDelivery=true" in B2B config mbean (b2b-config.xml). This makes B2B to deliver the inbound sync request to be delivered to backend queue but the response to be sent to remote caller still has to come from Callout. 2. Oracle JCA JMS Adapter: On the initiator side, we have used JCA JMS Request/Reply pattern to send/receive the synchronous message from B2B. 3. Oracle BPEL: Exposes WS-SOAP Endpoint that takes payload as input and passes the same to B2B and returns the synchronous response of B2B as SOAP response. For outside world, it looks as if it is the synchronous web service endpoint but under the cover it uses JMS to trigger/initiate B2B to send and receive the synchronous response. 4. Composite application: All the components discussed above are wired in SOA composite application that helps to model a end-to-end synchronous use case. Here's the composite application sca_B2BSyncSample_rev1.0.jar, you may just deploy this to your AS11 SOA to make use of it. For any editing, you can just import the project in your JDEV under any SOA Application. Here are the composite application screenshots, Composite Application: BPEL With JCA JMS Adapter (Request/Reply):

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  • Can't log in to GNOME after upgrade (raring -> saucy)

    - by x-yuri
    I've just upgraded my ubuntu (raring to saucy) and I now can't log in to GNOME. As opposed to virtual consoles (Ctrl-Alt-F1, for example). I set it up to log in automatically. But it asks for password now. I type in the password, press Enter, the screen blinks and here I am again at the login screen. Then I looked into /var/log/Xorg.0.log: [ 33.956] Initializing built-in extension DRI2 [ 33.956] (II) LoadModule: "glx" [ 33.956] (II) Loading /usr/lib/xorg/modules/extensions/libglx.so [ 33.956] (II) Module glx: vendor="X.Org Foundation" [ 33.956] compiled for 1.14.3, module version = 1.0.0 [ 33.956] ABI class: X.Org Server Extension, version 7.0 [ 33.956] (==) AIGLX enabled [ 33.956] Loading extension GLX [ 33.956] (==) Matched fglrx as autoconfigured driver 0 [ 33.956] (==) Matched ati as autoconfigured driver 1 [ 33.956] (==) Matched fglrx as autoconfigured driver 2 [ 33.956] (==) Matched ati as autoconfigured driver 3 [ 33.956] (==) Matched vesa as autoconfigured driver 4 [ 33.956] (==) Matched modesetting as autoconfigured driver 5 [ 33.956] (==) Matched fbdev as autoconfigured driver 6 [ 33.956] (==) Assigned the driver to the xf86ConfigLayout [ 33.956] (II) LoadModule: "fglrx" [ 33.957] (WW) Warning, couldn't open module fglrx [ 33.957] (II) UnloadModule: "fglrx" [ 33.957] (II) Unloading fglrx [ 33.957] (EE) Failed to load module "fglrx" (module does not exist, 0) [ 33.957] (II) LoadModule: "ati" [ 33.957] (WW) Warning, couldn't open module ati [ 33.957] (II) UnloadModule: "ati" [ 33.957] (II) Unloading ati [ 33.957] (EE) Failed to load module "ati" (module does not exist, 0) [ 33.957] (II) LoadModule: "vesa" [ 33.957] (II) Loading /usr/lib/xorg/modules/drivers/vesa_drv.so [ 33.957] (II) Module vesa: vendor="X.Org Foundation" [ 33.957] compiled for 1.14.1, module version = 2.3.2 [ 33.957] Module class: X.Org Video Driver [ 33.957] ABI class: X.Org Video Driver, version 14.1 [ 33.957] (II) LoadModule: "modesetting" [ 33.957] (II) Loading /usr/lib/xorg/modules/drivers/modesetting_drv.so [ 33.957] (II) Module modesetting: vendor="X.Org Foundation" [ 33.957] compiled for 1.14.1, module version = 0.8.0 [ 33.957] Module class: X.Org Video Driver [ 33.957] ABI class: X.Org Video Driver, version 14.1 [ 33.957] (II) LoadModule: "fbdev" [ 33.957] (II) Loading /usr/lib/xorg/modules/drivers/fbdev_drv.so [ 33.958] (II) Module fbdev: vendor="X.Org Foundation" [ 33.958] compiled for 1.14.1, module version = 0.4.3 [ 33.958] Module class: X.Org Video Driver [ 33.958] ABI class: X.Org Video Driver, version 14.1 [ 33.958] (==) Matched fglrx as autoconfigured driver 0 [ 33.958] (==) Matched ati as autoconfigured driver 1 [ 33.958] (==) Matched fglrx as autoconfigured driver 2 [ 33.958] (==) Matched ati as autoconfigured driver 3 [ 33.958] (==) Matched vesa as autoconfigured driver 4 [ 33.958] (==) Matched modesetting as autoconfigured driver 5 [ 33.958] (==) Matched fbdev as autoconfigured driver 6 [ 33.958] (==) Assigned the driver to the xf86ConfigLayout [ 33.958] (II) LoadModule: "fglrx" [ 33.958] (WW) Warning, couldn't open module fglrx [ 33.958] (II) UnloadModule: "fglrx" [ 33.958] (II) Unloading fglrx [ 33.958] (EE) Failed to load module "fglrx" (module does not exist, 0) [ 33.958] (II) LoadModule: "ati" [ 33.958] (WW) Warning, couldn't open module ati [ 33.958] (II) UnloadModule: "ati" [ 33.958] (II) Unloading ati [ 33.958] (EE) Failed to load module "ati" (module does not exist, 0) [ 33.958] (II) LoadModule: "vesa" [ 33.958] (II) Loading /usr/lib/xorg/modules/drivers/vesa_drv.so [ 33.958] (II) Module vesa: vendor="X.Org Foundation" [ 33.958] compiled for 1.14.1, module version = 2.3.2 [ 33.958] Module class: X.Org Video Driver [ 33.958] ABI class: X.Org Video Driver, version 14.1 [ 33.958] (II) UnloadModule: "vesa" [ 33.958] (II) Unloading vesa [ 33.958] (II) Failed to load module "vesa" (already loaded, 0) [ 33.958] (II) LoadModule: "modesetting" [ 33.959] (II) Loading /usr/lib/xorg/modules/drivers/modesetting_drv.so [ 33.959] (II) Module modesetting: vendor="X.Org Foundation" [ 33.959] compiled for 1.14.1, module version = 0.8.0 [ 33.959] Module class: X.Org Video Driver [ 33.959] ABI class: X.Org Video Driver, version 14.1 [ 33.959] (II) UnloadModule: "modesetting" [ 33.959] (II) Unloading modesetting [ 33.959] (II) Failed to load module "modesetting" (already loaded, 0) [ 33.959] (II) LoadModule: "fbdev" [ 33.959] (II) Loading /usr/lib/xorg/modules/drivers/fbdev_drv.so [ 33.959] (II) Module fbdev: vendor="X.Org Foundation" [ 33.959] compiled for 1.14.1, module version = 0.4.3 [ 33.959] Module class: X.Org Video Driver [ 33.959] ABI class: X.Org Video Driver, version 14.1 [ 33.959] (II) UnloadModule: "fbdev" [ 33.959] (II) Unloading fbdev [ 33.959] (II) Failed to load module "fbdev" (already loaded, 0) [ 33.959] (II) VESA: driver for VESA chipsets: vesa [ 33.959] (II) modesetting: Driver for Modesetting Kernel Drivers: kms [ 33.959] (II) FBDEV: driver for framebuffer: fbdev [ 33.959] (++) using VT number 7 If I install fglrx, it reads: [ 37.152] Initializing built-in extension DRI2 [ 37.152] (II) LoadModule: "glx" [ 37.152] (II) Loading /usr/lib/x86_64-linux-gnu/xorg/extra-modules/modules/extensions/libglx.so [ 37.152] (II) Module glx: vendor="Advanced Micro Devices, Inc." [ 37.152] compiled for 6.9.0, module version = 1.0.0 [ 37.152] Loading extension GLX [ 37.153] (==) Matched fglrx as autoconfigured driver 0 [ 37.153] (==) Matched ati as autoconfigured driver 1 [ 37.153] (==) Matched vesa as autoconfigured driver 2 [ 37.153] (==) Matched modesetting as autoconfigured driver 3 [ 37.153] (==) Matched fbdev as autoconfigured driver 4 [ 37.153] (==) Assigned the driver to the xf86ConfigLayout [ 37.153] (II) LoadModule: "fglrx" [ 37.153] (II) Loading /usr/lib/x86_64-linux-gnu/xorg/extra-modules/modules/drivers/fglrx_drv.so [ 37.168] (II) Module fglrx: vendor="FireGL - AMD Technologies Inc." [ 37.168] compiled for 1.4.99.906, module version = 13.10.10 [ 37.168] Module class: X.Org Video Driver [ 37.168] (II) Loading sub module "fglrxdrm" [ 37.168] (II) LoadModule: "fglrxdrm" [ 37.168] (II) Loading /usr/lib/x86_64-linux-gnu/xorg/extra-modules/modules/linux/libfglrxdrm.so [ 37.169] (II) Module fglrxdrm: vendor="FireGL - AMD Technologies Inc." [ 37.169] compiled for 1.4.99.906, module version = 13.10.10 [ 37.169] (II) LoadModule: "ati" [ 37.169] (WW) Warning, couldn't open module ati [ 37.169] (II) UnloadModule: "ati" [ 37.169] (II) Unloading ati [ 37.169] (EE) Failed to load module "ati" (module does not exist, 0) [ 37.169] (II) LoadModule: "vesa" [ 37.169] (II) Loading /usr/lib/xorg/modules/drivers/vesa_drv.so [ 37.169] (II) Module vesa: vendor="X.Org Foundation" [ 37.169] compiled for 1.14.1, module version = 2.3.2 [ 37.169] Module class: X.Org Video Driver [ 37.169] ABI class: X.Org Video Driver, version 14.1 [ 37.169] (II) LoadModule: "modesetting" [ 37.170] (II) Loading /usr/lib/xorg/modules/drivers/modesetting_drv.so [ 37.170] (II) Module modesetting: vendor="X.Org Foundation" [ 37.170] compiled for 1.14.1, module version = 0.8.0 [ 37.170] Module class: X.Org Video Driver [ 37.170] ABI class: X.Org Video Driver, version 14.1 [ 37.170] (II) LoadModule: "fbdev" [ 37.170] (II) Loading /usr/lib/xorg/modules/drivers/fbdev_drv.so [ 37.170] (II) Module fbdev: vendor="X.Org Foundation" [ 37.170] compiled for 1.14.1, module version = 0.4.3 [ 37.170] Module class: X.Org Video Driver [ 37.170] ABI class: X.Org Video Driver, version 14.1 [ 37.170] (==) Matched fglrx as autoconfigured driver 0 [ 37.170] (==) Matched ati as autoconfigured driver 1 [ 37.170] (==) Matched vesa as autoconfigured driver 2 [ 37.170] (==) Matched modesetting as autoconfigured driver 3 [ 37.170] (==) Matched fbdev as autoconfigured driver 4 [ 37.170] (==) Assigned the driver to the xf86ConfigLayout [ 37.170] (II) LoadModule: "fglrx" [ 37.170] (II) Loading /usr/lib/x86_64-linux-gnu/xorg/extra-modules/modules/drivers/fglrx_drv.so [ 37.170] (II) Module fglrx: vendor="FireGL - AMD Technologies Inc." [ 37.170] compiled for 1.4.99.906, module version = 13.10.10 [ 37.170] Module class: X.Org Video Driver [ 37.170] (II) LoadModule: "ati" [ 37.170] (WW) Warning, couldn't open module ati [ 37.170] (II) UnloadModule: "ati" [ 37.171] (II) Unloading ati [ 37.171] (EE) Failed to load module "ati" (module does not exist, 0) [ 37.171] (II) LoadModule: "vesa" [ 37.171] (II) Loading /usr/lib/xorg/modules/drivers/vesa_drv.so [ 37.171] (II) Module vesa: vendor="X.Org Foundation" [ 37.171] compiled for 1.14.1, module version = 2.3.2 [ 37.171] Module class: X.Org Video Driver [ 37.171] ABI class: X.Org Video Driver, version 14.1 [ 37.171] (II) UnloadModule: "vesa" [ 37.171] (II) Unloading vesa [ 37.171] (II) Failed to load module "vesa" (already loaded, 0) [ 37.171] (II) LoadModule: "modesetting" [ 37.171] (II) Loading /usr/lib/xorg/modules/drivers/modesetting_drv.so [ 37.171] (II) Module modesetting: vendor="X.Org Foundation" [ 37.171] compiled for 1.14.1, module version = 0.8.0 [ 37.171] Module class: X.Org Video Driver [ 37.171] ABI class: X.Org Video Driver, version 14.1 [ 37.171] (II) UnloadModule: "modesetting" [ 37.171] (II) Unloading modesetting [ 37.171] (II) Failed to load module "modesetting" (already loaded, 0) [ 37.171] (II) LoadModule: "fbdev" [ 37.171] (II) Loading /usr/lib/xorg/modules/drivers/fbdev_drv.so [ 37.171] (II) Module fbdev: vendor="X.Org Foundation" [ 37.171] compiled for 1.14.1, module version = 0.4.3 [ 37.171] Module class: X.Org Video Driver [ 37.171] ABI class: X.Org Video Driver, version 14.1 [ 37.171] (II) UnloadModule: "fbdev" [ 37.171] (II) Unloading fbdev [ 37.171] (II) Failed to load module "fbdev" (already loaded, 0) [ 37.171] (II) AMD Proprietary Linux Driver Version Identifier:13.10.10 [ 37.171] (II) AMD Proprietary Linux Driver Release Identifier: UNSUPPORTED-13.101 [ 37.171] (II) AMD Proprietary Linux Driver Build Date: May 23 2013 15:49:35 [ 37.171] (II) VESA: driver for VESA chipsets: vesa [ 37.171] (II) modesetting: Driver for Modesetting Kernel Drivers: kms [ 37.171] (II) FBDEV: driver for framebuffer: fbdev [ 37.171] (++) using VT number 7 I did more installing/removing packages than that. There were a moment when it said: (EE) Failed to load /usr/lib64/xorg/modules/libglamoregl.so: /usr/lib64/xorg/modules/libglamoregl.so: undefined symbol: _glapi_tls_Context Also there is init: not found in ~/.xsession-errors: /usr/sbin/lightdm-session: 5: exec: init: not found Actually, I'm out of ideas. What about you? :)

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  • Code Structure / Level Design: Plants vs Zombies game level dissection

    - by lalan
    Hi Friends, I am interested in learning the class structure of Plants vs Zombies, particularly level design; for those who haven't played it - this video contains nice play-through: http://www.youtube.com/watch?v=89DfdOIJ4xw. How would I go ahead and design the code, mostly structure & classes, which allows for maximum flexibility & clean development? I am familiar with data driven design concepts, and would use events to handle most of dynamic behavior. Dissection at macro level: (Once every Level) Load tilemap, props, etc -- basically build the map (Once every Level) Camera Movement - might consider it as short cut-scene (Once every Level) Show Enemies you'll face during present level (Once every Level) Unit Selection Window/Panel - selection of defensive plants (Once every Level) Camera Movement - might consider it as short cut-scene (Once every Level) HUD Creation - based on unit selection (Level Loop) Enemy creation - based on types of zombies allowed (Level Loop) Sun/Resource generation (Level Loop) Show messages like 'huge wave of zombies coming', 'final wave' (Level Loop) Other unique events - Spawn gifts, money, tombstones, etc (Once every Level) Unlock new plant Potential game scripts: a) Level definitions: Level_1_1.xml, Level_1_2.xml, etc. Level_1_1.xml :: Sample script <map> <tilemap>tilemapFrontLawn</tilemap> <SpawnPoints> tiles where particular type of zombies (land vs water) may spawn</spawnPoints> <props> position, entity array -- lawnmower, </props> </map> <zombies> <... list of zombies who gonna attack by ids...> </zombies> <plants> <... list by plants which are available for defense by ids...> </plants> <progression> <ZombieWave name='first wave' spawnScript='zombieLightWave.lua' unlock='null'> <startMessages time=1.5>Ready</startMessages> <endMessages time=1.5>Huge wave of zombies incoming</endMessages> </ZombieWave> </progression> b) Entities definitions: .xmls containing zombies, plants, sun, lawnmower, coins, etc description. Potential classes: //LevelManager - Based on the level under play, it will load level script. Few of the // functions it may have: class LevelManager { public: bool load(string levelFileName); bool enter(); bool update(float deltatime); bool exit(); private: LevelData* mLevelData; } // LevelData - Contains the details of level loaded by LevelManager. class LevelData { private: string file; // array of camera,dialog,attackwaves, etc in active level LevelCutSceneCamera** mArrayCutSceneCamera; LevelCutSceneDialog** mArrayCutSceneDialog; LevelAttackWave** mArrayAttackWave; .... // which camera,dialog,attackwave is active in level uint mCursorCutSceneCamera; uint mCursorCutSceneDialog; uint mCursorAttackWave; public: // based on cursor, get the next camera,dialog,attackwave,etc in active level // return false/true based on failure/success bool nextCutSceneCamera(LevelCutSceneCamera**); bool nextCutSceneDialog(LevelCutSceneDialog**); } // LevelUnderPlay- LevelManager class LevelUnderPlay { private: LevelCutSceneCamera* mCutSceneCamera; LevelCutSceneDialog* mCutSceneDialog; LevelAttackWave* mAttackWave; Entities** mSelectedPlants; Entities** mAllowedZombies; bool isCutSceneCameraActive; public: bool enter(); bool update(float deltatime); bool exit(); } I am totally confused.. :( Does it make sense of using class composition (have flat class hierarchy) for managing levels. Is it a good idea to just add/remove/update sprites (or any drawable stuff) to current scene from LevelManager or LevelUnderPlay? If I want to make non-linear level design, how should I go ahead? Perhaps I would need a LevelProgression class, which would decide what to do based on decision tree. Any suggestions would be appreciated very much. Thank for your time, lalan

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  • Remote Debug Windows Azure Cloud Service

    - by Shaun
    Originally posted on: http://geekswithblogs.net/shaunxu/archive/2013/11/02/remote-debug-windows-azure-cloud-service.aspxOn the 22nd of October Microsoft Announced the new Windows Azure SDK 2.2. It introduced a lot of cool features but one of it shocked most, which is the remote debug support for Windows Azure Cloud Service (a.k.a. WACS).   Live Debug is Nightmare for Cloud Application When we are developing against public cloud, debug might be the most difficult task, especially after the application had been deployed. In order to minimize the debug effort, Microsoft provided local emulator for cloud service and storage once the Windows Azure platform was announced. By using local emulator developers could be able run their application on local machine with almost the same behavior as running on Windows Azure, and that could be debug easily and quickly. But when we deployed our application to Azure, we have to use log, diagnostic monitor to debug, which is very low efficient. Visual Studio 2012 introduced a new feature named "anonymous remote debug" which allows any workstation under any user could be able to attach the remote process. This is less secure comparing the authenticated remote debug but much easier and simpler to use. Now in Windows Azure SDK 2.2, we could be able to attach our application from our local machine to Windows Azure, and it's very easy.   How to Use Remote Debugger First, let's create a new Windows Azure Cloud Project in Visual Studio and selected ASP.NET Web Role. Then create an ASP.NET WebForm application. Then right click on the cloud project and select "publish". In the publish dialog we need to make sure the application will be built in debug mode, since .NET assembly cannot be debugged in release mode. I enabled Remote Desktop as I will log into the virtual machine later in this post. It's NOT necessary for remote debug. And selected "advanced settings" tab, make sure we checked "Enable Remote Debugger for all roles". In WACS, a cloud service could be able to have one or more roles and each role could be able to have one or more instances. The remote debugger will be enabled for all roles and all instances if we checked. Currently there's no way for us to specify which role(s) and which instance(s) to enable. Finally click "publish" button. In the windows azure activity window in Visual Studio we can find some information about remote debugger. To attache remote process would be easy. Open the "server explorer" window in Visual Studio and expand "cloud services" node, find the cloud service, role and instance we had just published and wanted to debug, right click on the instance and select "attach debugger". Then after a while (it's based on how fast our Internet connect to Windows Azure Data Center) the Visual Studio will be switched to debug mode. Let's add a breakpoint in the default web page's form load function and refresh the page in browser to see what's happen. We can see that the our application was stopped at the breakpoint. The call stack, watch features are all available to use. Now let's hit F5 to continue the step, then back to the browser we will find the page was rendered successfully.   What Under the Hood Remote debugger is a WACS plugin. When we checked the "enable remote debugger" in the publish dialog, Visual Studio will add two cloud configuration settings in the CSCFG file. Since they were appended when deployment, we cannot find in our project's CSCFG file. But if we opened the publish package we could find as below. At the same time, Visual Studio will generate a certificate and included into the package for remote debugger. If we went to the azure management portal we will find there will a certificate under our application which was created, uploaded by remote debugger plugin. Since I enabled Remote Desktop there will be two certificates in the screenshot below. The other one is for remote debugger. When our application was deployed, windows azure system will open related ports for remote debugger. As below you can see there are two new ports opened on my application. Finally, in our WACS virtual machine, windows azure system will copy the remote debug component based on which version of Visual Studio we are using and start. Our application then can be debugged remotely through the visual studio remote debugger. Below is the task manager on the virtual machine of my WACS application.   Summary In this post I demonstrated one of the feature introduced in Windows Azure SDK 2.2, which is Remote Debugger. It allows us to attach our application from local machine to windows azure virtual machine once it had been deployed. Remote debugger is powerful and easy to use, but it brings more security risk. And since it's only available for debug build this means the performance will be worse than release build. Hence we should only use this feature for staging test and bug fix (publish our beta version to azure staging slot), rather than for production.   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.

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  • SQL SERVER – SSMS: Database Consistency History Report

    - by Pinal Dave
    Doctor and Database The last place I like to visit is always a hospital. With the monsoon season starting, intermittent rains, it has become sort of a routine to get a cycle of fever every other year (seriously I hate it). So when I visit my doctor, it is always interesting in the way he quizzes me. The routine question of – “How many days have you had this?”, “Is there any pattern?”, “Did you drench in rain?”, “Do you have any other symptom?” and so on. The idea here is that the doctor wants to find any anomaly or a pattern that will guide him to a viral or bacterial type. Most of the time they get it based on experience and sometimes after a battery of tests. So if there is consistent behavior to your problem, there is always a solution out. SQL Server has its way to find if the server data / files are in consistent state using the DBCC commands. Back to SQL Server In real life, Database consistency check is one of the critical operations a DBA generally doesn’t give much priority. Many readers of my blogs have asked many times, how do we know if the database is consistent? How do I read output of DBCC CHECKDB and find if everything is right or not? My common answer to all of them is – look at the bottom of checkdb (or checktable) output and look for below line. CHECKDB found 0 allocation errors and 0 consistency errors in database ‘DatabaseName’. Above is a “good sign” because we are seeing zero allocation and zero consistency error. If you are seeing non-zero errors then there is some problem with the database. Sample output is shown as below: CHECKDB found 0 allocation errors and 2 consistency errors in database ‘DatabaseName’. repair_allow_data_loss is the minimum repair level for the errors found by DBCC CHECKDB (DatabaseName). If we see non-zero error then most of the time (not always) we get repair options depending on the level of corruption. There is risk involved with above option (repair_allow_data_loss), that is – we would lose the data. Sometimes the option would be repair_rebuild which is little safer. Though these options are available, it is important to find the root cause to the problem. In standard report, there is a report which can show the history of checkdb executed for the selected database. Since this is a database level report, we need to right click on database, click Reports, click Standard Reports and then choose “Database Consistency History” report. The information in this report is picked from default trace. If default trace is disabled or there is no checkdb run or information is not there in default trace (because it’s rolled over), we would get report like below. As we can see report says it very clearly: Currently, no execution history of CHECKDB is available or default trace is not enabled. To demonstrate, I have caused corruption in one of the database and did below steps. Run CheckDB so that errors are reported. Fix the corruption by losing the data using repair option Run CheckDB again to check if corruption is cleared. After that I have launched the report and below is what we would see. If you are lazy like me and don’t want to run the report manually for each database then below query would be handy to provide same report for all database. This query is runs behind the scenes by the report. All I have done is remove the filter for database name (at the last – highlighted). DECLARE @curr_tracefilename VARCHAR(500); DECLARE @base_tracefilename VARCHAR(500); DECLARE @indx INT; SELECT @curr_tracefilename = path FROM sys.traces WHERE is_default = 1; SET @curr_tracefilename = REVERSE(@curr_tracefilename); SELECT @indx  = PATINDEX('%\%', @curr_tracefilename) ; SET @curr_tracefilename = REVERSE(@curr_tracefilename); SET @base_tracefilename = LEFT( @curr_tracefilename,LEN(@curr_tracefilename) - @indx) + '\log.trc'; SELECT  SUBSTRING(CONVERT(NVARCHAR(MAX),TEXTData),36, PATINDEX('%executed%',TEXTData)-36) AS command ,       LoginName ,       StartTime ,       CONVERT(INT,SUBSTRING(CONVERT(NVARCHAR(MAX),TEXTData),PATINDEX('%found%',TEXTData) +6,PATINDEX('%errors %',TEXTData)-PATINDEX('%found%',TEXTData)-6)) AS errors ,       CONVERT(INT,SUBSTRING(CONVERT(NVARCHAR(MAX),TEXTData),PATINDEX('%repaired%',TEXTData) +9,PATINDEX('%errors.%',TEXTData)-PATINDEX('%repaired%',TEXTData)-9)) repaired ,       SUBSTRING(CONVERT(NVARCHAR(MAX),TEXTData),PATINDEX('%time:%',TEXTData)+6,PATINDEX('%hours%',TEXTData)-PATINDEX('%time:%',TEXTData)-6)+':'+SUBSTRING(CONVERT(NVARCHAR(MAX),TEXTData),PATINDEX('%hours%',TEXTData) +6,PATINDEX('%minutes%',TEXTData)-PATINDEX('%hours%',TEXTData)-6)+':'+SUBSTRING(CONVERT(NVARCHAR(MAX),TEXTData),PATINDEX('%minutes%',TEXTData) +8,PATINDEX('%seconds.%',TEXTData)-PATINDEX('%minutes%',TEXTData)-8) AS time FROM::fn_trace_gettable( @base_tracefilename, DEFAULT) WHERE EventClass = 22 AND SUBSTRING(TEXTData,36,12) = 'DBCC CHECKDB' -- AND DatabaseName = @DatabaseName; Don’t get worried about the logic above. All it is doing is reading the trace files, parsing below entry and getting out information for underlined words. DBCC CHECKDB (CorruptedDatabase) executed by sa found 2 errors and repaired 0 errors. Elapsed time: 0 hours 0 minutes 0 seconds.  Internal database snapshot has split point LSN = 00000029:00000030:0001 and first LSN = 00000029:00000020:0001. Hopefully now onwards you would run checkdb and understand the importance of it. As responsible DBAs I am sure you are already doing it, let me know how often do you actually run them on you production environment? Reference: Pinal Dave (http://blog.sqlauthority.com)Filed under: PostADay, SQL, SQL Authority, SQL Query, SQL Server, SQL Server Management Studio, SQL Tips and Tricks, T SQL Tagged: SQL Reports

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  • Parent Objects

    - by Ali Bahrami
    Support for Parent Objects was added in Solaris 11 Update 1. The following material is adapted from the PSARC arc case, and the Solaris Linker and Libraries Manual. A "plugin" is a shared object, usually loaded via dlopen(), that is used by a program in order to allow the end user to add functionality to the program. Examples of plugins include those used by web browsers (flash, acrobat, etc), as well as mdb and elfedit modules. The object that loads the plugin at runtime is called the "parent object". Unlike most object dependencies, the parent is not identified by name, but by its status as the object doing the load. Historically, building a good plugin is has been more complicated than it should be: A parent and its plugin usually share a 2-way dependency: The plugin provides one or more routines for the parent to call, and the parent supplies support routines for use by the plugin for things like memory allocation and error reporting. It is a best practice to build all objects, including plugins, with the -z defs option, in order to ensure that the object specifies all of its dependencies, and is self contained. However: The parent is usually an executable, which cannot be linked to via the usual library mechanisms provided by the link editor. Even if the parent is a shared object, which could be a normal library dependency to the plugin, it may be desirable to build plugins that can be used by more than one parent, in which case embedding a dependency NEEDED entry for one of the parents is undesirable. The usual way to build a high quality plugin with -z defs uses a special mapfile provided by the parent. This mapfile defines the parent routines, specifying the PARENT attribute (see example below). This works, but is inconvenient, and error prone. The symbol table in the parent already describes what it makes available to plugins — ideally the plugin would obtain that information directly rather than from a separate mapfile. The new -z parent option to ld allows a plugin to link to the parent and access the parent symbol table. This differs from a typical dependency: No NEEDED record is created. The relationship is recorded as a logical connection to the parent, rather than as an explicit object name However, it operates in the same manner as any other dependency in terms of making symbols available to the plugin. When the -z parent option is used, the link-editor records the basename of the parent object in the dynamic section, using the new tag DT_SUNW_PARENT. This is an informational tag, which is not used by the runtime linker to locate the parent, but which is available for diagnostic purposes. The ld(1) manpage documentation for the -z parent option is: -z parent=object Specifies a "parent object", which can be an executable or shared object, against which to link the output object. This option is typically used when creating "plugin" shared objects intended to be loaded by an executable at runtime via the dlopen() function. The symbol table from the parent object is used to satisfy references from the plugin object. The use of the -z parent option makes symbols from the object calling dlopen() available to the plugin. Example For this example, we use a main program, and a plugin. The parent provides a function named parent_callback() for the plugin to call. The plugin provides a function named plugin_func() to the parent: % cat main.c #include <stdio.h> #include <dlfcn.h> #include <link.h> void parent_callback(void) { printf("plugin_func() has called parent_callback()\n"); } int main(int argc, char **argv) { typedef void plugin_func_t(void); void *hdl; plugin_func_t *plugin_func; if (argc != 2) { fprintf(stderr, "usage: main plugin\n"); return (1); } if ((hdl = dlopen(argv[1], RTLD_LAZY)) == NULL) { fprintf(stderr, "unable to load plugin: %s\n", dlerror()); return (1); } plugin_func = (plugin_func_t *) dlsym(hdl, "plugin_func"); if (plugin_func == NULL) { fprintf(stderr, "unable to find plugin_func: %s\n", dlerror()); return (1); } (*plugin_func)(); return (0); } % cat plugin.c #include <stdio.h> extern void parent_callback(void); void plugin_func(void) { printf("parent has called plugin_func() from plugin.so\n"); parent_callback(); } Building this in the traditional manner, without -zdefs: % cc -o main main.c % cc -G -o plugin.so plugin.c % ./main ./plugin.so parent has called plugin_func() from plugin.so plugin_func() has called parent_callback() As noted above, when building any shared object, the -z defs option is recommended, in order to ensure that the object is self contained and specifies all of its dependencies. However, the use of -z defs prevents the plugin object from linking due to the unsatisfied symbol from the parent object: % cc -zdefs -G -o plugin.so plugin.c Undefined first referenced symbol in file parent_callback plugin.o ld: fatal: symbol referencing errors. No output written to plugin.so A mapfile can be used to specify to ld that the parent_callback symbol is supplied by the parent object. % cat plugin.mapfile $mapfile_version 2 SYMBOL_SCOPE { global: parent_callback { FLAGS = PARENT }; }; % cc -zdefs -Mplugin.mapfile -G -o plugin.so plugin.c However, the -z parent option to ld is the most direct solution to this problem, allowing the plugin to actually link against the parent object, and obtain the available symbols from it. An added benefit of using -z parent instead of a mapfile, is that the name of the parent object is recorded in the dynamic section of the plugin, and can be displayed by the file utility: % cc -zdefs -zparent=main -G -o plugin.so plugin.c % elfdump -d plugin.so | grep PARENT [0] SUNW_PARENT 0xcc main % file plugin.so plugin.so: ELF 32-bit LSB dynamic lib 80386 Version 1, parent main, dynamically linked, not stripped % ./main ./plugin.so parent has called plugin_func() from plugin.so plugin_func() has called parent_callback() We can also observe this in elfedit plugins on Solaris systems running Solaris 11 Update 1 or newer: % file /usr/lib/elfedit/dyn.so /usr/lib/elfedit/dyn.so: ELF 32-bit LSB dynamic lib 80386 Version 1, parent elfedit, dynamically linked, not stripped, no debugging information available Related Other Work The GNU ld has an option named --just-symbols that can be used in a similar manner: --just-symbols=filename Read symbol names and their addresses from filename, but do not relocate it or include it in the output. This allows your output file to refer symbolically to absolute locations of memory defined in other programs. You may use this option more than once. -z parent is a higher level operation aimed specifically at simplifying the construction of high quality plugins. Although it employs the same operation, it differs from --just symbols in 2 significant ways: There can only be one parent. The parent is recorded in the created object, and can be displayed by 'file', or other similar tools.

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  • View Docs and PDFs Directly in Google Chrome

    - by Matthew Guay
    Would you like to view documents, presentations, and PDFs directly in Google Chrome?  Here’s a handy extension that makes Google Docs your default online viewer so don’t have to download the file first. Getting Started By default, when you come across a PDF or other common document file online in Google Chrome, you’ll have to download the file and open it in a separate application. It’d be much easier to simply view online documents directly in Chrome.  To do this, head over to the Docs PDF/PowerPoint Viewer page on the Chrome Extensions site (link below), and click Install to add it to your browser. Click Install to confirm that you want to install this extension. Extensions don’t run by default in Incognito mode, so if you’d like to always view documents directly in Chrome, open the Extensions page and check Allow this extension to run in incognito. Now, when you click a link for a document online, such as a .docx file from Word, it will open in the Google Docs viewer. These documents usually render in their original full-quality.  You can zoom in and out to see exactly what you want, or search within the document.  Or, if it doesn’t look correct, you can click the Download link in the top left to save the original document to your computer and open it in Office.   Even complex PDF render very nicely.  Do note that Docs will keep downloading the document as you’re reading it, so if you jump to the middle of a document it may look blurry at first but will quickly clear up. You can even view famous presentations online without opening them in PowerPoint.  Note that this will only display the slides themselves, but if you’re looking for information you likely don’t need the slideshow effects anyway.   Adobe Reader Conflicts If you already have Adobe Acrobat or Adobe Reader installed on your computer, PDF files may open with the Adobe plugin.  If you’d prefer to read your PDFs with the Docs PDF Viewer, then you need to disable the Adobe plugin.  Enter the following in your Address Bar to open your Chrome Plugins page: chrome://plugins/ and then click Disable underneath the Adobe Acrobat plugin. Now your PDFs will always open with the Docs viewer instead. Performance Who hasn’t been frustrated by clicking a link to a PDF file, only to have your browser pause for several minutes while Adobe Reader struggles to download and display the file?  Google Chrome’s default behavior of simply downloading the files and letting you open them is hardly more helpful.  This extension takes away both of these problems, since it renders the documents on Google’s servers.  Most documents opened fairly quickly in our tests, and we were able to read large PDFs only seconds after clicking their link.  Also, the Google Docs viewer rendered the documents much better than the HTML version in Google’s cache. Google Docs did seem to have problem on some files, and we saw error messages on several documents we tried to open.  If you encounter this, click the Download link in the top left corner to download the file and view it from your desktop instead. Conclusion Google Docs has improved over the years, and now it offers fairly good rendering even on more complex documents.  This extension can make your browsing easier, and help documents and PDFs feel more like part of the Internet.  And, since the documents are rendered on Google’s servers, it’s often faster to preview large files than to download them to your computer. Link Download the Docs PDF/PowerPoint Viewer extension from Google Similar Articles Productive Geek Tips Integrate Google Docs with Outlook the Easy WayGoogle Image Search Quick FixView the Time & Date in Chrome When Hiding Your TaskbarView Maps and Get Directions in Google ChromeHow To Export Documents from Google Docs to Your Computer TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Xobni Plus for Outlook All My Movies 5.9 CloudBerry Online Backup 1.5 for Windows Home Server Snagit 10 How to Forecast Weather, without Gadgets Outlook Tools, one stop tweaking for any Outlook version Zoofs, find the most popular tweeted YouTube videos Video preview of new Windows Live Essentials 21 Cursor Packs for XP, Vista & 7 Map the Stars with Stellarium

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  • The Case of the Extra Page: Rendering Reporting Services as PDF

    - by smisner
    I had to troubleshoot a problem with a mysterious extra page appearing in a PDF this week. My first thought was that it was likely to caused by one of the most common problems that people encounter when developing reports that eventually get rendered as PDF is getting blank pages inserted into the PDF document. The cause of the blank pages is usually related to sizing. You can learn more at Understanding Pagination in Reporting Services in Books Online. When designing a report, you have to be really careful with the layout of items in the body. As you move items around, the body will expand to accommodate the space you're using and you might eventually tighten everything back up again, but the body doesn't automatically collapse. One of my favorite things to do in Reporting Services 2005 - which I dubbed the "vacu-pack" method - was to just erase the size property of the Body and let it auto-calculate the new size, squeezing out all the extra space. Alas, that method no longer works beginning with Reporting Services 2008. Even when you make sure the body size is as small as possible (with no unnecessary extra space along the top, bottom, left, or right side of the body), it's important to calculate the body size plus header plus footer plus the margins and ensure that the calculated height and width do not exceed the report's height and width (shown as the page in the illustration above). This won't matter if users always render reports online, but they'll get extra pages in a PDF document if the report's height and width are smaller than the calculate space. Beginning the Investigation In the situation that I was troubleshooting, I checked the properties: Item Property Value Body Height 6.25in   Width 10.5in Page Header Height 1in Page Footer Height 0.25in Report Left Margin 0.1in   Right Margin 0.1in   Top Margin 0.05in   Bottom Margin 0.05in   Page Size - Height 8.5in   Page Size - Width 11in So I calculated the total width using Body Width + Left Margin + Right Margin and came up with a value of 10.7 inches. And then I calculated the total height using Body Height + Page Header Height + Page Footer Height + Top Margin + Bottom Margin and got 7.6 inches. Well, page sizing couldn't be the reason for the extra page in my report because 10.7 inches is smaller than the report's width of 11 inches and 7.6 inches is smaller than the report's height of 8.5 inches. I had to look elsewhere to find the culprit. Conducting the Third Degree My next thought was to focus on the rendering size of the items in the report. I've adapted my problem to use the Adventure Works database. At the top of the report are two charts, and then below each chart is a rectangle that contains a table. In the real-life scenario, there were some graphics present as a background for the tables which fit within the rectangles that were about 3 inches high so the visual space of the rectangles matched the visual space of the charts - also about 3 inches high. But there was also a huge amount of white space at the bottom of the page, and as I mentioned at the beginning of this post, a second page which was blank except for the footer that appeared at the bottom. Placing a textbox beneath the rectangles to see if they would appear on the first page resulted the textbox's appearance on the second page. For some reason, the rectangles wanted a buffer zone beneath them. What's going on? Taking the Suspect into Custody My next step was to see what was really going on with the rectangle. The graphic appeared to be correctly sized, but the behavior in the report indicated the rectangle was growing. So I added a border to the rectangle to see what it was doing. When I added borders, I could see that the size of each rectangle was growing to accommodate the table it contains. The rectangle on the right is slightly larger than the one on the left because the table on the right contains an extra row. The rectangle is trying to preserve the whitespace that appears in the layout, as shown below. Closing the Case Now that I knew what the problem was, what could I do about it? Because of the graphic in the rectangle (not shown), I couldn't eliminate the use of the rectangles and just show the tables. But fortunately, there is a report property that comes to the rescue: ConsumeContainerWhitespace (accessible only in the Properties window). I set the value of this property to True. Problem solved. Now the rectangles remain fixed at the configured size and don't grow vertically to preserve the whitespace. Case closed.

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  • Mocking successive calls of similar type via sequential mocking

    - by mehfuzh
    In this post , i show how you can benefit from  sequential mocking feature[In JustMock] for setting up expectations with successive calls of same type.  To start let’s first consider the following dummy database and entity class. public class Person {     public virtual string Name { get; set; }     public virtual int Age { get; set; } }   public interface IDataBase {     T Get<T>(); } Now, our test goal is to return different entity for successive calls on IDataBase.Get<T>(). By default, the behavior in JustMock is override , which is similar to other popular mocking tools. By override it means that the tool will consider always the latest user setup. Therefore, the first example will return the latest entity every-time and will fail in line #12: Person person1 = new Person { Age = 30, Name = "Kosev" }; Person person2 = new Person { Age = 80, Name = "Mihail" };   var database = Mock.Create<IDataBase>();   Queue<Person> queue = new Queue<Person>();   Mock.Arrange(() => database.Get<Person>()).Returns(() => queue.Dequeue()); Mock.Arrange(() => database.Get<Person>()).Returns(person2);   // this will fail Assert.Equal(person1.GetHashCode(), database.Get<Person>().GetHashCode());   Assert.Equal(person2.GetHashCode(), database.Get<Person>().GetHashCode()); We can solve it the following way using a Queue and that removes the item from bottom on each call: Person person1 = new Person { Age = 30, Name = "Kosev" }; Person person2 = new Person { Age = 80, Name = "Mihail" };   var database = Mock.Create<IDataBase>();   Queue<Person> queue = new Queue<Person>();   queue.Enqueue(person1); queue.Enqueue(person2);   Mock.Arrange(() => database.Get<Person>()).Returns(queue.Dequeue());   Assert.Equal(person1.GetHashCode(), database.Get<Person>().GetHashCode()); Assert.Equal(person2.GetHashCode(), database.Get<Person>().GetHashCode()); This will ensure that right entity is returned but this is not an elegant solution. So, in JustMock we introduced a  new option that lets you set up your expectations sequentially. Like: Person person1 = new Person { Age = 30, Name = "Kosev" }; Person person2 = new Person { Age = 80, Name = "Mihail" };   var database = Mock.Create<IDataBase>();   Mock.Arrange(() => database.Get<Person>()).Returns(person1).InSequence(); Mock.Arrange(() => database.Get<Person>()).Returns(person2).InSequence();   Assert.Equal(person1.GetHashCode(), database.Get<Person>().GetHashCode()); Assert.Equal(person2.GetHashCode(), database.Get<Person>().GetHashCode()); The  “InSequence” modifier will tell the mocking tool to return the expected result as in the order it is specified by user. The solution though pretty simple and but neat(to me) and way too simpler than using a collection to solve this type of cases. Hope that helps P.S. The example shown in my blog is using interface don’t require a profiler  and you can even use a notepad and build it referencing Telerik.JustMock.dll, run it with GUI tools and it will work. But this feature also applies to concrete methods that includes JM profiler and can be implemented for more complex scenarios.

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  • ct.sym steals the ASM class

    - by Geertjan
    Some mild consternation on the Twittersphere yesterday. Marcus Lagergren not being able to find the ASM classes in JDK 8 in NetBeans IDE: And there's no such problem in Eclipse (and apparently in IntelliJ IDEA). Help, does NetBeans (despite being incredibly awesome) suck, after all? The truth of the matter is that there's something called "ct.sym" in the JDK. When javac is compiling code, it doesn't link against rt.jar. Instead, it uses a special symbol file lib/ct.sym with class stubs. Internal JDK classes are not put in that symbol file, since those are internal classes. You shouldn't want to use them, at all. However, what if you're Marcus Lagergren who DOES need these classes? I.e., he's working on the internal JDK classes and hence needs to have access to them. Fair enough that the general Java population can't access those classes, since they're internal implementation classes that could be changed anytime and one wouldn't want all unknown clients of those classes to start breaking once changes are made to the implementation, i.e., this is the rt.jar's internal class protection mechanism. But, again, we're now Marcus Lagergen and not the general Java population. For the solution, read Jan Lahoda, NetBeans Java Editor guru, here: https://netbeans.org/bugzilla/show_bug.cgi?id=186120 In particular, take note of this: AFAIK, the ct.sym is new in JDK6. It contains stubs for all classes that existed in JDK5 (for compatibility with existing programs that would use private JDK classes), but does not contain implementation classes that were introduced in JDK6 (only API classes). This is to prevent application developers to accidentally use JDK's private classes (as such applications would be unportable and may not run on future versions of JDK). Note that this is not really a NB thing - this is the behavior of javac from the JDK. I do not know about any way to disable this except deleting ct.sym or the option mentioned above. Regarding loading the classes: JVM uses two classpath's: classpath and bootclasspath. rt.jar is on the bootclasspath and has precedence over anything on the "custom" classpath, which is used by the application. The usual way to override classes on bootclasspath is to start the JVM with "-Xbootclasspath/p:" option, which prepends the given jars (and presumably also directories) to bootclasspath. Hence, let's take the first option, the simpler one, and simply delete the "ct.sym" file. Again, only because we need to work with those internal classes as developers of the JDK, not because we want to hack our way around "ct.sym", which would mean you'd not have portable code at the end of the day. Go to the JDK 8 lib folder and you'll find the file: Delete it. Start NetBeans IDE again, either on JDK 7 or JDK 8, doesn't make a difference for these purposes, create a new Java application (or use an existing one), make sure you have set the JDK above as the JDK of the application, and hey presto: The above obviously assumes you have a build of JDK 8 that actually includes the ASM package. And below you can see that not only are the classes found but my build succeeded, even though I'm using internal JDK classes. The yellow markings in the sidebar mean that the classes are imported but not used in the code, where normally, if I hadn't removed "ct.sym", I would have seen red error marking instead, and the code wouldn't have compiled. Note: I've tried setting "-XDignore.symbol.file" in "netbeans.conf" and in other places, but so far haven't got that to work. Simply deleting the "ct.sym" file (or back it up somewhere and put it back when needed) is quite clearly the most straightforward solution. Ultimately, if you want to be able to use those internal classes while still having portable code, do you know what you need to do? You need to create a JDK bug report stating that you need an internal class to be added to "ct.sym". Probably you'll get a motivation back stating WHY that internal class isn't supposed to be used externally. There must be a reason why those classes aren't available for external usage, otherwise they would have been added to "ct.sym". So, now the only remaining question is why the Eclipse compiler doesn't hide the internal JDK classes. Apparently the Eclipse compiler ignores the "ct.sym" file. In other words, at the end of the day, far from being a bug in NetBeans... we have now found a (pretty enormous, I reckon) bug in Eclipse. The Eclipse compiler does not protect you from using internal JDK classes and the code that you create in Eclipse may not work with future releases of the JDK, since the JDK team is simply going to be changing those classes that are not found in the "ct.sym" file while assuming (correctly, thanks to the presence of "ct.sym" mechanism) that no code in the world, other than JDK code, is tied to those classes.

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  • The Data Scientist

    - by BuckWoody
    A new term - well, perhaps not that new - has come up and I’m actually very excited about it. The term is Data Scientist, and since it’s new, it’s fairly undefined. I’ll explain what I think it means, and why I’m excited about it. In general, I’ve found the term deals at its most basic with analyzing data. Of course, we all do that, and the term itself in that definition is redundant. There is no science that I know of that does not work with analyzing lots of data. But the term seems to refer to more than the common practices of looking at data visually, putting it in a spreadsheet or report, or even using simple coding to examine data sets. The term Data Scientist (as far as I can make out this early in it’s use) is someone who has a strong understanding of data sources, relevance (statistical and otherwise) and processing methods as well as front-end displays of large sets of complicated data. Some - but not all - Business Intelligence professionals have these skills. In other cases, senior developers, database architects or others fill these needs, but in my experience, many lack the strong mathematical skills needed to make these choices properly. I’ve divided the knowledge base for someone that would wear this title into three large segments. It remains to be seen if a given Data Scientist would be responsible for knowing all these areas or would specialize. There are pretty high requirements on the math side, specifically in graduate-degree level statistics, but in my experience a company will only have a few of these folks, so they are expected to know quite a bit in each of these areas. Persistence The first area is finding, cleaning and storing the data. In some cases, no cleaning is done prior to storage - it’s just identified and the cleansing is done in a later step. This area is where the professional would be able to tell if a particular data set should be stored in a Relational Database Management System (RDBMS), across a set of key/value pair storage (NoSQL) or in a file system like HDFS (part of the Hadoop landscape) or other methods. Or do you examine the stream of data without storing it in another system at all? This is an important decision - it’s a foundation choice that deals not only with a lot of expense of purchasing systems or even using Cloud Computing (PaaS, SaaS or IaaS) to source it, but also the skillsets and other resources needed to care and feed the system for a long time. The Data Scientist sets something into motion that will probably outlast his or her career at a company or organization. Often these choices are made by senior developers, database administrators or architects in a company. But sometimes each of these has a certain bias towards making a decision one way or another. The Data Scientist would examine these choices in light of the data itself, starting perhaps even before the business requirements are created. The business may not even be aware of all the strategic and tactical data sources that they have access to. Processing Once the decision is made to store the data, the next set of decisions are based around how to process the data. An RDBMS scales well to a certain level, and provides a high degree of ACID compliance as well as offering a well-known set-based language to work with this data. In other cases, scale should be spread among multiple nodes (as in the case of Hadoop landscapes or NoSQL offerings) or even across a Cloud provider like Windows Azure Table Storage. In fact, in many cases - most of the ones I’m dealing with lately - the data should be split among multiple types of processing environments. This is a newer idea. Many data professionals simply pick a methodology (RDBMS with Star Schemas, NoSQL, etc.) and put all data there, regardless of its shape, processing needs and so on. A Data Scientist is familiar not only with the various processing methods, but how they work, so that they can choose the right one for a given need. This is a huge time commitment, hence the need for a dedicated title like this one. Presentation This is where the need for a Data Scientist is most often already being filled, sometimes with more or less success. The latest Business Intelligence systems are quite good at allowing you to create amazing graphics - but it’s the data behind the graphics that are the most important component of truly effective displays. This is where the mathematics requirement of the Data Scientist title is the most unforgiving. In fact, someone without a good foundation in statistics is not a good candidate for creating reports. Even a basic level of statistics can be dangerous. Anyone who works in analyzing data will tell you that there are multiple errors possible when data just seems right - and basic statistics bears out that you’re on the right track - that are only solvable when you understanding why the statistical formula works the way it does. And there are lots of ways of presenting data. Sometimes all you need is a “yes” or “no” answer that can only come after heavy analysis work. In that case, a simple e-mail might be all the reporting you need. In others, complex relationships and multiple components require a deep understanding of the various graphical methods of presenting data. Knowing which kind of chart, color, graphic or shape conveys a particular datum best is essential knowledge for the Data Scientist. Why I’m excited I love this area of study. I like math, stats, and computing technologies, but it goes beyond that. I love what data can do - how it can help an organization. I’ve been fortunate enough in my professional career these past two decades to work with lots of folks who perform this role at companies from aerospace to medical firms, from manufacturing to retail. Interestingly, the size of the company really isn’t germane here. I worked with one very small bio-tech (cryogenics) company that worked deeply with analysis of complex interrelated data. So  watch this space. No, I’m not leaving Azure or distributed computing or Microsoft. In fact, I think I’m perfectly situated to investigate this role further. We have a huge set of tools, from RDBMS to Hadoop to allow me to explore. And I’m happy to share what I learn along the way.

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  • Windows 8 Live Accounts and the actual Windows Account

    - by Rick Strahl
    As if Windows Security wasn't confusing enough, in Windows 8 we get thrown yet another curve ball with Windows Live accounts to logon. When I set up my Windows 8 machine I originally set it up with a 'real', non-live account that I always use on my Windows machines. I did this mainly so I have a matching account for resources around my home and intranet network so I could log on to network resources properly. At some point later I decided to set up Windows Live security just to see how changes things. Windows wants you to use Windows Live Windows 8 logins are required in order for the Windows RT account info to work. Not that I care - since installing Windows 8 I've maybe spent 10 minutes with Windows RT because - well it's pretty freaking sucky on the desktop. From shitty apps to mis-managed screen real estate I can't say that there's anything compelling there to date, but then I haven't looked that hard either. Anyway… I set up the Windows Live account to see if that changes things. It does - I do get all my live logins to work from Live Account so that Twitter and Facebook posts and pictures and calendars all show up on live tiles on the start screen and in the actual apps. That's nice-ish, but hardly that exciting given that all of the apps tied to those live tiles are average at best. And it would have been nice if all of this could be done without being forced into running with a Windows Live User Account - this all feels like strong-arming you into moving into Microsofts walled garden… and that's probably what it's meant to do. Who am I? The real problem to me though is that these Windows Live and raw Windows User accounts are a bit unpredictable especially when it comes to developer information about the account and which credentials to use. So for example Windows reports folder security like this: Notice it's showing my Windows Live account. Now if I go to Edit and try to add my Windows user account (rstrahl) it'll just automatically show up as the live account. On the other hand though the underlying system sees everything as my real Windows account. After I switched to a Windows Live login account and I have to login to Windows with my Live account, what do you suppose this returns?Console.WriteLine(Environment.UserName); It returns my raw Windows user account (rstrahl). All my permissions, all my actual settings and the desktop console altogether run under that account. If I look in TaskManager (or Process Explorer for me) I see: Everything running on the desktop shell with my login running under my Windows user account. I suppose it makes sense, but where is that association happening? When I switched to a Windows Live account, nowhere did I associate my real account with the Live account - it just happened. And looking through the account configuration dialogs I can't find any reference to the raw Windows account. Other than switching back I see no mention anywhere of the raw Windows account - everything refers to the Live account. Right then, clear as potato soup! So this is who you really are! The problem is that in some situations this schizophrenic account behavior gets a bit weird. Today I was running a local Web application in IIS that uses Windows Authentication - I tried to log-in with my real Windows account login because that's what I'm used to using with WINDOWS freaking Authentication through IIS. But… it failed. I checked my IIS settings, my apps login settings and I just could not for the life of me get into the site with my Windows username. That is until I finally realized that I should try using my Windows Live credentials instead. And that worked. So now in this Windows Authentication dialog I had to type in my Live ID and password, which is - just weird. Then in IIS if I look at a Trace page (or in my case my app's Status page) I see that the logged on account is - my Windows user account. What's really annoying about this is that in some places it uses the live account in other places it uses my Windows account. If I remote desktop into my Web server online - I have to use the local authentication dialog but I have to put in my real Windows credentials not the Live account. Oh yes, it's all so terribly intuitive and logical… So in summary, when you log on with a Live account you are actually mapped to an underlying Windows user. In any application if you check the user name it'll be the underlying user account (not sure what happens in a Windows RT app or even what mechanism is used there to get the user name info).  When logging on to local machine resource with user name and password you have to use your Live IDs even if the permissions on the resources are mapped to your underlying Windows account. Easy enough I suppose, but still not exactly intuitive behavior…© Rick Strahl, West Wind Technologies, 2005-2012Posted in Windows   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); })();

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  • Silverlight Cream for January 11, 2011 -- #1024

    - by Dave Campbell
    1,000 blogposts is quite a few, but to die-hard geeks, 1000 isn't the number... 1K is the number, and today is my 1K blogpost! I've been working up to this for at least 11 months. Way back at MIX10, I approached some vendors about an idea I had. A month ago I contacted them and others, and everyone I contacted was very generous and supportive of my idea. My idea was not to run a contest, but blog as normal, and whoever ended up on my 1K post would get some swag... and I set a cut-off at 13 posts. So... blogging normally, I had some submittals, and then ran my normal process to pick up the next posts until I hit a total of 13. To provide a distribution channel for the swag, everyone on the list, please send me your snail mail (T-shirts) and email (licenses) addresses as soon as possible.   I'd like to thank the following generous sponsors for their contributions to my fun (in alphabetic order): and Rachel Hawley for contributing 4 Silverlight control sets First Floor Software and Koen Zwikstra for contributing 13 licenses for Silverlight Spy and Sara Faatz/Jason Beres for contributing 13 licenses for Silverlight Data Visualization controls and Svetla Stoycheva for contributing T-Shirts for everyone on the post and Ina Tontcheva for contributing 13 licenses for RadControls for Silverlight + RadControls for Windows Phone and Charlene Kozlan for contributing 1 combopack standard, 2 DataGrid for Silverlight, and 2 Listbox for Silverlight Standard And now finally...in this Issue: Nigel Sampson, Jeremy Likness, Dan Wahlin, Kunal Chowdhurry, Alex Knight, Wei-Meng Lee, Michael Crump, Jesse Liberty, Peter Kuhn, Michael Washington, Tau Sick, Max Paulousky, Damian Schenkelman Above the Fold: Silverlight: "Demystifying Silverlight Dependency Properties" Dan Wahlin WP7: "Using Windows Phone Gestures as Triggers" Nigel Sampson Expression Blend: "PathListBox: making data look cool" Alex Knight From SilverlightCream.com: Using Windows Phone Gestures as Triggers Nigel Sampson blogged about WP7 Gestures, the Toolkit, and using Gestures as Triggers, and actually makes it looks simple :) Jounce Part 9: Static and Dynamic Module Management Jeremy Likness has episode 9 of his explanation of his MVVM framework, Jounce, up... and a big discussion of Modules and Module Management from a Jounce perspective. Demystifying Silverlight Dependency Properties Dan Wahlin takes a page from one of his teaching opportunities, and shares his knowledge of Dependency Properties with us... beginning with what they are, defining them in code, and demonstrating their use. Customizing Silverlight ChildWindow Style using Blend Kunal Chowdhurry has a great post up about getting your Child Windows to match the look & feel of the rest of youra app... plus a bunch of Blend goodness thrown in. PathListBox: making data look cool File this post by Alex Knight in the 'holy crap' file along with the others in this series! ... just check out that cool Ticker Style Path ListBox at the top of the blog... too cool! Web Access in Windows Phone 7 Apps Wei-Meng Lee has the 3rd part of his series on WP7 development up and in this one is discussing Web Access... I mean *discussing* it... tons of detail, code, and explanation... great post. Prevent your Silverlight XAP file from caching in your browser. Michael Crump helps relieve stress on Silverlight developers everywhere by exploring how to avoid caching of your XAP in the browser... (WPFS) MVVM Light Toolkit: Soup To Nuts Part I Jesse Liberty continues his Windows Phone from Scratch series with a new segment exploring Laurent Bugnion's MVVMLight Toolkit beginning with acquiring and installing the toolkit, then proceeds to discuss linking the View and ViewModel, the ViewModel Locator, and page navigation. Silverlight: Making a DateTimePicker Peter Kuhn attacks a problem that crops up on the forums a lot -- a DateTimePicker control for Silverlight... following the "It's so simple to build one yourself" advice, he did so, and provides the code for all of us! Windows Phone 7 Animated Button Press Michael Washington took exception to button presses that gave no visual feedback and produced a behavior that does just that. Using TweetSharp in a Windows Phone 7 app Tau Sick demonstrates using TweetSharp to put a twitter feed into a WP7 app, as he did in "Hangover Helper"... all the instructions from getting Tweeetshaprt to the code necessary. Bindable Application Bar Extensions for Windows Phone 7 Max Paulousky has a post discussing some real extensions to the ApplicationBar for WP7.. he begins with a bindable application bar by Nicolas Humann that I've missed, probably because his blog is in French... and extends it to allow using DelegateCommand. How to: Load Prism modules packaged in a separate XAP file in an OOB application Damian Schenkelman posts about Prism, AppModules in separate XAPs and running OOB... if you've tried this, you know it's a hassle.. Damian has the solution. Stay in the 'Light! Twitter SilverlightNews | Twitter WynApse | WynApse.com | Tagged Posts | SilverlightCream Join me @ SilverlightCream | Phoenix Silverlight User Group Technorati Tags: Silverlight    Silverlight 3    Silverlight 4    Windows Phone MIX10

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  • Calculated Columns in Entity Framework Code First Migrations

    - by David Paquette
    I had a couple people ask me about calculated properties / columns in Entity Framework this week.  The question was, is there a way to specify a property in my C# class that is the result of some calculation involving 2 properties of the same class.  For example, in my database, I store a FirstName and a LastName column and I would like a FullName property that is computed from the FirstName and LastName columns.  My initial answer was: 1: public string FullName 2: { 3: get { return string.Format("{0} {1}", FirstName, LastName); } 4: } Of course, this works fine, but this does not give us the ability to write queries using the FullName property.  For example, this query: 1: var users = context.Users.Where(u => u.FullName.Contains("anan")); Would result in the following NotSupportedException: The specified type member 'FullName' is not supported in LINQ to Entities. Only initializers, entity members, and entity navigation properties are supported. It turns out there is a way to support this type of behavior with Entity Framework Code First Migrations by making use of Computed Columns in SQL Server.  While there is no native support for computed columns in Code First Migrations, we can manually configure our migration to use computed columns. Let’s start by defining our C# classes and DbContext: 1: public class UserProfile 2: { 3: public int Id { get; set; } 4: 5: public string FirstName { get; set; } 6: public string LastName { get; set; } 7: 8: [DatabaseGenerated(DatabaseGeneratedOption.Computed)] 9: public string FullName { get; private set; } 10: } 11: 12: public class UserContext : DbContext 13: { 14: public DbSet<UserProfile> Users { get; set; } 15: } The DatabaseGenerated attribute is needed on our FullName property.  This is a hint to let Entity Framework Code First know that the database will be computing this property for us. Next, we need to run 2 commands in the Package Manager Console.  First, run Enable-Migrations to enable Code First Migrations for the UserContext.  Next, run Add-Migration Initial to create an initial migration.  This will create a migration that creates the UserProfile table with 3 columns: FirstName, LastName, and FullName.  This is where we need to make a small change.  Instead of allowing Code First Migrations to create the FullName property, we will manually add that column as a computed column. 1: public partial class Initial : DbMigration 2: { 3: public override void Up() 4: { 5: CreateTable( 6: "dbo.UserProfiles", 7: c => new 8: { 9: Id = c.Int(nullable: false, identity: true), 10: FirstName = c.String(), 11: LastName = c.String(), 12: //FullName = c.String(), 13: }) 14: .PrimaryKey(t => t.Id); 15: Sql("ALTER TABLE dbo.UserProfiles ADD FullName AS FirstName + ' ' + LastName"); 16: } 17: 18: 19: public override void Down() 20: { 21: DropTable("dbo.UserProfiles"); 22: } 23: } Finally, run the Update-Database command.  Now we can query for Users using the FullName property and that query will be executed on the database server.  However, we encounter another potential problem. Since the FullName property is calculated by the database, it will get out of sync on the object side as soon as we make a change to the FirstName or LastName property.  Luckily, we can have the best of both worlds here by also adding the calculation back to the getter on the FullName property: 1: [DatabaseGenerated(DatabaseGeneratedOption.Computed)] 2: public string FullName 3: { 4: get { return FirstName + " " + LastName; } 5: private set 6: { 7: //Just need this here to trick EF 8: } 9: } Now we can both query for Users using the FullName property and we also won’t need to worry about the FullName property being out of sync with the FirstName and LastName properties.  When we run this code: 1: using(UserContext context = new UserContext()) 2: { 3: UserProfile userProfile = new UserProfile {FirstName = "Chanandler", LastName = "Bong"}; 4: 5: Console.WriteLine("Before saving: " + userProfile.FullName); 6: 7: context.Users.Add(userProfile); 8: context.SaveChanges(); 9:  10: Console.WriteLine("After saving: " + userProfile.FullName); 11:  12: UserProfile chanandler = context.Users.First(u => u.FullName == "Chanandler Bong"); 13: Console.WriteLine("After reading: " + chanandler.FullName); 14:  15: chanandler.FirstName = "Chandler"; 16: chanandler.LastName = "Bing"; 17:  18: Console.WriteLine("After changing: " + chanandler.FullName); 19:  20: } We get this output: It took a bit of work, but finally Chandler’s TV Guide can be delivered to the right person. The obvious downside to this implementation is that the FullName calculation is duplicated in the database and in the UserProfile class. This sample was written using Visual Studio 2012 and Entity Framework 5. Download the source code here.

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  • How do I prove or disprove "god" objects are wrong?

    - by honestduane
    Problem Summary: Long story short, I inherited a code base and an development team I am not allowed to replace and the use of God Objects is a big issue. Going forward, I want to have us re-factor things but I am getting push-back from the teams who want to do everything with God Objects "because its easier" and this means I would not be allowed to re-factor. I pushed back citing my years of dev experience, that I'm the new boss who was hired to know these things, etc, and so did the third party offshore companies account sales rep, and this is now at the executive level and my meeting is tomorrow and I want to go in with a lot of technical ammo to advocate best practices because I feel it will be cheaper in the long run (And I personally feel that is what the third party is worried about) for the company. My issue is from a technical level, I know its good long term but I'm having trouble with the ultra short term and 6 months term, and while its something I "know" I cant prove it with references and cited resources outside of one person (Robert C. Martin, aka Uncle Bob), as that is what I am being asked to do as I have been told having data from one person and only one person (Robert C Martin) is not good enough of an argument. Question: What are some resources I can cite directly (Title, year published, page number, quote) by well known experts in the field that explicitly say this use of "God" Objects/Classes/Systems is bad (or good, since we are looking for the most technically valid solution)? Research I have already done: I have a number of books here and I have searched their indexes for the use of the words "god object" and "god class". I found that oddly its almost never used and the copy of the GoF book I have for example, never uses it (At least according to the index in front of me) but I have found it in 2 books per the below, but I want more I can use. I checked the Wikipedia page for "God Object" and its currently a stub with little reference links so although I personally agree with that it says, It doesn't have much I can use in an environment where personal experience is not considered valid. The book cited is also considered too old to be valid by the people I am debating these technical points with as the argument they are making is that "it was once thought to be bad but nobody could prove it, and now modern software says "god" objects are good to use". I personally believe that this statement is incorrect, but I want to prove the truth, whatever it is. In Robert C Martin's "Agile Principles, Patterns, and Practices in C#" (ISBN: 0-13-185725-8, hardcover) where on page 266 it states "Everybody knows that god classes are a bad idea. We don't want to concentrate all the intelligence of a system into a single object or a single function. One of the goals of OOD is the partitioning and distribution of behavior into many classes and many function." -- And then goes on to say sometimes its better to use God Classes anyway sometimes (Citing micro-controllers as an example). In Robert C Martin's "Clean Code: A Handbook of Agile Software Craftsmanship" page 136 (And only this page) talks about the "God class" and calls it out as a prime example of a violation of the "classes should be small" rule he uses to promote the Single Responsibility Principle" starting on on page 138. The problem I have is all my references and citations come from the same person (Robert C. Martin), and am from the same single person/source. I am being told that because he is just one guy, my desire to not use "God Classes" is invalid and not accepted as a standard best practice in the software industry. Is this true? Am I doing things wrong from a technical perspective by trying to keep to the teaching of Uncle Bob? God Objects and Object Oriented Programming and Design: The more I think of this the more I think this is more something you learn when you study OOP and its never explicitly called out; Its implicit to good design is my thinking (Feel free to correct me, please, as I want to learn), The problem is I "know" this, but but not everybody does, so in this case its not considered a valid argument because I am effectively calling it out as universal truth when in fact most people are statistically ignorant of it since statistically most people are not programmers. Conclusion: I am at a loss on what to search for to get the best additional results to cite, since they are making a technical claim and I want to know the truth and be able to prove it with citations like a real engineer/scientist, even if I am biased against god objects due to my personal experience with code that used them. Any assistance or citations would be deeply appreciated.

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