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  • An Honest look at SharePoint Web Services

    - by juanlarios
    INTRODUCTION If you are a SharePoint developer you know that there are two basic ways to develop against SharePoint. 1) The object Model 2) Web services. SharePoint object model has the advantage of being quite rich. Anything you can do through the SharePoint UI as an administrator or end user, you can do through the object model. In fact everything that is done through the UI is done through the object model behind the scenes. The major disadvantage to getting at SharePoint this way is that the code needs to run on the server. This means that all web parts, event receivers, features, etc… all of this is code that is deployed to the server. The second way to get to SharePoint is through the built in web services. There are many articles on how to manipulate web services, how to authenticate to them and interact with them. The basic idea is that a remote application or process can contact SharePoint through a web service. Lots has been written about how great these web services are. This article is written to document the limitations, some of the issues and frustrations with working with SharePoint built in web services. Ultimately, for the tasks I was given to , SharePoint built in web services did not suffice. My evaluation of SharePoint built in services was compared against creating my own WCF Services to do what I needed. The current project I'm working on right now involved several "integration points". A remote application, installed on a separate server was to contact SharePoint and perform an task or operation. So I decided to start up Visual Studio and built a DLL and basically have 2 layers of logic. An integration layer and a data layer. A good friend of mine pointed me to SOLID principles and referred me to some videos and tutorials about it. I decided to implement the methodology (although a lot of the principles are common sense and I already incorporated in my coding practices). I was to deliver this dll to the application team and they would simply call the methods exposed by this dll and voila! it would do some task or operation in SharePoint. SOLUTION My integration layer implemented an interface that defined some of the basic integration tasks that I was to put together. My data layer was about the same, it implemented an interface with some of the tasks that I was going to develop. This gave me the opportunity to develop different data layers, ultimately different ways to get at SharePoint if I needed to. This is a classic SOLID principle. In this case it proved to be quite helpful because I wrote one data layer completely implementing SharePoint built in Web Services and another implementing my own WCF Service that I wrote. I should mention there is another layer underneath the data layer. In referencing SharePoint or WCF services in my visual studio project I created a class for every web service call. So for example, if I used List.asx. I created a class called "DocumentRetreival" this class would do the grunt work to connect to the correct URL, It would perform the basic operation of contacting the service and so on. If I used a view.asmx, I implemented a class called "ViewRetrieval" with the same idea as the last class but it would now interact with all he operations in view.asmx. This gave my data layer the ability to perform multiple calls without really worrying about some of the grunt work each class performs. This again, is a classic SOLID principle. So, in order to compare them side by side we can look at both data layers and with is involved in each. Lets take a look at the "Create Project" task or operation. The integration point is described as , "dll is to provide a way to create a project in SharePoint". Projects , in this case are basically document libraries. I am to implement a way in which a remote application can create a document library in SharePoint. Easy enough right? Use the list.asmx Web service in SharePoint. So here we go! Lets take a look at the code. I added the List.asmx web service reference to my project and this is the class that contacts it:  class DocumentRetrieval     {         private ListsSoapClient _service;      d   private bool _impersonation;         public DocumentRetrieval(bool impersonation, string endpt)         {             _service = new ListsSoapClient();             this.SetEndPoint(string.Format("{0}/{1}", endpt, ConfigurationManager.AppSettings["List"]));             _impersonation = impersonation;             if (_impersonation)             {                 _service.ClientCredentials.Windows.ClientCredential.Password = ConfigurationManager.AppSettings["password"];                 _service.ClientCredentials.Windows.ClientCredential.UserName = ConfigurationManager.AppSettings["username"];                 _service.ClientCredentials.Windows.AllowedImpersonationLevel =                     System.Security.Principal.TokenImpersonationLevel.Impersonation;             }     private void SetEndPoint(string p)          {             _service.Endpoint.Address = new EndpointAddress(p);          }          /// <summary>         /// Creates a document library with specific name and templateID         /// </summary>         /// <param name="listName">New list name</param>         /// <param name="templateID">Template ID</param>         /// <returns></returns>         public XmlElement CreateLibrary(string listName, int templateID, ref ExceptionContract exContract)         {             XmlDocument sample = new XmlDocument();             XmlElement viewCol = sample.CreateElement("Empty");             try             {                 _service.Open();                 viewCol = _service.AddList(listName, "", templateID);             }             catch (Exception ex)             {                 exContract = new ExceptionContract("DocumentRetrieval/CreateLibrary", ex.GetType(), "Connection Error", ex.StackTrace, ExceptionContract.ExceptionCode.error);                             }finally             {                 _service.Close();             }                                      return viewCol;         } } There was a lot more in this class (that I am not including) because i was reusing the grunt work and making other operations with LIst.asmx, For example, updating content types, changing or configuring lists or document libraries. One of the first things I noticed about working with the built in services is that you are really at the mercy of what is available to you. Before creating a document library (Project) I wanted to expose a IsProjectExisting method. This way the integration or data layer could recognize if a library already exists. Well there is no service call or method available to do that check. So this is what I wrote:   public bool DocLibExists(string listName, ref ExceptionContract exContract)         {             try             {                 var allLists = _service.GetListCollection();                                return allLists.ChildNodes.OfType<XmlElement>().ToList().Exists(x => x.Attributes["Title"].Value ==listName);             }             catch (Exception ex)             {                 exContract = new ExceptionContract("DocumentRetrieval/GetList/GetListWSCall", ex.GetType(), "Unable to Retrieve List Collection", ex.StackTrace, ExceptionContract.ExceptionCode.error);             }             return false;         } This really just gets an XMLElement with all the lists. It was then up to me to sift through the clutter and noise and see if Document library already existed. This took a little bit of getting used to. Now instead of working with code, you are working with XMLElement response format from web service. I wrote a LINQ query to go through and find if the attribute "Title" existed and had a value of the listname then it would return True, if not False. I didn't particularly like working this way. Dealing with XMLElement responses and then having to manipulate it to get at the exact data I was looking for. Once the check for the DocLibExists, was done, I would either create the document library or send back an error indicating the document library already existed. Now lets examine the code that actually creates the document library. It does what you are really after, it creates a document library. Notice how the template ID is really an integer. Every document library template in SharePoint has an ID associated with it. Document libraries, Image Library, Custom List, Project Tasks, etc… they all he a unique integer associated with it. Well, that's great but the client came back to me and gave me some specifics that each "project" or document library, should have. They specified they had 3 types of projects. Each project would have unique views, about 10 views for each project. Each Project specified unique configurations (auditing, versioning, content types, etc…) So what turned out to be a simple implementation of creating a document library as a repository for a project, turned out to be quite involved.  The first thing I thought of was to create a template for document library. There are other ways you can do this too. Using the web Service call, you could configure views, versioning, even content types, etc… the only catch is, you have to be working quite extensively with CAML. I am not fond of CAML. I can do it and work with it, I just don't like doing it. It is quite touchy and at times it is quite tough to understand where errors were made with CAML statements. Working with Web Services and CAML proved to be quite annoying. The service call would return a generic error message that did not particularly point me to a CAML statement syntax error, or even a CAML error. I was not sure if it was a security , performance or code based issue. It was quite tough to work with. At times it was difficult to work with because of the way SharePoint handles metadata. There are "Names", "Display Name", and "StaticName" fields. It was quite tough to understand at times, which one to use. So it took a lot of trial and error. There are tools that can help with CAML generation. There is also now intellisense for CAML statements in Visual Studio that might help but ultimately I'm not fond of CAML with Web Services.   So I decided on the template. So my plan was to create create a document library, configure it accordingly and then use The Template Builder that comes with the SharePoint SDK. This tool allows you to create site templates, list template etc… It is quite interesting because it does not generate an STP file, it actually generates an xml definition and a feature you can activate and make that template available on a site or site collection. The first issue I experienced with this is that one of the specifications to this template was that the "All Documents" view was to have 2 web parts on it. Well, it turns out that using the template builder , it did not include the web parts as part of the list template definition it generated. It backed up the settings, the views, the content types but not the custom web parts. I still decided to try this even without the web parts on the page. This new template defined a new Document library definition with a unique ID. The problem was that the service call accepts an int but it only has access to the built in library int definitions. Any new ones added or created will not be available to create. So this made it impossible for me to approach the problem this way.     I should also mention that one of the nice features about SharePoint is the ability to create list templates, back them up and then create lists based on that template. It can all be done by end user administrators. These templates are quite unique because they are saved as an STP file and not an xml definition. I also went this route and tried to see if there was another service call where I could create a document library based no given template name. Nope! none.      After some thinking I decide to implement a WCF service to do this creation for me. I was quite certain that the object model would allow me to create document libraries base on a template in which an ID was required and also templates saved as STP files. Now I don't want to bother with posting the code to contact WCF service because it's self explanatory, but I will post the code that I used to create a list with custom template. public ServiceResult CreateProject(string name, string templateName, string projectId)         {             string siteurl = SPContext.Current.Site.Url;             Guid webguid = SPContext.Current.Web.ID;                        using (SPSite site = new SPSite(siteurl))             {                 using (SPWeb rootweb = site.RootWeb)                 {                     SPListTemplateCollection temps = site.GetCustomListTemplates(rootweb);                     ProcessWeb(siteurl, webguid, web => Act_CreateProject(web, name, templateName, projectId, temps));                 }//SpWeb             }//SPSite              return _globalResult;                   }         private void Act_CreateProject(SPWeb targetsite, string name, string templateName, string projectId, SPListTemplateCollection temps) {                         var temp = temps.Cast<SPListTemplate>().FirstOrDefault(x => x.Name.Equals(templateName));             if (temp != null)             {                             try                 {                                         Guid listGuid = targetsite.Lists.Add(name, "", temp);                     SPList newList = targetsite.Lists[listGuid];                     _globalResult = new ServiceResult(true, "Success", "Success");                 }                 catch (Exception ex)                 {                     _globalResult = new ServiceResult(false, (string.IsNullOrEmpty(ex.Message) ? "None" : ex.Message + " " + templateName), ex.StackTrace.ToString());                 }                                       }        private void ProcessWeb(string siteurl, Guid webguid, Action<SPWeb> action) {                        using (SPSite sitecollection = new SPSite(siteurl)) {                 using (SPWeb web = sitecollection.AllWebs[webguid]) {                     action(web);                 }                     }                  } This code is actually some of the code I implemented for the service. there was a lot more I did on Project Creation which I will cover in my next blog post. I implemented an ACTION method to process the web. This allowed me to properly dispose the SPWEb and SPSite objects and not rewrite this code over and over again. So I implemented a WCF service to create projects for me, this allowed me to do a lot more than just create a document library with a template, it now gave me the flexibility to do just about anything the client wanted at project creation. Once this was implemented , the client came back to me and said, "we reference all our projects with ID's in our application. we want SharePoint to do the same". This has been something I have been doing for a little while now but I do hope that SharePoint 2010 can have more of an answer to this and address it properly. I have been adding metadata to SPWebs through property bag. I believe I have blogged about it before. This time it required metadata added to a document library. No problem!!! I also mentioned these web parts that were to go on the "All Documents" View. I took the opportunity to configure them to the appropriate settings. There were two settings that needed to be set on these web parts. One of them was a Project ID configured in the webpart properties. The following code enhances and replaces the "Act_CreateProject " method above:  private void Act_CreateProject(SPWeb targetsite, string name, string templateName, string projectId, SPListTemplateCollection temps) {                         var temp = temps.Cast<SPListTemplate>().FirstOrDefault(x => x.Name.Equals(templateName));             if (temp != null)             {                 SPLimitedWebPartManager wpmgr = null;                               try                 {                                         Guid listGuid = targetsite.Lists.Add(name, "", temp);                     SPList newList = targetsite.Lists[listGuid];                     SPFolder rootFolder = newList.RootFolder;                     rootFolder.Properties.Add(KEY, projectId);                     rootFolder.Update();                     if (rootFolder.ParentWeb != targetsite)                         rootFolder.ParentWeb.Dispose();                     if (!templateName.Contains("Natural"))                     {                         SPView alldocumentsview = newList.Views.Cast<SPView>().FirstOrDefault(x => x.Title.Equals(ALLDOCUMENTS));                         SPFile alldocfile = targetsite.GetFile(alldocumentsview.ServerRelativeUrl);                         wpmgr = alldocfile.GetLimitedWebPartManager(PersonalizationScope.Shared);                         ConfigureWebPart(wpmgr, projectId, CUSTOMWPNAME);                                              alldocfile.Update();                     }                                        if (newList.ParentWeb != targetsite)                         newList.ParentWeb.Dispose();                     _globalResult = new ServiceResult(true, "Success", "Success");                 }                 catch (Exception ex)                 {                     _globalResult = new ServiceResult(false, (string.IsNullOrEmpty(ex.Message) ? "None" : ex.Message + " " + templateName), ex.StackTrace.ToString());                 }                 finally                 {                     if (wpmgr != null)                     {                         wpmgr.Web.Dispose();                         wpmgr.Dispose();                     }                 }             }                         }       private void ConfigureWebPart(SPLimitedWebPartManager mgr, string prjId, string webpartname)         {             var wp = mgr.WebParts.Cast<System.Web.UI.WebControls.WebParts.WebPart>().FirstOrDefault(x => x.DisplayTitle.Equals(webpartname));             if (wp != null)             {                           (wp as ListRelationshipWebPart.ListRelationshipWebPart).ProjectID = prjId;                 mgr.SaveChanges(wp);             }         }   This Shows you how I was able to set metadata on the document library. It has to be added to the RootFolder of the document library, Unfortunately, the SPList does not have a Property bag that I can add a key\value pair to. It has to be done on the root folder. Now everything in the integration will reference projects by ID's and will not care about names. My, "DocLibExists" will now need to be changed because a web service is not set up to look at property bags.  I had to write another method on the Service to do the equivalent but with ID's instead of names.  The second thing you will notice about the code is the use of the Webpartmanager. I have seen several examples online, and also read a lot about memory leaks, The above code does not produce memory leaks. The web part manager creates an SPWeb, so just dispose it like I did. CONCLUSION This is a long long post so I will stop here for now, I will continue with more comparisons and limitations in my next post. My conclusion for this example is that Web Services will do the trick if you can suffer through CAML and if you are doing some simple operations. For Everything else, there's WCF! **** fireI apologize for the disorganization of this post, I was on a bus on a 12 hour trip to IOWA while I wrote it, I was half asleep and half awake, hopefully it makes enough sense to someone.

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  • Advanced TSQL Tuning: Why Internals Knowledge Matters

    - by Paul White
    There is much more to query tuning than reducing logical reads and adding covering nonclustered indexes.  Query tuning is not complete as soon as the query returns results quickly in the development or test environments.  In production, your query will compete for memory, CPU, locks, I/O and other resources on the server.  Today’s entry looks at some tuning considerations that are often overlooked, and shows how deep internals knowledge can help you write better TSQL. As always, we’ll need some example data.  In fact, we are going to use three tables today, each of which is structured like this: Each table has 50,000 rows made up of an INTEGER id column and a padding column containing 3,999 characters in every row.  The only difference between the three tables is in the type of the padding column: the first table uses CHAR(3999), the second uses VARCHAR(MAX), and the third uses the deprecated TEXT type.  A script to create a database with the three tables and load the sample data follows: USE master; GO IF DB_ID('SortTest') IS NOT NULL DROP DATABASE SortTest; GO CREATE DATABASE SortTest COLLATE LATIN1_GENERAL_BIN; GO ALTER DATABASE SortTest MODIFY FILE ( NAME = 'SortTest', SIZE = 3GB, MAXSIZE = 3GB ); GO ALTER DATABASE SortTest MODIFY FILE ( NAME = 'SortTest_log', SIZE = 256MB, MAXSIZE = 1GB, FILEGROWTH = 128MB ); GO ALTER DATABASE SortTest SET ALLOW_SNAPSHOT_ISOLATION OFF ; ALTER DATABASE SortTest SET AUTO_CLOSE OFF ; ALTER DATABASE SortTest SET AUTO_CREATE_STATISTICS ON ; ALTER DATABASE SortTest SET AUTO_SHRINK OFF ; ALTER DATABASE SortTest SET AUTO_UPDATE_STATISTICS ON ; ALTER DATABASE SortTest SET AUTO_UPDATE_STATISTICS_ASYNC ON ; ALTER DATABASE SortTest SET PARAMETERIZATION SIMPLE ; ALTER DATABASE SortTest SET READ_COMMITTED_SNAPSHOT OFF ; ALTER DATABASE SortTest SET MULTI_USER ; ALTER DATABASE SortTest SET RECOVERY SIMPLE ; USE SortTest; GO CREATE TABLE dbo.TestCHAR ( id INTEGER IDENTITY (1,1) NOT NULL, padding CHAR(3999) NOT NULL,   CONSTRAINT [PK dbo.TestCHAR (id)] PRIMARY KEY CLUSTERED (id), ) ; CREATE TABLE dbo.TestMAX ( id INTEGER IDENTITY (1,1) NOT NULL, padding VARCHAR(MAX) NOT NULL,   CONSTRAINT [PK dbo.TestMAX (id)] PRIMARY KEY CLUSTERED (id), ) ; CREATE TABLE dbo.TestTEXT ( id INTEGER IDENTITY (1,1) NOT NULL, padding TEXT NOT NULL,   CONSTRAINT [PK dbo.TestTEXT (id)] PRIMARY KEY CLUSTERED (id), ) ; -- ============= -- Load TestCHAR (about 3s) -- ============= INSERT INTO dbo.TestCHAR WITH (TABLOCKX) ( padding ) SELECT padding = REPLICATE(CHAR(65 + (Data.n % 26)), 3999) FROM ( SELECT TOP (50000) n = ROW_NUMBER() OVER (ORDER BY (SELECT 0)) - 1 FROM master.sys.columns C1, master.sys.columns C2, master.sys.columns C3 ORDER BY n ASC ) AS Data ORDER BY Data.n ASC ; -- ============ -- Load TestMAX (about 3s) -- ============ INSERT INTO dbo.TestMAX WITH (TABLOCKX) ( padding ) SELECT CONVERT(VARCHAR(MAX), padding) FROM dbo.TestCHAR ORDER BY id ; -- ============= -- Load TestTEXT (about 5s) -- ============= INSERT INTO dbo.TestTEXT WITH (TABLOCKX) ( padding ) SELECT CONVERT(TEXT, padding) FROM dbo.TestCHAR ORDER BY id ; -- ========== -- Space used -- ========== -- EXECUTE sys.sp_spaceused @objname = 'dbo.TestCHAR'; EXECUTE sys.sp_spaceused @objname = 'dbo.TestMAX'; EXECUTE sys.sp_spaceused @objname = 'dbo.TestTEXT'; ; CHECKPOINT ; That takes around 15 seconds to run, and shows the space allocated to each table in its output: To illustrate the points I want to make today, the example task we are going to set ourselves is to return a random set of 150 rows from each table.  The basic shape of the test query is the same for each of the three test tables: SELECT TOP (150) T.id, T.padding FROM dbo.Test AS T ORDER BY NEWID() OPTION (MAXDOP 1) ; Test 1 – CHAR(3999) Running the template query shown above using the TestCHAR table as the target, we find that the query takes around 5 seconds to return its results.  This seems slow, considering that the table only has 50,000 rows.  Working on the assumption that generating a GUID for each row is a CPU-intensive operation, we might try enabling parallelism to see if that speeds up the response time.  Running the query again (but without the MAXDOP 1 hint) on a machine with eight logical processors, the query now takes 10 seconds to execute – twice as long as when run serially. Rather than attempting further guesses at the cause of the slowness, let’s go back to serial execution and add some monitoring.  The script below monitors STATISTICS IO output and the amount of tempdb used by the test query.  We will also run a Profiler trace to capture any warnings generated during query execution. DECLARE @read BIGINT, @write BIGINT ; SELECT @read = SUM(num_of_bytes_read), @write = SUM(num_of_bytes_written) FROM tempdb.sys.database_files AS DBF JOIN sys.dm_io_virtual_file_stats(2, NULL) AS FS ON FS.file_id = DBF.file_id WHERE DBF.type_desc = 'ROWS' ; SET STATISTICS IO ON ; SELECT TOP (150) TC.id, TC.padding FROM dbo.TestCHAR AS TC ORDER BY NEWID() OPTION (MAXDOP 1) ; SET STATISTICS IO OFF ; SELECT tempdb_read_MB = (SUM(num_of_bytes_read) - @read) / 1024. / 1024., tempdb_write_MB = (SUM(num_of_bytes_written) - @write) / 1024. / 1024., internal_use_MB = ( SELECT internal_objects_alloc_page_count / 128.0 FROM sys.dm_db_task_space_usage WHERE session_id = @@SPID ) FROM tempdb.sys.database_files AS DBF JOIN sys.dm_io_virtual_file_stats(2, NULL) AS FS ON FS.file_id = DBF.file_id WHERE DBF.type_desc = 'ROWS' ; Let’s take a closer look at the statistics and query plan generated from this: Following the flow of the data from right to left, we see the expected 50,000 rows emerging from the Clustered Index Scan, with a total estimated size of around 191MB.  The Compute Scalar adds a column containing a random GUID (generated from the NEWID() function call) for each row.  With this extra column in place, the size of the data arriving at the Sort operator is estimated to be 192MB. Sort is a blocking operator – it has to examine all of the rows on its input before it can produce its first row of output (the last row received might sort first).  This characteristic means that Sort requires a memory grant – memory allocated for the query’s use by SQL Server just before execution starts.  In this case, the Sort is the only memory-consuming operator in the plan, so it has access to the full 243MB (248,696KB) of memory reserved by SQL Server for this query execution. Notice that the memory grant is significantly larger than the expected size of the data to be sorted.  SQL Server uses a number of techniques to speed up sorting, some of which sacrifice size for comparison speed.  Sorts typically require a very large number of comparisons, so this is usually a very effective optimization.  One of the drawbacks is that it is not possible to exactly predict the sort space needed, as it depends on the data itself.  SQL Server takes an educated guess based on data types, sizes, and the number of rows expected, but the algorithm is not perfect. In spite of the large memory grant, the Profiler trace shows a Sort Warning event (indicating that the sort ran out of memory), and the tempdb usage monitor shows that 195MB of tempdb space was used – all of that for system use.  The 195MB represents physical write activity on tempdb, because SQL Server strictly enforces memory grants – a query cannot ‘cheat’ and effectively gain extra memory by spilling to tempdb pages that reside in memory.  Anyway, the key point here is that it takes a while to write 195MB to disk, and this is the main reason that the query takes 5 seconds overall. If you are wondering why using parallelism made the problem worse, consider that eight threads of execution result in eight concurrent partial sorts, each receiving one eighth of the memory grant.  The eight sorts all spilled to tempdb, resulting in inefficiencies as the spilled sorts competed for disk resources.  More importantly, there are specific problems at the point where the eight partial results are combined, but I’ll cover that in a future post. CHAR(3999) Performance Summary: 5 seconds elapsed time 243MB memory grant 195MB tempdb usage 192MB estimated sort set 25,043 logical reads Sort Warning Test 2 – VARCHAR(MAX) We’ll now run exactly the same test (with the additional monitoring) on the table using a VARCHAR(MAX) padding column: DECLARE @read BIGINT, @write BIGINT ; SELECT @read = SUM(num_of_bytes_read), @write = SUM(num_of_bytes_written) FROM tempdb.sys.database_files AS DBF JOIN sys.dm_io_virtual_file_stats(2, NULL) AS FS ON FS.file_id = DBF.file_id WHERE DBF.type_desc = 'ROWS' ; SET STATISTICS IO ON ; SELECT TOP (150) TM.id, TM.padding FROM dbo.TestMAX AS TM ORDER BY NEWID() OPTION (MAXDOP 1) ; SET STATISTICS IO OFF ; SELECT tempdb_read_MB = (SUM(num_of_bytes_read) - @read) / 1024. / 1024., tempdb_write_MB = (SUM(num_of_bytes_written) - @write) / 1024. / 1024., internal_use_MB = ( SELECT internal_objects_alloc_page_count / 128.0 FROM sys.dm_db_task_space_usage WHERE session_id = @@SPID ) FROM tempdb.sys.database_files AS DBF JOIN sys.dm_io_virtual_file_stats(2, NULL) AS FS ON FS.file_id = DBF.file_id WHERE DBF.type_desc = 'ROWS' ; This time the query takes around 8 seconds to complete (3 seconds longer than Test 1).  Notice that the estimated row and data sizes are very slightly larger, and the overall memory grant has also increased very slightly to 245MB.  The most marked difference is in the amount of tempdb space used – this query wrote almost 391MB of sort run data to the physical tempdb file.  Don’t draw any general conclusions about VARCHAR(MAX) versus CHAR from this – I chose the length of the data specifically to expose this edge case.  In most cases, VARCHAR(MAX) performs very similarly to CHAR – I just wanted to make test 2 a bit more exciting. MAX Performance Summary: 8 seconds elapsed time 245MB memory grant 391MB tempdb usage 193MB estimated sort set 25,043 logical reads Sort warning Test 3 – TEXT The same test again, but using the deprecated TEXT data type for the padding column: DECLARE @read BIGINT, @write BIGINT ; SELECT @read = SUM(num_of_bytes_read), @write = SUM(num_of_bytes_written) FROM tempdb.sys.database_files AS DBF JOIN sys.dm_io_virtual_file_stats(2, NULL) AS FS ON FS.file_id = DBF.file_id WHERE DBF.type_desc = 'ROWS' ; SET STATISTICS IO ON ; SELECT TOP (150) TT.id, TT.padding FROM dbo.TestTEXT AS TT ORDER BY NEWID() OPTION (MAXDOP 1, RECOMPILE) ; SET STATISTICS IO OFF ; SELECT tempdb_read_MB = (SUM(num_of_bytes_read) - @read) / 1024. / 1024., tempdb_write_MB = (SUM(num_of_bytes_written) - @write) / 1024. / 1024., internal_use_MB = ( SELECT internal_objects_alloc_page_count / 128.0 FROM sys.dm_db_task_space_usage WHERE session_id = @@SPID ) FROM tempdb.sys.database_files AS DBF JOIN sys.dm_io_virtual_file_stats(2, NULL) AS FS ON FS.file_id = DBF.file_id WHERE DBF.type_desc = 'ROWS' ; This time the query runs in 500ms.  If you look at the metrics we have been checking so far, it’s not hard to understand why: TEXT Performance Summary: 0.5 seconds elapsed time 9MB memory grant 5MB tempdb usage 5MB estimated sort set 207 logical reads 596 LOB logical reads Sort warning SQL Server’s memory grant algorithm still underestimates the memory needed to perform the sorting operation, but the size of the data to sort is so much smaller (5MB versus 193MB previously) that the spilled sort doesn’t matter very much.  Why is the data size so much smaller?  The query still produces the correct results – including the large amount of data held in the padding column – so what magic is being performed here? TEXT versus MAX Storage The answer lies in how columns of the TEXT data type are stored.  By default, TEXT data is stored off-row in separate LOB pages – which explains why this is the first query we have seen that records LOB logical reads in its STATISTICS IO output.  You may recall from my last post that LOB data leaves an in-row pointer to the separate storage structure holding the LOB data. SQL Server can see that the full LOB value is not required by the query plan until results are returned, so instead of passing the full LOB value down the plan from the Clustered Index Scan, it passes the small in-row structure instead.  SQL Server estimates that each row coming from the scan will be 79 bytes long – 11 bytes for row overhead, 4 bytes for the integer id column, and 64 bytes for the LOB pointer (in fact the pointer is rather smaller – usually 16 bytes – but the details of that don’t really matter right now). OK, so this query is much more efficient because it is sorting a very much smaller data set – SQL Server delays retrieving the LOB data itself until after the Sort starts producing its 150 rows.  The question that normally arises at this point is: Why doesn’t SQL Server use the same trick when the padding column is defined as VARCHAR(MAX)? The answer is connected with the fact that if the actual size of the VARCHAR(MAX) data is 8000 bytes or less, it is usually stored in-row in exactly the same way as for a VARCHAR(8000) column – MAX data only moves off-row into LOB storage when it exceeds 8000 bytes.  The default behaviour of the TEXT type is to be stored off-row by default, unless the ‘text in row’ table option is set suitably and there is room on the page.  There is an analogous (but opposite) setting to control the storage of MAX data – the ‘large value types out of row’ table option.  By enabling this option for a table, MAX data will be stored off-row (in a LOB structure) instead of in-row.  SQL Server Books Online has good coverage of both options in the topic In Row Data. The MAXOOR Table The essential difference, then, is that MAX defaults to in-row storage, and TEXT defaults to off-row (LOB) storage.  You might be thinking that we could get the same benefits seen for the TEXT data type by storing the VARCHAR(MAX) values off row – so let’s look at that option now.  This script creates a fourth table, with the VARCHAR(MAX) data stored off-row in LOB pages: CREATE TABLE dbo.TestMAXOOR ( id INTEGER IDENTITY (1,1) NOT NULL, padding VARCHAR(MAX) NOT NULL,   CONSTRAINT [PK dbo.TestMAXOOR (id)] PRIMARY KEY CLUSTERED (id), ) ; EXECUTE sys.sp_tableoption @TableNamePattern = N'dbo.TestMAXOOR', @OptionName = 'large value types out of row', @OptionValue = 'true' ; SELECT large_value_types_out_of_row FROM sys.tables WHERE [schema_id] = SCHEMA_ID(N'dbo') AND name = N'TestMAXOOR' ; INSERT INTO dbo.TestMAXOOR WITH (TABLOCKX) ( padding ) SELECT SPACE(0) FROM dbo.TestCHAR ORDER BY id ; UPDATE TM WITH (TABLOCK) SET padding.WRITE (TC.padding, NULL, NULL) FROM dbo.TestMAXOOR AS TM JOIN dbo.TestCHAR AS TC ON TC.id = TM.id ; EXECUTE sys.sp_spaceused @objname = 'dbo.TestMAXOOR' ; CHECKPOINT ; Test 4 – MAXOOR We can now re-run our test on the MAXOOR (MAX out of row) table: DECLARE @read BIGINT, @write BIGINT ; SELECT @read = SUM(num_of_bytes_read), @write = SUM(num_of_bytes_written) FROM tempdb.sys.database_files AS DBF JOIN sys.dm_io_virtual_file_stats(2, NULL) AS FS ON FS.file_id = DBF.file_id WHERE DBF.type_desc = 'ROWS' ; SET STATISTICS IO ON ; SELECT TOP (150) MO.id, MO.padding FROM dbo.TestMAXOOR AS MO ORDER BY NEWID() OPTION (MAXDOP 1, RECOMPILE) ; SET STATISTICS IO OFF ; SELECT tempdb_read_MB = (SUM(num_of_bytes_read) - @read) / 1024. / 1024., tempdb_write_MB = (SUM(num_of_bytes_written) - @write) / 1024. / 1024., internal_use_MB = ( SELECT internal_objects_alloc_page_count / 128.0 FROM sys.dm_db_task_space_usage WHERE session_id = @@SPID ) FROM tempdb.sys.database_files AS DBF JOIN sys.dm_io_virtual_file_stats(2, NULL) AS FS ON FS.file_id = DBF.file_id WHERE DBF.type_desc = 'ROWS' ; TEXT Performance Summary: 0.3 seconds elapsed time 245MB memory grant 0MB tempdb usage 193MB estimated sort set 207 logical reads 446 LOB logical reads No sort warning The query runs very quickly – slightly faster than Test 3, and without spilling the sort to tempdb (there is no sort warning in the trace, and the monitoring query shows zero tempdb usage by this query).  SQL Server is passing the in-row pointer structure down the plan and only looking up the LOB value on the output side of the sort. The Hidden Problem There is still a huge problem with this query though – it requires a 245MB memory grant.  No wonder the sort doesn’t spill to tempdb now – 245MB is about 20 times more memory than this query actually requires to sort 50,000 records containing LOB data pointers.  Notice that the estimated row and data sizes in the plan are the same as in test 2 (where the MAX data was stored in-row). The optimizer assumes that MAX data is stored in-row, regardless of the sp_tableoption setting ‘large value types out of row’.  Why?  Because this option is dynamic – changing it does not immediately force all MAX data in the table in-row or off-row, only when data is added or actually changed.  SQL Server does not keep statistics to show how much MAX or TEXT data is currently in-row, and how much is stored in LOB pages.  This is an annoying limitation, and one which I hope will be addressed in a future version of the product. So why should we worry about this?  Excessive memory grants reduce concurrency and may result in queries waiting on the RESOURCE_SEMAPHORE wait type while they wait for memory they do not need.  245MB is an awful lot of memory, especially on 32-bit versions where memory grants cannot use AWE-mapped memory.  Even on a 64-bit server with plenty of memory, do you really want a single query to consume 0.25GB of memory unnecessarily?  That’s 32,000 8KB pages that might be put to much better use. The Solution The answer is not to use the TEXT data type for the padding column.  That solution happens to have better performance characteristics for this specific query, but it still results in a spilled sort, and it is hard to recommend the use of a data type which is scheduled for removal.  I hope it is clear to you that the fundamental problem here is that SQL Server sorts the whole set arriving at a Sort operator.  Clearly, it is not efficient to sort the whole table in memory just to return 150 rows in a random order. The TEXT example was more efficient because it dramatically reduced the size of the set that needed to be sorted.  We can do the same thing by selecting 150 unique keys from the table at random (sorting by NEWID() for example) and only then retrieving the large padding column values for just the 150 rows we need.  The following script implements that idea for all four tables: SET STATISTICS IO ON ; WITH TestTable AS ( SELECT * FROM dbo.TestCHAR ), TopKeys AS ( SELECT TOP (150) id FROM TestTable ORDER BY NEWID() ) SELECT T1.id, T1.padding FROM TestTable AS T1 WHERE T1.id = ANY (SELECT id FROM TopKeys) OPTION (MAXDOP 1) ; WITH TestTable AS ( SELECT * FROM dbo.TestMAX ), TopKeys AS ( SELECT TOP (150) id FROM TestTable ORDER BY NEWID() ) SELECT T1.id, T1.padding FROM TestTable AS T1 WHERE T1.id IN (SELECT id FROM TopKeys) OPTION (MAXDOP 1) ; WITH TestTable AS ( SELECT * FROM dbo.TestTEXT ), TopKeys AS ( SELECT TOP (150) id FROM TestTable ORDER BY NEWID() ) SELECT T1.id, T1.padding FROM TestTable AS T1 WHERE T1.id IN (SELECT id FROM TopKeys) OPTION (MAXDOP 1) ; WITH TestTable AS ( SELECT * FROM dbo.TestMAXOOR ), TopKeys AS ( SELECT TOP (150) id FROM TestTable ORDER BY NEWID() ) SELECT T1.id, T1.padding FROM TestTable AS T1 WHERE T1.id IN (SELECT id FROM TopKeys) OPTION (MAXDOP 1) ; SET STATISTICS IO OFF ; All four queries now return results in much less than a second, with memory grants between 6 and 12MB, and without spilling to tempdb.  The small remaining inefficiency is in reading the id column values from the clustered primary key index.  As a clustered index, it contains all the in-row data at its leaf.  The CHAR and VARCHAR(MAX) tables store the padding column in-row, so id values are separated by a 3999-character column, plus row overhead.  The TEXT and MAXOOR tables store the padding values off-row, so id values in the clustered index leaf are separated by the much-smaller off-row pointer structure.  This difference is reflected in the number of logical page reads performed by the four queries: Table 'TestCHAR' logical reads 25511 lob logical reads 000 Table 'TestMAX'. logical reads 25511 lob logical reads 000 Table 'TestTEXT' logical reads 00412 lob logical reads 597 Table 'TestMAXOOR' logical reads 00413 lob logical reads 446 We can increase the density of the id values by creating a separate nonclustered index on the id column only.  This is the same key as the clustered index, of course, but the nonclustered index will not include the rest of the in-row column data. CREATE UNIQUE NONCLUSTERED INDEX uq1 ON dbo.TestCHAR (id); CREATE UNIQUE NONCLUSTERED INDEX uq1 ON dbo.TestMAX (id); CREATE UNIQUE NONCLUSTERED INDEX uq1 ON dbo.TestTEXT (id); CREATE UNIQUE NONCLUSTERED INDEX uq1 ON dbo.TestMAXOOR (id); The four queries can now use the very dense nonclustered index to quickly scan the id values, sort them by NEWID(), select the 150 ids we want, and then look up the padding data.  The logical reads with the new indexes in place are: Table 'TestCHAR' logical reads 835 lob logical reads 0 Table 'TestMAX' logical reads 835 lob logical reads 0 Table 'TestTEXT' logical reads 686 lob logical reads 597 Table 'TestMAXOOR' logical reads 686 lob logical reads 448 With the new index, all four queries use the same query plan (click to enlarge): Performance Summary: 0.3 seconds elapsed time 6MB memory grant 0MB tempdb usage 1MB sort set 835 logical reads (CHAR, MAX) 686 logical reads (TEXT, MAXOOR) 597 LOB logical reads (TEXT) 448 LOB logical reads (MAXOOR) No sort warning I’ll leave it as an exercise for the reader to work out why trying to eliminate the Key Lookup by adding the padding column to the new nonclustered indexes would be a daft idea Conclusion This post is not about tuning queries that access columns containing big strings.  It isn’t about the internal differences between TEXT and MAX data types either.  It isn’t even about the cool use of UPDATE .WRITE used in the MAXOOR table load.  No, this post is about something else: Many developers might not have tuned our starting example query at all – 5 seconds isn’t that bad, and the original query plan looks reasonable at first glance.  Perhaps the NEWID() function would have been blamed for ‘just being slow’ – who knows.  5 seconds isn’t awful – unless your users expect sub-second responses – but using 250MB of memory and writing 200MB to tempdb certainly is!  If ten sessions ran that query at the same time in production that’s 2.5GB of memory usage and 2GB hitting tempdb.  Of course, not all queries can be rewritten to avoid large memory grants and sort spills using the key-lookup technique in this post, but that’s not the point either. The point of this post is that a basic understanding of execution plans is not enough.  Tuning for logical reads and adding covering indexes is not enough.  If you want to produce high-quality, scalable TSQL that won’t get you paged as soon as it hits production, you need a deep understanding of execution plans, and as much accurate, deep knowledge about SQL Server as you can lay your hands on.  The advanced database developer has a wide range of tools to use in writing queries that perform well in a range of circumstances. By the way, the examples in this post were written for SQL Server 2008.  They will run on 2005 and demonstrate the same principles, but you won’t get the same figures I did because 2005 had a rather nasty bug in the Top N Sort operator.  Fair warning: if you do decide to run the scripts on a 2005 instance (particularly the parallel query) do it before you head out for lunch… This post is dedicated to the people of Christchurch, New Zealand. © 2011 Paul White email: @[email protected] twitter: @SQL_Kiwi

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  • The Incremental Architect&acute;s Napkin &ndash; #3 &ndash; Make Evolvability inevitable

    - by Ralf Westphal
    Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/06/04/the-incremental-architectacutes-napkin-ndash-3-ndash-make-evolvability-inevitable.aspxThe easier something to measure the more likely it will be produced. Deviations between what is and what should be can be readily detected. That´s what automated acceptance tests are for. That´s what sprint reviews in Scrum are for. It´s no small wonder our software looks like it looks. It has all the traits whose conformance with requirements can easily be measured. And it´s lacking traits which cannot easily be measured. Evolvability (or Changeability) is such a trait. If an operation is correct, if an operation if fast enough, that can be checked very easily. But whether Evolvability is high or low, that cannot be checked by taking a measure or two. Evolvability might correlate with certain traits, e.g. number of lines of code (LOC) per function or Cyclomatic Complexity or test coverage. But there is no threshold value signalling “evolvability too low”; also Evolvability is hardly tangible for the customer. Nevertheless Evolvability is of great importance - at least in the long run. You can get away without much of it for a short time. Eventually, though, it´s needed like any other requirement. Or even more. Because without Evolvability no other requirement can be implemented. Evolvability is the foundation on which all else is build. Such fundamental importance is in stark contrast with its immeasurability. To compensate this, Evolvability must be put at the very center of software development. It must become the hub around everything else revolves. Since we cannot measure Evolvability, though, we cannot start watching it more. Instead we need to establish practices to keep it high (enough) at all times. Chefs have known that for long. That´s why everybody in a restaurant kitchen is constantly seeing after cleanliness. Hygiene is important as is to have clean tools at standardized locations. Only then the health of the patrons can be guaranteed and production efficiency is constantly high. Still a kitchen´s level of cleanliness is easier to measure than software Evolvability. That´s why important practices like reviews, pair programming, or TDD are not enough, I guess. What we need to keep Evolvability in focus and high is… to continually evolve. Change must not be something to avoid but too embrace. To me that means the whole change cycle from requirement analysis to delivery needs to be gone through more often. Scrum´s sprints of 4, 2 even 1 week are too long. Kanban´s flow of user stories across is too unreliable; it takes as long as it takes. Instead we should fix the cycle time at 2 days max. I call that Spinning. No increment must take longer than from this morning until tomorrow evening to finish. Then it should be acceptance checked by the customer (or his/her representative, e.g. a Product Owner). For me there are several resasons for such a fixed and short cycle time for each increment: Clear expectations Absolute estimates (“This will take X days to complete.”) are near impossible in software development as explained previously. Too much unplanned research and engineering work lurk in every feature. And then pervasive interruptions of work by peers and management. However, the smaller the scope the better our absolute estimates become. That´s because we understand better what really are the requirements and what the solution should look like. But maybe more importantly the shorter the timespan the more we can control how we use our time. So much can happen over the course of a week and longer timespans. But if push comes to shove I can block out all distractions and interruptions for a day or possibly two. That´s why I believe we can give rough absolute estimates on 3 levels: Noon Tonight Tomorrow Think of a meeting with a Product Owner at 8:30 in the morning. If she asks you, how long it will take you to implement a user story or bug fix, you can say, “It´ll be fixed by noon.”, or you can say, “I can manage to implement it until tonight before I leave.”, or you can say, “You´ll get it by tomorrow night at latest.” Yes, I believe all else would be naive. If you´re not confident to get something done by tomorrow night (some 34h from now) you just cannot reliably commit to any timeframe. That means you should not promise anything, you should not even start working on the issue. So when estimating use these four categories: Noon, Tonight, Tomorrow, NoClue - with NoClue meaning the requirement needs to be broken down further so each aspect can be assigned to one of the first three categories. If you like absolute estimates, here you go. But don´t do deep estimates. Don´t estimate dozens of issues; don´t think ahead (“Issue A is a Tonight, then B will be a Tomorrow, after that it´s C as a Noon, finally D is a Tonight - that´s what I´ll do this week.”). Just estimate so Work-in-Progress (WIP) is 1 for everybody - plus a small number of buffer issues. To be blunt: Yes, this makes promises impossible as to what a team will deliver in terms of scope at a certain date in the future. But it will give a Product Owner a clear picture of what to pull for acceptance feedback tonight and tomorrow. Trust through reliability Our trade is lacking trust. Customers don´t trust software companies/departments much. Managers don´t trust developers much. I find that perfectly understandable in the light of what we´re trying to accomplish: delivering software in the face of uncertainty by means of material good production. Customers as well as managers still expect software development to be close to production of houses or cars. But that´s a fundamental misunderstanding. Software development ist development. It´s basically research. As software developers we´re constantly executing experiments to find out what really provides value to users. We don´t know what they need, we just have mediated hypothesises. That´s why we cannot reliably deliver on preposterous demands. So trust is out of the window in no time. If we switch to delivering in short cycles, though, we can regain trust. Because estimates - explicit or implicit - up to 32 hours at most can be satisfied. I´d say: reliability over scope. It´s more important to reliably deliver what was promised then to cover a lot of requirement area. So when in doubt promise less - but deliver without delay. Deliver on scope (Functionality and Quality); but also deliver on Evolvability, i.e. on inner quality according to accepted principles. Always. Trust will be the reward. Less complexity of communication will follow. More goodwill buffer will follow. So don´t wait for some Kanban board to show you, that flow can be improved by scheduling smaller stories. You don´t need to learn that the hard way. Just start with small batch sizes of three different sizes. Fast feedback What has been finished can be checked for acceptance. Why wait for a sprint of several weeks to end? Why let the mental model of the issue and its solution dissipate? If you get final feedback after one or two weeks, you hardly remember what you did and why you did it. Resoning becomes hard. But more importantly youo probably are not in the mood anymore to go back to something you deemed done a long time ago. It´s boring, it´s frustrating to open up that mental box again. Learning is harder the longer it takes from event to feedback. Effort can be wasted between event (finishing an issue) and feedback, because other work might go in the wrong direction based on false premises. Checking finished issues for acceptance is the most important task of a Product Owner. It´s even more important than planning new issues. Because as long as work started is not released (accepted) it´s potential waste. So before starting new work better make sure work already done has value. By putting the emphasis on acceptance rather than planning true pull is established. As long as planning and starting work is more important, it´s a push process. Accept a Noon issue on the same day before leaving. Accept a Tonight issue before leaving today or first thing tomorrow morning. Accept a Tomorrow issue tomorrow night before leaving or early the day after tomorrow. After acceptance the developer(s) can start working on the next issue. Flexibility As if reliability/trust and fast feedback for less waste weren´t enough economic incentive, there is flexibility. After each issue the Product Owner can change course. If on Monday morning feature slices A, B, C, D, E were important and A, B, C were scheduled for acceptance by Monday evening and Tuesday evening, the Product Owner can change her mind at any time. Maybe after A got accepted she asks for continuation with D. But maybe, just maybe, she has gotten a completely different idea by then. Maybe she wants work to continue on F. And after B it´s neither D nor E, but G. And after G it´s D. With Spinning every 32 hours at latest priorities can be changed. And nothing is lost. Because what got accepted is of value. It provides an incremental value to the customer/user. Or it provides internal value to the Product Owner as increased knowledge/decreased uncertainty. I find such reactivity over commitment economically very benefical. Why commit a team to some workload for several weeks? It´s unnecessary at beast, and inflexible and wasteful at worst. If we cannot promise delivery of a certain scope on a certain date - which is what customers/management usually want -, we can at least provide them with unpredecented flexibility in the face of high uncertainty. Where the path is not clear, cannot be clear, make small steps so you´re able to change your course at any time. Premature completion Customers/management are used to premeditating budgets. They want to know exactly how much to pay for a certain amount of requirements. That´s understandable. But it does not match with the nature of software development. We should know that by now. Maybe there´s somewhere in the world some team who can consistently deliver on scope, quality, and time, and budget. Great! Congratulations! I, however, haven´t seen such a team yet. Which does not mean it´s impossible, but I think it´s nothing I can recommend to strive for. Rather I´d say: Don´t try this at home. It might hurt you one way or the other. However, what we can do, is allow customers/management stop work on features at any moment. With spinning every 32 hours a feature can be declared as finished - even though it might not be completed according to initial definition. I think, progress over completion is an important offer software development can make. Why think in terms of completion beyond a promise for the next 32 hours? Isn´t it more important to constantly move forward? Step by step. We´re not running sprints, we´re not running marathons, not even ultra-marathons. We´re in the sport of running forever. That makes it futile to stare at the finishing line. The very concept of a burn-down chart is misleading (in most cases). Whoever can only think in terms of completed requirements shuts out the chance for saving money. The requirements for a features mostly are uncertain. So how does a Product Owner know in the first place, how much is needed. Maybe more than specified is needed - which gets uncovered step by step with each finished increment. Maybe less than specified is needed. After each 4–32 hour increment the Product Owner can do an experient (or invite users to an experiment) if a particular trait of the software system is already good enough. And if so, she can switch the attention to a different aspect. In the end, requirements A, B, C then could be finished just 70%, 80%, and 50%. What the heck? It´s good enough - for now. 33% money saved. Wouldn´t that be splendid? Isn´t that a stunning argument for any budget-sensitive customer? You can save money and still get what you need? Pull on practices So far, in addition to more trust, more flexibility, less money spent, Spinning led to “doing less” which also means less code which of course means higher Evolvability per se. Last but not least, though, I think Spinning´s short acceptance cycles have one more effect. They excert pull-power on all sorts of practices known for increasing Evolvability. If, for example, you believe high automated test coverage helps Evolvability by lowering the fear of inadverted damage to a code base, why isn´t 90% of the developer community practicing automated tests consistently? I think, the answer is simple: Because they can do without. Somehow they manage to do enough manual checks before their rare releases/acceptance checks to ensure good enough correctness - at least in the short term. The same goes for other practices like component orientation, continuous build/integration, code reviews etc. None of that is compelling, urgent, imperative. Something else always seems more important. So Evolvability principles and practices fall through the cracks most of the time - until a project hits a wall. Then everybody becomes desperate; but by then (re)gaining Evolvability has become as very, very difficult and tedious undertaking. Sometimes up to the point where the existence of a project/company is in danger. With Spinning that´s different. If you´re practicing Spinning you cannot avoid all those practices. With Spinning you very quickly realize you cannot deliver reliably even on your 32 hour promises. Spinning thus is pulling on developers to adopt principles and practices for Evolvability. They will start actively looking for ways to keep their delivery rate high. And if not, management will soon tell them to do that. Because first the Product Owner then management will notice an increasing difficulty to deliver value within 32 hours. There, finally there emerges a way to measure Evolvability: The more frequent developers tell the Product Owner there is no way to deliver anything worth of feedback until tomorrow night, the poorer Evolvability is. Don´t count the “WTF!”, count the “No way!” utterances. In closing For sustainable software development we need to put Evolvability first. Functionality and Quality must not rule software development but be implemented within a framework ensuring (enough) Evolvability. Since Evolvability cannot be measured easily, I think we need to put software development “under pressure”. Software needs to be changed more often, in smaller increments. Each increment being relevant to the customer/user in some way. That does not mean each increment is worthy of shipment. It´s sufficient to gain further insight from it. Increments primarily serve the reduction of uncertainty, not sales. Sales even needs to be decoupled from this incremental progress. No more promises to sales. No more delivery au point. Rather sales should look at a stream of accepted increments (or incremental releases) and scoup from that whatever they find valuable. Sales and marketing need to realize they should work on what´s there, not what might be possible in the future. But I digress… In my view a Spinning cycle - which is not easy to reach, which requires practice - is the core practice to compensate the immeasurability of Evolvability. From start to finish of each issue in 32 hours max - that´s the challenge we need to accept if we´re serious increasing Evolvability. Fortunately higher Evolvability is not the only outcome of Spinning. Customer/management will like the increased flexibility and “getting more bang for the buck”.

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  • ANTS Memory Profiler 7.0 Review

    - by Michael B. McLaughlin
    (This is my first review as a part of the GeeksWithBlogs.net Influencers program. It’s a program in which I (and the others who have been selected for it) get the opportunity to check out new products and services and write reviews about them. We don’t get paid for this, but we do generally get to keep a copy of the software or retain an account for some period of time on the service that we review. In this case I received a copy of Red Gate Software’s ANTS Memory Profiler 7.0, which was released in January. I don’t have any upgrade rights nor is my review guided, restrained, influenced, or otherwise controlled by Red Gate or anyone else. But I do get to keep the software license. I will always be clear about what I received whenever I do a review – I leave it up to you to decide whether you believe I can be objective. I believe I can be. If I used something and really didn’t like it, keeping a copy of it wouldn’t be worth anything to me. In that case though, I would simply uninstall/deactivate/whatever the software or service and tell the company what I didn’t like about it so they could (hopefully) make it better in the future. I don’t think it’d be polite to write up a terrible review, nor do I think it would be a particularly good use of my time. There are people who get paid for a living to review things, so I leave it to them to tell you what they think is bad and why. I’ll only spend my time telling you about things I think are good.) Overview of Common .NET Memory Problems When coming to land of managed memory from the wilds of unmanaged code, it’s easy to say to one’s self, “Wow! Now I never have to worry about memory problems again!” But this simply isn’t true. Managed code environments, such as .NET, make many, many things easier. You will never have to worry about memory corruption due to a bad pointer, for example (unless you’re working with unsafe code, of course). But managed code has its own set of memory concerns. For example, failing to unsubscribe from events when you are done with them leaves the publisher of an event with a reference to the subscriber. If you eliminate all your own references to the subscriber, then that memory is effectively lost since the GC won’t delete it because of the publishing object’s reference. When the publishing object itself becomes subject to garbage collection then you’ll get that memory back finally, but that could take a very long time depending of the life of the publisher. Another common source of resource leaks is failing to properly release unmanaged resources. When writing a class that contains members that hold unmanaged resources (e.g. any of the Stream-derived classes, IsolatedStorageFile, most classes ending in “Reader” or “Writer”), you should always implement IDisposable, making sure to use a properly written Dispose method. And when you are using an instance of a class that implements IDisposable, you should always make sure to use a 'using' statement in order to ensure that the object’s unmanaged resources are disposed of properly. (A ‘using’ statement is a nicer, cleaner looking, and easier to use version of a try-finally block. The compiler actually translates it as though it were a try-finally block. Note that Code Analysis warning 2202 (CA2202) will often be triggered by nested using blocks. A properly written dispose method ensures that it only runs once such that calling dispose multiple times should not be a problem. Nonetheless, CA2202 exists and if you want to avoid triggering it then you should write your code such that only the innermost IDisposable object uses a ‘using’ statement, with any outer code making use of appropriate try-finally blocks instead). Then, of course, there are situations where you are operating in a memory-constrained environment or else you want to limit or even eliminate allocations within a certain part of your program (e.g. within the main game loop of an XNA game) in order to avoid having the GC run. On the Xbox 360 and Windows Phone 7, for example, for every 1 MB of heap allocations you make, the GC runs; the added time of a GC collection can cause a game to drop frames or run slowly thereby making it look bad. Eliminating allocations (or else minimizing them and calling an explicit Collect at an appropriate time) is a common way of avoiding this (the other way is to simplify your heap so that the GC’s latency is low enough not to cause performance issues). ANTS Memory Profiler 7.0 When the opportunity to review Red Gate’s recently released ANTS Memory Profiler 7.0 arose, I jumped at it. In order to review it, I was given a free copy (which does not include upgrade rights for future versions) which I am allowed to keep. For those of you who are familiar with ANTS Memory Profiler, you can find a list of new features and enhancements here. If you are an experienced .NET developer who is familiar with .NET memory management issues, ANTS Memory Profiler is great. More importantly still, if you are new to .NET development or you have no experience or limited experience with memory profiling, ANTS Memory Profiler is awesome. From the very beginning, it guides you through the process of memory profiling. If you’re experienced and just want dive in however, it doesn’t get in your way. The help items GAHSFLASHDAJLDJA are well designed and located right next to the UI controls so that they are easy to find without being intrusive. When you first launch it, it presents you with a “Getting Started” screen that contains links to “Memory profiling video tutorials”, “Strategies for memory profiling”, and the “ANTS Memory Profiler forum”. I’m normally the kind of person who looks at a screen like that only to find the “Don’t show this again” checkbox. Since I was doing a review, though, I decided I should examine them. I was pleasantly surprised. The overview video clocks in at three minutes and fifty seconds. It begins by showing you how to get started profiling an application. It explains that profiling is done by taking memory snapshots periodically while your program is running and then comparing them. ANTS Memory Profiler (I’m just going to call it “ANTS MP” from here) analyzes these snapshots in the background while your application is running. It briefly mentions a new feature in Version 7, a new API that give you the ability to trigger snapshots from within your application’s source code (more about this below). You can also, and this is the more common way you would do it, take a memory snapshot at any time from within the ANTS MP window by clicking the “Take Memory Snapshot” button in the upper right corner. The overview video goes on to demonstrate a basic profiling session on an application that pulls information from a database and displays it. It shows how to switch which snapshots you are comparing, explains the different sections of the Summary view and what they are showing, and proceeds to show you how to investigate memory problems using the “Instance Categorizer” to track the path from an object (or set of objects) to the GC’s root in order to find what things along the path are holding a reference to it/them. For a set of objects, you can then click on it and get the “Instance List” view. This displays all of the individual objects (including their individual sizes, values, etc.) of that type which share the same path to the GC root. You can then click on one of the objects to generate an “Instance Retention Graph” view. This lets you track directly up to see the reference chain for that individual object. In the overview video, it turned out that there was an event handler which was holding on to a reference, thereby keeping a large number of strings that should have been freed in memory. Lastly the video shows the “Class List” view, which lets you dig in deeply to find problems that might not have been clear when following the previous workflow. Once you have at least one memory snapshot you can begin analyzing. The main interface is in the “Analysis” tab. You can also switch to the “Session Overview” tab, which gives you several bar charts highlighting basic memory data about the snapshots you’ve taken. If you hover over the individual bars (and the individual colors in bars that have more than one), you will see a detailed text description of what the bar is representing visually. The Session Overview is good for a quick summary of memory usage and information about the different heaps. You are going to spend most of your time in the Analysis tab, but it’s good to remember that the Session Overview is there to give you some quick feedback on basic memory usage stats. As described above in the summary of the overview video, there is a certain natural workflow to the Analysis tab. You’ll spin up your application and take some snapshots at various times such as before and after clicking a button to open a window or before and after closing a window. Taking these snapshots lets you examine what is happening with memory. You would normally expect that a lot of memory would be freed up when closing a window or exiting a document. By taking snapshots before and after performing an action like that you can see whether or not the memory is really being freed. If you already know an area that’s giving you trouble, you can run your application just like normal until just before getting to that part and then you can take a few strategic snapshots that should help you pin down the problem. Something the overview didn’t go into is how to use the “Filters” section at the bottom of ANTS MP together with the Class List view in order to narrow things down. The video tutorials page has a nice 3 minute intro video called “How to use the filters”. It’s a nice introduction and covers some of the basics. I’m going to cover a bit more because I think they’re a really neat, really helpful feature. Large programs can bring up thousands of classes. Even simple programs can instantiate far more classes than you might realize. In a basic .NET 4 WPF application for example (and when I say basic, I mean just MainWindow.xaml with a button added to it), the unfiltered Class List view will have in excess of 1000 classes (my simple test app had anywhere from 1066 to 1148 classes depending on which snapshot I was using as the “Current” snapshot). This is amazing in some ways as it shows you how in stark detail just how immensely powerful the WPF framework is. But hunting through 1100 classes isn’t productive, no matter how cool it is that there are that many classes instantiated and doing all sorts of awesome things. Let’s say you wanted to examine just the classes your application contains source code for (in my simple example, that would be the MainWindow and App). Under “Basic Filters”, click on “Classes with source” under “Show only…”. Voilà. Down from 1070 classes in the snapshot I was using as “Current” to 2 classes. If you then click on a class’s name, it will show you (to the right of the class name) two little icon buttons. Hover over them and you will see that you can click one to view the Instance Categorizer for the class and another to view the Instance List for the class. You can also show classes based on which heap they live on. If you chose both a Baseline snapshot and a Current snapshot then you can use the “Comparing snapshots” filters to show only: “New objects”; “Surviving objects”; “Survivors in growing classes”; or “Zombie objects” (if you aren’t sure what one of these means, you can click the helpful “?” in a green circle icon to bring up a popup that explains them and provides context). Remember that your selection(s) under the “Show only…” heading will still apply, so you should update those selections to make sure you are seeing the view you want. There are also links under the “What is my memory problem?” heading that can help you diagnose the problems you are seeing including one for “I don’t know which kind I have” for situations where you know generally that your application has some problems but aren’t sure what the behavior you have been seeing (OutOfMemoryExceptions, continually growing memory usage, larger memory use than expected at certain points in the program). The Basic Filters are not the only filters there are. “Filter by Object Type” gives you the ability to filter by: “Objects that are disposable”; “Objects that are/are not disposed”; “Objects that are/are not GC roots” (GC roots are things like static variables); and “Objects that implement _______”. “Objects that implement” is particularly neat. Once you check the box, you can then add one or more classes and interfaces that an object must implement in order to survive the filtering. Lastly there is “Filter by Reference”, which gives you the option to pare down the list based on whether an object is “Kept in memory exclusively by” a particular item, a class/interface, or a namespace; whether an object is “Referenced by” one or more of those choices; and whether an object is “Never referenced by” one or more of those choices. Remember that filtering is cumulative, so anything you had set in one of the filter sections still remains in effect unless and until you go back and change it. There’s quite a bit more to ANTS MP – it’s a very full featured product – but I think I touched on all of the most significant pieces. You can use it to debug: a .NET executable; an ASP.NET web application (running on IIS); an ASP.NET web application (running on Visual Studio’s built-in web development server); a Silverlight 4 browser application; a Windows service; a COM+ server; and even something called an XBAP (local XAML browser application). You can also attach to a .NET 4 process to profile an application that’s already running. The startup screen also has a large number of “Charting Options” that let you adjust which statistics ANTS MP should collect. The default selection is a good, minimal set. It’s worth your time to browse through the charting options to examine other statistics that may also help you diagnose a particular problem. The more statistics ANTS MP collects, the longer it will take to collect statistics. So just turning everything on is probably a bad idea. But the option to selectively add in additional performance counters from the extensive list could be a very helpful thing for your memory profiling as it lets you see additional data that might provide clues about a particular problem that has been bothering you. ANTS MP integrates very nicely with all versions of Visual Studio that support plugins (i.e. all of the non-Express versions). Just note that if you choose “Profile Memory” from the “ANTS” menu that it will launch profiling for whichever project you have set as the Startup project. One quick tip from my experience so far using ANTS MP: if you want to properly understand your memory usage in an application you’ve written, first create an “empty” version of the type of project you are going to profile (a WPF application, an XNA game, etc.) and do a quick profiling session on that so that you know the baseline memory usage of the framework itself. By “empty” I mean just create a new project of that type in Visual Studio then compile it and run it with profiling – don’t do anything special or add in anything (except perhaps for any external libraries you’re planning to use). The first thing I tried ANTS MP out on was a demo XNA project of an editor that I’ve been working on for quite some time that involves a custom extension to XNA’s content pipeline. The first time I ran it and saw the unmanaged memory usage I was convinced I had some horrible bug that was creating extra copies of texture data (the demo project didn’t have a lot of texture data so when I saw a lot of unmanaged memory I instantly figured I was doing something wrong). Then I thought to run an empty project through and when I saw that the amount of unmanaged memory was virtually identical, it dawned on me that the CLR itself sits in unmanaged memory and that (thankfully) there was nothing wrong with my code! Quite a relief. Earlier, when discussing the overview video, I mentioned the API that lets you take snapshots from within your application. I gave it a quick trial and it’s very easy to integrate and make use of and is a really nice addition (especially for projects where you want to know what, if any, allocations there are in a specific, complicated section of code). The only concern I had was that if I hadn’t watched the overview video I might never have known it existed. Even then it took me five minutes of hunting around Red Gate’s website before I found the “Taking snapshots from your code" article that explains what DLL you need to add as a reference and what method of what class you should call in order to take an automatic snapshot (including the helpful warning to wrap it in a try-catch block since, under certain circumstances, it can raise an exception, such as trying to call it more than 5 times in 30 seconds. The difficulty in discovering and then finding information about the automatic snapshots API was one thing I thought could use improvement. Another thing I think would make it even better would be local copies of the webpages it links to. Although I’m generally always connected to the internet, I imagine there are more than a few developers who aren’t or who are behind very restrictive firewalls. For them (and for me, too, if my internet connection happens to be down), it would be nice to have those documents installed locally or to have the option to download an additional “documentation” package that would add local copies. Another thing that I wish could be easier to manage is the Filters area. Finding and setting individual filters is very easy as is understanding what those filter do. And breaking it up into three sections (basic, by object, and by reference) makes sense. But I could easily see myself running a long profiling session and forgetting that I had set some filter a long while earlier in a different filter section and then spending quite a bit of time trying to figure out why some problem that was clearly visible in the data wasn’t showing up in, e.g. the instance list before remembering to check all the filters for that one setting that was only culling a few things from view. Some sort of indicator icon next to the filter section names that appears you have at least one filter set in that area would be a nice visual clue to remind me that “oh yeah, I told it to only show objects on the Gen 2 heap! That’s why I’m not seeing those instances of the SuperMagic class!” Something that would be nice (but that Red Gate cannot really do anything about) would be if this could be used in Windows Phone 7 development. If Microsoft and Red Gate could work together to make this happen (even if just on the WP7 emulator), that would be amazing. Especially given the memory constraints that apps and games running on mobile devices need to work within, a good memory profiler would be a phenomenally helpful tool. If anyone at Microsoft reads this, it’d be really great if you could make something like that happen. Perhaps even a (subsidized) custom version just for WP7 development. (For XNA games, of course, you can create a Windows version of the game and use ANTS MP on the Windows version in order to get a better picture of your memory situation. For Silverlight on WP7, though, there’s quite a bit of educated guess work and WeakReference creation followed by forced collections in order to find the source of a memory problem.) The only other thing I found myself wanting was a “Back” button. Between my Windows Phone 7, Zune, and other things, I’ve grown very used to having a “back stack” that lets me just navigate back to where I came from. The ANTS MP interface is surprisingly easy to use given how much it lets you do, and once you start using it for any amount of time, you learn all of the different areas such that you know where to go. And it does remember the state of the areas you were previously in, of course. So if you go to, e.g., the Instance Retention Graph from the Class List and then return back to the Class List, it will remember which class you had selected and all that other state information. Still, a “Back” button would be a welcome addition to a future release. Bottom Line ANTS Memory Profiler is not an inexpensive tool. But my time is valuable. I can easily see ANTS MP saving me enough time tracking down memory problems to justify it on a cost basis. More importantly to me, knowing what is happening memory-wise in my programs and having the confidence that my code doesn’t have any hidden time bombs in it that will cause it to OOM if I leave it running for longer than I do when I spin it up real quickly for debugging or just to see how a new feature looks and feels is a good feeling. It’s a feeling that I like having and want to continue to have. I got the current version for free in order to review it. Having done so, I’ve now added it to my must-have tools and will gladly lay out the money for the next version when it comes out. It has a 14 day free trial, so if you aren’t sure if it’s right for you or if you think it seems interesting but aren’t really sure if it’s worth shelling out the money for it, give it a try.

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  • Informed TDD &ndash; Kata &ldquo;To Roman Numerals&rdquo;

    - by Ralf Westphal
    Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/05/28/informed-tdd-ndash-kata-ldquoto-roman-numeralsrdquo.aspxIn a comment on my article on what I call Informed TDD (ITDD) reader gustav asked how this approach would apply to the kata “To Roman Numerals”. And whether ITDD wasn´t a violation of TDD´s principle of leaving out “advanced topics like mocks”. I like to respond with this article to his questions. There´s more to say than fits into a commentary. Mocks and TDD I don´t see in how far TDD is avoiding or opposed to mocks. TDD and mocks are orthogonal. TDD is about pocess, mocks are about structure and costs. Maybe by moving forward in tiny red+green+refactor steps less need arises for mocks. But then… if the functionality you need to implement requires “expensive” resource access you can´t avoid using mocks. Because you don´t want to constantly run all your tests against the real resource. True, in ITDD mocks seem to be in almost inflationary use. That´s not what you usually see in TDD demonstrations. However, there´s a reason for that as I tried to explain. I don´t use mocks as proxies for “expensive” resource. Rather they are stand-ins for functionality not yet implemented. They allow me to get a test green on a high level of abstraction. That way I can move forward in a top-down fashion. But if you think of mocks as “advanced” or if you don´t want to use a tool like JustMock, then you don´t need to use mocks. You just need to stand the sight of red tests for a little longer ;-) Let me show you what I mean by that by doing a kata. ITDD for “To Roman Numerals” gustav asked for the kata “To Roman Numerals”. I won´t explain the requirements again. You can find descriptions and TDD demonstrations all over the internet, like this one from Corey Haines. Now here is, how I would do this kata differently. 1. Analyse A demonstration of TDD should never skip the analysis phase. It should be made explicit. The requirements should be formalized and acceptance test cases should be compiled. “Formalization” in this case to me means describing the API of the required functionality. “[D]esign a program to work with Roman numerals” like written in this “requirement document” is not enough to start software development. Coding should only begin, if the interface between the “system under development” and its context is clear. If this interface is not readily recognizable from the requirements, it has to be developed first. Exploration of interface alternatives might be in order. It might be necessary to show several interface mock-ups to the customer – even if that´s you fellow developer. Designing the interface is a task of it´s own. It should not be mixed with implementing the required functionality behind the interface. Unfortunately, though, this happens quite often in TDD demonstrations. TDD is used to explore the API and implement it at the same time. To me that´s a violation of the Single Responsibility Principle (SRP) which not only should hold for software functional units but also for tasks or activities. In the case of this kata the API fortunately is obvious. Just one function is needed: string ToRoman(int arabic). And it lives in a class ArabicRomanConversions. Now what about acceptance test cases? There are hardly any stated in the kata descriptions. Roman numerals are explained, but no specific test cases from the point of view of a customer. So I just “invent” some acceptance test cases by picking roman numerals from a wikipedia article. They are supposed to be just “typical examples” without special meaning. Given the acceptance test cases I then try to develop an understanding of the problem domain. I´ll spare you that. The domain is trivial and is explain in almost all kata descriptions. How roman numerals are built is not difficult to understand. What´s more difficult, though, might be to find an efficient solution to convert into them automatically. 2. Solve The usual TDD demonstration skips a solution finding phase. Like the interface exploration it´s mixed in with the implementation. But I don´t think this is how it should be done. I even think this is not how it really works for the people demonstrating TDD. They´re simplifying their true software development process because they want to show a streamlined TDD process. I doubt this is helping anybody. Before you code you better have a plan what to code. This does not mean you have to do “Big Design Up-Front”. It just means: Have a clear picture of the logical solution in your head before you start to build a physical solution (code). Evidently such a solution can only be as good as your understanding of the problem. If that´s limited your solution will be limited, too. Fortunately, in the case of this kata your understanding does not need to be limited. Thus the logical solution does not need to be limited or preliminary or tentative. That does not mean you need to know every line of code in advance. It just means you know the rough structure of your implementation beforehand. Because it should mirror the process described by the logical or conceptual solution. Here´s my solution approach: The arabic “encoding” of numbers represents them as an ordered set of powers of 10. Each digit is a factor to multiply a power of ten with. The “encoding” 123 is the short form for a set like this: {1*10^2, 2*10^1, 3*10^0}. And the number is the sum of the set members. The roman “encoding” is different. There is no base (like 10 for arabic numbers), there are just digits of different value, and they have to be written in descending order. The “encoding” XVI is short for [10, 5, 1]. And the number is still the sum of the members of this list. The roman “encoding” thus is simpler than the arabic. Each “digit” can be taken at face value. No multiplication with a base required. But what about IV which looks like a contradiction to the above rule? It is not – if you accept roman “digits” not to be limited to be single characters only. Usually I, V, X, L, C, D, M are viewed as “digits”, and IV, IX etc. are viewed as nuisances preventing a simple solution. All looks different, though, once IV, IX etc. are taken as “digits”. Then MCMLIV is just a sum: M+CM+L+IV which is 1000+900+50+4. Whereas before it would have been understood as M-C+M+L-I+V – which is more difficult because here some “digits” get subtracted. Here´s the list of roman “digits” with their values: {1, I}, {4, IV}, {5, V}, {9, IX}, {10, X}, {40, XL}, {50, L}, {90, XC}, {100, C}, {400, CD}, {500, D}, {900, CM}, {1000, M} Since I take IV, IX etc. as “digits” translating an arabic number becomes trivial. I just need to find the values of the roman “digits” making up the number, e.g. 1954 is made up of 1000, 900, 50, and 4. I call those “digits” factors. If I move from the highest factor (M=1000) to the lowest (I=1) then translation is a two phase process: Find all the factors Translate the factors found Compile the roman representation Translation is just a look-up. Finding, though, needs some calculation: Find the highest remaining factor fitting in the value Remember and subtract it from the value Repeat with remaining value and remaining factors Please note: This is just an algorithm. It´s not code, even though it might be close. Being so close to code in my solution approach is due to the triviality of the problem. In more realistic examples the conceptual solution would be on a higher level of abstraction. With this solution in hand I finally can do what TDD advocates: find and prioritize test cases. As I can see from the small process description above, there are two aspects to test: Test the translation Test the compilation Test finding the factors Testing the translation primarily means to check if the map of factors and digits is comprehensive. That´s simple, even though it might be tedious. Testing the compilation is trivial. Testing factor finding, though, is a tad more complicated. I can think of several steps: First check, if an arabic number equal to a factor is processed correctly (e.g. 1000=M). Then check if an arabic number consisting of two consecutive factors (e.g. 1900=[M,CM]) is processed correctly. Then check, if a number consisting of the same factor twice is processed correctly (e.g. 2000=[M,M]). Finally check, if an arabic number consisting of non-consecutive factors (e.g. 1400=[M,CD]) is processed correctly. I feel I can start an implementation now. If something becomes more complicated than expected I can slow down and repeat this process. 3. Implement First I write a test for the acceptance test cases. It´s red because there´s no implementation even of the API. That´s in conformance with “TDD lore”, I´d say: Next I implement the API: The acceptance test now is formally correct, but still red of course. This will not change even now that I zoom in. Because my goal is not to most quickly satisfy these tests, but to implement my solution in a stepwise manner. That I do by “faking” it: I just “assume” three functions to represent the transformation process of my solution: My hypothesis is that those three functions in conjunction produce correct results on the API-level. I just have to implement them correctly. That´s what I´m trying now – one by one. I start with a simple “detail function”: Translate(). And I start with all the test cases in the obvious equivalence partition: As you can see I dare to test a private method. Yes. That´s a white box test. But as you´ll see it won´t make my tests brittle. It serves a purpose right here and now: it lets me focus on getting one aspect of my solution right. Here´s the implementation to satisfy the test: It´s as simple as possible. Right how TDD wants me to do it: KISS. Now for the second equivalence partition: translating multiple factors. (It´a pattern: if you need to do something repeatedly separate the tests for doing it once and doing it multiple times.) In this partition I just need a single test case, I guess. Stepping up from a single translation to multiple translations is no rocket science: Usually I would have implemented the final code right away. Splitting it in two steps is just for “educational purposes” here. How small your implementation steps are is a matter of your programming competency. Some “see” the final code right away before their mental eye – others need to work their way towards it. Having two tests I find more important. Now for the next low hanging fruit: compilation. It´s even simpler than translation. A single test is enough, I guess. And normally I would not even have bothered to write that one, because the implementation is so simple. I don´t need to test .NET framework functionality. But again: if it serves the educational purpose… Finally the most complicated part of the solution: finding the factors. There are several equivalence partitions. But still I decide to write just a single test, since the structure of the test data is the same for all partitions: Again, I´m faking the implementation first: I focus on just the first test case. No looping yet. Faking lets me stay on a high level of abstraction. I can write down the implementation of the solution without bothering myself with details of how to actually accomplish the feat. That´s left for a drill down with a test of the fake function: There are two main equivalence partitions, I guess: either the first factor is appropriate or some next. The implementation seems easy. Both test cases are green. (Of course this only works on the premise that there´s always a matching factor. Which is the case since the smallest factor is 1.) And the first of the equivalence partitions on the higher level also is satisfied: Great, I can move on. Now for more than a single factor: Interestingly not just one test becomes green now, but all of them. Great! You might say, then I must have done not the simplest thing possible. And I would reply: I don´t care. I did the most obvious thing. But I also find this loop very simple. Even simpler than a recursion of which I had thought briefly during the problem solving phase. And by the way: Also the acceptance tests went green: Mission accomplished. At least functionality wise. Now I´ve to tidy up things a bit. TDD calls for refactoring. Not uch refactoring is needed, because I wrote the code in top-down fashion. I faked it until I made it. I endured red tests on higher levels while lower levels weren´t perfected yet. But this way I saved myself from refactoring tediousness. At the end, though, some refactoring is required. But maybe in a different way than you would expect. That´s why I rather call it “cleanup”. First I remove duplication. There are two places where factors are defined: in Translate() and in Find_factors(). So I factor the map out into a class constant. Which leads to a small conversion in Find_factors(): And now for the big cleanup: I remove all tests of private methods. They are scaffolding tests to me. They only have temporary value. They are brittle. Only acceptance tests need to remain. However, I carry over the single “digit” tests from Translate() to the acceptance test. I find them valuable to keep, since the other acceptance tests only exercise a subset of all roman “digits”. This then is my final test class: And this is the final production code: Test coverage as reported by NCrunch is 100%: Reflexion Is this the smallest possible code base for this kata? Sure not. You´ll find more concise solutions on the internet. But LOC are of relatively little concern – as long as I can understand the code quickly. So called “elegant” code, however, often is not easy to understand. The same goes for KISS code – especially if left unrefactored, as it is often the case. That´s why I progressed from requirements to final code the way I did. I first understood and solved the problem on a conceptual level. Then I implemented it top down according to my design. I also could have implemented it bottom-up, since I knew some bottom of the solution. That´s the leaves of the functional decomposition tree. Where things became fuzzy, since the design did not cover any more details as with Find_factors(), I repeated the process in the small, so to speak: fake some top level, endure red high level tests, while first solving a simpler problem. Using scaffolding tests (to be thrown away at the end) brought two advantages: Encapsulation of the implementation details was not compromised. Naturally private methods could stay private. I did not need to make them internal or public just to be able to test them. I was able to write focused tests for small aspects of the solution. No need to test everything through the solution root, the API. The bottom line thus for me is: Informed TDD produces cleaner code in a systematic way. It conforms to core principles of programming: Single Responsibility Principle and/or Separation of Concerns. Distinct roles in development – being a researcher, being an engineer, being a craftsman – are represented as different phases. First find what, what there is. Then devise a solution. Then code the solution, manifest the solution in code. Writing tests first is a good practice. But it should not be taken dogmatic. And above all it should not be overloaded with purposes. And finally: moving from top to bottom through a design produces refactored code right away. Clean code thus almost is inevitable – and not left to a refactoring step at the end which is skipped often for different reasons.   PS: Yes, I have done this kata several times. But that has only an impact on the time needed for phases 1 and 2. I won´t skip them because of that. And there are no shortcuts during implementation because of that.

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  • Responsive Design for your ADF Faces Web Applications

    - by Shay Shmeltzer
    Responsive web applications are a common pattern for designing web pages that adjust their UI based on the device that access them. With the increase in the number of ADF applications that are being accessed from mobile phones and tablet we are getting more and more questions around this topic. Steven Davelaar wrote a comprehensive article covering key concepts in this area that you can find here. The article focuses on what I would refer to as server adaptive application, where the server adapts the UI it generates based on the device that is accessing the server. However there is one more technique that is not covered in that article and can be used with Oracle ADF - it is CSS manipulation on the client that can achieve responsive design. I'll cover this technique in this blog entry. The main advantage of this technique is that the UI manipulation does not require the server to send over a new UI when a change is needed. This for example allows your page to change immediately when you change the orientation of your device. (By the way this example was developed for one of the seminars in the upcoming Oracle ADF OTN Virtual Developer Day). In the demo that you'll see below you'll see a single page that changes the way it is displayed based on the orientation of the device. Here is the page with the tablet in landscape and portrait: To achieve this I'm using a CSS media query in my page template that changes the display property of a couple of style classes that are used in my page. The media query has this format: @media screen and (max-width:700px) {            .narrow {                display: inline;            }            .wide {                display: none;            }            .adjustFont {                font-size: small;            }            .icon-home {                font-size: 24px;            }        } This changes the properties of the same styleClasses that are defined in my application's skin. Here is a quick demo video that shows you the full application and explains how it works. For those looking to replicate this, here are the basic files: skin1.css @charset "UTF-8";/**ADFFaces_Skin_File / DO NOT REMOVE**/@namespace af "http://xmlns.oracle.com/adf/faces/rich";@namespace dvt "http://xmlns.oracle.com/dss/adf/faces";.wide {    display: inline;}.narrow {    display: none;}.adjustFont {    font-size: large;}.icon-home {        font-family: 'UIShellUGH';    -webkit-font-smoothing: antialiased;        font-size: 36px;        color: #ffa000;} pageTemplate: <?xml version='1.0' encoding='UTF-8'?><af:pageTemplateDef xmlns:af="http://xmlns.oracle.com/adf/faces/rich" var="attrs" definition="private"                    xmlns:afc="http://xmlns.oracle.com/adf/faces/rich/component">    <af:xmlContent>        <afc:component>            <afc:description>A template that will work on phones and desktop</afc:description>            <afc:display-name>ResponsiveTemplate</afc:display-name>            <afc:facet>                <afc:facet-name>main</afc:facet-name>            </afc:facet>        </afc:component>    </af:xmlContent>    <meta name="viewport" content="width=device-width, initial-scale=1"/>    <af:resource type="css">@media screen and (max-width:700px) {            .narrow {                display: inline;            }            .wide {                display: none;            }            .adjustFont {                font-size: small;            }            .icon-home {                font-size: 24px;            }        }@font-face {            font-family: 'UIShellUGH';            src: url(data:application/x-font-woff;charset=utf-8;base64,d09GRk9UVE8AA..removed code here...AzV6b1g==)format('truetype');            font-weight: normal;            font-style: normal;        }    </af:resource>    <af:panelGroupLayout id="pt_pgl4" layout="vertical" styleClass="sizeStyle">        <af:panelGridLayout id="pt_pgl1">            <af:gridRow marginTop="5px" height="40px" id="pt_gr1">                <af:gridCell marginStart="5px" width="100%" marginEnd="5px" id="pt_gc1">                    <af:panelGroupLayout id="pt_pgl3" halign="center" layout="horizontal">                        <af:outputText value="h" id="ot2" styleClass="icon-home"/>                        <af:outputText value="HR System" id="ot3" styleClass="adjustFont"/>                    </af:panelGroupLayout>                </af:gridCell>            </af:gridRow>            <af:gridRow marginTop="5px" height="auto" id="pt_gr2">                <af:gridCell marginStart="5px" width="100%" marginEnd="5px" id="pt_gc2" halign="stretch">                    <af:panelGroupLayout id="pt_pgl2" layout="scroll">                        <af:facetRef facetName="main"/>                    </af:panelGroupLayout>                </af:gridCell>            </af:gridRow>            <af:gridRow marginTop="5px" height="20px" marginBottom="5px" id="pt_gr3">                <af:gridCell marginStart="5px" width="100%" marginEnd="5px" id="pt_gc3">                    <af:panelGroupLayout id="pt_pgl5" layout="vertical" halign="center">                        <af:separator id="pt_s1"/>                        <af:outputText value="Copyright Oracle Corp. 2013" id="pt_ot1" styleClass="adjustFont"/>                    </af:panelGroupLayout>                </af:gridCell>            </af:gridRow>        </af:panelGridLayout>    </af:panelGroupLayout></af:pageTemplateDef> Example from the page:                         <af:gridRow id="gr3">                            <af:gridCell id="gc7" columnSpan="2">                                <af:panelGroupLayout id="pgl8" styleClass="narrow">                                    <af:link text="Menu" id="l1">                                        <af:showPopupBehavior triggerType="action" popupId="p1" align="afterEnd"/>                                    </af:link>                                </af:panelGroupLayout>                                <af:panelGroupLayout id="pgl7" styleClass="wide">                                    <af:navigationPane id="np1" hint="buttons">                                        <af:commandNavigationItem text="Departments" id="cni1"/>                                        <af:commandNavigationItem text="Employees" id="cni2"/>                                        <af:commandNavigationItem text="Salaries" id="cni3"/>                                        <af:commandNavigationItem text="Jobs" id="cni4"/>                                        <af:commandNavigationItem text="Services" id="cni5"/>                                        <af:commandNavigationItem text="Support" id="cni6"/>                                        <af:commandNavigationItem text="Help" id="cni7"/>                                    </af:navigationPane>                                </af:panelGroupLayout>                            </af:gridCell>                        </af:gridRow>

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  • How to display a JSON error message?

    - by Tiny Giant Studios
    I'm currently developing a tumblr theme and have built a jQuery JSON thingamabob that uses the Tumblr API to do the following: The user would click on the "post type" link (e.g. Video Posts), at which stage jQuery would use JSON to grab all the posts that's related to that type and then dynamically display them in a designated area. Now everything works absolutely peachy, except that with Tumblr being Tumblr and their servers taking a knock every now and then, the Tumblr API thingy is sometimes offline. Now I can't foresee when this function will be down, which is why I want to display some generic error message if JSON (for whatever reason) was unable to load the post. You'll see I've already written some code to show an error message when jQuery can't find any posts related to that post type BUT it doesn't cover any server errors. Note: I sometimes get this error: Failed to load resource: the server responded with a status of 503 (Service Temporarily Unavailable) It is for this 503 Error message that I need to write some code, but I'm slightly clueless :) Here's the jQuery JSON code: $('ul.right li').find('a').click(function() { var postType = this.className; var count = 0; byCategory(postType); return false; function byCategory(postType, callback) { $.getJSON('{URL}/api/read/json?type=' + postType + '&callback=?', function(data) { var article = []; $.each(data.posts, function(i, item) { // i = index // item = data for a particular post switch(item.type) { case 'photo': article[i] = '<div class="post_wrap"><div class="photo" style="padding-bottom:5px;">' + '<a href="' + item.url + '" title="{Title}" class="type_icon"><img src="http://static.tumblr.com/ewjv7ap/XSTldh6ds/photo_icon.png" alt="type_icon"/></a>' + '<a href="' + item.url + '" title="{Title}"><img src="' + item['photo-url-500'] + '"alt="image" /></a></div></div>'; count = 1; break; case 'video': article[i] = '<div class="post_wrap"><div class="video" style="padding-bottom:5px;">' + '<a href="' + item.url + '" title="{Title}" class="type_icon">' + '<img src="http://static.tumblr.com/ewjv7ap/nuSldhclv/video_icon.png" alt="type_icon"/></a>' + '<span style="margin: auto;">' + item['video-player'] + '</span>' + '</div></div>'; count = 1; break; case 'audio': if (use_IE == true) { article[i] = '<div class="post_wrap"><div class="regular">' + '<a href="' + item.url + '" title="{Title}" class="type_icon"><img src="http://static.tumblr.com/ewjv7ap/R50ldh5uj/audio_icon.png" alt="type_icon"/></a>' + '<h3><a href="' + item.url + '">' + item['id3-artist'] +' - ' + item['id3-title'] + '</a></h3>' + '</div></div>'; } else { article[i] = '<div class="post_wrap"><div class="regular">' + '<a href="' + item.url + '" title="{Title}" class="type_icon"><img src="http://static.tumblr.com/ewjv7ap/R50ldh5uj/audio_icon.png" alt="type_icon"/></a>' + '<h3><a href="' + item.url + '">' + item['id3-artist'] +' - ' + item['id3-title'] + '</a></h3><div class="player">' + item['audio-player'] + '</div>' + '</div></div>'; }; count = 1; break; case 'regular': article[i] = '<div class="post_wrap"><div class="regular">' + '<a href="' + item.url + '" title="{Title}" class="type_icon"><img src="http://static.tumblr.com/ewjv7ap/dwxldhck1/regular_icon.png" alt="type_icon"/></a><h3><a href="' + item.url + '">' + item['regular-title'] + '</a></h3><div class="description_container">' + item['regular-body'] + '</div></div></div>'; count = 1; break; case 'quote': article[i] = '<div class="post_wrap"><div class="quote">' + '<a href="' + item.url + '" title="{Title}" class="type_icon"><img src="http://static.tumblr.com/ewjv7ap/loEldhcpr/quote_icon.png" alt="type_icon"/></a><blockquote><h3><a href="' + item.url + '" title="{Title}">' + item['quote-text'] + '</a></h3></blockquote><cite>- ' + item['quote-source'] + '</cite></div></div>'; count = 1; break; case 'conversation': article[i] = '<div class="post_wrap"><div class="chat">' + '<a href="' + item.url + '" title="{Title}" class="type_icon"><img src="http://static.tumblr.com/ewjv7ap/MVuldhcth/conversation_icon.png" alt="type_icon"/></a><h3><a href="' + item.url + '">' + item['conversation-title'] + '</a></h3></div></div>'; count = 1; break; case 'link': article[i] = '<div class="post_wrap"><div class="link">' + '<a href="' + item.url + '" title="{Title}" class="type_icon"><img src="http://static.tumblr.com/ewjv7ap/EQGldhc30/link_icon.png" alt="type_icon"/></a><h3><a href="' + item['link-url'] + '" target="_blank">' + item['link-text'] + '</a></h3></div></div>'; count = 1; break; default: alert('No Entries Found.'); }; }) // end each if (!(count == 0)) { $('#content_right') .hide('fast') .html('<div class="first_div"><span class="left_corner"></span><span class="right_corner"></span><h2>Displaying ' + postType + ' Posts Only</h2></div>' + article.join('')) .slideDown('fast') } else { $('#content_right') .hide('fast') .html('<div class="first_div"><span class="left_corner"></span><span class="right_corner"></span><h2>Hmmm, currently there are no ' + postType + ' posts to display</h2></div>') .slideDown('fast') } // end getJSON }); // end byCategory } }); If you'd like to see the demo in action, check out Elegantem but do note that everything might work absolutely fine for you (or not), depending on Tumblr's temperament.

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  • How to find an entry-level job after you already have a graduate degree?

    - by Uri
    Note: I asked this question in early 2009. A couple of months later, I found a great job. I've previously updated this question with some tips for whoever ends up in a similar situation, and now cleaned it up a little for the benefit of the fresh batch of graduates. Original post: In my early 20s I abandoned a great C++ development career path in a major company to go to graduate school and get a research masters (3 years). I did another year in industrial research, and then moved to the US to attend graduate school again, getting another masters and a Ph.D in software engineering from a top school (another 6 years down the drain). I was coding the whole way throughout my degrees (core Java and Eclipse plug-ins) and working on research related to software engineering (usability of APIs). I ended up graduating the year of the recession, with a son on the way and the prospects of no healthcare. Academic jobs and industrial research jobs are quite scarce. Initially, I was naive, thinking that with my background, I could easily find a coding job. Big mistake. It turns out that I'm in a complicated position. Entry level positions are usually offered to college undergraduates. I attended my school's career fairs, but you could immediately see signs of Ph.D. aversion and overqualification issues. Some of the recruiters I spoke with explicitly told me that they wanted 20 year olds with clean slates, and some were looking for interns since they are in various forms of hiring freezes. I managed to get a couple of interviews from these career fairs and through recruiters. However, since I've been out of school for a long time and programming primarily in Java, I am also no longer proficient in C/C++ and the usual range of college-level interview questions that everyone uses. I had no problems with this when I was 19 and interviewing for my first job since a lot of what you do in C is manipulate pointers and I was coding C++ for fun and for school. Later I was routinely doing pointer manipulation on the job, and during my first masters taught college courses with data structures and C++. But even though I remember many properties of C++ well, it's been close to ten years since I regularly used C++ and pointers. As a Java developer I rarely had to work at this level, but experience in OOD and in writing good maintainable code is meaningless for C++ interviews. Reading books as a refresh and looking at sample code did not do the trick. I also looked at mid-to-senior level Java positions, but most of them focused on J2EE APIs rather than on core Java and required a certain number of years in industrial positions. Coding research tools and prior C++ experience doesn't count. So that sends me back to entry-level jobs that are posted through job-boards, and these are not common (mostly they are Monster junk), and small companies are even less likely to answer a Ph.D. compared to the giants who participate in top-10 career fairs. Even worse, in many companies initial screening is done by HR folks who really don't want to deal with anything anomalous like a Ph.D. Any tips on how I should approach this intractable position? For example, what should I write in cover letters? Note that while immigration is not an issue for me, I cannot go freelance as I need the benefits (and in particular group health insurance). During my studies I had no time to contribute to open-source projects or maintain a popular blog, so even if I invested in that now there would be no immediate benefit. Updates: In the two months after posting this I received several offers to work as a core Java developer in the financial industry and accepted one from a firm where I am working to this day. For those who find themselves in similar situations, here are my tips: Give up on trying to find an entry level positions. You can't undo time. Accept the fact that there is Ph.D. discrimination in the job market (some might say rightfully so). It is legal to discriminate based on education. No point fighting it. The most important tip is to focus on the language you are comfortable with. The sad truth about programming in a particular language is that it is not like riding a bike. If you haven't used a language in the last few years, and can't actually apply it routinely (not just as a refresher) before you start your search, it is going to be very difficult to do well in an interview. Now that I'm interviewing others, I routinely see it in folks with a mixed C++/Java background. We maintain "a shadow" of the old language but end up with a weird mix that makes it hard to interview on either. Entry-level folks are at an advantage here since they usually have one language. Memory can help you do great in a screening interview, but without recent day-to-day experience, code tests will be difficult. Despite the supposed relation, core Java programming and J2EE programming are two different things with different skillsets. If you come from academia, you likely have very little J2EE experience and may find it hard to get accepted for a J2EE job. J2EE jobs seem to have a larger list of acronyms in their requirements. In addition, from interviewing J2EE developers it seems that for many there is a focus on mastering specific APIs and architectures, whereas core Java development tends to be secondary. In the same way that I can no longer manipulate pointers well, a J2EE developer may have difficulties doing low level Java manipulation. This puts you at a relative advantage in competing for core Java jobs! If you are able to work for startups (in terms of family life and stability) or migrate to startup-rich areas such as the west coast, you can find many exciting opportunities where advanced degrees are a benefit. I've since been approached by several startups, although I had to decline. Work through a recruiter if possible. They have direct contacts with the hiring parties, allowing you to "stand out". It is better to get a clear yes/no confirmation from a recruiter on whether a company might be interested in interviewing you, than it is to send your resume and hope that someone will ever see it. Recruiters are also a great way of bypassing HR. However, also beware of recruiters. They have a vested interest and will go to various shady practices and pressure tactics. To find a good recruiter, talk to a friend who declined a job offer he got through a recruiter. A good recruiter, to me, is measured in how they handle that. Interview for the jobs that require your core strength. If you're rusty or entirely unfamiliar with a technology around which the job revolves, you're probably not a good match. Yes, you probably have the talent to master them, but most companies would want "instant gratification". I got my offers from companies that wanted core Java developer. I didn't do well on places that wanted advance C++ because I am too rusty and not up to date on recent libraries. I also didn't hear from companies that wanted lots of J2EE experience, and that's ok. Finding companies that want core Java without web is harder, but exists in specific industries (e.g., finance, defense). This requires a lot more legwork in terms of search, but these jobs do exist. There are different interview styles. Some companies focus on puzzles, some companies focus on algorithms, and some companies focus on design and coding skills. I had the most success in places where the questions were the most related to the function I would have been performing. Pick companies accordingly as well.

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  • The Incremental Architect&rsquo;s Napkin - #5 - Design functions for extensibility and readability

    - by Ralf Westphal
    Originally posted on: http://geekswithblogs.net/theArchitectsNapkin/archive/2014/08/24/the-incremental-architectrsquos-napkin---5---design-functions-for.aspx The functionality of programs is entered via Entry Points. So what we´re talking about when designing software is a bunch of functions handling the requests represented by and flowing in through those Entry Points. Designing software thus consists of at least three phases: Analyzing the requirements to find the Entry Points and their signatures Designing the functionality to be executed when those Entry Points get triggered Implementing the functionality according to the design aka coding I presume, you´re familiar with phase 1 in some way. And I guess you´re proficient in implementing functionality in some programming language. But in my experience developers in general are not experienced in going through an explicit phase 2. “Designing functionality? What´s that supposed to mean?” you might already have thought. Here´s my definition: To design functionality (or functional design for short) means thinking about… well, functions. You find a solution for what´s supposed to happen when an Entry Point gets triggered in terms of functions. A conceptual solution that is, because those functions only exist in your head (or on paper) during this phase. But you may have guess that, because it´s “design” not “coding”. And here is, what functional design is not: It´s not about logic. Logic is expressions (e.g. +, -, && etc.) and control statements (e.g. if, switch, for, while etc.). Also I consider calling external APIs as logic. It´s equally basic. It´s what code needs to do in order to deliver some functionality or quality. Logic is what´s doing that needs to be done by software. Transformations are either done through expressions or API-calls. And then there is alternative control flow depending on the result of some expression. Basically it´s just jumps in Assembler, sometimes to go forward (if, switch), sometimes to go backward (for, while, do). But calling your own function is not logic. It´s not necessary to produce any outcome. Functionality is not enhanced by adding functions (subroutine calls) to your code. Nor is quality increased by adding functions. No performance gain, no higher scalability etc. through functions. Functions are not relevant to functionality. Strange, isn´t it. What they are important for is security of investment. By introducing functions into our code we can become more productive (re-use) and can increase evolvability (higher unterstandability, easier to keep code consistent). That´s no small feat, however. Evolvable code can hardly be overestimated. That´s why to me functional design is so important. It´s at the core of software development. To sum this up: Functional design is on a level of abstraction above (!) logical design or algorithmic design. Functional design is only done until you get to a point where each function is so simple you are very confident you can easily code it. Functional design an logical design (which mostly is coding, but can also be done using pseudo code or flow charts) are complementary. Software needs both. If you start coding right away you end up in a tangled mess very quickly. Then you need back out through refactoring. Functional design on the other hand is bloodless without actual code. It´s just a theory with no experiments to prove it. But how to do functional design? An example of functional design Let´s assume a program to de-duplicate strings. The user enters a number of strings separated by commas, e.g. a, b, a, c, d, b, e, c, a. And the program is supposed to clear this list of all doubles, e.g. a, b, c, d, e. There is only one Entry Point to this program: the user triggers the de-duplication by starting the program with the string list on the command line C:\>deduplicate "a, b, a, c, d, b, e, c, a" a, b, c, d, e …or by clicking on a GUI button. This leads to the Entry Point function to get called. It´s the program´s main function in case of the batch version or a button click event handler in the GUI version. That´s the physical Entry Point so to speak. It´s inevitable. What then happens is a three step process: Transform the input data from the user into a request. Call the request handler. Transform the output of the request handler into a tangible result for the user. Or to phrase it a bit more generally: Accept input. Transform input into output. Present output. This does not mean any of these steps requires a lot of effort. Maybe it´s just one line of code to accomplish it. Nevertheless it´s a distinct step in doing the processing behind an Entry Point. Call it an aspect or a responsibility - and you will realize it most likely deserves a function of its own to satisfy the Single Responsibility Principle (SRP). Interestingly the above list of steps is already functional design. There is no logic, but nevertheless the solution is described - albeit on a higher level of abstraction than you might have done yourself. But it´s still on a meta-level. The application to the domain at hand is easy, though: Accept string list from command line De-duplicate Present de-duplicated strings on standard output And this concrete list of processing steps can easily be transformed into code:static void Main(string[] args) { var input = Accept_string_list(args); var output = Deduplicate(input); Present_deduplicated_string_list(output); } Instead of a big problem there are three much smaller problems now. If you think each of those is trivial to implement, then go for it. You can stop the functional design at this point. But maybe, just maybe, you´re not so sure how to go about with the de-duplication for example. Then just implement what´s easy right now, e.g.private static string Accept_string_list(string[] args) { return args[0]; } private static void Present_deduplicated_string_list( string[] output) { var line = string.Join(", ", output); Console.WriteLine(line); } Accept_string_list() contains logic in the form of an API-call. Present_deduplicated_string_list() contains logic in the form of an expression and an API-call. And then repeat the functional design for the remaining processing step. What´s left is the domain logic: de-duplicating a list of strings. How should that be done? Without any logic at our disposal during functional design you´re left with just functions. So which functions could make up the de-duplication? Here´s a suggestion: De-duplicate Parse the input string into a true list of strings. Register each string in a dictionary/map/set. That way duplicates get cast away. Transform the data structure into a list of unique strings. Processing step 2 obviously was the core of the solution. That´s where real creativity was needed. That´s the core of the domain. But now after this refinement the implementation of each step is easy again:private static string[] Parse_string_list(string input) { return input.Split(',') .Select(s => s.Trim()) .ToArray(); } private static Dictionary<string,object> Compile_unique_strings(string[] strings) { return strings.Aggregate( new Dictionary<string, object>(), (agg, s) => { agg[s] = null; return agg; }); } private static string[] Serialize_unique_strings( Dictionary<string,object> dict) { return dict.Keys.ToArray(); } With these three additional functions Main() now looks like this:static void Main(string[] args) { var input = Accept_string_list(args); var strings = Parse_string_list(input); var dict = Compile_unique_strings(strings); var output = Serialize_unique_strings(dict); Present_deduplicated_string_list(output); } I think that´s very understandable code: just read it from top to bottom and you know how the solution to the problem works. It´s a mirror image of the initial design: Accept string list from command line Parse the input string into a true list of strings. Register each string in a dictionary/map/set. That way duplicates get cast away. Transform the data structure into a list of unique strings. Present de-duplicated strings on standard output You can even re-generate the design by just looking at the code. Code and functional design thus are always in sync - if you follow some simple rules. But about that later. And as a bonus: all the functions making up the process are small - which means easy to understand, too. So much for an initial concrete example. Now it´s time for some theory. Because there is method to this madness ;-) The above has only scratched the surface. Introducing Flow Design Functional design starts with a given function, the Entry Point. Its goal is to describe the behavior of the program when the Entry Point is triggered using a process, not an algorithm. An algorithm consists of logic, a process on the other hand consists just of steps or stages. Each processing step transforms input into output or a side effect. Also it might access resources, e.g. a printer, a database, or just memory. Processing steps thus can rely on state of some sort. This is different from Functional Programming, where functions are supposed to not be stateful and not cause side effects.[1] In its simplest form a process can be written as a bullet point list of steps, e.g. Get data from user Output result to user Transform data Parse data Map result for output Such a compilation of steps - possibly on different levels of abstraction - often is the first artifact of functional design. It can be generated by a team in an initial design brainstorming. Next comes ordering the steps. What should happen first, what next etc.? Get data from user Parse data Transform data Map result for output Output result to user That´s great for a start into functional design. It´s better than starting to code right away on a given function using TDD. Please get me right: TDD is a valuable practice. But it can be unnecessarily hard if the scope of a functionn is too large. But how do you know beforehand without investing some thinking? And how to do this thinking in a systematic fashion? My recommendation: For any given function you´re supposed to implement first do a functional design. Then, once you´re confident you know the processing steps - which are pretty small - refine and code them using TDD. You´ll see that´s much, much easier - and leads to cleaner code right away. For more information on this approach I call “Informed TDD” read my book of the same title. Thinking before coding is smart. And writing down the solution as a bunch of functions possibly is the simplest thing you can do, I´d say. It´s more according to the KISS (Keep It Simple, Stupid) principle than returning constants or other trivial stuff TDD development often is started with. So far so good. A simple ordered list of processing steps will do to start with functional design. As shown in the above example such steps can easily be translated into functions. Moving from design to coding thus is simple. However, such a list does not scale. Processing is not always that simple to be captured in a list. And then the list is just text. Again. Like code. That means the design is lacking visuality. Textual representations need more parsing by your brain than visual representations. Plus they are limited in their “dimensionality”: text just has one dimension, it´s sequential. Alternatives and parallelism are hard to encode in text. In addition the functional design using numbered lists lacks data. It´s not visible what´s the input, output, and state of the processing steps. That´s why functional design should be done using a lightweight visual notation. No tool is necessary to draw such designs. Use pen and paper; a flipchart, a whiteboard, or even a napkin is sufficient. Visualizing processes The building block of the functional design notation is a functional unit. I mostly draw it like this: Something is done, it´s clear what goes in, it´s clear what comes out, and it´s clear what the processing step requires in terms of state or hardware. Whenever input flows into a functional unit it gets processed and output is produced and/or a side effect occurs. Flowing data is the driver of something happening. That´s why I call this approach to functional design Flow Design. It´s about data flow instead of control flow. Control flow like in algorithms is of no concern to functional design. Thinking about control flow simply is too low level. Once you start with control flow you easily get bogged down by tons of details. That´s what you want to avoid during design. Design is supposed to be quick, broad brush, abstract. It should give overview. But what about all the details? As Robert C. Martin rightly said: “Programming is abot detail”. Detail is a matter of code. Once you start coding the processing steps you designed you can worry about all the detail you want. Functional design does not eliminate all the nitty gritty. It just postpones tackling them. To me that´s also an example of the SRP. Function design has the responsibility to come up with a solution to a problem posed by a single function (Entry Point). And later coding has the responsibility to implement the solution down to the last detail (i.e. statement, API-call). TDD unfortunately mixes both responsibilities. It´s just coding - and thereby trying to find detailed implementations (green phase) plus getting the design right (refactoring). To me that´s one reason why TDD has failed to deliver on its promise for many developers. Using functional units as building blocks of functional design processes can be depicted very easily. Here´s the initial process for the example problem: For each processing step draw a functional unit and label it. Choose a verb or an “action phrase” as a label, not a noun. Functional design is about activities, not state or structure. Then make the output of an upstream step the input of a downstream step. Finally think about the data that should flow between the functional units. Write the data above the arrows connecting the functional units in the direction of the data flow. Enclose the data description in brackets. That way you can clearly see if all flows have already been specified. Empty brackets mean “no data is flowing”, but nevertheless a signal is sent. A name like “list” or “strings” in brackets describes the data content. Use lower case labels for that purpose. A name starting with an upper case letter like “String” or “Customer” on the other hand signifies a data type. If you like, you also can combine descriptions with data types by separating them with a colon, e.g. (list:string) or (strings:string[]). But these are just suggestions from my practice with Flow Design. You can do it differently, if you like. Just be sure to be consistent. Flows wired-up in this manner I call one-dimensional (1D). Each functional unit just has one input and/or one output. A functional unit without an output is possible. It´s like a black hole sucking up input without producing any output. Instead it produces side effects. A functional unit without an input, though, does make much sense. When should it start to work? What´s the trigger? That´s why in the above process even the first processing step has an input. If you like, view such 1D-flows as pipelines. Data is flowing through them from left to right. But as you can see, it´s not always the same data. It get´s transformed along its passage: (args) becomes a (list) which is turned into (strings). The Principle of Mutual Oblivion A very characteristic trait of flows put together from function units is: no functional units knows another one. They are all completely independent of each other. Functional units don´t know where their input is coming from (or even when it´s gonna arrive). They just specify a range of values they can process. And they promise a certain behavior upon input arriving. Also they don´t know where their output is going. They just produce it in their own time independent of other functional units. That means at least conceptually all functional units work in parallel. Functional units don´t know their “deployment context”. They now nothing about the overall flow they are place in. They are just consuming input from some upstream, and producing output for some downstream. That makes functional units very easy to test. At least as long as they don´t depend on state or resources. I call this the Principle of Mutual Oblivion (PoMO). Functional units are oblivious of others as well as an overall context/purpose. They are just parts of a whole focused on a single responsibility. How the whole is built, how a larger goal is achieved, is of no concern to the single functional units. By building software in such a manner, functional design interestingly follows nature. Nature´s building blocks for organisms also follow the PoMO. The cells forming your body do not know each other. Take a nerve cell “controlling” a muscle cell for example:[2] The nerve cell does not know anything about muscle cells, let alone the specific muscel cell it is “attached to”. Likewise the muscle cell does not know anything about nerve cells, let a lone a specific nerve cell “attached to” it. Saying “the nerve cell is controlling the muscle cell” thus only makes sense when viewing both from the outside. “Control” is a concept of the whole, not of its parts. Control is created by wiring-up parts in a certain way. Both cells are mutually oblivious. Both just follow a contract. One produces Acetylcholine (ACh) as output, the other consumes ACh as input. Where the ACh is going, where it´s coming from neither cell cares about. Million years of evolution have led to this kind of division of labor. And million years of evolution have produced organism designs (DNA) which lead to the production of these different cell types (and many others) and also to their co-location. The result: the overall behavior of an organism. How and why this happened in nature is a mystery. For our software, though, it´s clear: functional and quality requirements needs to be fulfilled. So we as developers have to become “intelligent designers” of “software cells” which we put together to form a “software organism” which responds in satisfying ways to triggers from it´s environment. My bet is: If nature gets complex organisms working by following the PoMO, who are we to not apply this recipe for success to our much simpler “machines”? So my rule is: Wherever there is functionality to be delivered, because there is a clear Entry Point into software, design the functionality like nature would do it. Build it from mutually oblivious functional units. That´s what Flow Design is about. In that way it´s even universal, I´d say. Its notation can also be applied to biology: Never mind labeling the functional units with nouns. That´s ok in Flow Design. You´ll do that occassionally for functional units on a higher level of abstraction or when their purpose is close to hardware. Getting a cockroach to roam your bedroom takes 1,000,000 nerve cells (neurons). Getting the de-duplication program to do its job just takes 5 “software cells” (functional units). Both, though, follow the same basic principle. Translating functional units into code Moving from functional design to code is no rocket science. In fact it´s straightforward. There are two simple rules: Translate an input port to a function. Translate an output port either to a return statement in that function or to a function pointer visible to that function. The simplest translation of a functional unit is a function. That´s what you saw in the above example. Functions are mutually oblivious. That why Functional Programming likes them so much. It makes them composable. Which is the reason, nature works according to the PoMO. Let´s be clear about one thing: There is no dependency injection in nature. For all of an organism´s complexity no DI container is used. Behavior is the result of smooth cooperation between mutually oblivious building blocks. Functions will often be the adequate translation for the functional units in your designs. But not always. Take for example the case, where a processing step should not always produce an output. Maybe the purpose is to filter input. Here the functional unit consumes words and produces words. But it does not pass along every word flowing in. Some words are swallowed. Think of a spell checker. It probably should not check acronyms for correctness. There are too many of them. Or words with no more than two letters. Such words are called “stop words”. In the above picture the optionality of the output is signified by the astrisk outside the brackets. It means: Any number of (word) data items can flow from the functional unit for each input data item. It might be none or one or even more. This I call a stream of data. Such behavior cannot be translated into a function where output is generated with return. Because a function always needs to return a value. So the output port is translated into a function pointer or continuation which gets passed to the subroutine when called:[3]void filter_stop_words( string word, Action<string> onNoStopWord) { if (...check if not a stop word...) onNoStopWord(word); } If you want to be nitpicky you might call such a function pointer parameter an injection. And technically you´re right. Conceptually, though, it´s not an injection. Because the subroutine is not functionally dependent on the continuation. Firstly continuations are procedures, i.e. subroutines without a return type. Remember: Flow Design is about unidirectional data flow. Secondly the name of the formal parameter is chosen in a way as to not assume anything about downstream processing steps. onNoStopWord describes a situation (or event) within the functional unit only. Translating output ports into function pointers helps keeping functional units mutually oblivious in cases where output is optional or produced asynchronically. Either pass the function pointer to the function upon call. Or make it global by putting it on the encompassing class. Then it´s called an event. In C# that´s even an explicit feature.class Filter { public void filter_stop_words( string word) { if (...check if not a stop word...) onNoStopWord(word); } public event Action<string> onNoStopWord; } When to use a continuation and when to use an event dependens on how a functional unit is used in flows and how it´s packed together with others into classes. You´ll see examples further down the Flow Design road. Another example of 1D functional design Let´s see Flow Design once more in action using the visual notation. How about the famous word wrap kata? Robert C. Martin has posted a much cited solution including an extensive reasoning behind his TDD approach. So maybe you want to compare it to Flow Design. The function signature given is:string WordWrap(string text, int maxLineLength) {...} That´s not an Entry Point since we don´t see an application with an environment and users. Nevertheless it´s a function which is supposed to provide a certain functionality. The text passed in has to be reformatted. The input is a single line of arbitrary length consisting of words separated by spaces. The output should consist of one or more lines of a maximum length specified. If a word is longer than a the maximum line length it can be split in multiple parts each fitting in a line. Flow Design Let´s start by brainstorming the process to accomplish the feat of reformatting the text. What´s needed? Words need to be assembled into lines Words need to be extracted from the input text The resulting lines need to be assembled into the output text Words too long to fit in a line need to be split Does sound about right? I guess so. And it shows a kind of priority. Long words are a special case. So maybe there is a hint for an incremental design here. First let´s tackle “average words” (words not longer than a line). Here´s the Flow Design for this increment: The the first three bullet points turned into functional units with explicit data added. As the signature requires a text is transformed into another text. See the input of the first functional unit and the output of the last functional unit. In between no text flows, but words and lines. That´s good to see because thereby the domain is clearly represented in the design. The requirements are talking about words and lines and here they are. But note the asterisk! It´s not outside the brackets but inside. That means it´s not a stream of words or lines, but lists or sequences. For each text a sequence of words is output. For each sequence of words a sequence of lines is produced. The asterisk is used to abstract from the concrete implementation. Like with streams. Whether the list of words gets implemented as an array or an IEnumerable is not important during design. It´s an implementation detail. Does any processing step require further refinement? I don´t think so. They all look pretty “atomic” to me. And if not… I can always backtrack and refine a process step using functional design later once I´ve gained more insight into a sub-problem. Implementation The implementation is straightforward as you can imagine. The processing steps can all be translated into functions. Each can be tested easily and separately. Each has a focused responsibility. And the process flow becomes just a sequence of function calls: Easy to understand. It clearly states how word wrapping works - on a high level of abstraction. And it´s easy to evolve as you´ll see. Flow Design - Increment 2 So far only texts consisting of “average words” are wrapped correctly. Words not fitting in a line will result in lines too long. Wrapping long words is a feature of the requested functionality. Whether it´s there or not makes a difference to the user. To quickly get feedback I decided to first implement a solution without this feature. But now it´s time to add it to deliver the full scope. Fortunately Flow Design automatically leads to code following the Open Closed Principle (OCP). It´s easy to extend it - instead of changing well tested code. How´s that possible? Flow Design allows for extension of functionality by inserting functional units into the flow. That way existing functional units need not be changed. The data flow arrow between functional units is a natural extension point. No need to resort to the Strategy Pattern. No need to think ahead where extions might need to be made in the future. I just “phase in” the remaining processing step: Since neither Extract words nor Reformat know of their environment neither needs to be touched due to the “detour”. The new processing step accepts the output of the existing upstream step and produces data compatible with the existing downstream step. Implementation - Increment 2 A trivial implementation checking the assumption if this works does not do anything to split long words. The input is just passed on: Note how clean WordWrap() stays. The solution is easy to understand. A developer looking at this code sometime in the future, when a new feature needs to be build in, quickly sees how long words are dealt with. Compare this to Robert C. Martin´s solution:[4] How does this solution handle long words? Long words are not even part of the domain language present in the code. At least I need considerable time to understand the approach. Admittedly the Flow Design solution with the full implementation of long word splitting is longer than Robert C. Martin´s. At least it seems. Because his solution does not cover all the “word wrap situations” the Flow Design solution handles. Some lines would need to be added to be on par, I guess. But even then… Is a difference in LOC that important as long as it´s in the same ball park? I value understandability and openness for extension higher than saving on the last line of code. Simplicity is not just less code, it´s also clarity in design. But don´t take my word for it. Try Flow Design on larger problems and compare for yourself. What´s the easier, more straightforward way to clean code? And keep in mind: You ain´t seen all yet ;-) There´s more to Flow Design than described in this chapter. In closing I hope I was able to give you a impression of functional design that makes you hungry for more. To me it´s an inevitable step in software development. Jumping from requirements to code does not scale. And it leads to dirty code all to quickly. Some thought should be invested first. Where there is a clear Entry Point visible, it´s functionality should be designed using data flows. Because with data flows abstraction is possible. For more background on why that´s necessary read my blog article here. For now let me point out to you - if you haven´t already noticed - that Flow Design is a general purpose declarative language. It´s “programming by intention” (Shalloway et al.). Just write down how you think the solution should work on a high level of abstraction. This breaks down a large problem in smaller problems. And by following the PoMO the solutions to those smaller problems are independent of each other. So they are easy to test. Or you could even think about getting them implemented in parallel by different team members. Flow Design not only increases evolvability, but also helps becoming more productive. All team members can participate in functional design. This goes beyon collective code ownership. We´re talking collective design/architecture ownership. Because with Flow Design there is a common visual language to talk about functional design - which is the foundation for all other design activities.   PS: If you like what you read, consider getting my ebook “The Incremental Architekt´s Napkin”. It´s where I compile all the articles in this series for easier reading. I like the strictness of Function Programming - but I also find it quite hard to live by. And it certainly is not what millions of programmers are used to. Also to me it seems, the real world is full of state and side effects. So why give them such a bad image? That´s why functional design takes a more pragmatic approach. State and side effects are ok for processing steps - but be sure to follow the SRP. Don´t put too much of it into a single processing step. ? Image taken from www.physioweb.org ? My code samples are written in C#. C# sports typed function pointers called delegates. Action is such a function pointer type matching functions with signature void someName(T t). Other languages provide similar ways to work with functions as first class citizens - even Java now in version 8. I trust you find a way to map this detail of my translation to your favorite programming language. I know it works for Java, C++, Ruby, JavaScript, Python, Go. And if you´re using a Functional Programming language it´s of course a no brainer. ? Taken from his blog post “The Craftsman 62, The Dark Path”. ?

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  • Configuring Fed Authentication Methods in OIF / IdP

    - by Damien Carru
    In this article, I will provide examples on how to configure OIF/IdP to map OAM Authentication Schemes to Federation Authentication Methods, based on the concepts introduced in my previous entry. I will show examples for the three protocols supported by OIF: SAML 2.0 SSO SAML 1.1 SSO OpenID 2.0 Enjoy the reading! Configuration As I mentioned in my previous article, mapping Federation Authentication Methods to OAM Authentication Schemes is protocol dependent, since the methods are defined in the various protocols (SAML 2.0, SAML 1.1, OpenID 2.0). As such, the WLST commands to set those mappings will involve: Either the SP Partner Profile and affect all Partners referencing that profile, which do not override the Federation Authentication Method to OAM Authentication Scheme mappings Or the SP Partner entry, which will only affect the SP Partner It is important to note that if an SP Partner is configured to define one or more Federation Authentication Method to OAM Authentication Scheme mappings, then all the mappings defined in the SP Partner Profile will be ignored. WLST Commands The two OIF WLST commands that can be used to define mapping Federation Authentication Methods to OAM Authentication Schemes are: addSPPartnerProfileAuthnMethod() to define a mapping on an SP Partner Profile, taking as parameters: The name of the SP Partner Profile The Federation Authentication Method The OAM Authentication Scheme name addSPPartnerAuthnMethod() to define a mapping on an SP Partner , taking as parameters: The name of the SP Partner The Federation Authentication Method The OAM Authentication Scheme name Note: I will discuss in a subsequent article the other parameters of those commands. In the next sections, I will show examples on how to use those methods: For SAML 2.0, I will configure the SP Partner Profile, that will apply all the mappings to SP Partners referencing this profile, unless they override mapping definition For SAML 1.1, I will configure the SP Partner. For OpenID 2.0, I will configure the SP/RP Partner SAML 2.0 Test Setup In this setup, OIF is acting as an IdP and is integrated with a remote SAML 2.0 SP partner identified by AcmeSP. In this test, I will perform Federation SSO with OIF/IdP configured to: Use LDAPScheme as the Authentication Scheme Use BasicScheme as the Authentication Scheme Map BasicSessionScheme  to  the urn:oasis:names:tc:SAML:2.0:ac:classes:Password Federation Authentication Method Use OAMLDAPPluginAuthnScheme as the Authentication Scheme Map OAMLDAPPluginAuthnScheme to  the urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport Federation Authentication Method LDAPScheme as Authentication Scheme Using the OOTB settings regarding user authentication in OAM, the user will be challenged via a FORM based login page based on the LDAPScheme. Also the default Federation Authentication Method mappings configuration maps only the urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport to LDAPScheme (also marked as the default scheme used for authentication), FAAuthScheme, BasicScheme and BasicFAScheme. After authentication via FORM, OIF/IdP would issue an Assertion similar to: <samlp:Response ...>    <saml:Issuer ...>https://idp.com/oam/fed</saml:Issuer>    <samlp:Status>        <samlp:StatusCode Value="urn:oasis:names:tc:SAML:2.0:status:Success"/>    </samlp:Status>    <saml:Assertion ...>        <saml:Issuer ...>https://idp.com/oam/fed</saml:Issuer>        <dsig:Signature>            ...        </dsig:Signature>        <saml:Subject>            <saml:NameID ...>[email protected]</saml:NameID>            <saml:SubjectConfirmation Method="urn:oasis:names:tc:SAML:2.0:cm:bearer">                <saml:SubjectConfirmationData .../>            </saml:SubjectConfirmation>        </saml:Subject>        <saml:Conditions ...>            <saml:AudienceRestriction>                <saml:Audience>https://acme.com/sp</saml:Audience>            </saml:AudienceRestriction>        </saml:Conditions>        <saml:AuthnStatement AuthnInstant="2014-03-21T20:53:55Z" SessionIndex="id-6i-Dm0yB-HekG6cejktwcKIFMzYE8Yrmqwfd0azz" SessionNotOnOrAfter="2014-03-21T21:53:55Z">            <saml:AuthnContext>                <saml:AuthnContextClassRef>                   urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport                </saml:AuthnContextClassRef>            </saml:AuthnContext>        </saml:AuthnStatement>    </saml:Assertion></samlp:Response> BasicScheme as Authentication Scheme For this test, I will switch the default Authentication Scheme for the SP Partner Profile to BasicScheme instead of LDAPScheme. I will use the OIF WLST setSPPartnerProfileDefaultScheme() command and specify which scheme to be used as the default for the SP Partner Profile referenced by AcmeSP (which is saml20-sp-partner-profile in this case: getFedPartnerProfile("AcmeSP", "sp") ): Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the setSPPartnerProfileDefaultScheme() command:setSPPartnerProfileDefaultScheme("saml20-sp-partner-profile", "BasicScheme") Exit the WLST environment:exit() The user will now be challenged via HTTP Basic Authentication defined in the BasicScheme for AcmeSP. Also, as noted earlier, the default Federation Authentication Method mappings configuration maps only the urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport to LDAPScheme (also marked as the default scheme used for authentication), FAAuthScheme, BasicScheme and BasicFAScheme. After authentication via HTTP Basic Authentication, OIF/IdP would issue an Assertion similar to: <samlp:Response ...>    <saml:Issuer ...>https://idp.com/oam/fed</saml:Issuer>    <samlp:Status>        <samlp:StatusCode Value="urn:oasis:names:tc:SAML:2.0:status:Success"/>    </samlp:Status>    <saml:Assertion ...>        <saml:Issuer ...>https://idp.com/oam/fed</saml:Issuer>        <dsig:Signature>            ...        </dsig:Signature>        <saml:Subject>            <saml:NameID ...>[email protected]</saml:NameID>            <saml:SubjectConfirmation Method="urn:oasis:names:tc:SAML:2.0:cm:bearer">                <saml:SubjectConfirmationData .../>            </saml:SubjectConfirmation>        </saml:Subject>        <saml:Conditions ...>            <saml:AudienceRestriction>                <saml:Audience>https://acme.com/sp</saml:Audience>            </saml:AudienceRestriction>        </saml:Conditions>        <saml:AuthnStatement AuthnInstant="2014-03-21T20:53:55Z" SessionIndex="id-6i-Dm0yB-HekG6cejktwcKIFMzYE8Yrmqwfd0azz" SessionNotOnOrAfter="2014-03-21T21:53:55Z">            <saml:AuthnContext>                <saml:AuthnContextClassRef>                   urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport                </saml:AuthnContextClassRef>            </saml:AuthnContext>        </saml:AuthnStatement>    </saml:Assertion></samlp:Response> Mapping BasicScheme To change the Federation Authentication Method mapping for the BasicScheme to urn:oasis:names:tc:SAML:2.0:ac:classes:Password instead of urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport for the saml20-sp-partner-profile SAML 2.0 SP Partner Profile (the profile to which my AcmeSP Partner is bound to), I will execute the addSPPartnerProfileAuthnMethod() method: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the addSPPartnerProfileAuthnMethod() command:addSPPartnerProfileAuthnMethod("saml20-sp-partner-profile", "urn:oasis:names:tc:SAML:2.0:ac:classes:Password", "BasicScheme") Exit the WLST environment:exit() After authentication via HTTP Basic Authentication, OIF/IdP would now issue an Assertion similar to (see that the AuthnContextClassRef was changed from PasswordProtectedTransport to Password): <samlp:Response ...>    <saml:Issuer ...>https://idp.com/oam/fed</saml:Issuer>    <samlp:Status>        <samlp:StatusCode Value="urn:oasis:names:tc:SAML:2.0:status:Success"/>    </samlp:Status>    <saml:Assertion ...>        <saml:Issuer ...>https://idp.com/oam/fed</saml:Issuer>        <dsig:Signature>            ...        </dsig:Signature>        <saml:Subject>            <saml:NameID ...>[email protected]</saml:NameID>            <saml:SubjectConfirmation Method="urn:oasis:names:tc:SAML:2.0:cm:bearer">                <saml:SubjectConfirmationData .../>            </saml:SubjectConfirmation>        </saml:Subject>        <saml:Conditions ...>            <saml:AudienceRestriction>                <saml:Audience>https://acme.com/sp</saml:Audience>            </saml:AudienceRestriction>        </saml:Conditions>        <saml:AuthnStatement AuthnInstant="2014-03-21T20:53:55Z" SessionIndex="id-6i-Dm0yB-HekG6cejktwcKIFMzYE8Yrmqwfd0azz" SessionNotOnOrAfter="2014-03-21T21:53:55Z">            <saml:AuthnContext>                <saml:AuthnContextClassRef>                   urn:oasis:names:tc:SAML:2.0:ac:classes:Password                </saml:AuthnContextClassRef>            </saml:AuthnContext>        </saml:AuthnStatement>    </saml:Assertion></samlp:Response> OAMLDAPPluginAuthnScheme as Authentication Scheme For this test, I will switch the default Authentication Scheme for the SP Partner Profile to OAMLDAPPluginAuthnScheme instead of BasicScheme. I will use the OIF WLST setSPPartnerProfileDefaultScheme() command and specify which scheme to be used as the default for the SP Partner Profile referenced by AcmeSP (which is saml20-sp-partner-profile in this case: getFedPartnerProfile("AcmeSP", "sp") ): Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the setSPPartnerProfileDefaultScheme() command:setSPPartnerProfileDefaultScheme("saml20-sp-partner-profile", "OAMLDAPPluginAuthnScheme") Exit the WLST environment:exit() The user will now be challenged via FORM defined in the OAMLDAPPluginAuthnScheme for AcmeSP. Contrarily to LDAPScheme and BasicScheme, the OAMLDAPPluginAuthnScheme is not mapped by default to any Federation Authentication Methods. As such, OIF/IdP will not be able to find a Federation Authentication Method and will set the method in the SAML Assertion to the OAM Authentication Scheme name. After authentication via FORM, OIF/IdP would issue an Assertion similar to (see the AuthnContextClassRef set to OAMLDAPPluginAuthnScheme): <samlp:Response ...>    <saml:Issuer ...>https://idp.com/oam/fed</saml:Issuer>    <samlp:Status>        <samlp:StatusCode Value="urn:oasis:names:tc:SAML:2.0:status:Success"/>    </samlp:Status>    <saml:Assertion ...>        <saml:Issuer ...>https://idp.com/oam/fed</saml:Issuer>        <dsig:Signature>            ...        </dsig:Signature>        <saml:Subject>            <saml:NameID ...>[email protected]</saml:NameID>            <saml:SubjectConfirmation Method="urn:oasis:names:tc:SAML:2.0:cm:bearer">                <saml:SubjectConfirmationData .../>            </saml:SubjectConfirmation>        </saml:Subject>        <saml:Conditions ...>            <saml:AudienceRestriction>                <saml:Audience>https://acme.com/sp</saml:Audience>            </saml:AudienceRestriction>        </saml:Conditions>        <saml:AuthnStatement AuthnInstant="2014-03-21T20:53:55Z" SessionIndex="id-6i-Dm0yB-HekG6cejktwcKIFMzYE8Yrmqwfd0azz" SessionNotOnOrAfter="2014-03-21T21:53:55Z">            <saml:AuthnContext>                <saml:AuthnContextClassRef> OAMLDAPPluginAuthnScheme                </saml:AuthnContextClassRef>            </saml:AuthnContext>        </saml:AuthnStatement>    </saml:Assertion></samlp:Response> Mapping OAMLDAPPluginAuthnScheme To add the OAMLDAPPluginAuthnScheme  to the Federation Authentication Method urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport mapping, I will execute the addSPPartnerProfileAuthnMethod() method: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the addSPPartnerProfileAuthnMethod() command:addSPPartnerProfileAuthnMethod("saml20-sp-partner-profile", "urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport", "OAMLDAPPluginAuthnScheme") Exit the WLST environment:exit() After authentication via FORM, OIF/IdP would now issue an Assertion similar to (see that the method was changed from OAMLDAPPluginAuthnScheme to PasswordProtectedTransport): <samlp:Response ...>    <saml:Issuer ...>https://idp.com/oam/fed</saml:Issuer>    <samlp:Status>        <samlp:StatusCode Value="urn:oasis:names:tc:SAML:2.0:status:Success"/>    </samlp:Status>    <saml:Assertion ...>        <saml:Issuer ...>https://idp.com/oam/fed</saml:Issuer>        <dsig:Signature>            ...        </dsig:Signature>        <saml:Subject>            <saml:NameID ...>[email protected]</saml:NameID>            <saml:SubjectConfirmation Method="urn:oasis:names:tc:SAML:2.0:cm:bearer">                <saml:SubjectConfirmationData .../>            </saml:SubjectConfirmation>        </saml:Subject>        <saml:Conditions ...>            <saml:AudienceRestriction>                <saml:Audience>https://acme.com/sp</saml:Audience>            </saml:AudienceRestriction>        </saml:Conditions>        <saml:AuthnStatement AuthnInstant="2014-03-21T20:53:55Z" SessionIndex="id-6i-Dm0yB-HekG6cejktwcKIFMzYE8Yrmqwfd0azz" SessionNotOnOrAfter="2014-03-21T21:53:55Z">            <saml:AuthnContext>                <saml:AuthnContextClassRef>                   urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport                </saml:AuthnContextClassRef>            </saml:AuthnContext>        </saml:AuthnStatement>    </saml:Assertion></samlp:Response> SAML 1.1 Test Setup In this setup, OIF is acting as an IdP and is integrated with a remote SAML 1.1 SP partner identified by AcmeSP. In this test, I will perform Federation SSO with OIF/IdP configured to: Use LDAPScheme as the Authentication Scheme Use OAMLDAPPluginAuthnScheme as the Authentication Scheme Map OAMLDAPPluginAuthnScheme to  the urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport Federation Authentication Method Use LDAPScheme as the Authentication Scheme Map LDAPScheme to  the urn:oasis:names:tc:SAML:2.0:ac:classes:PasswordProtectedTransport Federation Authentication Method LDAPScheme as Authentication Scheme Using the OOTB settings regarding user authentication in OAM, the user will be challenged via a FORM based login page based on the LDAPScheme. Also the default Federation Authentication Method mappings configuration maps only the urn:oasis:names:tc:SAML:1.0:am:password to LDAPScheme (also marked as the default scheme used for authentication), FAAuthScheme, BasicScheme and BasicFAScheme. After authentication via FORM, OIF/IdP would issue an Assertion similar to: <samlp:Response ...>    <samlp:Status>        <samlp:StatusCode Value="samlp:Success"/>    </samlp:Status>    <saml:Assertion Issuer="https://idp.com/oam/fed" ...>        <saml:Conditions ...>            <saml:AudienceRestriction>                <saml:Audience>https://acme.com/sp/ssov11</saml:Audience>            </saml:AudienceRestriction>        </saml:Conditions>        <saml:AuthnStatement AuthenticationInstant="2014-03-21T20:53:55Z" AuthenticationMethod="urn:oasis:names:tc:SAML:1.0:am:password">            <saml:Subject>                <saml:NameIdentifier ...>[email protected]</saml:NameIdentifier>                <saml:SubjectConfirmation>                   <saml:ConfirmationMethod>                       urn:oasis:names:tc:SAML:1.0:cm:bearer                   </saml:ConfirmationMethod>                </saml:SubjectConfirmation>            </saml:Subject>        </saml:AuthnStatement>        <dsig:Signature>            ...        </dsig:Signature>    </saml:Assertion></samlp:Response> OAMLDAPPluginAuthnScheme as Authentication Scheme For this test, I will switch the default Authentication Scheme for the SP Partner to OAMLDAPPluginAuthnScheme instead of LDAPScheme. I will use the OIF WLST setSPPartnerDefaultScheme() command and specify which scheme to be used as the default for the SP Partner: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the setSPPartnerDefaultScheme() command:setSPPartnerDefaultScheme("AcmeSP", "OAMLDAPPluginAuthnScheme") Exit the WLST environment:exit() The user will be challenged via FORM defined in the OAMLDAPPluginAuthnScheme for AcmeSP. Contrarily to LDAPScheme, the OAMLDAPPluginAuthnScheme is not mapped by default to any Federation Authentication Methods (in the SP Partner Profile). As such, OIF/IdP will not be able to find a Federation Authentication Method and will set the method in the SAML Assertion to the OAM Authentication Scheme name. After authentication via FORM, OIF/IdP would issue an Assertion similar to (see the AuthenticationMethod set to OAMLDAPPluginAuthnScheme): <samlp:Response ...>    <samlp:Status>        <samlp:StatusCode Value="samlp:Success"/>    </samlp:Status>    <saml:Assertion Issuer="https://idp.com/oam/fed" ...>        <saml:Conditions ...>            <saml:AudienceRestriction>                <saml:Audience>https://acme.com/sp/ssov11</saml:Audience>            </saml:AudienceRestriction>        </saml:Conditions>        <saml:AuthnStatement AuthenticationInstant="2014-03-21T20:53:55Z" AuthenticationMethod="OAMLDAPPluginAuthnScheme">            <saml:Subject>                <saml:NameIdentifier ...>[email protected]</saml:NameIdentifier>                <saml:SubjectConfirmation>                   <saml:ConfirmationMethod>                       urn:oasis:names:tc:SAML:1.0:cm:bearer                   </saml:ConfirmationMethod>                </saml:SubjectConfirmation>            </saml:Subject>        </saml:AuthnStatement>        <dsig:Signature>            ...        </dsig:Signature>    </saml:Assertion></samlp:Response> Mapping OAMLDAPPluginAuthnScheme To map the OAMLDAPPluginAuthnScheme  to the Federation Authentication Method urn:oasis:names:tc:SAML:1.0:am:password for this SP Partner only, I will execute the addSPPartnerAuthnMethod() method: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the addSPPartnerAuthnMethod() command:addSPPartnerAuthnMethod("AcmeSP", "urn:oasis:names:tc:SAML:1.0:am:password", "OAMLDAPPluginAuthnScheme") Exit the WLST environment:exit() After authentication via FORM, OIF/IdP would now issue an Assertion similar to (see that the method was changed from OAMLDAPPluginAuthnScheme to password): <samlp:Response ...>    <samlp:Status>        <samlp:StatusCode Value="samlp:Success"/>    </samlp:Status>    <saml:Assertion Issuer="https://idp.com/oam/fed" ...>        <saml:Conditions ...>            <saml:AudienceRestriction>                <saml:Audience>https://acme.com/sp/ssov11</saml:Audience>            </saml:AudienceRestriction>        </saml:Conditions>        <saml:AuthnStatement AuthenticationInstant="2014-03-21T20:53:55Z" AuthenticationMethod="urn:oasis:names:tc:SAML:1.0:am:password">            <saml:Subject>                <saml:NameIdentifier ...>[email protected]</saml:NameIdentifier>                <saml:SubjectConfirmation>                   <saml:ConfirmationMethod>                       urn:oasis:names:tc:SAML:1.0:cm:bearer                   </saml:ConfirmationMethod>                </saml:SubjectConfirmation>            </saml:Subject>        </saml:AuthnStatement>        <dsig:Signature>            ...        </dsig:Signature>    </saml:Assertion></samlp:Response> LDAPScheme as Authentication Scheme I will now show that by defining a Federation Authentication Mapping at the Partner level, this now ignores all mappings defined at the SP Partner Profile level. For this test, I will switch the default Authentication Scheme for this SP Partner back to LDAPScheme, and the Assertion issued by OIF/IdP will not be able to map this LDAPScheme to a Federation Authentication Method anymore, since A Federation Authentication Method mapping is defined at the SP Partner level and thus the mappings defined at the SP Partner Profile are ignored The LDAPScheme is not listed in the mapping at the Partner level I will use the OIF WLST setSPPartnerDefaultScheme() command and specify which scheme to be used as the default for this SP Partner: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the setSPPartnerDefaultScheme() command:setSPPartnerDefaultScheme("AcmeSP", "LDAPScheme") Exit the WLST environment:exit() After authentication via FORM, OIF/IdP would issue an Assertion similar to (see the AuthenticationMethod set to LDAPScheme): <samlp:Response ...>    <samlp:Status>        <samlp:StatusCode Value="samlp:Success"/>    </samlp:Status>    <saml:Assertion Issuer="https://idp.com/oam/fed" ...>        <saml:Conditions ...>            <saml:AudienceRestriction>                <saml:Audience>https://acme.com/sp/ssov11</saml:Audience>            </saml:AudienceRestriction>        </saml:Conditions>        <saml:AuthnStatement AuthenticationInstant="2014-03-21T20:53:55Z" AuthenticationMethod="LDAPScheme">            <saml:Subject>                <saml:NameIdentifier ...>[email protected]</saml:NameIdentifier>                <saml:SubjectConfirmation>                   <saml:ConfirmationMethod>                       urn:oasis:names:tc:SAML:1.0:cm:bearer                   </saml:ConfirmationMethod>                </saml:SubjectConfirmation>            </saml:Subject>        </saml:AuthnStatement>        <dsig:Signature>            ...        </dsig:Signature>    </saml:Assertion></samlp:Response> Mapping LDAPScheme at Partner Level To fix this issue, we will need to add the LDAPScheme  to the Federation Authentication Method urn:oasis:names:tc:SAML:1.0:am:password mapping for this SP Partner only. I will execute the addSPPartnerAuthnMethod() method: Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the addSPPartnerAuthnMethod() command:addSPPartnerAuthnMethod("AcmeSP", "urn:oasis:names:tc:SAML:1.0:am:password", "LDAPScheme") Exit the WLST environment:exit() After authentication via FORM, OIF/IdP would now issue an Assertion similar to (see that the method was changed from LDAPScheme to password): <samlp:Response ...>    <samlp:Status>        <samlp:StatusCode Value="samlp:Success"/>    </samlp:Status>    <saml:Assertion Issuer="https://idp.com/oam/fed" ...>        <saml:Conditions ...>            <saml:AudienceRestriction>                <saml:Audience>https://acme.com/sp/ssov11</saml:Audience>            </saml:AudienceRestriction>        </saml:Conditions>        <saml:AuthnStatement AuthenticationInstant="2014-03-21T20:53:55Z" AuthenticationMethod="urn:oasis:names:tc:SAML:1.0:am:password">            <saml:Subject>                <saml:NameIdentifier ...>[email protected]</saml:NameIdentifier>                <saml:SubjectConfirmation>                   <saml:ConfirmationMethod>                       urn:oasis:names:tc:SAML:1.0:cm:bearer                   </saml:ConfirmationMethod>                </saml:SubjectConfirmation>            </saml:Subject>        </saml:AuthnStatement>        <dsig:Signature>            ...        </dsig:Signature>    </saml:Assertion></samlp:Response> OpenID 2.0 In the OpenID 2.0 flows, the RP must request use of PAPE, in order for OIF/IdP/OP to include PAPE information. For OpenID 2.0, the configuration will involve mapping a list of OpenID 2.0 policies to a list of Authentication Schemes. The WLST command will take a list of policies, delimited by the ',' character, instead of SAML 2.0 or SAML 1.1 where a single Federation Authentication Method had to be specified. Test Setup In this setup, OIF is acting as an IdP/OP and is integrated with a remote OpenID 2.0 SP/RP partner identified by AcmeRP. In this test, I will perform Federation SSO with OIF/IdP configured to: Use LDAPScheme as the Authentication Scheme Map LDAPScheme to  the http://schemas.openid.net/pape/policies/2007/06/phishing-resistant and http://openid-policies/password-protected policies Federation Authentication Methods (the second one is a custom for this use case) LDAPScheme as Authentication Scheme Using the OOTB settings regarding user authentication in OAM, the user will be challenged via a FORM based login page based on the LDAPScheme. No Federation Authentication Method is defined OOTB for OpenID 2.0, so if the IdP/OP issue an SSO response with a PAPE Response element, it will specify the scheme name instead of Federation Authentication Methods After authentication via FORM, OIF/IdP would issue an SSO Response similar to: https://acme.com/openid?refid=id-9PKVXZmRxAeDYcgLqPm36ClzOMA-&openid.ns=http%3A%2F%2Fspecs.openid.net%2Fauth%2F2.0&openid.mode=id_res&openid.op_endpoint=https%3A%2F%2Fidp.com%2Fopenid&openid.claimed_id=https%3A%2F%2Fidp.com%2Fopenid%3Fid%3Did-38iCmmlAVEXPsFjnFVKArfn5RIiF75D5doorhEgqqPM%3D&openid.identity=https%3A%2F%2Fidp.com%2Fopenid%3Fid%3Did-38iCmmlAVEXPsFjnFVKArfn5RIiF75D5doorhEgqqPM%3D&openid.return_to=https%3A%2F%2Facme.com%2Fopenid%3Frefid%3Did-9PKVXZmRxAeDYcgLqPm36ClzOMA-&openid.response_nonce=2014-03-24T19%3A20%3A06Zid-YPa2kTNNFftZkgBb460jxJGblk2g--iNwPpDI7M1&openid.assoc_handle=id-6a5S6zhAKaRwQNUnjTKROREdAGSjWodG1el4xyz3&openid.ns.ax=http%3A%2F%2Fopenid.net%2Fsrv%2Fax%2F1.0&openid.ax.mode=fetch_response&openid.ax.type.attr0=http%3A%2F%2Fsession%2Fcount&openid.ax.value.attr0=1&openid.ax.type.attr1=http%3A%2F%2Fopenid.net%2Fschema%2FnamePerson%2Ffriendly&openid.ax.value.attr1=My+name+is+Bobby+Smith&openid.ax.type.attr2=http%3A%2F%2Fschemas.openid.net%2Fax%2Fapi%2Fuser_id&openid.ax.value.attr2=bob&openid.ax.type.attr3=http%3A%2F%2Faxschema.org%2Fcontact%2Femail&openid.ax.value.attr3=bob%40oracle.com&openid.ax.type.attr4=http%3A%2F%2Fsession%2Fipaddress&openid.ax.value.attr4=10.145.120.253&openid.ns.pape=http%3A%2F%2Fspecs.openid.net%2Fextensions%2Fpape%2F1.0&openid.pape.auth_time=2014-03-24T19%3A20%3A05Z&openid.pape.auth_policies=LDAPScheme&openid.signed=op_endpoint%2Cclaimed_id%2Cidentity%2Creturn_to%2Cresponse_nonce%2Cassoc_handle%2Cns.ax%2Cax.mode%2Cax.type.attr0%2Cax.value.attr0%2Cax.type.attr1%2Cax.value.attr1%2Cax.type.attr2%2Cax.value.attr2%2Cax.type.attr3%2Cax.value.attr3%2Cax.type.attr4%2Cax.value.attr4%2Cns.pape%2Cpape.auth_time%2Cpape.auth_policies&openid.sig=mYMgbGYSs22l8e%2FDom9NRPw15u8%3D Mapping LDAPScheme To map the LDAP Scheme to the http://schemas.openid.net/pape/policies/2007/06/phishing-resistant and http://openid-policies/password-protected policies Federation Authentication Methods, I will execute the addSPPartnerAuthnMethod() method (the policies will be comma separated): Enter the WLST environment by executing:$IAM_ORACLE_HOME/common/bin/wlst.sh Connect to the WLS Admin server:connect() Navigate to the Domain Runtime branch:domainRuntime() Execute the addSPPartnerAuthnMethod() command:addSPPartnerAuthnMethod("AcmeRP", "http://schemas.openid.net/pape/policies/2007/06/phishing-resistant,http://openid-policies/password-protected", "LDAPScheme") Exit the WLST environment:exit() After authentication via FORM, OIF/IdP would now issue an Assertion similar to (see that the method was changed from LDAPScheme to the two policies): https://acme.com/openid?refid=id-9PKVXZmRxAeDYcgLqPm36ClzOMA-&openid.ns=http%3A%2F%2Fspecs.openid.net%2Fauth%2F2.0&openid.mode=id_res&openid.op_endpoint=https%3A%2F%2Fidp.com%2Fopenid&openid.claimed_id=https%3A%2F%2Fidp.com%2Fopenid%3Fid%3Did-38iCmmlAVEXPsFjnFVKArfn5RIiF75D5doorhEgqqPM%3D&openid.identity=https%3A%2F%2Fidp.com%2Fopenid%3Fid%3Did-38iCmmlAVEXPsFjnFVKArfn5RIiF75D5doorhEgqqPM%3D&openid.return_to=https%3A%2F%2Facme.com%2Fopenid%3Frefid%3Did-9PKVXZmRxAeDYcgLqPm36ClzOMA-&openid.response_nonce=2014-03-24T19%3A20%3A06Zid-YPa2kTNNFftZkgBb460jxJGblk2g--iNwPpDI7M1&openid.assoc_handle=id-6a5S6zhAKaRwQNUnjTKROREdAGSjWodG1el4xyz3&openid.ns.ax=http%3A%2F%2Fopenid.net%2Fsrv%2Fax%2F1.0&openid.ax.mode=fetch_response&openid.ax.type.attr0=http%3A%2F%2Fsession%2Fcount&openid.ax.value.attr0=1&openid.ax.type.attr1=http%3A%2F%2Fopenid.net%2Fschema%2FnamePerson%2Ffriendly&openid.ax.value.attr1=My+name+is+Bobby+Smith&openid.ax.type.attr2=http%3A%2F%2Fschemas.openid.net%2Fax%2Fapi%2Fuser_id&openid.ax.value.attr2=bob&openid.ax.type.attr3=http%3A%2F%2Faxschema.org%2Fcontact%2Femail&openid.ax.value.attr3=bob%40oracle.com&openid.ax.type.attr4=http%3A%2F%2Fsession%2Fipaddress&openid.ax.value.attr4=10.145.120.253&openid.ns.pape=http%3A%2F%2Fspecs.openid.net%2Fextensions%2Fpape%2F1.0&openid.pape.auth_time=2014-03-24T19%3A20%3A05Z&openid.pape.auth_policies=http%3A%2F%2Fschemas.openid.net%2Fpape%2Fpolicies%2F2007%2F06%2Fphishing-resistant+http%3A%2F%2Fopenid-policies%2Fpassword-protected&openid.signed=op_endpoint%2Cclaimed_id%2Cidentity%2Creturn_to%2Cresponse_nonce%2Cassoc_handle%2Cns.ax%2Cax.mode%2Cax.type.attr0%2Cax.value.attr0%2Cax.type.attr1%2Cax.value.attr1%2Cax.type.attr2%2Cax.value.attr2%2Cax.type.attr3%2Cax.value.attr3%2Cax.type.attr4%2Cax.value.attr4%2Cns.pape%2Cpape.auth_time%2Cpape.auth_policies&openid.sig=mYMgbGYSs22l8e%2FDom9NRPw15u8%3D In the next article, I will cover how OIF/IdP can be configured so that an SP can request a specific Federation Authentication Method to challenge the user during Federation SSO.Cheers,Damien Carru

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  • Table sorting & pagination with jQuery and Razor in ASP.NET MVC

    - by hajan
    Introduction jQuery enjoys living inside pages which are built on top of ASP.NET MVC Framework. The ASP.NET MVC is a place where things are organized very well and it is quite hard to make them dirty, especially because the pattern enforces you on purity (you can still make it dirty if you want so ;) ). We all know how easy is to build a HTML table with a header row, footer row and table rows showing some data. With ASP.NET MVC we can do this pretty easy, but, the result will be pure HTML table which only shows data, but does not includes sorting, pagination or some other advanced features that we were used to have in the ASP.NET WebForms GridView. Ok, there is the WebGrid MVC Helper, but what if we want to make something from pure table in our own clean style? In one of my recent projects, I’ve been using the jQuery tablesorter and tablesorter.pager plugins that go along. You don’t need to know jQuery to make this work… You need to know little CSS to create nice design for your table, but of course you can use mine from the demo… So, what you will see in this blog is how to attach this plugin to your pure html table and a div for pagination and make your table with advanced sorting and pagination features.   Demo Project Resources The resources I’m using for this demo project are shown in the following solution explorer window print screen: Content/images – folder that contains all the up/down arrow images, pagination buttons etc. You can freely replace them with your own, but keep the names the same if you don’t want to change anything in the CSS we will built later. Content/Site.css – The main css theme, where we will add the theme for our table too Controllers/HomeController.cs – The controller I’m using for this project Models/Person.cs – For this demo, I’m using Person.cs class Scripts – jquery-1.4.4.min.js, jquery.tablesorter.js, jquery.tablesorter.pager.js – required script to make the magic happens Views/Home/Index.cshtml – Index view (razor view engine) the other items are not important for the demo. ASP.NET MVC 1. Model In this demo I use only one Person class which defines Person entity with several properties. You can use your own model, maybe one which will access data from database or any other resource. Person.cs public class Person {     public string Name { get; set; }     public string Surname { get; set; }     public string Email { get; set; }     public int? Phone { get; set; }     public DateTime? DateAdded { get; set; }     public int? Age { get; set; }     public Person(string name, string surname, string email,         int? phone, DateTime? dateadded, int? age)     {         Name = name;         Surname = surname;         Email = email;         Phone = phone;         DateAdded = dateadded;         Age = age;     } } 2. View In our example, we have only one Index.chtml page where Razor View engine is used. Razor view engine is my favorite for ASP.NET MVC because it’s very intuitive, fluid and keeps your code clean. 3. Controller Since this is simple example with one page, we use one HomeController.cs where we have two methods, one of ActionResult type (Index) and another GetPeople() used to create and return list of people. HomeController.cs public class HomeController : Controller {     //     // GET: /Home/     public ActionResult Index()     {         ViewBag.People = GetPeople();         return View();     }     public List<Person> GetPeople()     {         List<Person> listPeople = new List<Person>();                  listPeople.Add(new Person("Hajan", "Selmani", "[email protected]", 070070070,DateTime.Now, 25));                     listPeople.Add(new Person("Straight", "Dean", "[email protected]", 123456789, DateTime.Now.AddDays(-5), 35));         listPeople.Add(new Person("Karsen", "Livia", "[email protected]", 46874651, DateTime.Now.AddDays(-2), 31));         listPeople.Add(new Person("Ringer", "Anne", "[email protected]", null, DateTime.Now, null));         listPeople.Add(new Person("O'Leary", "Michael", "[email protected]", 32424344, DateTime.Now, 44));         listPeople.Add(new Person("Gringlesby", "Anne", "[email protected]", null, DateTime.Now.AddDays(-9), 18));         listPeople.Add(new Person("Locksley", "Stearns", "[email protected]", 2135345, DateTime.Now, null));         listPeople.Add(new Person("DeFrance", "Michel", "[email protected]", 235325352, DateTime.Now.AddDays(-18), null));         listPeople.Add(new Person("White", "Johnson", null, null, DateTime.Now.AddDays(-22), 55));         listPeople.Add(new Person("Panteley", "Sylvia", null, 23233223, DateTime.Now.AddDays(-1), 32));         listPeople.Add(new Person("Blotchet-Halls", "Reginald", null, 323243423, DateTime.Now, 26));         listPeople.Add(new Person("Merr", "South", "[email protected]", 3232442, DateTime.Now.AddDays(-5), 85));         listPeople.Add(new Person("MacFeather", "Stearns", "[email protected]", null, DateTime.Now, null));         return listPeople;     } }   TABLE CSS/HTML DESIGN Now, lets start with the implementation. First of all, lets create the table structure and the main CSS. 1. HTML Structure @{     Layout = null;     } <!DOCTYPE html> <html> <head>     <title>ASP.NET & jQuery</title>     <!-- referencing styles, scripts and writing custom js scripts will go here --> </head> <body>     <div>         <table class="tablesorter">             <thead>                 <tr>                     <th> value </th>                 </tr>             </thead>             <tbody>                 <tr>                     <td>value</td>                 </tr>             </tbody>             <tfoot>                 <tr>                     <th> value </th>                 </tr>             </tfoot>         </table>         <div id="pager">                      </div>     </div> </body> </html> So, this is the main structure you need to create for each of your tables where you want to apply the functionality we will create. Of course the scripts are referenced once ;). As you see, our table has class tablesorter and also we have a div with id pager. In the next steps we will use both these to create the needed functionalities. The complete Index.cshtml coded to get the data from controller and display in the page is: <body>     <div>         <table class="tablesorter">             <thead>                 <tr>                     <th>Name</th>                     <th>Surname</th>                     <th>Email</th>                     <th>Phone</th>                     <th>Date Added</th>                 </tr>             </thead>             <tbody>                 @{                     foreach (var p in ViewBag.People)                     {                                 <tr>                         <td>@p.Name</td>                         <td>@p.Surname</td>                         <td>@p.Email</td>                         <td>@p.Phone</td>                         <td>@p.DateAdded</td>                     </tr>                     }                 }             </tbody>             <tfoot>                 <tr>                     <th>Name</th>                     <th>Surname</th>                     <th>Email</th>                     <th>Phone</th>                     <th>Date Added</th>                 </tr>             </tfoot>         </table>         <div id="pager" style="position: none;">             <form>             <img src="@Url.Content("~/Content/images/first.png")" class="first" />             <img src="@Url.Content("~/Content/images/prev.png")" class="prev" />             <input type="text" class="pagedisplay" />             <img src="@Url.Content("~/Content/images/next.png")" class="next" />             <img src="@Url.Content("~/Content/images/last.png")" class="last" />             <select class="pagesize">                 <option selected="selected" value="5">5</option>                 <option value="10">10</option>                 <option value="20">20</option>                 <option value="30">30</option>                 <option value="40">40</option>             </select>             </form>         </div>     </div> </body> So, mainly the structure is the same. I have added @Razor code to create table with data retrieved from the ViewBag.People which has been filled with data in the home controller. 2. CSS Design The CSS code I’ve created is: /* DEMO TABLE */ body {     font-size: 75%;     font-family: Verdana, Tahoma, Arial, "Helvetica Neue", Helvetica, Sans-Serif;     color: #232323;     background-color: #fff; } table { border-spacing:0; border:1px solid gray;} table.tablesorter thead tr .header {     background-image: url(images/bg.png);     background-repeat: no-repeat;     background-position: center right;     cursor: pointer; } table.tablesorter tbody td {     color: #3D3D3D;     padding: 4px;     background-color: #FFF;     vertical-align: top; } table.tablesorter tbody tr.odd td {     background-color:#F0F0F6; } table.tablesorter thead tr .headerSortUp {     background-image: url(images/asc.png); } table.tablesorter thead tr .headerSortDown {     background-image: url(images/desc.png); } table th { width:150px;            border:1px outset gray;            background-color:#3C78B5;            color:White;            cursor:pointer; } table thead th:hover { background-color:Yellow; color:Black;} table td { width:150px; border:1px solid gray;} PAGINATION AND SORTING Now, when everything is ready and we have the data, lets make pagination and sorting functionalities 1. jQuery Scripts referencing <link href="@Url.Content("~/Content/Site.css")" rel="stylesheet" type="text/css" /> <script src="@Url.Content("~/Scripts/jquery-1.4.4.min.js")" type="text/javascript"></script> <script src="@Url.Content("~/Scripts/jquery.tablesorter.js")" type="text/javascript"></script> <script src="@Url.Content("~/Scripts/jquery.tablesorter.pager.js")" type="text/javascript"></script> 2. jQuery Sorting and Pagination script   <script type="text/javascript">     $(function () {         $("table.tablesorter").tablesorter({ widthFixed: true, sortList: [[0, 0]] })         .tablesorterPager({ container: $("#pager"), size: $(".pagesize option:selected").val() });     }); </script> So, with only two lines of code, I’m using both tablesorter and tablesorterPager plugins, giving some options to both these. Options added: tablesorter - widthFixed: true – gives fixed width of the columns tablesorter - sortList[[0,0]] – An array of instructions for per-column sorting and direction in the format: [[columnIndex, sortDirection], ... ] where columnIndex is a zero-based index for your columns left-to-right and sortDirection is 0 for Ascending and 1 for Descending. A valid argument that sorts ascending first by column 1 and then column 2 looks like: [[0,0],[1,0]] (source: http://tablesorter.com/docs/) tablesorterPager – container: $(“#pager”) – tells the pager container, the div with id pager in our case. tablesorterPager – size: the default size of each page, where I get the default value selected, so if you put selected to any other of the options in your select list, you will have this number of rows as default per page for the table too. END RESULTS 1. Table once the page is loaded (default results per page is 5 and is automatically sorted by 1st column as sortList is specified) 2. Sorted by Phone Descending 3. Changed pagination to 10 items per page 4. Sorted by Phone and Name (use SHIFT to sort on multiple columns) 5. Sorted by Date Added 6. Page 3, 5 items per page   ADDITIONAL ENHANCEMENTS We can do additional enhancements to the table. We can make search for each column. I will cover this in one of my next blogs. Stay tuned. DEMO PROJECT You can download demo project source code from HERE.CONCLUSION Once you finish with the demo, run your page and open the source code. You will be amazed of the purity of your code.Working with pagination in client side can be very useful. One of the benefits is performance, but if you have thousands of rows in your tables, you will get opposite result when talking about performance. Hence, sometimes it is nice idea to make pagination on back-end. So, the compromise between both approaches would be best to combine both of them. I use at most up to 500 rows on client-side and once the user reach the last page, we can trigger ajax postback which can get the next 500 rows using server-side pagination of the same data. I would like to recommend the following blog post http://weblogs.asp.net/gunnarpeipman/archive/2010/09/14/returning-paged-results-from-repositories-using-pagedresult-lt-t-gt.aspx, which will help you understand how to return page results from repository. I hope this was helpful post for you. Wait for my next posts ;). Please do let me know your feedback. Best Regards, Hajan

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  • Why should you choose Oracle WebLogic 12c instead of JBoss EAP 6?

    - by Ricardo Ferreira
    In this post, I will cover some technical differences between Oracle WebLogic 12c and JBoss EAP 6, which was released a couple days ago from Red Hat. This article claims to help you in the evaluation of key points that you should consider when choosing for an Java EE application server. In the following sections, I will present to you some important aspects that most customers ask us when they are seriously evaluating for an middleware infrastructure, specially if you are considering JBoss for some reason. I would suggest that you keep the following question in mind while you are reading the points: "Why should I choose JBoss instead of WebLogic?" 1) Multi Datacenter Deployment and Clustering - D/R ("Disaster & Recovery") architecture support is embedded on the WebLogic Server 12c product. JBoss EAP 6 on the other hand has no direct D/R support included, Red Hat relies on third-part tools with higher prices. When you consider a middleware solution to host your business critical application, you should worry with every architectural aspect that are related with the solution. Fail-over support is one little aspect of a truly reliable solution. If you do not worry about D/R, your solution will not be reliable. Having said that, with Red Hat and JBoss EAP 6, you have this extra cost that will increase considerably the total cost of ownership of the solution. As we commonly hear from analysts, open-source are not so cheaper when you start seeing the big picture. - WebLogic Server 12c supports advanced LAN clustering, detection of death servers and have a common alert framework. JBoss EAP 6 on the other hand has limited LAN clustering support with no server death detection. They do not generate any alerts when servers goes down (only if you buy JBoss ON which is a separated technology, but until now does not support JBoss EAP 6) and manual intervention are required when servers goes down. In most cases, admin people must rely on "kill -9", "tail -f someFile.log" and "ps ax | grep java" commands to manage failures and clustering anomalies. - WebLogic Server 12c supports the concept of Node Manager, which is a separated process that runs on the physical | virtual servers that allows extend the administration of the cluster to WebLogic managed servers that are often distributed across multiple machines and geographic locations. JBoss EAP 6 on the other hand has no equivalent technology. Whole server instances must be managed individually. - WebLogic Server 12c Node Manager supports Coherence to boost performance when managing servers. JBoss EAP 6 on the other hand has no similar technology. There is no way to coordinate JBoss and infiniband instances provided by JBoss using high throughput and low latency protocols like InfiniBand. The Node Manager feature also allows another very important feature that JBoss EAP lacks: secure the administration. When using WebLogic Node Manager, all the administration tasks are sent to the managed servers in a secure tunel protected by a certificate, which means that the transport layer that separates the WebLogic administration console from the managed servers are secured by SSL. - WebLogic Server 12c are now integrated with OTD ("Oracle Traffic Director") which is a web server technology derived from the former Sun iPlanet Web Server. This software complements the web server support offered by OHS ("Oracle HTTP Server"). Using OTD, WebLogic instances are load-balanced by a high powerful software that knows how to handle SDP ("Socket Direct Protocol") over InfiniBand, which boost performance when used with engineered systems technologies like Oracle Exalogic Elastic Cloud. JBoss EAP 6 on the other hand only offers support to Apache Web Server with custom modules created to deal with JBoss clusters, but only across standard TCP/IP networks.  2) Application and Runtime Diagnostics - WebLogic Server 12c have diagnostics capabilities embedded on the server called WLDF ("WebLogic Diagnostic Framework") so there is no need to rely on third-part tools. JBoss EAP 6 on the other hand has no diagnostics capabilities. Their only diagnostics tool is the log generated by the application server. Admin people are encouraged to analyse thousands of log lines to find out what is going on. - WebLogic Server 12c complement WLDF with JRockit MC ("Mission Control"), which provides to administrators and developers a complete insight about the JVM performance, behavior and possible bottlenecks. WebLogic Server 12c also have an classloader analysis tool embedded, and even a log analyzer tool that enables administrators and developers to view logs of multiple servers at the same time. JBoss EAP 6 on the other hand relies on third-part tools to do something similar. Again, only log searching are offered to find out whats going on. - WebLogic Server 12c offers end-to-end traceability and monitoring available through Oracle EM ("Enterprise Manager"), including monitoring of business transactions that flows through web servers, ESBs, application servers and database servers, all of this with high deep JVM analysis and diagnostics. JBoss EAP 6 on the other hand, even using JBoss ON ("Operations Network"), which is a separated technology, does not support those features. Red Hat relies on third-part tools to provide direct Oracle database traceability across JVMs. One of those tools are Oracle EM for non-Oracle middleware that manage JBoss, Tomcat, Websphere and IIS transparently. - WebLogic Server 12c with their JRockit support offers a tool called JRockit Flight Recorder, which can give developers a complete visibility of a certain period of application production monitoring with zero extra overhead. This automatic recording allows you to deep analyse threads latency, memory leaks, thread contention, resource utilization, stack overflow damages and GC ("Garbage Collection") cycles, to observe in real time stop-the-world phenomenons, generational, reference count and parallel collects and mutator threads analysis. JBoss EAP 6 don't even dream to support something similar, even because they don't have their own JVM. 3) Application Server Administration - WebLogic Server 12c offers a complete administration console complemented with scripting and macro-like recording capabilities. A single WebLogic console can managed up to hundreds of WebLogic servers belonging to the same domain. JBoss EAP 6 on the other hand has a limited console and provides a XML centric administration. JBoss, after ten years, started the development of a rudimentary centralized administration that still leave a lot of administration tasks aside, so admin people and developers must touch scripts and XML configuration files for most advanced and even simple administration tasks. This lead applications to error prone and risky deployments. Even using JBoss ON, JBoss EAP are not able to offer decent administration features for admin people which must be high skilled in JBoss internal architecture and its managing capabilities. - Oracle EM is available to manage multiple domains, databases, application servers, operating systems and virtualization, with a complete end-to-end visibility. JBoss ON does not provide management capabilities across the complete architecture, only basic monitoring. Even deployment must be done aside JBoss ON which does no integrate well with others softwares than JBoss. Until now, JBoss ON does not supports JBoss EAP 6, so even their minimal support for JBoss are not available for JBoss EAP 6 leaving customers uncovered and subject to high skilled JBoss admin people. - WebLogic Server 12c has the same administration model whatever is the topology selected by the customer. JBoss EAP 6 on the other hand differentiates between two operational models: standalone-mode and domain-mode, that are not consistent with each other. Depending on the mode used, the administration skill is different. - WebLogic Server 12c has no point-of-failures processes, and it does not need to define any specialized server. Domain model in WebLogic is available for years (at least ten years or more) and is production proven. JBoss EAP 6 on the other hand needs special processes to garantee JBoss integrity, the PC ("Process-Controller") and the HC ("Host-Controller"). Different from WebLogic, the domain model in JBoss is quite new (one year at tops) of maturity, and need to mature considerably until start doing things like WebLogic domain model does. - WebLogic Server 12c supports parallel deployment model which enables some artifacts being deployed at the same time. JBoss EAP 6 on the other hand does not have any similar feature. Every deployment are done atomically in the containers. This means that if you have a huge EAR (an EAR of 120 MB of size for instance) and deploy onto JBoss EAP 6, this EAR will take some minutes in order to starting accept thread requests. The same EAR deployed onto WebLogic Server 12c will reduce the deployment time at least in 2X compared to JBoss. 4) Support and Upgrades - WebLogic Server 12c has patch management available. JBoss EAP 6 on the other hand has no patch management available, each JBoss EAP instance should be patched manually. To achieve such feature, you need to buy a separated technology called JBoss ON ("Operations Network") that manage this type of stuff. But until now, JBoss ON does not support JBoss EAP 6 so, in practice, JBoss EAP 6 does not have this feature. - WebLogic Server 12c supports previuous WebLogic domains without any reconfiguration since its kernel is robust and mature since its creation in 1995. JBoss EAP 6 on the other hand has a proven lack of supportability between JBoss AS 4, 5, 6 and 7. Different kernels and messaging engines were implemented in JBoss stack in the last five years reveling their incapacity to create a well architected and proven middleware technology. - WebLogic Server 12c has patch prescription based on customer configuration. JBoss EAP 6 on the other hand has no such capability. People need to create ticket supports and have their installations revised by Red Hat support guys to gain some patch prescription from them. - Oracle WebLogic Server independent of the version has 8 years of support of new patches and has lifetime release of existing patches beyond that. JBoss EAP 6 on the other hand provides patches for a specific application server version up to 5 years after the release date. JBoss EAP 4 and previous versions had only 4 years. A good question that Red Hat will argue to answer is: "what happens when you find issues after year 5"?  5) RAC ("Real Application Clusters") Support - WebLogic Server 12c ships with a specific JDBC driver to leverage Oracle RAC clustering capabilities (Fast-Application-Notification, Transaction Affinity, Fast-Connection-Failover, etc). Oracle JDBC thin driver are also available. JBoss EAP 6 on the other hand ships only the standard Oracle JDBC thin driver. Load balancing with Oracle RAC are not supported. Manual intervention in case of planned or unplanned RAC downtime are necessary. In JBoss EAP 6, situation does not reestablish automatically after downtime. - WebLogic Server 12c has a feature called Active GridLink for Oracle RAC which provides up to 3X performance on OLTP applications. This seamless integration between WebLogic and Oracle database enable more value added to critical business applications leveraging their investments in Oracle database technology and Oracle middleware. JBoss EAP 6 on the other hand has no performance gains at all, even when admin people implement some kind of connection-pooling tuning. - WebLogic Server 12c also supports transaction and web session affinity to the Oracle RAC, which provides aditional gains of performance. This is particularly interesting if you are creating a reliable solution that are distributed not only in an LAN cluster, but into a different data center. JBoss EAP 6 on the other hand has no such support. 6) Standards and Technology Support - WebLogic Server 12c is fully Java EE 6 compatible and production ready since december of 2011. JBoss EAP 6 on the other hand became fully compatible with Java EE 6 only in the community version after three months, and production ready only in a few days considering that this article was written in June of 2012. Red Hat says that they are the masters of innovation and technology proliferation, but compared with Oracle and even other proprietary vendors like IBM, they historically speaking are lazy to deliver the most newest technologies and standards adherence. - Oracle is the steward of Java, driving innovation into the platform from commercial and open-source vendors. Red Hat on the other hand does not have its own JVM and relies on third-part JVMs to complete their application server offer. 95% of Red Hat customers are using Oracle HotSpot as JVM, which means that without Oracle involvement, their support are limited exclusively to the application server layer and we all know that most problems are happens in the JVM layer. - WebLogic Server 12c supports natively JDK 7, which empower developers to explore the maximum of the Java platform productivity when writing code. This feature differentiate WebLogic from others application servers (except GlassFish that are also managed by Oracle) because the usage of JDK 7 introduce such remarkable productivity features like the "try-with-resources" enhancement, catching multiple exceptions with one try block, Strings in the switch statements, JVM improvements in terms of JDBC, I/O, networking, security, concurrency and of course, the most important feature of Java 7: native support for multiple non-Java languages. More features regarding JDK 7 can be found here. JBoss EAP 6 on the other hand does not support JDK 7 officially, they comment in their community version that "Java SE 7 can be used with JBoss 7" which does not gives you any guarantees of enterprise support for JDK 7. - Oracle WebLogic Server 12c supports integration with Spring framework allowing Spring applications to use WebLogic special transaction manager, exposing bean interfaces to WebLogic MBeans to take advantage of all WebLogic monitoring and administration advantages. JBoss EAP 6 on the other hand has no special integration with Spring. In fact, Red Hat offers a suspicious package called "JBoss Web Platform" that in theory supports Spring, but in practice this package does not offers any special integration. It is just a facility for Red Hat customers to have support from both JBoss and Spring technology using the same customer support. 7) Lightweight Development - Oracle WebLogic Server 12c and Oracle GlassFish are completely integrated and can share applications without any modifications. Starting with the 12c version, WebLogic now understands natively GlassFish deployment descriptors and specific configurations in order to offer you a truly and reliable migration path from a community Java EE application server to a enterprise middleware product like WebLogic. JBoss EAP 6 on the other hand has no support to natively reuse an existing (or still in development) application from JBoss AS community server. Users of JBoss suffer of critical issues during deployment time that includes: changing the libraries and dependencies of the application, patching the DTD or XSD deployment descriptors, refactoring of the application layers due classloading issues and anomalies, rebuilding of persistence, business and web layers due issues with "usage of the certified version of an certain dependency" or "frameworks that Red Hat potentially does not recommend" etc. If you have the culture or enterprise IT directive of developing Java EE applications using community middleware to in a certain future, transition to enterprise (supported by a vendor) middleware, Oracle WebLogic plus Oracle GlassFish offers you a more sustainable solution. - WebLogic Server 12c has a very light ZIP distribution (less than 165 MB). JBoss EAP 6 ZIP size is around 130 MB, together with JBoss ON you have more 100 MB resulting in a higher download footprint. This is particularly interesting if you plan to use automated setup of application server instances (for example, to rapidly setup a development or staging environment) using Maven or Hudson. - WebLogic Server 12c has a complete integration with Maven allowing developers to setup WebLogic domains with few commands. Tasks like downloading WebLogic, installation, domain creation, data sources deployment are completely integrated. JBoss EAP 6 on the other hand has a limited offer integration with those tools.  - WebLogic Server 12c has a startup mode called WLX that turns-off EJB, JMS and JCA containers leaving enabled only the web container with Java EE 6 web profile. JBoss EAP 6 on the other hand has no such feature, you need to disable manually the containers that you do not want to use. - WebLogic Server 12c supports fastswap, which enables you to change classes without redeployment. This is particularly interesting if you are developing patches for the application that is already deployed and you do not want to redeploy the entire application. This is the same behavior that most application servers offers to JSP pages, but with WebLogic Server 12c, you have the same feature for Java classes in general. JBoss EAP 6 on the other hand has no such support. Even JBoss EAP 5 does not support this until now. 8) JMS and Messaging - WebLogic Server 12c has a proven and high scalable JMS implementation since its initial release in 1995. JBoss EAP 6 on the other hand has a still immature technology called HornetQ, which was introduced in JBoss EAP 5 replacing everything that was implemented in the previous versions. Red Hat loves to introduce new technologies across JBoss versions, playing around with customers and their investments. And when they are asked about why they have changed the implementation and caused such a mess, their answer is always: "the previous implementation was inadequate and not aligned with the community strategy so we are creating a new a improved one". This Red Hat practice leads to uncomfortable investments that in a near future (sometimes less than a year) will be affected in someway. - WebLogic Server 12c has troubleshooting and monitoring features included on the WebLogic console and WLDF. JBoss EAP 6 on the other hand has no direct monitoring on the console, activity is reflected only on the logs, no debug logs available in case of JMS issues. - WebLogic Server 12c has extremely good performance and scalability. JBoss EAP 6 on the other hand has a JMS storage mechanism relying on Oracle database or MySQL. This means that if an issue in production happens and Red Hat affirms that an performance issue is happening due to database problems, they will not support you on the performance issue. They will orient you to call Oracle instead. - WebLogic Server 12c supports messaging enterprise features like SAF ("Store and Forward"), Distributed Queues/Topics and Foreign JMS providers support that leverage JMS implementations without compromise developer code making things completely transparent. JBoss EAP 6 on the other hand do not even dream to support such features. 9) Caching and Grid - Coherence, which is the leading and most mature data grid technology from Oracle, is available since early 2000 and was integrated with WebLogic in 2009. Coherence and WebLogic clusters can be both managed from WebLogic administrative console. Even Node Manager supports Coherence. JBoss on the other hand discontinued JBoss Cache, which was their caching implementation just like they did with the messaging implementation (JBossMQ) which was a issue for long term customers. JBoss EAP 6 ships InfiniSpan version 1.0 which is immature and lack a proven record of successful cases and reliability. - WebLogic Server 12c has a feature called ActiveCache which uses Coherence to, without any code changes, replicate HTTP sessions from both WebLogic and other application servers like JBoss, Tomcat, Websphere, GlassFish and even Microsoft IIS. JBoss EAP 6 on the other hand does have such support and even when they do in the future, they probably will support only their own application server. - Coherence can be used to manage both L1 and L2 cache levels, providing support to Oracle TopLink and others JPA compliant implementations, even Hibernate. JBoss EAP 6 and Infinispan on the other hand supports only Hibernate. And most important of all: Infinispan does not have any successful case of L1 or L2 caching level support using Hibernate, which lead us to reflect about its viability. 10) Performance - WebLogic Server 12c is certified with Oracle Exalogic Elastic Cloud and can run unchanged applications at this engineered system. This approach can benefit customers from Exalogic optimization's of both kernel and JVM layers to boost performance in terms of 10X for web, OLTP, JMS and grid applications. JBoss EAP 6 on the other hand has no investment on engineered systems: customers do not have the choice to deploy on a Java ultra fast system if their project becomes relevant and performance issues are detected. - WebLogic Server 12c maintains a performance gain across each new release: starting on WebLogic 5.1, the overall performance gain has been close to 4X, which close to a 20% gain release by release. JBoss on the other hand does not provide SPECJAppServer or SPECJEnterprise performance benchmarks. Their so called "performance gains" remains hidden in their customer environments, which lead us to think if it is true or not since we will never get access to those environments. - WebLogic Server 12c has industry performance benchmarks with submissions across platforms and configurations leading SPECJ. Oracle WebLogic leads SPECJAppServer performance in multiple categories, fitting all customer topologies like: dual-node, single-node, multi-node and multi-node with RAC. JBoss... again, does not provide any SPECJAppServer performance benchmarks. - WebLogic Server 12c has a feature called work manager which allows your application to embrace new performance levels based on critical resource utilization of the CPUs usage. Work managers prioritizes work and allocates threads based on an execution model that takes into account administrator-defined parameters and actual run-time performance and throughput. JBoss EAP 6 on the other hand has no compared feature and probably they never will. Not supporting such feature like work managers, JBoss EAP 6 forces admin people and specially developers to uncover performance gains in a intrusive way, rewriting the code and doing performance refactorings. 11) Professional Services Support - WebLogic Server 12c and any other technology sold by Oracle give customers the possibility of hire OCS ("Oracle Consulting Services") to manage critical scenarios, deployment assistance of new applications, high skilled consultancy of architecture, best practices and people allocation together with customer teams. All OCS services are available without any restrictions, having the customer bought software from Oracle or just starting their implementation before any acquisition. JBoss EAP 6 or Red Hat to be more specifically, only offers professional services if you buy subscriptions from them. If you are developing a new critical application for your business and need the help of Red Hat for a serious issue or architecture decision, they will probably say: "OK... I can help you but after you buy subscriptions from me". Red Hat also does not allows their professional services consultants to manage environments that uses community based software. They will probably force you to first buy a subscription, download their "enterprise" version and them, optionally hire their consultants. - Oracle provides you our university to educate your team into our technologies, including of course specialized trainings of WebLogic application server. At any time and location, you can hire Oracle to train your team so you get trustful knowledge according to your specific needs. Certifications for the products are also available if your technical people desire to differentiate themselves as professionals. Red Hat on the other hand have a limited pool of resources to train your team in their technologies. Basically they are selling training and certification for RHEL ("Red Hat Enterprise Linux") but if you demand more specialized training in JBoss middleware, they will probably connect you to some "certified" partner localized training since they are apparently discontinuing their education center, at least here in Brazil. They were not able to reproduce their success with RHEL education to their middleware division since they need first sell the subscriptions to after gives you specialized training. And again, they only offer you specialized training based on their enterprise version (EAP in the case of JBoss) which means that the courses will be a quite outdated. There are reports of developers that took official training's from Red Hat at this year (2012) and in a certain JBoss advanced course, Red Hat supposedly covered JBossMQ as the messaging subsystem, and even the printed material provided was based on JBossMQ since the training was created for JBoss EAP 4.3. 12) Encouraging Transparency without Ulterior Motives - WebLogic Server 12c like any other software from Oracle can be downloaded any time from anywhere, you should only possess an OTN ("Oracle Technology Network") credential and you can download any enterprise software how many times you want. And is not some kind of "trial" version. It is the official binaries that will be running for ever in your data center. Oracle does not encourages the usage of "specific versions" of our software. The binaries you buy from Oracle are the same binaries anyone in the world could download and use for testing and personal education. JBoss EAP 6 on the other hand are not available for download unless you buy a subscription and get access to the Red Hat enterprise repositories. If you need to test, learn or just start creating your application using Red Hat's middleware software, you should download it from the community website. You are not allowed to download the enterprise version that, according to Red Hat are more secure, reliable and robust. But no one of us want to start the development of a software with an unsecured, unreliable and not scalable middleware right? So what you do? You are "invited" by Red Hat to buy subscriptions from them to get access to the "cool" version of the software. - WebLogic Server 12c prices are publicly available in the Oracle website. If you want to know right now how much WebLogic will cost to your organization, just click here and get access to our price list. In the case of WebLogic, check out the "US Oracle Technology Commercial Price List". Oracle also encourages you to get in touch with a sales representative to discuss discounts that would make possible the investment into our technology. But you are not required to do this, only if you are interested in buying our technology or maybe you want to discuss some discount scenarios. JBoss EAP 6 on the other hand does not have its cost publicly available in Red Hat's website or in any other media, at least is not so easy to get such information. The only link you will possibly find in their website is a "Contact a Sales Representative" link. This is not a very good relationship between an customer and an vendor. This is not an example of transparency, mainly when the software are sold as open. In this situations, customers expects to see the software prices publicly available, so they can have the chance to decide, based on the existing features of the software, if the cost is fair or not. Conclusion Oracle WebLogic is the most mature, secure, reliable and scalable Java EE application server of the market, and have a proven record of success around the globe to prove it's majority. Don't lose the chance to discover today how WebLogic could fit your needs and sustain your global IT middleware strategy, no matter if your strategy are completely based on the Cloud or not.

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  • What&rsquo;s New in ASP.NET 4.0 Part Two: WebForms and Visual Studio Enhancements

    - by Rick Strahl
    In the last installment I talked about the core changes in the ASP.NET runtime that I’ve been taking advantage of. In this column, I’ll cover the changes to the Web Forms engine and some of the cool improvements in Visual Studio that make Web and general development easier. WebForms The WebForms engine is the area that has received most significant changes in ASP.NET 4.0. Probably the most widely anticipated features are related to managing page client ids and of ViewState on WebForm pages. Take Control of Your ClientIDs Unique ClientID generation in ASP.NET has been one of the most complained about “features” in ASP.NET. Although there’s a very good technical reason for these unique generated ids - they guarantee unique ids for each and every server control on a page - these unique and generated ids often get in the way of client-side JavaScript development and CSS styling as it’s often inconvenient and fragile to work with the long, generated ClientIDs. In ASP.NET 4.0 you can now specify an explicit client id mode on each control or each naming container parent control to control how client ids are generated. By default, ASP.NET generates mangled client ids for any control contained in a naming container (like a Master Page, or a User Control for example). The key to ClientID management in ASP.NET 4.0 are the new ClientIDMode and ClientIDRowSuffix properties. ClientIDMode supports four different ClientID generation settings shown below. For the following examples, imagine that you have a Textbox control named txtName inside of a master page control container on a WebForms page. <%@Page Language="C#"      MasterPageFile="~/Site.Master"     CodeBehind="WebForm2.aspx.cs"     Inherits="WebApplication1.WebForm2"  %> <asp:Content ID="content"  ContentPlaceHolderID="content"               runat="server"               ClientIDMode="Static" >       <asp:TextBox runat="server" ID="txtName" /> </asp:Content> The four available ClientIDMode values are: AutoID This is the existing behavior in ASP.NET 1.x-3.x where full naming container munging takes place. <input name="ctl00$content$txtName" type="text"        id="ctl00_content_txtName" /> This should be familiar to any ASP.NET developer and results in fairly unpredictable client ids that can easily change if the containership hierarchy changes. For example, removing the master page changes the name in this case, so if you were to move a block of script code that works against the control to a non-Master page, the script code immediately breaks. Static This option is the most deterministic setting that forces the control’s ClientID to use its ID value directly. No naming container naming at all is applied and you end up with clean client ids: <input name="ctl00$content$txtName"         type="text" id="txtName" /> Note that the name property which is used for postback variables to the server still is munged, but the ClientID property is displayed simply as the ID value that you have assigned to the control. This option is what most of us want to use, but you have to be clear on that because it can potentially cause conflicts with other controls on the page. If there are several instances of the same naming container (several instances of the same user control for example) there can easily be a client id naming conflict. Note that if you assign Static to a data-bound control, like a list child control in templates, you do not get unique ids either, so for list controls where you rely on unique id for child controls, you’ll probably want to use Predictable rather than Static. I’ll write more on this a little later when I discuss ClientIDRowSuffix. Predictable The previous two values are pretty self-explanatory. Predictable however, requires some explanation. To me at least it’s not in the least bit predictable. MSDN defines this value as follows: This algorithm is used for controls that are in data-bound controls. The ClientID value is generated by concatenating the ClientID value of the parent naming container with the ID value of the control. If the control is a data-bound control that generates multiple rows, the value of the data field specified in the ClientIDRowSuffix property is added at the end. For the GridView control, multiple data fields can be specified. If the ClientIDRowSuffix property is blank, a sequential number is added at the end instead of a data-field value. Each segment is separated by an underscore character (_). The key that makes this value a bit confusing is that it relies on the parent NamingContainer’s ClientID to build its own ClientID value. This effectively means that the value is not predictable at all but rather very tightly coupled to the parent naming container’s ClientIDMode setting. For my simple textbox example, if the ClientIDMode property of the parent naming container (Page in this case) is set to “Predictable” you’ll get this: <input name="ctl00$content$txtName" type="text"         id="content_txtName" /> which gives an id that based on walking up to the currently active naming container (the MasterPage content container) and starting the id formatting from there downward. Think of this as a semi unique name that’s guaranteed unique only for the naming container. If, on the other hand, the Page is set to “AutoID” you get the following with Predictable on txtName: <input name="ctl00$content$txtName" type="text"         id="ctl00_content_txtName" /> The latter is effectively the same as if you specified AutoID because it inherits the AutoID naming from the Page and Content Master Page control of the page. But again - predictable behavior always depends on the parent naming container and how it generates its id, so the id may not always be exactly the same as the AutoID generated value because somewhere in the NamingContainer chain the ClientIDMode setting may be set to a different value. For example, if you had another naming container in the middle that was set to Static you’d end up effectively with an id that starts with the NamingContainers id rather than the whole ctl000_content munging. The most common use for Predictable is likely to be for data-bound controls, which results in each data bound item getting a unique ClientID. Unfortunately, even here the behavior can be very unpredictable depending on which data-bound control you use - I found significant differences in how template controls in a GridView behave from those that are used in a ListView control. For example, GridView creates clean child ClientIDs, while ListView still has a naming container in the ClientID, presumably because of the template container on which you can’t set ClientIDMode. Predictable is useful, but only if all naming containers down the chain use this setting. Otherwise you’re right back to the munged ids that are pretty unpredictable. Another property, ClientIDRowSuffix, can be used in combination with ClientIDMode of Predictable to force a suffix onto list client controls. For example: <asp:GridView runat="server" ID="gvItems"              AutoGenerateColumns="false"             ClientIDMode="Static"              ClientIDRowSuffix="Id">     <Columns>     <asp:TemplateField>         <ItemTemplate>             <asp:Label runat="server" id="txtName"                        Text='<%# Eval("Name") %>'                   ClientIDMode="Predictable"/>         </ItemTemplate>     </asp:TemplateField>     <asp:TemplateField>         <ItemTemplate>         <asp:Label runat="server" id="txtId"                     Text='<%# Eval("Id") %>'                     ClientIDMode="Predictable" />         </ItemTemplate>     </asp:TemplateField>     </Columns>  </asp:GridView> generates client Ids inside of a column in the master page described earlier: <td>     <span id="txtName_0">Rick</span> </td> where the value after the underscore is the ClientIDRowSuffix field - in this case “Id” of the item data bound to the control. Note that all of the child controls require ClientIDMode=”Predictable” in order for the ClientIDRowSuffix to be applied, and the parent GridView controls need to be set to Static either explicitly or via Naming Container inheritance to give these simple names. It’s a bummer that ClientIDRowSuffix doesn’t work with Static to produce this automatically. Another real problem is that other controls process the ClientIDMode differently. For example, a ListView control processes the Predictable ClientIDMode differently and produces the following with the Static ListView and Predictable child controls: <span id="ctrl0_txtName_0">Rick</span> I couldn’t even figure out a way using ClientIDMode to get a simple ID that also uses a suffix short of falling back to manually generated ids using <%= %> expressions instead. Given the inconsistencies inside of list controls using <%= %>, ids for the ListView might not be a bad idea anyway. Inherit The final setting is Inherit, which is the default for all controls except Page. This means that controls by default inherit the parent naming container’s ClientIDMode setting. For more detailed information on ClientID behavior and different scenarios you can check out a blog post of mine on this subject: http://www.west-wind.com/weblog/posts/54760.aspx. ClientID Enhancements Summary The ClientIDMode property is a welcome addition to ASP.NET 4.0. To me this is probably the most useful WebForms feature as it allows me to generate clean IDs simply by setting ClientIDMode="Static" on either the page or inside of Web.config (in the Pages section) which applies the setting down to the entire page which is my 95% scenario. For the few cases when it matters - for list controls and inside of multi-use user controls or custom server controls) - I can use Predictable or even AutoID to force controls to unique names. For application-level page development, this is easy to accomplish and provides maximum usability for working with client script code against page controls. ViewStateMode Another area of large criticism for WebForms is ViewState. ViewState is used internally by ASP.NET to persist page-level changes to non-postback properties on controls as pages post back to the server. It’s a useful mechanism that works great for the overall mechanics of WebForms, but it can also cause all sorts of overhead for page operation as ViewState can very quickly get out of control and consume huge amounts of bandwidth in your page content. ViewState can also wreak havoc with client-side scripting applications that modify control properties that are tracked by ViewState, which can produce very unpredictable results on a Postback after client-side updates. Over the years in my own development, I’ve often turned off ViewState on pages to reduce overhead. Yes, you lose some functionality, but you can easily implement most of the common functionality in non-ViewState workarounds. Relying less on heavy ViewState controls and sticking with simpler controls or raw HTML constructs avoids getting around ViewState problems. In ASP.NET 3.x and prior, it wasn’t easy to control ViewState - you could turn it on or off and if you turned it off at the page or web.config level, you couldn’t turn it back on for specific controls. In short, it was an all or nothing approach. With ASP.NET 4.0, the new ViewStateMode property gives you more control. It allows you to disable ViewState globally either on the page or web.config level and then turn it back on for specific controls that might need it. ViewStateMode only works when EnableViewState="true" on the page or web.config level (which is the default). You can then use ViewStateMode of Disabled, Enabled or Inherit to control the ViewState settings on the page. If you’re shooting for minimal ViewState usage, the ideal situation is to set ViewStateMode to disabled on the Page or web.config level and only turn it back on particular controls: <%@Page Language="C#"      CodeBehind="WebForm2.aspx.cs"     Inherits="Westwind.WebStore.WebForm2"        ClientIDMode="Static"                ViewStateMode="Disabled"     EnableViewState="true"  %> <!-- this control has viewstate  --> <asp:TextBox runat="server" ID="txtName"  ViewStateMode="Enabled" />       <!-- this control has no viewstate - it inherits  from parent container --> <asp:TextBox runat="server" ID="txtAddress" /> Note that the EnableViewState="true" at the Page level isn’t required since it’s the default, but it’s important that the value is true. ViewStateMode has no effect if EnableViewState="false" at the page level. The main benefit of ViewStateMode is that it allows you to more easily turn off ViewState for most of the page and enable only a few key controls that might need it. For me personally, this is a perfect combination as most of my WebForm apps can get away without any ViewState at all. But some controls - especially third party controls - often don’t work well without ViewState enabled, and now it’s much easier to selectively enable controls rather than the old way, which required you to pretty much turn off ViewState for all controls that you didn’t want ViewState on. Inline HTML Encoding HTML encoding is an important feature to prevent cross-site scripting attacks in data entered by users on your site. In order to make it easier to create HTML encoded content, ASP.NET 4.0 introduces a new Expression syntax using <%: %> to encode string values. The encoding expression syntax looks like this: <%: "<script type='text/javascript'>" +     "alert('Really?');</script>" %> which produces properly encoded HTML: &lt;script type=&#39;text/javascript&#39; &gt;alert(&#39;Really?&#39;);&lt;/script&gt; Effectively this is a shortcut to: <%= HttpUtility.HtmlEncode( "<script type='text/javascript'>" + "alert('Really?');</script>") %> Of course the <%: %> syntax can also evaluate expressions just like <%= %> so the more common scenario applies this expression syntax against data your application is displaying. Here’s an example displaying some data model values: <%: Model.Address.Street %> This snippet shows displaying data from your application’s data store or more importantly, from data entered by users. Anything that makes it easier and less verbose to HtmlEncode text is a welcome addition to avoid potential cross-site scripting attacks. Although I listed Inline HTML Encoding here under WebForms, anything that uses the WebForms rendering engine including ASP.NET MVC, benefits from this feature. ScriptManager Enhancements The ASP.NET ScriptManager control in the past has introduced some nice ways to take programmatic and markup control over script loading, but there were a number of shortcomings in this control. The ASP.NET 4.0 ScriptManager has a number of improvements that make it easier to control script loading and addresses a few of the shortcomings that have often kept me from using the control in favor of manual script loading. The first is the AjaxFrameworkMode property which finally lets you suppress loading the ASP.NET AJAX runtime. Disabled doesn’t load any ASP.NET AJAX libraries, but there’s also an Explicit mode that lets you pick and choose the library pieces individually and reduce the footprint of ASP.NET AJAX script included if you are using the library. There’s also a new EnableCdn property that forces any script that has a new WebResource attribute CdnPath property set to a CDN supplied URL. If the script has this Attribute property set to a non-null/empty value and EnableCdn is enabled on the ScriptManager, that script will be served from the specified CdnPath. [assembly: WebResource(    "Westwind.Web.Resources.ww.jquery.js",    "application/x-javascript",    CdnPath =  "http://mysite.com/scripts/ww.jquery.min.js")] Cool, but a little too static for my taste since this value can’t be changed at runtime to point at a debug script as needed, for example. Assembly names for loading scripts from resources can now be simple names rather than fully qualified assembly names, which make it less verbose to reference scripts from assemblies loaded from your bin folder or the assembly reference area in web.config: <asp:ScriptManager runat="server" id="Id"          EnableCdn="true"         AjaxFrameworkMode="disabled">     <Scripts>         <asp:ScriptReference          Name="Westwind.Web.Resources.ww.jquery.js"         Assembly="Westwind.Web" />     </Scripts>        </asp:ScriptManager> The ScriptManager in 4.0 also supports script combining via the CompositeScript tag, which allows you to very easily combine scripts into a single script resource served via ASP.NET. Even nicer: You can specify the URL that the combined script is served with. Check out the following script manager markup that combines several static file scripts and a script resource into a single ASP.NET served resource from a static URL (allscripts.js): <asp:ScriptManager runat="server" id="Id"          EnableCdn="true"         AjaxFrameworkMode="disabled">     <CompositeScript          Path="~/scripts/allscripts.js">         <Scripts>             <asp:ScriptReference                    Path="~/scripts/jquery.js" />             <asp:ScriptReference                    Path="~/scripts/ww.jquery.js" />             <asp:ScriptReference            Name="Westwind.Web.Resources.editors.js"                 Assembly="Westwind.Web" />         </Scripts>     </CompositeScript> </asp:ScriptManager> When you render this into HTML, you’ll see a single script reference in the page: <script src="scripts/allscripts.debug.js"          type="text/javascript"></script> All you need to do to make this work is ensure that allscripts.js and allscripts.debug.js exist in the scripts folder of your application - they can be empty but the file has to be there. This is pretty cool, but you want to be real careful that you use unique URLs for each combination of scripts you combine or else browser and server caching will easily screw you up royally. The script manager also allows you to override native ASP.NET AJAX scripts now as any script references defined in the Scripts section of the ScriptManager trump internal references. So if you want custom behavior or you want to fix a possible bug in the core libraries that normally are loaded from resources, you can now do this simply by referencing the script resource name in the Name property and pointing at System.Web for the assembly. Not a common scenario, but when you need it, it can come in real handy. Still, there are a number of shortcomings in this control. For one, the ScriptManager and ClientScript APIs still have no common entry point so control developers are still faced with having to check and support both APIs to load scripts so that controls can work on pages that do or don’t have a ScriptManager on the page. The CdnUrl is static and compiled in, which is very restrictive. And finally, there’s still no control over where scripts get loaded on the page - ScriptManager still injects scripts into the middle of the HTML markup rather than in the header or optionally the footer. This, in turn, means there is little control over script loading order, which can be problematic for control developers. MetaDescription, MetaKeywords Page Properties There are also a number of additional Page properties that correspond to some of the other features discussed in this column: ClientIDMode, ClientTarget and ViewStateMode. Another minor but useful feature is that you can now directly access the MetaDescription and MetaKeywords properties on the Page object to set the corresponding meta tags programmatically. Updating these values programmatically previously required either <%= %> expressions in the page markup or dynamic insertion of literal controls into the page. You can now just set these properties programmatically on the Page object in any Control derived class on the page or the Page itself: Page.MetaKeywords = "ASP.NET,4.0,New Features"; Page.MetaDescription = "This article discusses the new features in ASP.NET 4.0"; Note, that there’s no corresponding ASP.NET tag for the HTML Meta element, so the only way to specify these values in markup and access them is via the @Page tag: <%@Page Language="C#"      CodeBehind="WebForm2.aspx.cs"     Inherits="Westwind.WebStore.WebForm2"      ClientIDMode="Static"                MetaDescription="Article that discusses what's                      new in ASP.NET 4.0"     MetaKeywords="ASP.NET,4.0,New Features" %> Nothing earth shattering but quite convenient. Visual Studio 2010 Enhancements for Web Development For Web development there are also a host of editor enhancements in Visual Studio 2010. Some of these are not Web specific but they are useful for Web developers in general. Text Editors Throughout Visual Studio 2010, the text editors have all been updated to a new core engine based on WPF which provides some interesting new features for various code editors including the nice ability to zoom in and out with Ctrl-MouseWheel to quickly change the size of text. There are many more API options to control the editor and although Visual Studio 2010 doesn’t yet use many of these features, we can look forward to enhancements in add-ins and future editor updates from the various language teams that take advantage of the visual richness that WPF provides to editing. On the negative side, I’ve noticed that occasionally the code editor and especially the HTML and JavaScript editors will lose the ability to use various navigation keys like arrows, back and delete keys, which requires closing and reopening the documents at times. This issue seems to be well documented so I suspect this will be addressed soon with a hotfix or within the first service pack. Overall though, the code editors work very well, especially given that they were re-written completely using WPF, which was one of my big worries when I first heard about the complete redesign of the editors. Multi-Targeting Visual Studio now targets all versions of the .NET framework from 2.0 forward. You can use Visual Studio 2010 to work on your ASP.NET 2, 3.0 and 3.5 applications which is a nice way to get your feet wet with the new development environment without having to make changes to existing applications. It’s nice to have one tool to work in for all the different versions. Multi-Monitor Support One cool feature of Visual Studio 2010 is the ability to drag windows out of the Visual Studio environment and out onto the desktop including onto another monitor easily. Since Web development often involves working with a host of designers at the same time - visual designer, HTML markup window, code behind and JavaScript editor - it’s really nice to be able to have a little more screen real estate to work on each of these editors. Microsoft made a welcome change in the environment. IntelliSense Snippets for HTML and JavaScript Editors The HTML and JavaScript editors now finally support IntelliSense scripts to create macro-based template expansions that have been in the core C# and Visual Basic code editors since Visual Studio 2005. Snippets allow you to create short XML-based template definitions that can act as static macros or real templates that can have replaceable values that can be embedded into the expanded text. The XML syntax for these snippets is straight forward and it’s pretty easy to create custom snippets manually. You can easily create snippets using XML and store them in your custom snippets folder (C:\Users\rstrahl\Documents\Visual Studio 2010\Code Snippets\Visual Web Developer\My HTML Snippets and My JScript Snippets), but it helps to use one of the third-party tools that exist to simplify the process for you. I use SnippetEditor, by Bill McCarthy, which makes short work of creating snippets interactively (http://snippeteditor.codeplex.com/). Note: You may have to manually add the Visual Studio 2010 User specific Snippet folders to this tool to see existing ones you’ve created. Code snippets are some of the biggest time savers and HTML editing more than anything deals with lots of repetitive tasks that lend themselves to text expansion. Visual Studio 2010 includes a slew of built-in snippets (that you can also customize!) and you can create your own very easily. If you haven’t done so already, I encourage you to spend a little time examining your coding patterns and find the repetitive code that you write and convert it into snippets. I’ve been using CodeRush for this for years, but now you can do much of the basic expansion natively for HTML and JavaScript snippets. jQuery Integration Is Now Native jQuery is a popular JavaScript library and recently Microsoft has recently stated that it will become the primary client-side scripting technology to drive higher level script functionality in various ASP.NET Web projects that Microsoft provides. In Visual Studio 2010, the default full project template includes jQuery as part of a new project including the support files that provide IntelliSense (-vsdoc files). IntelliSense support for jQuery is now also baked into Visual Studio 2010, so unlike Visual Studio 2008 which required a separate download, no further installs are required for a rich IntelliSense experience with jQuery. Summary ASP.NET 4.0 brings many useful improvements to the platform, but thankfully most of the changes are incremental changes that don’t compromise backwards compatibility and they allow developers to ease into the new features one feature at a time. None of the changes in ASP.NET 4.0 or Visual Studio 2010 are monumental or game changers. The bigger features are language and .NET Framework changes that are also optional. This ASP.NET and tools release feels more like fine tuning and getting some long-standing kinks worked out of the platform. It shows that the ASP.NET team is dedicated to paying attention to community feedback and responding with changes to the platform and development environment based on this feedback. If you haven’t gotten your feet wet with ASP.NET 4.0 and Visual Studio 2010, there’s no reason not to give it a shot now - the ASP.NET 4.0 platform is solid and Visual Studio 2010 works very well for a brand new release. Check it out. © Rick Strahl, West Wind Technologies, 2005-2010Posted in ASP.NET  

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  • Silverlight Recruiting Application Part 5 - Jobs Module / View

    Now we starting getting into a more code-heavy portion of this series, thankfully though this means the groundwork is all set for the most part and after adding the modules we will have a complete application that can be provided with full source. The Jobs module will have two concerns- adding and maintaining jobs that can then be broadcast out to the website. How they are displayed on the site will be handled by our admin system (which will just poll from this common database), so we aren't too concerned with that, but rather with getting the information into the system and allowing the backend administration/HR users to keep things up to date. Since there is a fair bit of information that we want to display, we're going to move editing to a separate view so we can get all that information in an easy-to-use spot. With all the files created for this module, the project looks something like this: And now... on to the code. XAML for the Job Posting View All we really need for the Job Posting View is a RadGridView and a few buttons. This will let us both show off records and perform operations on the records without much hassle. That XAML is going to look something like this: 01.<Grid x:Name="LayoutRoot" 02.Background="White"> 03.<Grid.RowDefinitions> 04.<RowDefinition Height="30" /> 05.<RowDefinition /> 06.</Grid.RowDefinitions> 07.<StackPanel Orientation="Horizontal"> 08.<Button x:Name="xAddRecordButton" 09.Content="Add Job" 10.Width="120" 11.cal:Click.Command="{Binding AddRecord}" 12.telerik:StyleManager.Theme="Windows7" /> 13.<Button x:Name="xEditRecordButton" 14.Content="Edit Job" 15.Width="120" 16.cal:Click.Command="{Binding EditRecord}" 17.telerik:StyleManager.Theme="Windows7" /> 18.</StackPanel> 19.<telerikGrid:RadGridView x:Name="xJobsGrid" 20.Grid.Row="1" 21.IsReadOnly="True" 22.AutoGenerateColumns="False" 23.ColumnWidth="*" 24.RowDetailsVisibilityMode="VisibleWhenSelected" 25.ItemsSource="{Binding MyJobs}" 26.SelectedItem="{Binding SelectedJob, Mode=TwoWay}" 27.command:SelectedItemChangedEventClass.Command="{Binding SelectedItemChanged}"> 28.<telerikGrid:RadGridView.Columns> 29.<telerikGrid:GridViewDataColumn Header="Job Title" 30.DataMemberBinding="{Binding JobTitle}" 31.UniqueName="JobTitle" /> 32.<telerikGrid:GridViewDataColumn Header="Location" 33.DataMemberBinding="{Binding Location}" 34.UniqueName="Location" /> 35.<telerikGrid:GridViewDataColumn Header="Resume Required" 36.DataMemberBinding="{Binding NeedsResume}" 37.UniqueName="NeedsResume" /> 38.<telerikGrid:GridViewDataColumn Header="CV Required" 39.DataMemberBinding="{Binding NeedsCV}" 40.UniqueName="NeedsCV" /> 41.<telerikGrid:GridViewDataColumn Header="Overview Required" 42.DataMemberBinding="{Binding NeedsOverview}" 43.UniqueName="NeedsOverview" /> 44.<telerikGrid:GridViewDataColumn Header="Active" 45.DataMemberBinding="{Binding IsActive}" 46.UniqueName="IsActive" /> 47.</telerikGrid:RadGridView.Columns> 48.</telerikGrid:RadGridView> 49.</Grid> I'll explain what's happening here by line numbers: Lines 11 and 16: Using the same type of click commands as we saw in the Menu module, we tie the button clicks to delegate commands in the viewmodel. Line 25: The source for the jobs will be a collection in the viewmodel. Line 26: We also bind the selected item to a public property from the viewmodel for use in code. Line 27: We've turned the event into a command so we can handle it via code in the viewmodel. So those first three probably make sense to you as far as Silverlight/WPF binding magic is concerned, but for line 27... This actually comes from something I read onDamien Schenkelman's blog back in the day for creating an attached behavior from any event. So, any time you see me using command:Whatever.Command, the backing for it is actually something like this: SelectedItemChangedEventBehavior.cs: 01.public class SelectedItemChangedEventBehavior : CommandBehaviorBase<Telerik.Windows.Controls.DataControl> 02.{ 03.public SelectedItemChangedEventBehavior(DataControl element) 04.: base(element) 05.{ 06.element.SelectionChanged += new EventHandler<SelectionChangeEventArgs>(element_SelectionChanged); 07.} 08.void element_SelectionChanged(object sender, SelectionChangeEventArgs e) 09.{ 10.// We'll only ever allow single selection, so will only need item index 0 11.base.CommandParameter = e.AddedItems[0]; 12.base.ExecuteCommand(); 13.} 14.} SelectedItemChangedEventClass.cs: 01.public class SelectedItemChangedEventClass 02.{ 03.#region The Command Stuff 04.public static ICommand GetCommand(DependencyObject obj) 05.{ 06.return (ICommand)obj.GetValue(CommandProperty); 07.} 08.public static void SetCommand(DependencyObject obj, ICommand value) 09.{ 10.obj.SetValue(CommandProperty, value); 11.} 12.public static readonly DependencyProperty CommandProperty = 13.DependencyProperty.RegisterAttached("Command", typeof(ICommand), 14.typeof(SelectedItemChangedEventClass), new PropertyMetadata(OnSetCommandCallback)); 15.public static void OnSetCommandCallback(DependencyObject dependencyObject, DependencyPropertyChangedEventArgs e) 16.{ 17.DataControl element = dependencyObject as DataControl; 18.if (element != null) 19.{ 20.SelectedItemChangedEventBehavior behavior = GetOrCreateBehavior(element); 21.behavior.Command = e.NewValue as ICommand; 22.} 23.} 24.#endregion 25.public static SelectedItemChangedEventBehavior GetOrCreateBehavior(DataControl element) 26.{ 27.SelectedItemChangedEventBehavior behavior = element.GetValue(SelectedItemChangedEventBehaviorProperty) as SelectedItemChangedEventBehavior; 28.if (behavior == null) 29.{ 30.behavior = new SelectedItemChangedEventBehavior(element); 31.element.SetValue(SelectedItemChangedEventBehaviorProperty, behavior); 32.} 33.return behavior; 34.} 35.public static SelectedItemChangedEventBehavior GetSelectedItemChangedEventBehavior(DependencyObject obj) 36.{ 37.return (SelectedItemChangedEventBehavior)obj.GetValue(SelectedItemChangedEventBehaviorProperty); 38.} 39.public static void SetSelectedItemChangedEventBehavior(DependencyObject obj, SelectedItemChangedEventBehavior value) 40.{ 41.obj.SetValue(SelectedItemChangedEventBehaviorProperty, value); 42.} 43.public static readonly DependencyProperty SelectedItemChangedEventBehaviorProperty = 44.DependencyProperty.RegisterAttached("SelectedItemChangedEventBehavior", 45.typeof(SelectedItemChangedEventBehavior), typeof(SelectedItemChangedEventClass), null); 46.} These end up looking very similar from command to command, but in a nutshell you create a command based on any event, determine what the parameter for it will be, then execute. It attaches via XAML and ties to a DelegateCommand in the viewmodel, so you get the full event experience (since some controls get a bit event-rich for added functionality). Simple enough, right? Viewmodel for the Job Posting View The Viewmodel is going to need to handle all events going back and forth, maintaining interactions with the data we are using, and both publishing and subscribing to events. Rather than breaking this into tons of little pieces, I'll give you a nice view of the entire viewmodel and then hit up the important points line-by-line: 001.public class JobPostingViewModel : ViewModelBase 002.{ 003.private readonly IEventAggregator eventAggregator; 004.private readonly IRegionManager regionManager; 005.public DelegateCommand<object> AddRecord { get; set; } 006.public DelegateCommand<object> EditRecord { get; set; } 007.public DelegateCommand<object> SelectedItemChanged { get; set; } 008.public RecruitingContext context; 009.private QueryableCollectionView _myJobs; 010.public QueryableCollectionView MyJobs 011.{ 012.get { return _myJobs; } 013.} 014.private QueryableCollectionView _selectionJobActionHistory; 015.public QueryableCollectionView SelectedJobActionHistory 016.{ 017.get { return _selectionJobActionHistory; } 018.} 019.private JobPosting _selectedJob; 020.public JobPosting SelectedJob 021.{ 022.get { return _selectedJob; } 023.set 024.{ 025.if (value != _selectedJob) 026.{ 027._selectedJob = value; 028.NotifyChanged("SelectedJob"); 029.} 030.} 031.} 032.public SubscriptionToken editToken = new SubscriptionToken(); 033.public SubscriptionToken addToken = new SubscriptionToken(); 034.public JobPostingViewModel(IEventAggregator eventAgg, IRegionManager regionmanager) 035.{ 036.// set Unity items 037.this.eventAggregator = eventAgg; 038.this.regionManager = regionmanager; 039.// load our context 040.context = new RecruitingContext(); 041.this._myJobs = new QueryableCollectionView(context.JobPostings); 042.context.Load(context.GetJobPostingsQuery()); 043.// set command events 044.this.AddRecord = new DelegateCommand<object>(this.AddNewRecord); 045.this.EditRecord = new DelegateCommand<object>(this.EditExistingRecord); 046.this.SelectedItemChanged = new DelegateCommand<object>(this.SelectedRecordChanged); 047.SetSubscriptions(); 048.} 049.#region DelegateCommands from View 050.public void AddNewRecord(object obj) 051.{ 052.this.eventAggregator.GetEvent<AddJobEvent>().Publish(true); 053.} 054.public void EditExistingRecord(object obj) 055.{ 056.if (_selectedJob == null) 057.{ 058.this.eventAggregator.GetEvent<NotifyUserEvent>().Publish("No job selected."); 059.} 060.else 061.{ 062.this._myJobs.EditItem(this._selectedJob); 063.this.eventAggregator.GetEvent<EditJobEvent>().Publish(this._selectedJob); 064.} 065.} 066.public void SelectedRecordChanged(object obj) 067.{ 068.if (obj.GetType() == typeof(ActionHistory)) 069.{ 070.// event bubbles up so we don't catch items from the ActionHistory grid 071.} 072.else 073.{ 074.JobPosting job = obj as JobPosting; 075.GrabHistory(job.PostingID); 076.} 077.} 078.#endregion 079.#region Subscription Declaration and Events 080.public void SetSubscriptions() 081.{ 082.EditJobCompleteEvent editComplete = eventAggregator.GetEvent<EditJobCompleteEvent>(); 083.if (editToken != null) 084.editComplete.Unsubscribe(editToken); 085.editToken = editComplete.Subscribe(this.EditCompleteEventHandler); 086.AddJobCompleteEvent addComplete = eventAggregator.GetEvent<AddJobCompleteEvent>(); 087.if (addToken != null) 088.addComplete.Unsubscribe(addToken); 089.addToken = addComplete.Subscribe(this.AddCompleteEventHandler); 090.} 091.public void EditCompleteEventHandler(bool complete) 092.{ 093.if (complete) 094.{ 095.JobPosting thisJob = _myJobs.CurrentEditItem as JobPosting; 096.this._myJobs.CommitEdit(); 097.this.context.SubmitChanges((s) => 098.{ 099.ActionHistory myAction = new ActionHistory(); 100.myAction.PostingID = thisJob.PostingID; 101.myAction.Description = String.Format("Job '{0}' has been edited by {1}", thisJob.JobTitle, "default user"); 102.myAction.TimeStamp = DateTime.Now; 103.eventAggregator.GetEvent<AddActionEvent>().Publish(myAction); 104.} 105., null); 106.} 107.else 108.{ 109.this._myJobs.CancelEdit(); 110.} 111.this.MakeMeActive(this.regionManager, "MainRegion", "JobPostingsView"); 112.} 113.public void AddCompleteEventHandler(JobPosting job) 114.{ 115.if (job == null) 116.{ 117.// do nothing, new job add cancelled 118.} 119.else 120.{ 121.this.context.JobPostings.Add(job); 122.this.context.SubmitChanges((s) => 123.{ 124.ActionHistory myAction = new ActionHistory(); 125.myAction.PostingID = job.PostingID; 126.myAction.Description = String.Format("Job '{0}' has been added by {1}", job.JobTitle, "default user"); 127.myAction.TimeStamp = DateTime.Now; 128.eventAggregator.GetEvent<AddActionEvent>().Publish(myAction); 129.} 130., null); 131.} 132.this.MakeMeActive(this.regionManager, "MainRegion", "JobPostingsView"); 133.} 134.#endregion 135.public void GrabHistory(int postID) 136.{ 137.context.ActionHistories.Clear(); 138._selectionJobActionHistory = new QueryableCollectionView(context.ActionHistories); 139.context.Load(context.GetHistoryForJobQuery(postID)); 140.} Taking it from the top, we're injecting an Event Aggregator and Region Manager for use down the road and also have the public DelegateCommands (just like in the Menu module). We also grab a reference to our context, which we'll obviously need for data, then set up a few fields with public properties tied to them. We're also setting subscription tokens, which we have not yet seen but I will get into below. The AddNewRecord (50) and EditExistingRecord (54) methods should speak for themselves for functionality, the one thing of note is we're sending events off to the Event Aggregator which some module, somewhere will take care of. Since these aren't entirely relying on one another, the Jobs View doesn't care if anyone is listening, but it will publish AddJobEvent (52), NotifyUserEvent (58) and EditJobEvent (63)regardless. Don't mind the GrabHistory() method so much, that is just grabbing history items (visibly being created in the SubmitChanges callbacks), and adding them to the database. Every action will trigger a history event, so we'll know who modified what and when, just in case. ;) So where are we at? Well, if we click to Add a job, we publish an event, if we edit a job, we publish an event with the selected record (attained through the magic of binding). Where is this all going though? To the Viewmodel, of course! XAML for the AddEditJobView This is pretty straightforward except for one thing, noted below: 001.<Grid x:Name="LayoutRoot" 002.Background="White"> 003.<Grid x:Name="xEditGrid" 004.Margin="10" 005.validationHelper:ValidationScope.Errors="{Binding Errors}"> 006.<Grid.Background> 007.<LinearGradientBrush EndPoint="0.5,1" 008.StartPoint="0.5,0"> 009.<GradientStop Color="#FFC7C7C7" 010.Offset="0" /> 011.<GradientStop Color="#FFF6F3F3" 012.Offset="1" /> 013.</LinearGradientBrush> 014.</Grid.Background> 015.<Grid.RowDefinitions> 016.<RowDefinition Height="40" /> 017.<RowDefinition Height="40" /> 018.<RowDefinition Height="40" /> 019.<RowDefinition Height="100" /> 020.<RowDefinition Height="100" /> 021.<RowDefinition Height="100" /> 022.<RowDefinition Height="40" /> 023.<RowDefinition Height="40" /> 024.<RowDefinition Height="40" /> 025.</Grid.RowDefinitions> 026.<Grid.ColumnDefinitions> 027.<ColumnDefinition Width="150" /> 028.<ColumnDefinition Width="150" /> 029.<ColumnDefinition Width="300" /> 030.<ColumnDefinition Width="100" /> 031.</Grid.ColumnDefinitions> 032.<!-- Title --> 033.<TextBlock Margin="8" 034.Text="{Binding AddEditString}" 035.TextWrapping="Wrap" 036.Grid.Column="1" 037.Grid.ColumnSpan="2" 038.FontSize="16" /> 039.<!-- Data entry area--> 040. 041.<TextBlock Margin="8,0,0,0" 042.Style="{StaticResource LabelTxb}" 043.Grid.Row="1" 044.Text="Job Title" 045.VerticalAlignment="Center" /> 046.<TextBox x:Name="xJobTitleTB" 047.Margin="0,8" 048.Grid.Column="1" 049.Grid.Row="1" 050.Text="{Binding activeJob.JobTitle, Mode=TwoWay, NotifyOnValidationError=True, ValidatesOnExceptions=True}" 051.Grid.ColumnSpan="2" /> 052.<TextBlock Margin="8,0,0,0" 053.Grid.Row="2" 054.Text="Location" 055.d:LayoutOverrides="Height" 056.VerticalAlignment="Center" /> 057.<TextBox x:Name="xLocationTB" 058.Margin="0,8" 059.Grid.Column="1" 060.Grid.Row="2" 061.Text="{Binding activeJob.Location, Mode=TwoWay, NotifyOnValidationError=True, ValidatesOnExceptions=True}" 062.Grid.ColumnSpan="2" /> 063. 064.<TextBlock Margin="8,11,8,0" 065.Grid.Row="3" 066.Text="Description" 067.TextWrapping="Wrap" 068.VerticalAlignment="Top" /> 069. 070.<TextBox x:Name="xDescriptionTB" 071.Height="84" 072.TextWrapping="Wrap" 073.ScrollViewer.VerticalScrollBarVisibility="Auto" 074.Grid.Column="1" 075.Grid.Row="3" 076.Text="{Binding activeJob.Description, Mode=TwoWay, NotifyOnValidationError=True, ValidatesOnExceptions=True}" 077.Grid.ColumnSpan="2" /> 078.<TextBlock Margin="8,11,8,0" 079.Grid.Row="4" 080.Text="Requirements" 081.TextWrapping="Wrap" 082.VerticalAlignment="Top" /> 083. 084.<TextBox x:Name="xRequirementsTB" 085.Height="84" 086.TextWrapping="Wrap" 087.ScrollViewer.VerticalScrollBarVisibility="Auto" 088.Grid.Column="1" 089.Grid.Row="4" 090.Text="{Binding activeJob.Requirements, Mode=TwoWay, NotifyOnValidationError=True, ValidatesOnExceptions=True}" 091.Grid.ColumnSpan="2" /> 092.<TextBlock Margin="8,11,8,0" 093.Grid.Row="5" 094.Text="Qualifications" 095.TextWrapping="Wrap" 096.VerticalAlignment="Top" /> 097. 098.<TextBox x:Name="xQualificationsTB" 099.Height="84" 100.TextWrapping="Wrap" 101.ScrollViewer.VerticalScrollBarVisibility="Auto" 102.Grid.Column="1" 103.Grid.Row="5" 104.Text="{Binding activeJob.Qualifications, Mode=TwoWay, NotifyOnValidationError=True, ValidatesOnExceptions=True}" 105.Grid.ColumnSpan="2" /> 106.<!-- Requirements Checkboxes--> 107. 108.<CheckBox x:Name="xResumeRequiredCB" Margin="8,8,8,15" 109.Content="Resume Required" 110.Grid.Row="6" 111.Grid.ColumnSpan="2" 112.IsChecked="{Binding activeJob.NeedsResume, Mode=TwoWay, NotifyOnValidationError=True, ValidatesOnExceptions=True}"/> 113. 114.<CheckBox x:Name="xCoverletterRequiredCB" Margin="8,8,8,15" 115.Content="Cover Letter Required" 116.Grid.Column="2" 117.Grid.Row="6" 118.IsChecked="{Binding activeJob.NeedsCV, Mode=TwoWay, NotifyOnValidationError=True, ValidatesOnExceptions=True}"/> 119. 120.<CheckBox x:Name="xOverviewRequiredCB" Margin="8,8,8,15" 121.Content="Overview Required" 122.Grid.Row="7" 123.Grid.ColumnSpan="2" 124.IsChecked="{Binding activeJob.NeedsOverview, Mode=TwoWay, NotifyOnValidationError=True, ValidatesOnExceptions=True}"/> 125. 126.<CheckBox x:Name="xJobActiveCB" Margin="8,8,8,15" 127.Content="Job is Active" 128.Grid.Column="2" 129.Grid.Row="7" 130.IsChecked="{Binding activeJob.IsActive, Mode=TwoWay, NotifyOnValidationError=True, ValidatesOnExceptions=True}"/> 131. 132.<!-- Buttons --> 133. 134.<Button x:Name="xAddEditButton" Margin="8,8,0,10" 135.Content="{Binding AddEditButtonString}" 136.cal:Click.Command="{Binding AddEditCommand}" 137.Grid.Column="2" 138.Grid.Row="8" 139.HorizontalAlignment="Left" 140.Width="125" 141.telerik:StyleManager.Theme="Windows7" /> 142. 143.<Button x:Name="xCancelButton" HorizontalAlignment="Right" 144.Content="Cancel" 145.cal:Click.Command="{Binding CancelCommand}" 146.Margin="0,8,8,10" 147.Width="125" 148.Grid.Column="2" 149.Grid.Row="8" 150.telerik:StyleManager.Theme="Windows7" /> 151.</Grid> 152.</Grid> The 'validationHelper:ValidationScope' line may seem odd. This is a handy little trick for catching current and would-be validation errors when working in this whole setup. This all comes from an approach found on theJoy Of Code blog, although it looks like the story for this will be changing slightly with new advances in SL4/WCF RIA Services, so this section can definitely get an overhaul a little down the road. The code is the fun part of all this, so let us see what's happening under the hood. Viewmodel for the AddEditJobView We are going to see some of the same things happening here, so I'll skip over the repeat info and get right to the good stuff: 001.public class AddEditJobViewModel : ViewModelBase 002.{ 003.private readonly IEventAggregator eventAggregator; 004.private readonly IRegionManager regionManager; 005. 006.public RecruitingContext context; 007. 008.private JobPosting _activeJob; 009.public JobPosting activeJob 010.{ 011.get { return _activeJob; } 012.set 013.{ 014.if (_activeJob != value) 015.{ 016._activeJob = value; 017.NotifyChanged("activeJob"); 018.} 019.} 020.} 021. 022.public bool isNewJob; 023. 024.private string _addEditString; 025.public string AddEditString 026.{ 027.get { return _addEditString; } 028.set 029.{ 030.if (_addEditString != value) 031.{ 032._addEditString = value; 033.NotifyChanged("AddEditString"); 034.} 035.} 036.} 037. 038.private string _addEditButtonString; 039.public string AddEditButtonString 040.{ 041.get { return _addEditButtonString; } 042.set 043.{ 044.if (_addEditButtonString != value) 045.{ 046._addEditButtonString = value; 047.NotifyChanged("AddEditButtonString"); 048.} 049.} 050.} 051. 052.public SubscriptionToken addJobToken = new SubscriptionToken(); 053.public SubscriptionToken editJobToken = new SubscriptionToken(); 054. 055.public DelegateCommand<object> AddEditCommand { get; set; } 056.public DelegateCommand<object> CancelCommand { get; set; } 057. 058.private ObservableCollection<ValidationError> _errors = new ObservableCollection<ValidationError>(); 059.public ObservableCollection<ValidationError> Errors 060.{ 061.get { return _errors; } 062.} 063. 064.private ObservableCollection<ValidationResult> _valResults = new ObservableCollection<ValidationResult>(); 065.public ObservableCollection<ValidationResult> ValResults 066.{ 067.get { return this._valResults; } 068.} 069. 070.public AddEditJobViewModel(IEventAggregator eventAgg, IRegionManager regionmanager) 071.{ 072.// set Unity items 073.this.eventAggregator = eventAgg; 074.this.regionManager = regionmanager; 075. 076.context = new RecruitingContext(); 077. 078.AddEditCommand = new DelegateCommand<object>(this.AddEditJobCommand); 079.CancelCommand = new DelegateCommand<object>(this.CancelAddEditCommand); 080. 081.SetSubscriptions(); 082.} 083. 084.#region Subscription Declaration and Events 085. 086.public void SetSubscriptions() 087.{ 088.AddJobEvent addJob = this.eventAggregator.GetEvent<AddJobEvent>(); 089. 090.if (addJobToken != null) 091.addJob.Unsubscribe(addJobToken); 092. 093.addJobToken = addJob.Subscribe(this.AddJobEventHandler); 094. 095.EditJobEvent editJob = this.eventAggregator.GetEvent<EditJobEvent>(); 096. 097.if (editJobToken != null) 098.editJob.Unsubscribe(editJobToken); 099. 100.editJobToken = editJob.Subscribe(this.EditJobEventHandler); 101.} 102. 103.public void AddJobEventHandler(bool isNew) 104.{ 105.this.activeJob = null; 106.this.activeJob = new JobPosting(); 107.this.activeJob.IsActive = true; // We assume that we want a new job to go up immediately 108.this.isNewJob = true; 109.this.AddEditString = "Add New Job Posting"; 110.this.AddEditButtonString = "Add Job"; 111. 112.MakeMeActive(this.regionManager, "MainRegion", "AddEditJobView"); 113.} 114. 115.public void EditJobEventHandler(JobPosting editJob) 116.{ 117.this.activeJob = null; 118.this.activeJob = editJob; 119.this.isNewJob = false; 120.this.AddEditString = "Edit Job Posting"; 121.this.AddEditButtonString = "Edit Job"; 122. 123.MakeMeActive(this.regionManager, "MainRegion", "AddEditJobView"); 124.} 125. 126.#endregion 127. 128.#region DelegateCommands from View 129. 130.public void AddEditJobCommand(object obj) 131.{ 132.if (this.Errors.Count > 0) 133.{ 134.List<string> errorMessages = new List<string>(); 135. 136.foreach (var valR in this.Errors) 137.{ 138.errorMessages.Add(valR.Exception.Message); 139.} 140. 141.this.eventAggregator.GetEvent<DisplayValidationErrorsEvent>().Publish(errorMessages); 142. 143.} 144.else if (!Validator.TryValidateObject(this.activeJob, new ValidationContext(this.activeJob, null, null), _valResults, true)) 145.{ 146.List<string> errorMessages = new List<string>(); 147. 148.foreach (var valR in this._valResults) 149.{ 150.errorMessages.Add(valR.ErrorMessage); 151.} 152. 153.this._valResults.Clear(); 154. 155.this.eventAggregator.GetEvent<DisplayValidationErrorsEvent>().Publish(errorMessages); 156.} 157.else 158.{ 159.if (this.isNewJob) 160.{ 161.this.eventAggregator.GetEvent<AddJobCompleteEvent>().Publish(this.activeJob); 162.} 163.else 164.{ 165.this.eventAggregator.GetEvent<EditJobCompleteEvent>().Publish(true); 166.} 167.} 168.} 169. 170.public void CancelAddEditCommand(object obj) 171.{ 172.if (this.isNewJob) 173.{ 174.this.eventAggregator.GetEvent<AddJobCompleteEvent>().Publish(null); 175.} 176.else 177.{ 178.this.eventAggregator.GetEvent<EditJobCompleteEvent>().Publish(false); 179.} 180.} 181. 182.#endregion 183.} 184.} We start seeing something new on line 103- the AddJobEventHandler will create a new job and set that to the activeJob item on the ViewModel. When this is all set, the view calls that familiar MakeMeActive method to activate itself. I made a bit of a management call on making views self-activate like this, but I figured it works for one reason. As I create this application, views may not exist that I have in mind, so after a view receives its 'ping' from being subscribed to an event, it prepares whatever it needs to do and then goes active. This way if I don't have 'edit' hooked up, I can click as the day is long on the main view and won't get lost in an empty region. Total personal preference here. :) Everything else should again be pretty straightforward, although I do a bit of validation checking in the AddEditJobCommand, which can either fire off an event back to the main view/viewmodel if everything is a success or sent a list of errors to our notification module, which pops open a RadWindow with the alerts if any exist. As a bonus side note, here's what my WCF RIA Services metadata looks like for handling all of the validation: private JobPostingMetadata() { } [StringLength(2500, ErrorMessage = "Description should be more than one and less than 2500 characters.", MinimumLength = 1)] [Required(ErrorMessage = "Description is required.")] public string Description; [Required(ErrorMessage="Active Status is Required")] public bool IsActive; [StringLength(100, ErrorMessage = "Posting title must be more than 3 but less than 100 characters.", MinimumLength = 3)] [Required(ErrorMessage = "Job Title is required.")] public bool JobTitle; [Required] public string Location; public bool NeedsCV; public bool NeedsOverview; public bool NeedsResume; public int PostingID; [Required(ErrorMessage="Qualifications are required.")] [StringLength(2500, ErrorMessage="Qualifications should be more than one and less than 2500 characters.", MinimumLength=1)] public string Qualifications; [StringLength(2500, ErrorMessage = "Requirements should be more than one and less than 2500 characters.", MinimumLength = 1)] [Required(ErrorMessage="Requirements are required.")] public string Requirements;   The RecruitCB Alternative See all that Xaml I pasted above? Those are now two pieces sitting in the JobsView.xaml file now. The only real difference is that the xEditGrid now sits in the same place as xJobsGrid, with visibility swapping out between the two for a quick switch. I also took out all the cal: and command: command references and replaced Button events with clicks and the Grid selection command replaced with a SelectedItemChanged event. Also, at the bottom of the xEditGrid after the last button, I add a ValidationSummary (with Visibility=Collapsed) to catch any errors that are popping up. Simple as can be, and leads to this being the single code-behind file: 001.public partial class JobsView : UserControl 002.{ 003.public RecruitingContext context; 004.public JobPosting activeJob; 005.public bool isNew; 006.private ObservableCollection<ValidationResult> _valResults = new ObservableCollection<ValidationResult>(); 007.public ObservableCollection<ValidationResult> ValResults 008.{ 009.get { return this._valResults; } 010.} 011.public JobsView() 012.{ 013.InitializeComponent(); 014.this.Loaded += new RoutedEventHandler(JobsView_Loaded); 015.} 016.void JobsView_Loaded(object sender, RoutedEventArgs e) 017.{ 018.context = new RecruitingContext(); 019.xJobsGrid.ItemsSource = context.JobPostings; 020.context.Load(context.GetJobPostingsQuery()); 021.} 022.private void xAddRecordButton_Click(object sender, RoutedEventArgs e) 023.{ 024.activeJob = new JobPosting(); 025.isNew = true; 026.xAddEditTitle.Text = "Add a Job Posting"; 027.xAddEditButton.Content = "Add"; 028.xEditGrid.DataContext = activeJob; 029.HideJobsGrid(); 030.} 031.private void xEditRecordButton_Click(object sender, RoutedEventArgs e) 032.{ 033.activeJob = xJobsGrid.SelectedItem as JobPosting; 034.isNew = false; 035.xAddEditTitle.Text = "Edit a Job Posting"; 036.xAddEditButton.Content = "Edit"; 037.xEditGrid.DataContext = activeJob; 038.HideJobsGrid(); 039.} 040.private void xAddEditButton_Click(object sender, RoutedEventArgs e) 041.{ 042.if (!Validator.TryValidateObject(this.activeJob, new ValidationContext(this.activeJob, null, null), _valResults, true)) 043.{ 044.List<string> errorMessages = new List<string>(); 045.foreach (var valR in this._valResults) 046.{ 047.errorMessages.Add(valR.ErrorMessage); 048.} 049.this._valResults.Clear(); 050.ShowErrors(errorMessages); 051.} 052.else if (xSummary.Errors.Count > 0) 053.{ 054.List<string> errorMessages = new List<string>(); 055.foreach (var err in xSummary.Errors) 056.{ 057.errorMessages.Add(err.Message); 058.} 059.ShowErrors(errorMessages); 060.} 061.else 062.{ 063.if (this.isNew) 064.{ 065.context.JobPostings.Add(activeJob); 066.context.SubmitChanges((s) => 067.{ 068.ActionHistory thisAction = new ActionHistory(); 069.thisAction.PostingID = activeJob.PostingID; 070.thisAction.Description = String.Format("Job '{0}' has been edited by {1}", activeJob.JobTitle, "default user"); 071.thisAction.TimeStamp = DateTime.Now; 072.context.ActionHistories.Add(thisAction); 073.context.SubmitChanges(); 074.}, null); 075.} 076.else 077.{ 078.context.SubmitChanges((s) => 079.{ 080.ActionHistory thisAction = new ActionHistory(); 081.thisAction.PostingID = activeJob.PostingID; 082.thisAction.Description = String.Format("Job '{0}' has been added by {1}", activeJob.JobTitle, "default user"); 083.thisAction.TimeStamp = DateTime.Now; 084.context.ActionHistories.Add(thisAction); 085.context.SubmitChanges(); 086.}, null); 087.} 088.ShowJobsGrid(); 089.} 090.} 091.private void xCancelButton_Click(object sender, RoutedEventArgs e) 092.{ 093.ShowJobsGrid(); 094.} 095.private void ShowJobsGrid() 096.{ 097.xAddEditRecordButtonPanel.Visibility = Visibility.Visible; 098.xEditGrid.Visibility = Visibility.Collapsed; 099.xJobsGrid.Visibility = Visibility.Visible; 100.} 101.private void HideJobsGrid() 102.{ 103.xAddEditRecordButtonPanel.Visibility = Visibility.Collapsed; 104.xJobsGrid.Visibility = Visibility.Collapsed; 105.xEditGrid.Visibility = Visibility.Visible; 106.} 107.private void ShowErrors(List<string> errorList) 108.{ 109.string nm = "Errors received: \n"; 110.foreach (string anerror in errorList) 111.nm += anerror + "\n"; 112.RadWindow.Alert(nm); 113.} 114.} The first 39 lines should be pretty familiar, not doing anything too unorthodox to get this up and running. Once we hit the xAddEditButton_Click on line 40, we're still doing pretty much the same things except instead of checking the ValidationHelper errors, we both run a check on the current activeJob object as well as check the ValidationSummary errors list. Once that is set, we again use the callback of context.SubmitChanges (lines 68 and 78) to create an ActionHistory which we will use to track these items down the line. That's all? Essentially... yes. If you look back through this post, most of the code and adventures we have taken were just to get things working in the MVVM/Prism setup. Since I have the whole 'module' self-contained in a single JobView+code-behind setup, I don't have to worry about things like sending events off into space for someone to pick up, communicating through an Infrastructure project, or even re-inventing events to be used with attached behaviors. Everything just kinda works, and again with much less code. Here's a picture of the MVVM and Code-behind versions on the Jobs and AddEdit views, but since the functionality is the same in both apps you still cannot tell them apart (for two-strike): Looking ahead, the Applicants module is effectively the same thing as the Jobs module, so most of the code is being cut-and-pasted back and forth with minor tweaks here and there. So that one is being taken care of by me behind the scenes. Next time, we get into a new world of fun- the interview scheduling module, which will pull from available jobs and applicants for each interview being scheduled, tying everything together with RadScheduler to the rescue. Did you know that DotNetSlackers also publishes .net articles written by top known .net Authors? We already have over 80 articles in several categories including Silverlight. Take a look: here.

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  • value types in the vm

    - by john.rose
    value types in the vm p.p1 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Times} p.p2 {margin: 0.0px 0.0px 14.0px 0.0px; font: 14.0px Times} p.p3 {margin: 0.0px 0.0px 12.0px 0.0px; font: 14.0px Times} p.p4 {margin: 0.0px 0.0px 15.0px 0.0px; font: 14.0px Times} p.p5 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Courier} p.p6 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Courier; min-height: 17.0px} p.p7 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Times; min-height: 18.0px} p.p8 {margin: 0.0px 0.0px 0.0px 36.0px; text-indent: -36.0px; font: 14.0px Times; min-height: 18.0px} p.p9 {margin: 0.0px 0.0px 12.0px 0.0px; font: 14.0px Times; min-height: 18.0px} p.p10 {margin: 0.0px 0.0px 12.0px 0.0px; font: 14.0px Times; color: #000000} li.li1 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Times} li.li7 {margin: 0.0px 0.0px 0.0px 0.0px; font: 14.0px Times; min-height: 18.0px} span.s1 {font: 14.0px Courier} span.s2 {color: #000000} span.s3 {font: 14.0px Courier; color: #000000} ol.ol1 {list-style-type: decimal} Or, enduring values for a changing world. Introduction A value type is a data type which, generally speaking, is designed for being passed by value in and out of methods, and stored by value in data structures. The only value types which the Java language directly supports are the eight primitive types. Java indirectly and approximately supports value types, if they are implemented in terms of classes. For example, both Integer and String may be viewed as value types, especially if their usage is restricted to avoid operations appropriate to Object. In this note, we propose a definition of value types in terms of a design pattern for Java classes, accompanied by a set of usage restrictions. We also sketch the relation of such value types to tuple types (which are a JVM-level notion), and point out JVM optimizations that can apply to value types. This note is a thought experiment to extend the JVM’s performance model in support of value types. The demonstration has two phases.  Initially the extension can simply use design patterns, within the current bytecode architecture, and in today’s Java language. But if the performance model is to be realized in practice, it will probably require new JVM bytecode features, changes to the Java language, or both.  We will look at a few possibilities for these new features. An Axiom of Value In the context of the JVM, a value type is a data type equipped with construction, assignment, and equality operations, and a set of typed components, such that, whenever two variables of the value type produce equal corresponding values for their components, the values of the two variables cannot be distinguished by any JVM operation. Here are some corollaries: A value type is immutable, since otherwise a copy could be constructed and the original could be modified in one of its components, allowing the copies to be distinguished. Changing the component of a value type requires construction of a new value. The equals and hashCode operations are strictly component-wise. If a value type is represented by a JVM reference, that reference cannot be successfully synchronized on, and cannot be usefully compared for reference equality. A value type can be viewed in terms of what it doesn’t do. We can say that a value type omits all value-unsafe operations, which could violate the constraints on value types.  These operations, which are ordinarily allowed for Java object types, are pointer equality comparison (the acmp instruction), synchronization (the monitor instructions), all the wait and notify methods of class Object, and non-trivial finalize methods. The clone method is also value-unsafe, although for value types it could be treated as the identity function. Finally, and most importantly, any side effect on an object (however visible) also counts as an value-unsafe operation. A value type may have methods, but such methods must not change the components of the value. It is reasonable and useful to define methods like toString, equals, and hashCode on value types, and also methods which are specifically valuable to users of the value type. Representations of Value Value types have two natural representations in the JVM, unboxed and boxed. An unboxed value consists of the components, as simple variables. For example, the complex number x=(1+2i), in rectangular coordinate form, may be represented in unboxed form by the following pair of variables: /*Complex x = Complex.valueOf(1.0, 2.0):*/ double x_re = 1.0, x_im = 2.0; These variables might be locals, parameters, or fields. Their association as components of a single value is not defined to the JVM. Here is a sample computation which computes the norm of the difference between two complex numbers: double distance(/*Complex x:*/ double x_re, double x_im,         /*Complex y:*/ double y_re, double y_im) {     /*Complex z = x.minus(y):*/     double z_re = x_re - y_re, z_im = x_im - y_im;     /*return z.abs():*/     return Math.sqrt(z_re*z_re + z_im*z_im); } A boxed representation groups component values under a single object reference. The reference is to a ‘wrapper class’ that carries the component values in its fields. (A primitive type can naturally be equated with a trivial value type with just one component of that type. In that view, the wrapper class Integer can serve as a boxed representation of value type int.) The unboxed representation of complex numbers is practical for many uses, but it fails to cover several major use cases: return values, array elements, and generic APIs. The two components of a complex number cannot be directly returned from a Java function, since Java does not support multiple return values. The same story applies to array elements: Java has no ’array of structs’ feature. (Double-length arrays are a possible workaround for complex numbers, but not for value types with heterogeneous components.) By generic APIs I mean both those which use generic types, like Arrays.asList and those which have special case support for primitive types, like String.valueOf and PrintStream.println. Those APIs do not support unboxed values, and offer some problems to boxed values. Any ’real’ JVM type should have a story for returns, arrays, and API interoperability. The basic problem here is that value types fall between primitive types and object types. Value types are clearly more complex than primitive types, and object types are slightly too complicated. Objects are a little bit dangerous to use as value carriers, since object references can be compared for pointer equality, and can be synchronized on. Also, as many Java programmers have observed, there is often a performance cost to using wrapper objects, even on modern JVMs. Even so, wrapper classes are a good starting point for talking about value types. If there were a set of structural rules and restrictions which would prevent value-unsafe operations on value types, wrapper classes would provide a good notation for defining value types. This note attempts to define such rules and restrictions. Let’s Start Coding Now it is time to look at some real code. Here is a definition, written in Java, of a complex number value type. @ValueSafe public final class Complex implements java.io.Serializable {     // immutable component structure:     public final double re, im;     private Complex(double re, double im) {         this.re = re; this.im = im;     }     // interoperability methods:     public String toString() { return "Complex("+re+","+im+")"; }     public List<Double> asList() { return Arrays.asList(re, im); }     public boolean equals(Complex c) {         return re == c.re && im == c.im;     }     public boolean equals(@ValueSafe Object x) {         return x instanceof Complex && equals((Complex) x);     }     public int hashCode() {         return 31*Double.valueOf(re).hashCode()                 + Double.valueOf(im).hashCode();     }     // factory methods:     public static Complex valueOf(double re, double im) {         return new Complex(re, im);     }     public Complex changeRe(double re2) { return valueOf(re2, im); }     public Complex changeIm(double im2) { return valueOf(re, im2); }     public static Complex cast(@ValueSafe Object x) {         return x == null ? ZERO : (Complex) x;     }     // utility methods and constants:     public Complex plus(Complex c)  { return new Complex(re+c.re, im+c.im); }     public Complex minus(Complex c) { return new Complex(re-c.re, im-c.im); }     public double abs() { return Math.sqrt(re*re + im*im); }     public static final Complex PI = valueOf(Math.PI, 0.0);     public static final Complex ZERO = valueOf(0.0, 0.0); } This is not a minimal definition, because it includes some utility methods and other optional parts.  The essential elements are as follows: The class is marked as a value type with an annotation. The class is final, because it does not make sense to create subclasses of value types. The fields of the class are all non-private and final.  (I.e., the type is immutable and structurally transparent.) From the supertype Object, all public non-final methods are overridden. The constructor is private. Beyond these bare essentials, we can observe the following features in this example, which are likely to be typical of all value types: One or more factory methods are responsible for value creation, including a component-wise valueOf method. There are utility methods for complex arithmetic and instance creation, such as plus and changeIm. There are static utility constants, such as PI. The type is serializable, using the default mechanisms. There are methods for converting to and from dynamically typed references, such as asList and cast. The Rules In order to use value types properly, the programmer must avoid value-unsafe operations.  A helpful Java compiler should issue errors (or at least warnings) for code which provably applies value-unsafe operations, and should issue warnings for code which might be correct but does not provably avoid value-unsafe operations.  No such compilers exist today, but to simplify our account here, we will pretend that they do exist. A value-safe type is any class, interface, or type parameter marked with the @ValueSafe annotation, or any subtype of a value-safe type.  If a value-safe class is marked final, it is in fact a value type.  All other value-safe classes must be abstract.  The non-static fields of a value class must be non-public and final, and all its constructors must be private. Under the above rules, a standard interface could be helpful to define value types like Complex.  Here is an example: @ValueSafe public interface ValueType extends java.io.Serializable {     // All methods listed here must get redefined.     // Definitions must be value-safe, which means     // they may depend on component values only.     List<? extends Object> asList();     int hashCode();     boolean equals(@ValueSafe Object c);     String toString(); } //@ValueSafe inherited from supertype: public final class Complex implements ValueType { … The main advantage of such a conventional interface is that (unlike an annotation) it is reified in the runtime type system.  It could appear as an element type or parameter bound, for facilities which are designed to work on value types only.  More broadly, it might assist the JVM to perform dynamic enforcement of the rules for value types. Besides types, the annotation @ValueSafe can mark fields, parameters, local variables, and methods.  (This is redundant when the type is also value-safe, but may be useful when the type is Object or another supertype of a value type.)  Working forward from these annotations, an expression E is defined as value-safe if it satisfies one or more of the following: The type of E is a value-safe type. E names a field, parameter, or local variable whose declaration is marked @ValueSafe. E is a call to a method whose declaration is marked @ValueSafe. E is an assignment to a value-safe variable, field reference, or array reference. E is a cast to a value-safe type from a value-safe expression. E is a conditional expression E0 ? E1 : E2, and both E1 and E2 are value-safe. Assignments to value-safe expressions and initializations of value-safe names must take their values from value-safe expressions. A value-safe expression may not be the subject of a value-unsafe operation.  In particular, it cannot be synchronized on, nor can it be compared with the “==” operator, not even with a null or with another value-safe type. In a program where all of these rules are followed, no value-type value will be subject to a value-unsafe operation.  Thus, the prime axiom of value types will be satisfied, that no two value type will be distinguishable as long as their component values are equal. More Code To illustrate these rules, here are some usage examples for Complex: Complex pi = Complex.valueOf(Math.PI, 0); Complex zero = pi.changeRe(0);  //zero = pi; zero.re = 0; ValueType vtype = pi; @SuppressWarnings("value-unsafe")   Object obj = pi; @ValueSafe Object obj2 = pi; obj2 = new Object();  // ok List<Complex> clist = new ArrayList<Complex>(); clist.add(pi);  // (ok assuming List.add param is @ValueSafe) List<ValueType> vlist = new ArrayList<ValueType>(); vlist.add(pi);  // (ok) List<Object> olist = new ArrayList<Object>(); olist.add(pi);  // warning: "value-unsafe" boolean z = pi.equals(zero); boolean z1 = (pi == zero);  // error: reference comparison on value type boolean z2 = (pi == null);  // error: reference comparison on value type boolean z3 = (pi == obj2);  // error: reference comparison on value type synchronized (pi) { }  // error: synch of value, unpredictable result synchronized (obj2) { }  // unpredictable result Complex qq = pi; qq = null;  // possible NPE; warning: “null-unsafe" qq = (Complex) obj;  // warning: “null-unsafe" qq = Complex.cast(obj);  // OK @SuppressWarnings("null-unsafe")   Complex empty = null;  // possible NPE qq = empty;  // possible NPE (null pollution) The Payoffs It follows from this that either the JVM or the java compiler can replace boxed value-type values with unboxed ones, without affecting normal computations.  Fields and variables of value types can be split into their unboxed components.  Non-static methods on value types can be transformed into static methods which take the components as value parameters. Some common questions arise around this point in any discussion of value types. Why burden the programmer with all these extra rules?  Why not detect programs automagically and perform unboxing transparently?  The answer is that it is easy to break the rules accidently unless they are agreed to by the programmer and enforced.  Automatic unboxing optimizations are tantalizing but (so far) unreachable ideal.  In the current state of the art, it is possible exhibit benchmarks in which automatic unboxing provides the desired effects, but it is not possible to provide a JVM with a performance model that assures the programmer when unboxing will occur.  This is why I’m writing this note, to enlist help from, and provide assurances to, the programmer.  Basically, I’m shooting for a good set of user-supplied “pragmas” to frame the desired optimization. Again, the important thing is that the unboxing must be done reliably, or else programmers will have no reason to work with the extra complexity of the value-safety rules.  There must be a reasonably stable performance model, wherein using a value type has approximately the same performance characteristics as writing the unboxed components as separate Java variables. There are some rough corners to the present scheme.  Since Java fields and array elements are initialized to null, value-type computations which incorporate uninitialized variables can produce null pointer exceptions.  One workaround for this is to require such variables to be null-tested, and the result replaced with a suitable all-zero value of the value type.  That is what the “cast” method does above. Generically typed APIs like List<T> will continue to manipulate boxed values always, at least until we figure out how to do reification of generic type instances.  Use of such APIs will elicit warnings until their type parameters (and/or relevant members) are annotated or typed as value-safe.  Retrofitting List<T> is likely to expose flaws in the present scheme, which we will need to engineer around.  Here are a couple of first approaches: public interface java.util.List<@ValueSafe T> extends Collection<T> { … public interface java.util.List<T extends Object|ValueType> extends Collection<T> { … (The second approach would require disjunctive types, in which value-safety is “contagious” from the constituent types.) With more transformations, the return value types of methods can also be unboxed.  This may require significant bytecode-level transformations, and would work best in the presence of a bytecode representation for multiple value groups, which I have proposed elsewhere under the title “Tuples in the VM”. But for starters, the JVM can apply this transformation under the covers, to internally compiled methods.  This would give a way to express multiple return values and structured return values, which is a significant pain-point for Java programmers, especially those who work with low-level structure types favored by modern vector and graphics processors.  The lack of multiple return values has a strong distorting effect on many Java APIs. Even if the JVM fails to unbox a value, there is still potential benefit to the value type.  Clustered computing systems something have copy operations (serialization or something similar) which apply implicitly to command operands.  When copying JVM objects, it is extremely helpful to know when an object’s identity is important or not.  If an object reference is a copied operand, the system may have to create a proxy handle which points back to the original object, so that side effects are visible.  Proxies must be managed carefully, and this can be expensive.  On the other hand, value types are exactly those types which a JVM can “copy and forget” with no downside. Array types are crucial to bulk data interfaces.  (As data sizes and rates increase, bulk data becomes more important than scalar data, so arrays are definitely accompanying us into the future of computing.)  Value types are very helpful for adding structure to bulk data, so a successful value type mechanism will make it easier for us to express richer forms of bulk data. Unboxing arrays (i.e., arrays containing unboxed values) will provide better cache and memory density, and more direct data movement within clustered or heterogeneous computing systems.  They require the deepest transformations, relative to today’s JVM.  There is an impedance mismatch between value-type arrays and Java’s covariant array typing, so compromises will need to be struck with existing Java semantics.  It is probably worth the effort, since arrays of unboxed value types are inherently more memory-efficient than standard Java arrays, which rely on dependent pointer chains. It may be sufficient to extend the “value-safe” concept to array declarations, and allow low-level transformations to change value-safe array declarations from the standard boxed form into an unboxed tuple-based form.  Such value-safe arrays would not be convertible to Object[] arrays.  Certain connection points, such as Arrays.copyOf and System.arraycopy might need additional input/output combinations, to allow smooth conversion between arrays with boxed and unboxed elements. Alternatively, the correct solution may have to wait until we have enough reification of generic types, and enough operator overloading, to enable an overhaul of Java arrays. Implicit Method Definitions The example of class Complex above may be unattractively complex.  I believe most or all of the elements of the example class are required by the logic of value types. If this is true, a programmer who writes a value type will have to write lots of error-prone boilerplate code.  On the other hand, I think nearly all of the code (except for the domain-specific parts like plus and minus) can be implicitly generated. Java has a rule for implicitly defining a class’s constructor, if no it defines no constructors explicitly.  Likewise, there are rules for providing default access modifiers for interface members.  Because of the highly regular structure of value types, it might be reasonable to perform similar implicit transformations on value types.  Here’s an example of a “highly implicit” definition of a complex number type: public class Complex implements ValueType {  // implicitly final     public double re, im;  // implicitly public final     //implicit methods are defined elementwise from te fields:     //  toString, asList, equals(2), hashCode, valueOf, cast     //optionally, explicit methods (plus, abs, etc.) would go here } In other words, with the right defaults, a simple value type definition can be a one-liner.  The observant reader will have noticed the similarities (and suitable differences) between the explicit methods above and the corresponding methods for List<T>. Another way to abbreviate such a class would be to make an annotation the primary trigger of the functionality, and to add the interface(s) implicitly: public @ValueType class Complex { … // implicitly final, implements ValueType (But to me it seems better to communicate the “magic” via an interface, even if it is rooted in an annotation.) Implicitly Defined Value Types So far we have been working with nominal value types, which is to say that the sequence of typed components is associated with a name and additional methods that convey the intention of the programmer.  A simple ordered pair of floating point numbers can be variously interpreted as (to name a few possibilities) a rectangular or polar complex number or Cartesian point.  The name and the methods convey the intended meaning. But what if we need a truly simple ordered pair of floating point numbers, without any further conceptual baggage?  Perhaps we are writing a method (like “divideAndRemainder”) which naturally returns a pair of numbers instead of a single number.  Wrapping the pair of numbers in a nominal type (like “QuotientAndRemainder”) makes as little sense as wrapping a single return value in a nominal type (like “Quotient”).  What we need here are structural value types commonly known as tuples. For the present discussion, let us assign a conventional, JVM-friendly name to tuples, roughly as follows: public class java.lang.tuple.$DD extends java.lang.tuple.Tuple {      double $1, $2; } Here the component names are fixed and all the required methods are defined implicitly.  The supertype is an abstract class which has suitable shared declarations.  The name itself mentions a JVM-style method parameter descriptor, which may be “cracked” to determine the number and types of the component fields. The odd thing about such a tuple type (and structural types in general) is it must be instantiated lazily, in response to linkage requests from one or more classes that need it.  The JVM and/or its class loaders must be prepared to spin a tuple type on demand, given a simple name reference, $xyz, where the xyz is cracked into a series of component types.  (Specifics of naming and name mangling need some tasteful engineering.) Tuples also seem to demand, even more than nominal types, some support from the language.  (This is probably because notations for non-nominal types work best as combinations of punctuation and type names, rather than named constructors like Function3 or Tuple2.)  At a minimum, languages with tuples usually (I think) have some sort of simple bracket notation for creating tuples, and a corresponding pattern-matching syntax (or “destructuring bind”) for taking tuples apart, at least when they are parameter lists.  Designing such a syntax is no simple thing, because it ought to play well with nominal value types, and also with pre-existing Java features, such as method parameter lists, implicit conversions, generic types, and reflection.  That is a task for another day. Other Use Cases Besides complex numbers and simple tuples there are many use cases for value types.  Many tuple-like types have natural value-type representations. These include rational numbers, point locations and pixel colors, and various kinds of dates and addresses. Other types have a variable-length ‘tail’ of internal values. The most common example of this is String, which is (mathematically) a sequence of UTF-16 character values. Similarly, bit vectors, multiple-precision numbers, and polynomials are composed of sequences of values. Such types include, in their representation, a reference to a variable-sized data structure (often an array) which (somehow) represents the sequence of values. The value type may also include ’header’ information. Variable-sized values often have a length distribution which favors short lengths. In that case, the design of the value type can make the first few values in the sequence be direct ’header’ fields of the value type. In the common case where the header is enough to represent the whole value, the tail can be a shared null value, or even just a null reference. Note that the tail need not be an immutable object, as long as the header type encapsulates it well enough. This is the case with String, where the tail is a mutable (but never mutated) character array. Field types and their order must be a globally visible part of the API.  The structure of the value type must be transparent enough to have a globally consistent unboxed representation, so that all callers and callees agree about the type and order of components  that appear as parameters, return types, and array elements.  This is a trade-off between efficiency and encapsulation, which is forced on us when we remove an indirection enjoyed by boxed representations.  A JVM-only transformation would not care about such visibility, but a bytecode transformation would need to take care that (say) the components of complex numbers would not get swapped after a redefinition of Complex and a partial recompile.  Perhaps constant pool references to value types need to declare the field order as assumed by each API user. This brings up the delicate status of private fields in a value type.  It must always be possible to load, store, and copy value types as coordinated groups, and the JVM performs those movements by moving individual scalar values between locals and stack.  If a component field is not public, what is to prevent hostile code from plucking it out of the tuple using a rogue aload or astore instruction?  Nothing but the verifier, so we may need to give it more smarts, so that it treats value types as inseparable groups of stack slots or locals (something like long or double). My initial thought was to make the fields always public, which would make the security problem moot.  But public is not always the right answer; consider the case of String, where the underlying mutable character array must be encapsulated to prevent security holes.  I believe we can win back both sides of the tradeoff, by training the verifier never to split up the components in an unboxed value.  Just as the verifier encapsulates the two halves of a 64-bit primitive, it can encapsulate the the header and body of an unboxed String, so that no code other than that of class String itself can take apart the values. Similar to String, we could build an efficient multi-precision decimal type along these lines: public final class DecimalValue extends ValueType {     protected final long header;     protected private final BigInteger digits;     public DecimalValue valueOf(int value, int scale) {         assert(scale >= 0);         return new DecimalValue(((long)value << 32) + scale, null);     }     public DecimalValue valueOf(long value, int scale) {         if (value == (int) value)             return valueOf((int)value, scale);         return new DecimalValue(-scale, new BigInteger(value));     } } Values of this type would be passed between methods as two machine words. Small values (those with a significand which fits into 32 bits) would be represented without any heap data at all, unless the DecimalValue itself were boxed. (Note the tension between encapsulation and unboxing in this case.  It would be better if the header and digits fields were private, but depending on where the unboxing information must “leak”, it is probably safer to make a public revelation of the internal structure.) Note that, although an array of Complex can be faked with a double-length array of double, there is no easy way to fake an array of unboxed DecimalValues.  (Either an array of boxed values or a transposed pair of homogeneous arrays would be reasonable fallbacks, in a current JVM.)  Getting the full benefit of unboxing and arrays will require some new JVM magic. Although the JVM emphasizes portability, system dependent code will benefit from using machine-level types larger than 64 bits.  For example, the back end of a linear algebra package might benefit from value types like Float4 which map to stock vector types.  This is probably only worthwhile if the unboxing arrays can be packed with such values. More Daydreams A more finely-divided design for dynamic enforcement of value safety could feature separate marker interfaces for each invariant.  An empty marker interface Unsynchronizable could cause suitable exceptions for monitor instructions on objects in marked classes.  More radically, a Interchangeable marker interface could cause JVM primitives that are sensitive to object identity to raise exceptions; the strangest result would be that the acmp instruction would have to be specified as raising an exception. @ValueSafe public interface ValueType extends java.io.Serializable,         Unsynchronizable, Interchangeable { … public class Complex implements ValueType {     // inherits Serializable, Unsynchronizable, Interchangeable, @ValueSafe     … It seems possible that Integer and the other wrapper types could be retro-fitted as value-safe types.  This is a major change, since wrapper objects would be unsynchronizable and their references interchangeable.  It is likely that code which violates value-safety for wrapper types exists but is uncommon.  It is less plausible to retro-fit String, since the prominent operation String.intern is often used with value-unsafe code. We should also reconsider the distinction between boxed and unboxed values in code.  The design presented above obscures that distinction.  As another thought experiment, we could imagine making a first class distinction in the type system between boxed and unboxed representations.  Since only primitive types are named with a lower-case initial letter, we could define that the capitalized version of a value type name always refers to the boxed representation, while the initial lower-case variant always refers to boxed.  For example: complex pi = complex.valueOf(Math.PI, 0); Complex boxPi = pi;  // convert to boxed myList.add(boxPi); complex z = myList.get(0);  // unbox Such a convention could perhaps absorb the current difference between int and Integer, double and Double. It might also allow the programmer to express a helpful distinction among array types. As said above, array types are crucial to bulk data interfaces, but are limited in the JVM.  Extending arrays beyond the present limitations is worth thinking about; for example, the Maxine JVM implementation has a hybrid object/array type.  Something like this which can also accommodate value type components seems worthwhile.  On the other hand, does it make sense for value types to contain short arrays?  And why should random-access arrays be the end of our design process, when bulk data is often sequentially accessed, and it might make sense to have heterogeneous streams of data as the natural “jumbo” data structure.  These considerations must wait for another day and another note. More Work It seems to me that a good sequence for introducing such value types would be as follows: Add the value-safety restrictions to an experimental version of javac. Code some sample applications with value types, including Complex and DecimalValue. Create an experimental JVM which internally unboxes value types but does not require new bytecodes to do so.  Ensure the feasibility of the performance model for the sample applications. Add tuple-like bytecodes (with or without generic type reification) to a major revision of the JVM, and teach the Java compiler to switch in the new bytecodes without code changes. A staggered roll-out like this would decouple language changes from bytecode changes, which is always a convenient thing. A similar investigation should be applied (concurrently) to array types.  In this case, it seems to me that the starting point is in the JVM: Add an experimental unboxing array data structure to a production JVM, perhaps along the lines of Maxine hybrids.  No bytecode or language support is required at first; everything can be done with encapsulated unsafe operations and/or method handles. Create an experimental JVM which internally unboxes value types but does not require new bytecodes to do so.  Ensure the feasibility of the performance model for the sample applications. Add tuple-like bytecodes (with or without generic type reification) to a major revision of the JVM, and teach the Java compiler to switch in the new bytecodes without code changes. That’s enough musing me for now.  Back to work!

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  • How John Got 15x Improvement Without Really Trying

    - by rchrd
    The following article was published on a Sun Microsystems website a number of years ago by John Feo. It is still useful and worth preserving. So I'm republishing it here.  How I Got 15x Improvement Without Really Trying John Feo, Sun Microsystems Taking ten "personal" program codes used in scientific and engineering research, the author was able to get from 2 to 15 times performance improvement easily by applying some simple general optimization techniques. Introduction Scientific research based on computer simulation depends on the simulation for advancement. The research can advance only as fast as the computational codes can execute. The codes' efficiency determines both the rate and quality of results. In the same amount of time, a faster program can generate more results and can carry out a more detailed simulation of physical phenomena than a slower program. Highly optimized programs help science advance quickly and insure that monies supporting scientific research are used as effectively as possible. Scientific computer codes divide into three broad categories: ISV, community, and personal. ISV codes are large, mature production codes developed and sold commercially. The codes improve slowly over time both in methods and capabilities, and they are well tuned for most vendor platforms. Since the codes are mature and complex, there are few opportunities to improve their performance solely through code optimization. Improvements of 10% to 15% are typical. Examples of ISV codes are DYNA3D, Gaussian, and Nastran. Community codes are non-commercial production codes used by a particular research field. Generally, they are developed and distributed by a single academic or research institution with assistance from the community. Most users just run the codes, but some develop new methods and extensions that feed back into the general release. The codes are available on most vendor platforms. Since these codes are younger than ISV codes, there are more opportunities to optimize the source code. Improvements of 50% are not unusual. Examples of community codes are AMBER, CHARM, BLAST, and FASTA. Personal codes are those written by single users or small research groups for their own use. These codes are not distributed, but may be passed from professor-to-student or student-to-student over several years. They form the primordial ocean of applications from which community and ISV codes emerge. Government research grants pay for the development of most personal codes. This paper reports on the nature and performance of this class of codes. Over the last year, I have looked at over two dozen personal codes from more than a dozen research institutions. The codes cover a variety of scientific fields, including astronomy, atmospheric sciences, bioinformatics, biology, chemistry, geology, and physics. The sources range from a few hundred lines to more than ten thousand lines, and are written in Fortran, Fortran 90, C, and C++. For the most part, the codes are modular, documented, and written in a clear, straightforward manner. They do not use complex language features, advanced data structures, programming tricks, or libraries. I had little trouble understanding what the codes did or how data structures were used. Most came with a makefile. Surprisingly, only one of the applications is parallel. All developers have access to parallel machines, so availability is not an issue. Several tried to parallelize their applications, but stopped after encountering difficulties. Lack of education and a perception that parallelism is difficult prevented most from trying. I parallelized several of the codes using OpenMP, and did not judge any of the codes as difficult to parallelize. Even more surprising than the lack of parallelism is the inefficiency of the codes. I was able to get large improvements in performance in a matter of a few days applying simple optimization techniques. Table 1 lists ten representative codes [names and affiliation are omitted to preserve anonymity]. Improvements on one processor range from 2x to 15.5x with a simple average of 4.75x. I did not use sophisticated performance tools or drill deep into the program's execution character as one would do when tuning ISV or community codes. Using only a profiler and source line timers, I identified inefficient sections of code and improved their performance by inspection. The changes were at a high level. I am sure there is another factor of 2 or 3 in each code, and more if the codes are parallelized. The study’s results show that personal scientific codes are running many times slower than they should and that the problem is pervasive. Computational scientists are not sloppy programmers; however, few are trained in the art of computer programming or code optimization. I found that most have a working knowledge of some programming language and standard software engineering practices; but they do not know, or think about, how to make their programs run faster. They simply do not know the standard techniques used to make codes run faster. In fact, they do not even perceive that such techniques exist. The case studies described in this paper show that applying simple, well known techniques can significantly increase the performance of personal codes. It is important that the scientific community and the Government agencies that support scientific research find ways to better educate academic scientific programmers. The inefficiency of their codes is so bad that it is retarding both the quality and progress of scientific research. # cacheperformance redundantoperations loopstructures performanceimprovement 1 x x 15.5 2 x 2.8 3 x x 2.5 4 x 2.1 5 x x 2.0 6 x 5.0 7 x 5.8 8 x 6.3 9 2.2 10 x x 3.3 Table 1 — Area of improvement and performance gains of 10 codes The remainder of the paper is organized as follows: sections 2, 3, and 4 discuss the three most common sources of inefficiencies in the codes studied. These are cache performance, redundant operations, and loop structures. Each section includes several examples. The last section summaries the work and suggests a possible solution to the issues raised. Optimizing cache performance Commodity microprocessor systems use caches to increase memory bandwidth and reduce memory latencies. Typical latencies from processor to L1, L2, local, and remote memory are 3, 10, 50, and 200 cycles, respectively. Moreover, bandwidth falls off dramatically as memory distances increase. Programs that do not use cache effectively run many times slower than programs that do. When optimizing for cache, the biggest performance gains are achieved by accessing data in cache order and reusing data to amortize the overhead of cache misses. Secondary considerations are prefetching, associativity, and replacement; however, the understanding and analysis required to optimize for the latter are probably beyond the capabilities of the non-expert. Much can be gained simply by accessing data in the correct order and maximizing data reuse. 6 out of the 10 codes studied here benefited from such high level optimizations. Array Accesses The most important cache optimization is the most basic: accessing Fortran array elements in column order and C array elements in row order. Four of the ten codes—1, 2, 4, and 10—got it wrong. Compilers will restructure nested loops to optimize cache performance, but may not do so if the loop structure is too complex, or the loop body includes conditionals, complex addressing, or function calls. In code 1, the compiler failed to invert a key loop because of complex addressing do I = 0, 1010, delta_x IM = I - delta_x IP = I + delta_x do J = 5, 995, delta_x JM = J - delta_x JP = J + delta_x T1 = CA1(IP, J) + CA1(I, JP) T2 = CA1(IM, J) + CA1(I, JM) S1 = T1 + T2 - 4 * CA1(I, J) CA(I, J) = CA1(I, J) + D * S1 end do end do In code 2, the culprit is conditionals do I = 1, N do J = 1, N If (IFLAG(I,J) .EQ. 0) then T1 = Value(I, J-1) T2 = Value(I-1, J) T3 = Value(I, J) T4 = Value(I+1, J) T5 = Value(I, J+1) Value(I,J) = 0.25 * (T1 + T2 + T5 + T4) Delta = ABS(T3 - Value(I,J)) If (Delta .GT. MaxDelta) MaxDelta = Delta endif enddo enddo I fixed both programs by inverting the loops by hand. Code 10 has three-dimensional arrays and triply nested loops. The structure of the most computationally intensive loops is too complex to invert automatically or by hand. The only practical solution is to transpose the arrays so that the dimension accessed by the innermost loop is in cache order. The arrays can be transposed at construction or prior to entering a computationally intensive section of code. The former requires all array references to be modified, while the latter is cost effective only if the cost of the transpose is amortized over many accesses. I used the second approach to optimize code 10. Code 5 has four-dimensional arrays and loops are nested four deep. For all of the reasons cited above the compiler is not able to restructure three key loops. Assume C arrays and let the four dimensions of the arrays be i, j, k, and l. In the original code, the index structure of the three loops is L1: for i L2: for i L3: for i for l for l for j for k for j for k for j for k for l So only L3 accesses array elements in cache order. L1 is a very complex loop—much too complex to invert. I brought the loop into cache alignment by transposing the second and fourth dimensions of the arrays. Since the code uses a macro to compute all array indexes, I effected the transpose at construction and changed the macro appropriately. The dimensions of the new arrays are now: i, l, k, and j. L3 is a simple loop and easily inverted. L2 has a loop-carried scalar dependence in k. By promoting the scalar name that carries the dependence to an array, I was able to invert the third and fourth subloops aligning the loop with cache. Code 5 is by far the most difficult of the four codes to optimize for array accesses; but the knowledge required to fix the problems is no more than that required for the other codes. I would judge this code at the limits of, but not beyond, the capabilities of appropriately trained computational scientists. Array Strides When a cache miss occurs, a line (64 bytes) rather than just one word is loaded into the cache. If data is accessed stride 1, than the cost of the miss is amortized over 8 words. Any stride other than one reduces the cost savings. Two of the ten codes studied suffered from non-unit strides. The codes represent two important classes of "strided" codes. Code 1 employs a multi-grid algorithm to reduce time to convergence. The grids are every tenth, fifth, second, and unit element. Since time to convergence is inversely proportional to the distance between elements, coarse grids converge quickly providing good starting values for finer grids. The better starting values further reduce the time to convergence. The downside is that grids of every nth element, n > 1, introduce non-unit strides into the computation. In the original code, much of the savings of the multi-grid algorithm were lost due to this problem. I eliminated the problem by compressing (copying) coarse grids into continuous memory, and rewriting the computation as a function of the compressed grid. On convergence, I copied the final values of the compressed grid back to the original grid. The savings gained from unit stride access of the compressed grid more than paid for the cost of copying. Using compressed grids, the loop from code 1 included in the previous section becomes do j = 1, GZ do i = 1, GZ T1 = CA(i+0, j-1) + CA(i-1, j+0) T4 = CA1(i+1, j+0) + CA1(i+0, j+1) S1 = T1 + T4 - 4 * CA1(i+0, j+0) CA(i+0, j+0) = CA1(i+0, j+0) + DD * S1 enddo enddo where CA and CA1 are compressed arrays of size GZ. Code 7 traverses a list of objects selecting objects for later processing. The labels of the selected objects are stored in an array. The selection step has unit stride, but the processing steps have irregular stride. A fix is to save the parameters of the selected objects in temporary arrays as they are selected, and pass the temporary arrays to the processing functions. The fix is practical if the same parameters are used in selection as in processing, or if processing comprises a series of distinct steps which use overlapping subsets of the parameters. Both conditions are true for code 7, so I achieved significant improvement by copying parameters to temporary arrays during selection. Data reuse In the previous sections, we optimized for spatial locality. It is also important to optimize for temporal locality. Once read, a datum should be used as much as possible before it is forced from cache. Loop fusion and loop unrolling are two techniques that increase temporal locality. Unfortunately, both techniques increase register pressure—as loop bodies become larger, the number of registers required to hold temporary values grows. Once register spilling occurs, any gains evaporate quickly. For multiprocessors with small register sets or small caches, the sweet spot can be very small. In the ten codes presented here, I found no opportunities for loop fusion and only two opportunities for loop unrolling (codes 1 and 3). In code 1, unrolling the outer and inner loop one iteration increases the number of result values computed by the loop body from 1 to 4, do J = 1, GZ-2, 2 do I = 1, GZ-2, 2 T1 = CA1(i+0, j-1) + CA1(i-1, j+0) T2 = CA1(i+1, j-1) + CA1(i+0, j+0) T3 = CA1(i+0, j+0) + CA1(i-1, j+1) T4 = CA1(i+1, j+0) + CA1(i+0, j+1) T5 = CA1(i+2, j+0) + CA1(i+1, j+1) T6 = CA1(i+1, j+1) + CA1(i+0, j+2) T7 = CA1(i+2, j+1) + CA1(i+1, j+2) S1 = T1 + T4 - 4 * CA1(i+0, j+0) S2 = T2 + T5 - 4 * CA1(i+1, j+0) S3 = T3 + T6 - 4 * CA1(i+0, j+1) S4 = T4 + T7 - 4 * CA1(i+1, j+1) CA(i+0, j+0) = CA1(i+0, j+0) + DD * S1 CA(i+1, j+0) = CA1(i+1, j+0) + DD * S2 CA(i+0, j+1) = CA1(i+0, j+1) + DD * S3 CA(i+1, j+1) = CA1(i+1, j+1) + DD * S4 enddo enddo The loop body executes 12 reads, whereas as the rolled loop shown in the previous section executes 20 reads to compute the same four values. In code 3, two loops are unrolled 8 times and one loop is unrolled 4 times. Here is the before for (k = 0; k < NK[u]; k++) { sum = 0.0; for (y = 0; y < NY; y++) { sum += W[y][u][k] * delta[y]; } backprop[i++]=sum; } and after code for (k = 0; k < KK - 8; k+=8) { sum0 = 0.0; sum1 = 0.0; sum2 = 0.0; sum3 = 0.0; sum4 = 0.0; sum5 = 0.0; sum6 = 0.0; sum7 = 0.0; for (y = 0; y < NY; y++) { sum0 += W[y][0][k+0] * delta[y]; sum1 += W[y][0][k+1] * delta[y]; sum2 += W[y][0][k+2] * delta[y]; sum3 += W[y][0][k+3] * delta[y]; sum4 += W[y][0][k+4] * delta[y]; sum5 += W[y][0][k+5] * delta[y]; sum6 += W[y][0][k+6] * delta[y]; sum7 += W[y][0][k+7] * delta[y]; } backprop[k+0] = sum0; backprop[k+1] = sum1; backprop[k+2] = sum2; backprop[k+3] = sum3; backprop[k+4] = sum4; backprop[k+5] = sum5; backprop[k+6] = sum6; backprop[k+7] = sum7; } for one of the loops unrolled 8 times. Optimizing for temporal locality is the most difficult optimization considered in this paper. The concepts are not difficult, but the sweet spot is small. Identifying where the program can benefit from loop unrolling or loop fusion is not trivial. Moreover, it takes some effort to get it right. Still, educating scientific programmers about temporal locality and teaching them how to optimize for it will pay dividends. Reducing instruction count Execution time is a function of instruction count. Reduce the count and you usually reduce the time. The best solution is to use a more efficient algorithm; that is, an algorithm whose order of complexity is smaller, that converges quicker, or is more accurate. Optimizing source code without changing the algorithm yields smaller, but still significant, gains. This paper considers only the latter because the intent is to study how much better codes can run if written by programmers schooled in basic code optimization techniques. The ten codes studied benefited from three types of "instruction reducing" optimizations. The two most prevalent were hoisting invariant memory and data operations out of inner loops. The third was eliminating unnecessary data copying. The nature of these inefficiencies is language dependent. Memory operations The semantics of C make it difficult for the compiler to determine all the invariant memory operations in a loop. The problem is particularly acute for loops in functions since the compiler may not know the values of the function's parameters at every call site when compiling the function. Most compilers support pragmas to help resolve ambiguities; however, these pragmas are not comprehensive and there is no standard syntax. To guarantee that invariant memory operations are not executed repetitively, the user has little choice but to hoist the operations by hand. The problem is not as severe in Fortran programs because in the absence of equivalence statements, it is a violation of the language's semantics for two names to share memory. Codes 3 and 5 are C programs. In both cases, the compiler did not hoist all invariant memory operations from inner loops. Consider the following loop from code 3 for (y = 0; y < NY; y++) { i = 0; for (u = 0; u < NU; u++) { for (k = 0; k < NK[u]; k++) { dW[y][u][k] += delta[y] * I1[i++]; } } } Since dW[y][u] can point to the same memory space as delta for one or more values of y and u, assignment to dW[y][u][k] may change the value of delta[y]. In reality, dW and delta do not overlap in memory, so I rewrote the loop as for (y = 0; y < NY; y++) { i = 0; Dy = delta[y]; for (u = 0; u < NU; u++) { for (k = 0; k < NK[u]; k++) { dW[y][u][k] += Dy * I1[i++]; } } } Failure to hoist invariant memory operations may be due to complex address calculations. If the compiler can not determine that the address calculation is invariant, then it can hoist neither the calculation nor the associated memory operations. As noted above, code 5 uses a macro to address four-dimensional arrays #define MAT4D(a,q,i,j,k) (double *)((a)->data + (q)*(a)->strides[0] + (i)*(a)->strides[3] + (j)*(a)->strides[2] + (k)*(a)->strides[1]) The macro is too complex for the compiler to understand and so, it does not identify any subexpressions as loop invariant. The simplest way to eliminate the address calculation from the innermost loop (over i) is to define a0 = MAT4D(a,q,0,j,k) before the loop and then replace all instances of *MAT4D(a,q,i,j,k) in the loop with a0[i] A similar problem appears in code 6, a Fortran program. The key loop in this program is do n1 = 1, nh nx1 = (n1 - 1) / nz + 1 nz1 = n1 - nz * (nx1 - 1) do n2 = 1, nh nx2 = (n2 - 1) / nz + 1 nz2 = n2 - nz * (nx2 - 1) ndx = nx2 - nx1 ndy = nz2 - nz1 gxx = grn(1,ndx,ndy) gyy = grn(2,ndx,ndy) gxy = grn(3,ndx,ndy) balance(n1,1) = balance(n1,1) + (force(n2,1) * gxx + force(n2,2) * gxy) * h1 balance(n1,2) = balance(n1,2) + (force(n2,1) * gxy + force(n2,2) * gyy)*h1 end do end do The programmer has written this loop well—there are no loop invariant operations with respect to n1 and n2. However, the loop resides within an iterative loop over time and the index calculations are independent with respect to time. Trading space for time, I precomputed the index values prior to the entering the time loop and stored the values in two arrays. I then replaced the index calculations with reads of the arrays. Data operations Ways to reduce data operations can appear in many forms. Implementing a more efficient algorithm produces the biggest gains. The closest I came to an algorithm change was in code 4. This code computes the inner product of K-vectors A(i) and B(j), 0 = i < N, 0 = j < M, for most values of i and j. Since the program computes most of the NM possible inner products, it is more efficient to compute all the inner products in one triply-nested loop rather than one at a time when needed. The savings accrue from reading A(i) once for all B(j) vectors and from loop unrolling. for (i = 0; i < N; i+=8) { for (j = 0; j < M; j++) { sum0 = 0.0; sum1 = 0.0; sum2 = 0.0; sum3 = 0.0; sum4 = 0.0; sum5 = 0.0; sum6 = 0.0; sum7 = 0.0; for (k = 0; k < K; k++) { sum0 += A[i+0][k] * B[j][k]; sum1 += A[i+1][k] * B[j][k]; sum2 += A[i+2][k] * B[j][k]; sum3 += A[i+3][k] * B[j][k]; sum4 += A[i+4][k] * B[j][k]; sum5 += A[i+5][k] * B[j][k]; sum6 += A[i+6][k] * B[j][k]; sum7 += A[i+7][k] * B[j][k]; } C[i+0][j] = sum0; C[i+1][j] = sum1; C[i+2][j] = sum2; C[i+3][j] = sum3; C[i+4][j] = sum4; C[i+5][j] = sum5; C[i+6][j] = sum6; C[i+7][j] = sum7; }} This change requires knowledge of a typical run; i.e., that most inner products are computed. The reasons for the change, however, derive from basic optimization concepts. It is the type of change easily made at development time by a knowledgeable programmer. In code 5, we have the data version of the index optimization in code 6. Here a very expensive computation is a function of the loop indices and so cannot be hoisted out of the loop; however, the computation is invariant with respect to an outer iterative loop over time. We can compute its value for each iteration of the computation loop prior to entering the time loop and save the values in an array. The increase in memory required to store the values is small in comparison to the large savings in time. The main loop in Code 8 is doubly nested. The inner loop includes a series of guarded computations; some are a function of the inner loop index but not the outer loop index while others are a function of the outer loop index but not the inner loop index for (j = 0; j < N; j++) { for (i = 0; i < M; i++) { r = i * hrmax; R = A[j]; temp = (PRM[3] == 0.0) ? 1.0 : pow(r, PRM[3]); high = temp * kcoeff * B[j] * PRM[2] * PRM[4]; low = high * PRM[6] * PRM[6] / (1.0 + pow(PRM[4] * PRM[6], 2.0)); kap = (R > PRM[6]) ? high * R * R / (1.0 + pow(PRM[4]*r, 2.0) : low * pow(R/PRM[6], PRM[5]); < rest of loop omitted > }} Note that the value of temp is invariant to j. Thus, we can hoist the computation for temp out of the loop and save its values in an array. for (i = 0; i < M; i++) { r = i * hrmax; TEMP[i] = pow(r, PRM[3]); } [N.B. – the case for PRM[3] = 0 is omitted and will be reintroduced later.] We now hoist out of the inner loop the computations invariant to i. Since the conditional guarding the value of kap is invariant to i, it behooves us to hoist the computation out of the inner loop, thereby executing the guard once rather than M times. The final version of the code is for (j = 0; j < N; j++) { R = rig[j] / 1000.; tmp1 = kcoeff * par[2] * beta[j] * par[4]; tmp2 = 1.0 + (par[4] * par[4] * par[6] * par[6]); tmp3 = 1.0 + (par[4] * par[4] * R * R); tmp4 = par[6] * par[6] / tmp2; tmp5 = R * R / tmp3; tmp6 = pow(R / par[6], par[5]); if ((par[3] == 0.0) && (R > par[6])) { for (i = 1; i <= imax1; i++) KAP[i] = tmp1 * tmp5; } else if ((par[3] == 0.0) && (R <= par[6])) { for (i = 1; i <= imax1; i++) KAP[i] = tmp1 * tmp4 * tmp6; } else if ((par[3] != 0.0) && (R > par[6])) { for (i = 1; i <= imax1; i++) KAP[i] = tmp1 * TEMP[i] * tmp5; } else if ((par[3] != 0.0) && (R <= par[6])) { for (i = 1; i <= imax1; i++) KAP[i] = tmp1 * TEMP[i] * tmp4 * tmp6; } for (i = 0; i < M; i++) { kap = KAP[i]; r = i * hrmax; < rest of loop omitted > } } Maybe not the prettiest piece of code, but certainly much more efficient than the original loop, Copy operations Several programs unnecessarily copy data from one data structure to another. This problem occurs in both Fortran and C programs, although it manifests itself differently in the two languages. Code 1 declares two arrays—one for old values and one for new values. At the end of each iteration, the array of new values is copied to the array of old values to reset the data structures for the next iteration. This problem occurs in Fortran programs not included in this study and in both Fortran 77 and Fortran 90 code. Introducing pointers to the arrays and swapping pointer values is an obvious way to eliminate the copying; but pointers is not a feature that many Fortran programmers know well or are comfortable using. An easy solution not involving pointers is to extend the dimension of the value array by 1 and use the last dimension to differentiate between arrays at different times. For example, if the data space is N x N, declare the array (N, N, 2). Then store the problem’s initial values in (_, _, 2) and define the scalar names new = 2 and old = 1. At the start of each iteration, swap old and new to reset the arrays. The old–new copy problem did not appear in any C program. In programs that had new and old values, the code swapped pointers to reset data structures. Where unnecessary coping did occur is in structure assignment and parameter passing. Structures in C are handled much like scalars. Assignment causes the data space of the right-hand name to be copied to the data space of the left-hand name. Similarly, when a structure is passed to a function, the data space of the actual parameter is copied to the data space of the formal parameter. If the structure is large and the assignment or function call is in an inner loop, then copying costs can grow quite large. While none of the ten programs considered here manifested this problem, it did occur in programs not included in the study. A simple fix is always to refer to structures via pointers. Optimizing loop structures Since scientific programs spend almost all their time in loops, efficient loops are the key to good performance. Conditionals, function calls, little instruction level parallelism, and large numbers of temporary values make it difficult for the compiler to generate tightly packed, highly efficient code. Conditionals and function calls introduce jumps that disrupt code flow. Users should eliminate or isolate conditionls to their own loops as much as possible. Often logical expressions can be substituted for if-then-else statements. For example, code 2 includes the following snippet MaxDelta = 0.0 do J = 1, N do I = 1, M < code omitted > Delta = abs(OldValue ? NewValue) if (Delta > MaxDelta) MaxDelta = Delta enddo enddo if (MaxDelta .gt. 0.001) goto 200 Since the only use of MaxDelta is to control the jump to 200 and all that matters is whether or not it is greater than 0.001, I made MaxDelta a boolean and rewrote the snippet as MaxDelta = .false. do J = 1, N do I = 1, M < code omitted > Delta = abs(OldValue ? NewValue) MaxDelta = MaxDelta .or. (Delta .gt. 0.001) enddo enddo if (MaxDelta) goto 200 thereby, eliminating the conditional expression from the inner loop. A microprocessor can execute many instructions per instruction cycle. Typically, it can execute one or more memory, floating point, integer, and jump operations. To be executed simultaneously, the operations must be independent. Thick loops tend to have more instruction level parallelism than thin loops. Moreover, they reduce memory traffice by maximizing data reuse. Loop unrolling and loop fusion are two techniques to increase the size of loop bodies. Several of the codes studied benefitted from loop unrolling, but none benefitted from loop fusion. This observation is not too surpising since it is the general tendency of programmers to write thick loops. As loops become thicker, the number of temporary values grows, increasing register pressure. If registers spill, then memory traffic increases and code flow is disrupted. A thick loop with many temporary values may execute slower than an equivalent series of thin loops. The biggest gain will be achieved if the thick loop can be split into a series of independent loops eliminating the need to write and read temporary arrays. I found such an occasion in code 10 where I split the loop do i = 1, n do j = 1, m A24(j,i)= S24(j,i) * T24(j,i) + S25(j,i) * U25(j,i) B24(j,i)= S24(j,i) * T25(j,i) + S25(j,i) * U24(j,i) A25(j,i)= S24(j,i) * C24(j,i) + S25(j,i) * V24(j,i) B25(j,i)= S24(j,i) * U25(j,i) + S25(j,i) * V25(j,i) C24(j,i)= S26(j,i) * T26(j,i) + S27(j,i) * U26(j,i) D24(j,i)= S26(j,i) * T27(j,i) + S27(j,i) * V26(j,i) C25(j,i)= S27(j,i) * S28(j,i) + S26(j,i) * U28(j,i) D25(j,i)= S27(j,i) * T28(j,i) + S26(j,i) * V28(j,i) end do end do into two disjoint loops do i = 1, n do j = 1, m A24(j,i)= S24(j,i) * T24(j,i) + S25(j,i) * U25(j,i) B24(j,i)= S24(j,i) * T25(j,i) + S25(j,i) * U24(j,i) A25(j,i)= S24(j,i) * C24(j,i) + S25(j,i) * V24(j,i) B25(j,i)= S24(j,i) * U25(j,i) + S25(j,i) * V25(j,i) end do end do do i = 1, n do j = 1, m C24(j,i)= S26(j,i) * T26(j,i) + S27(j,i) * U26(j,i) D24(j,i)= S26(j,i) * T27(j,i) + S27(j,i) * V26(j,i) C25(j,i)= S27(j,i) * S28(j,i) + S26(j,i) * U28(j,i) D25(j,i)= S27(j,i) * T28(j,i) + S26(j,i) * V28(j,i) end do end do Conclusions Over the course of the last year, I have had the opportunity to work with over two dozen academic scientific programmers at leading research universities. Their research interests span a broad range of scientific fields. Except for two programs that relied almost exclusively on library routines (matrix multiply and fast Fourier transform), I was able to improve significantly the single processor performance of all codes. Improvements range from 2x to 15.5x with a simple average of 4.75x. Changes to the source code were at a very high level. I did not use sophisticated techniques or programming tools to discover inefficiencies or effect the changes. Only one code was parallel despite the availability of parallel systems to all developers. Clearly, we have a problem—personal scientific research codes are highly inefficient and not running parallel. The developers are unaware of simple optimization techniques to make programs run faster. They lack education in the art of code optimization and parallel programming. I do not believe we can fix the problem by publishing additional books or training manuals. To date, the developers in questions have not studied the books or manual available, and are unlikely to do so in the future. Short courses are a possible solution, but I believe they are too concentrated to be much use. The general concepts can be taught in a three or four day course, but that is not enough time for students to practice what they learn and acquire the experience to apply and extend the concepts to their codes. Practice is the key to becoming proficient at optimization. I recommend that graduate students be required to take a semester length course in optimization and parallel programming. We would never give someone access to state-of-the-art scientific equipment costing hundreds of thousands of dollars without first requiring them to demonstrate that they know how to use the equipment. Yet the criterion for time on state-of-the-art supercomputers is at most an interesting project. Requestors are never asked to demonstrate that they know how to use the system, or can use the system effectively. A semester course would teach them the required skills. Government agencies that fund academic scientific research pay for most of the computer systems supporting scientific research as well as the development of most personal scientific codes. These agencies should require graduate schools to offer a course in optimization and parallel programming as a requirement for funding. About the Author John Feo received his Ph.D. in Computer Science from The University of Texas at Austin in 1986. After graduate school, Dr. Feo worked at Lawrence Livermore National Laboratory where he was the Group Leader of the Computer Research Group and principal investigator of the Sisal Language Project. In 1997, Dr. Feo joined Tera Computer Company where he was project manager for the MTA, and oversaw the programming and evaluation of the MTA at the San Diego Supercomputer Center. In 2000, Dr. Feo joined Sun Microsystems as an HPC application specialist. He works with university research groups to optimize and parallelize scientific codes. Dr. Feo has published over two dozen research articles in the areas of parallel parallel programming, parallel programming languages, and application performance.

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  • Node.js Adventure - Storage Services and Service Runtime

    - by Shaun
    When I described on how to host a Node.js application on Windows Azure, one of questions might be raised about how to consume the vary Windows Azure services, such as the storage, service bus, access control, etc.. Interact with windows azure services is available in Node.js through the Windows Azure Node.js SDK, which is a module available in NPM. In this post I would like to describe on how to use Windows Azure Storage (a.k.a. WAS) as well as the service runtime.   Consume Windows Azure Storage Let’s firstly have a look on how to consume WAS through Node.js. As we know in the previous post we can host Node.js application on Windows Azure Web Site (a.k.a. WAWS) as well as Windows Azure Cloud Service (a.k.a. WACS). In theory, WAWS is also built on top of WACS worker roles with some more features. Hence in this post I will only demonstrate for hosting in WACS worker role. The Node.js code can be used when consuming WAS when hosted on WAWS. But since there’s no roles in WAWS, the code for consuming service runtime mentioned in the next section cannot be used for WAWS node application. We can use the solution that I created in my last post. Alternatively we can create a new windows azure project in Visual Studio with a worker role, add the “node.exe” and “index.js” and install “express” and “node-sqlserver” modules, make all files as “Copy always”. In order to use windows azure services we need to have Windows Azure Node.js SDK, as knows as a module named “azure” which can be installed through NPM. Once we downloaded and installed, we need to include them in our worker role project and make them as “Copy always”. You can use my “Copy all always” tool mentioned in my last post to update the currently worker role project file. You can also find the source code of this tool here. The source code of Windows Azure SDK for Node.js can be found in its GitHub page. It contains two parts. One is a CLI tool which provides a cross platform command line package for Mac and Linux to manage WAWS and Windows Azure Virtual Machines (a.k.a. WAVM). The other is a library for managing and consuming vary windows azure services includes tables, blobs, queues, service bus and the service runtime. I will not cover all of them but will only demonstrate on how to use tables and service runtime information in this post. You can find the full document of this SDK here. Back to Visual Studio and open the “index.js”, let’s continue our application from the last post, which was working against Windows Azure SQL Database (a.k.a. WASD). The code should looks like this. 1: var express = require("express"); 2: var sql = require("node-sqlserver"); 3:  4: var connectionString = "Driver={SQL Server Native Client 10.0};Server=tcp:ac6271ya9e.database.windows.net,1433;Database=synctile;Uid=shaunxu@ac6271ya9e;Pwd={PASSWORD};Encrypt=yes;Connection Timeout=30;"; 5: var port = 80; 6:  7: var app = express(); 8:  9: app.configure(function () { 10: app.use(express.bodyParser()); 11: }); 12:  13: app.get("/", function (req, res) { 14: sql.open(connectionString, function (err, conn) { 15: if (err) { 16: console.log(err); 17: res.send(500, "Cannot open connection."); 18: } 19: else { 20: conn.queryRaw("SELECT * FROM [Resource]", function (err, results) { 21: if (err) { 22: console.log(err); 23: res.send(500, "Cannot retrieve records."); 24: } 25: else { 26: res.json(results); 27: } 28: }); 29: } 30: }); 31: }); 32:  33: app.get("/text/:key/:culture", function (req, res) { 34: sql.open(connectionString, function (err, conn) { 35: if (err) { 36: console.log(err); 37: res.send(500, "Cannot open connection."); 38: } 39: else { 40: var key = req.params.key; 41: var culture = req.params.culture; 42: var command = "SELECT * FROM [Resource] WHERE [Key] = '" + key + "' AND [Culture] = '" + culture + "'"; 43: conn.queryRaw(command, function (err, results) { 44: if (err) { 45: console.log(err); 46: res.send(500, "Cannot retrieve records."); 47: } 48: else { 49: res.json(results); 50: } 51: }); 52: } 53: }); 54: }); 55:  56: app.get("/sproc/:key/:culture", function (req, res) { 57: sql.open(connectionString, function (err, conn) { 58: if (err) { 59: console.log(err); 60: res.send(500, "Cannot open connection."); 61: } 62: else { 63: var key = req.params.key; 64: var culture = req.params.culture; 65: var command = "EXEC GetItem '" + key + "', '" + culture + "'"; 66: conn.queryRaw(command, function (err, results) { 67: if (err) { 68: console.log(err); 69: res.send(500, "Cannot retrieve records."); 70: } 71: else { 72: res.json(results); 73: } 74: }); 75: } 76: }); 77: }); 78:  79: app.post("/new", function (req, res) { 80: var key = req.body.key; 81: var culture = req.body.culture; 82: var val = req.body.val; 83:  84: sql.open(connectionString, function (err, conn) { 85: if (err) { 86: console.log(err); 87: res.send(500, "Cannot open connection."); 88: } 89: else { 90: var command = "INSERT INTO [Resource] VALUES ('" + key + "', '" + culture + "', N'" + val + "')"; 91: conn.queryRaw(command, function (err, results) { 92: if (err) { 93: console.log(err); 94: res.send(500, "Cannot retrieve records."); 95: } 96: else { 97: res.send(200, "Inserted Successful"); 98: } 99: }); 100: } 101: }); 102: }); 103:  104: app.listen(port); Now let’s create a new function, copy the records from WASD to table service. 1. Delete the table named “resource”. 2. Create a new table named “resource”. These 2 steps ensures that we have an empty table. 3. Load all records from the “resource” table in WASD. 4. For each records loaded from WASD, insert them into the table one by one. 5. Prompt to user when finished. In order to use table service we need the storage account and key, which can be found from the developer portal. Just select the storage account and click the Manage Keys button. Then create two local variants in our Node.js application for the storage account name and key. Since we need to use WAS we need to import the azure module. Also I created another variant stored the table name. In order to work with table service I need to create the storage client for table service. This is very similar as the Windows Azure SDK for .NET. As the code below I created a new variant named “client” and use “createTableService”, specified my storage account name and key. 1: var azure = require("azure"); 2: var storageAccountName = "synctile"; 3: var storageAccountKey = "/cOy9L7xysXOgPYU9FjDvjrRAhaMX/5tnOpcjqloPNDJYucbgTy7MOrAW7CbUg6PjaDdmyl+6pkwUnKETsPVNw=="; 4: var tableName = "resource"; 5: var client = azure.createTableService(storageAccountName, storageAccountKey); Now create a new function for URL “/was/init” so that we can trigger it through browser. Then in this function we will firstly load all records from WASD. 1: app.get("/was/init", function (req, res) { 2: // load all records from windows azure sql database 3: sql.open(connectionString, function (err, conn) { 4: if (err) { 5: console.log(err); 6: res.send(500, "Cannot open connection."); 7: } 8: else { 9: conn.queryRaw("SELECT * FROM [Resource]", function (err, results) { 10: if (err) { 11: console.log(err); 12: res.send(500, "Cannot retrieve records."); 13: } 14: else { 15: if (results.rows.length > 0) { 16: // begin to transform the records into table service 17: } 18: } 19: }); 20: } 21: }); 22: }); When we succeed loaded all records we can start to transform them into table service. First I need to recreate the table in table service. This can be done by deleting and creating the table through table client I had just created previously. 1: app.get("/was/init", function (req, res) { 2: // load all records from windows azure sql database 3: sql.open(connectionString, function (err, conn) { 4: if (err) { 5: console.log(err); 6: res.send(500, "Cannot open connection."); 7: } 8: else { 9: conn.queryRaw("SELECT * FROM [Resource]", function (err, results) { 10: if (err) { 11: console.log(err); 12: res.send(500, "Cannot retrieve records."); 13: } 14: else { 15: if (results.rows.length > 0) { 16: // begin to transform the records into table service 17: // recreate the table named 'resource' 18: client.deleteTable(tableName, function (error) { 19: client.createTableIfNotExists(tableName, function (error) { 20: if (error) { 21: error["target"] = "createTableIfNotExists"; 22: res.send(500, error); 23: } 24: else { 25: // transform the records 26: } 27: }); 28: }); 29: } 30: } 31: }); 32: } 33: }); 34: }); As you can see, the azure SDK provide its methods in callback pattern. In fact, almost all modules in Node.js use the callback pattern. For example, when I deleted a table I invoked “deleteTable” method, provided the name of the table and a callback function which will be performed when the table had been deleted or failed. Underlying, the azure module will perform the table deletion operation in POSIX async threads pool asynchronously. And once it’s done the callback function will be performed. This is the reason we need to nest the table creation code inside the deletion function. If we perform the table creation code after the deletion code then they will be invoked in parallel. Next, for each records in WASD I created an entity and then insert into the table service. Finally I send the response to the browser. Can you find a bug in the code below? I will describe it later in this post. 1: app.get("/was/init", function (req, res) { 2: // load all records from windows azure sql database 3: sql.open(connectionString, function (err, conn) { 4: if (err) { 5: console.log(err); 6: res.send(500, "Cannot open connection."); 7: } 8: else { 9: conn.queryRaw("SELECT * FROM [Resource]", function (err, results) { 10: if (err) { 11: console.log(err); 12: res.send(500, "Cannot retrieve records."); 13: } 14: else { 15: if (results.rows.length > 0) { 16: // begin to transform the records into table service 17: // recreate the table named 'resource' 18: client.deleteTable(tableName, function (error) { 19: client.createTableIfNotExists(tableName, function (error) { 20: if (error) { 21: error["target"] = "createTableIfNotExists"; 22: res.send(500, error); 23: } 24: else { 25: // transform the records 26: for (var i = 0; i < results.rows.length; i++) { 27: var entity = { 28: "PartitionKey": results.rows[i][1], 29: "RowKey": results.rows[i][0], 30: "Value": results.rows[i][2] 31: }; 32: client.insertEntity(tableName, entity, function (error) { 33: if (error) { 34: error["target"] = "insertEntity"; 35: res.send(500, error); 36: } 37: else { 38: console.log("entity inserted"); 39: } 40: }); 41: } 42: // send the 43: console.log("all done"); 44: res.send(200, "All done!"); 45: } 46: }); 47: }); 48: } 49: } 50: }); 51: } 52: }); 53: }); Now we can publish it to the cloud and have a try. But normally we’d better test it at the local emulator first. In Node.js SDK there are three build-in properties which provides the account name, key and host address for local storage emulator. We can use them to initialize our table service client. We also need to change the SQL connection string to let it use my local database. The code will be changed as below. 1: // windows azure sql database 2: //var connectionString = "Driver={SQL Server Native Client 10.0};Server=tcp:ac6271ya9e.database.windows.net,1433;Database=synctile;Uid=shaunxu@ac6271ya9e;Pwd=eszqu94XZY;Encrypt=yes;Connection Timeout=30;"; 3: // sql server 4: var connectionString = "Driver={SQL Server Native Client 11.0};Server={.};Database={Caspar};Trusted_Connection={Yes};"; 5:  6: var azure = require("azure"); 7: var storageAccountName = "synctile"; 8: var storageAccountKey = "/cOy9L7xysXOgPYU9FjDvjrRAhaMX/5tnOpcjqloPNDJYucbgTy7MOrAW7CbUg6PjaDdmyl+6pkwUnKETsPVNw=="; 9: var tableName = "resource"; 10: // windows azure storage 11: //var client = azure.createTableService(storageAccountName, storageAccountKey); 12: // local storage emulator 13: var client = azure.createTableService(azure.ServiceClient.DEVSTORE_STORAGE_ACCOUNT, azure.ServiceClient.DEVSTORE_STORAGE_ACCESS_KEY, azure.ServiceClient.DEVSTORE_TABLE_HOST); Now let’s run the application and navigate to “localhost:12345/was/init” as I hosted it on port 12345. We can find it transformed the data from my local database to local table service. Everything looks fine. But there is a bug in my code. If we have a look on the Node.js command window we will find that it sent response before all records had been inserted, which is not what I expected. The reason is that, as I mentioned before, Node.js perform all IO operations in non-blocking model. When we inserted the records we executed the table service insert method in parallel, and the operation of sending response was also executed in parallel, even though I wrote it at the end of my logic. The correct logic should be, when all entities had been copied to table service with no error, then I will send response to the browser, otherwise I should send error message to the browser. To do so I need to import another module named “async”, which helps us to coordinate our asynchronous code. Install the module and import it at the beginning of the code. Then we can use its “forEach” method for the asynchronous code of inserting table entities. The first argument of “forEach” is the array that will be performed. The second argument is the operation for each items in the array. And the third argument will be invoked then all items had been performed or any errors occurred. Here we can send our response to browser. 1: app.get("/was/init", function (req, res) { 2: // load all records from windows azure sql database 3: sql.open(connectionString, function (err, conn) { 4: if (err) { 5: console.log(err); 6: res.send(500, "Cannot open connection."); 7: } 8: else { 9: conn.queryRaw("SELECT * FROM [Resource]", function (err, results) { 10: if (err) { 11: console.log(err); 12: res.send(500, "Cannot retrieve records."); 13: } 14: else { 15: if (results.rows.length > 0) { 16: // begin to transform the records into table service 17: // recreate the table named 'resource' 18: client.deleteTable(tableName, function (error) { 19: client.createTableIfNotExists(tableName, function (error) { 20: if (error) { 21: error["target"] = "createTableIfNotExists"; 22: res.send(500, error); 23: } 24: else { 25: async.forEach(results.rows, 26: // transform the records 27: function (row, callback) { 28: var entity = { 29: "PartitionKey": row[1], 30: "RowKey": row[0], 31: "Value": row[2] 32: }; 33: client.insertEntity(tableName, entity, function (error) { 34: if (error) { 35: callback(error); 36: } 37: else { 38: console.log("entity inserted."); 39: callback(null); 40: } 41: }); 42: }, 43: // send reponse 44: function (error) { 45: if (error) { 46: error["target"] = "insertEntity"; 47: res.send(500, error); 48: } 49: else { 50: console.log("all done"); 51: res.send(200, "All done!"); 52: } 53: } 54: ); 55: } 56: }); 57: }); 58: } 59: } 60: }); 61: } 62: }); 63: }); Run it locally and now we can find the response was sent after all entities had been inserted. Query entities against table service is simple as well. Just use the “queryEntity” method from the table service client and providing the partition key and row key. We can also provide a complex query criteria as well, for example the code here. In the code below I queried an entity by the partition key and row key, and return the proper localization value in response. 1: app.get("/was/:key/:culture", function (req, res) { 2: var key = req.params.key; 3: var culture = req.params.culture; 4: client.queryEntity(tableName, culture, key, function (error, entity) { 5: if (error) { 6: res.send(500, error); 7: } 8: else { 9: res.json(entity); 10: } 11: }); 12: }); And then tested it on local emulator. Finally if we want to publish this application to the cloud we should change the database connection string and storage account. For more information about how to consume blob and queue service, as well as the service bus please refer to the MSDN page.   Consume Service Runtime As I mentioned above, before we published our application to the cloud we need to change the connection string and account information in our code. But if you had played with WACS you should have known that the service runtime provides the ability to retrieve configuration settings, endpoints and local resource information at runtime. Which means we can have these values defined in CSCFG and CSDEF files and then the runtime should be able to retrieve the proper values. For example we can add some role settings though the property window of the role, specify the connection string and storage account for cloud and local. And the can also use the endpoint which defined in role environment to our Node.js application. In Node.js SDK we can get an object from “azure.RoleEnvironment”, which provides the functionalities to retrieve the configuration settings and endpoints, etc.. In the code below I defined the connection string variants and then use the SDK to retrieve and initialize the table client. 1: var connectionString = ""; 2: var storageAccountName = ""; 3: var storageAccountKey = ""; 4: var tableName = ""; 5: var client; 6:  7: azure.RoleEnvironment.getConfigurationSettings(function (error, settings) { 8: if (error) { 9: console.log("ERROR: getConfigurationSettings"); 10: console.log(JSON.stringify(error)); 11: } 12: else { 13: console.log(JSON.stringify(settings)); 14: connectionString = settings["SqlConnectionString"]; 15: storageAccountName = settings["StorageAccountName"]; 16: storageAccountKey = settings["StorageAccountKey"]; 17: tableName = settings["TableName"]; 18:  19: console.log("connectionString = %s", connectionString); 20: console.log("storageAccountName = %s", storageAccountName); 21: console.log("storageAccountKey = %s", storageAccountKey); 22: console.log("tableName = %s", tableName); 23:  24: client = azure.createTableService(storageAccountName, storageAccountKey); 25: } 26: }); In this way we don’t need to amend the code for the configurations between local and cloud environment since the service runtime will take care of it. At the end of the code we will listen the application on the port retrieved from SDK as well. 1: azure.RoleEnvironment.getCurrentRoleInstance(function (error, instance) { 2: if (error) { 3: console.log("ERROR: getCurrentRoleInstance"); 4: console.log(JSON.stringify(error)); 5: } 6: else { 7: console.log(JSON.stringify(instance)); 8: if (instance["endpoints"] && instance["endpoints"]["nodejs"]) { 9: var endpoint = instance["endpoints"]["nodejs"]; 10: app.listen(endpoint["port"]); 11: } 12: else { 13: app.listen(8080); 14: } 15: } 16: }); But if we tested the application right now we will find that it cannot retrieve any values from service runtime. This is because by default, the entry point of this role was defined to the worker role class. In windows azure environment the service runtime will open a named pipeline to the entry point instance, so that it can connect to the runtime and retrieve values. But in this case, since the entry point was worker role and the Node.js was opened inside the role, the named pipeline was established between our worker role class and service runtime, so our Node.js application cannot use it. To fix this problem we need to open the CSDEF file under the azure project, add a new element named Runtime. Then add an element named EntryPoint which specify the Node.js command line. So that the Node.js application will have the connection to service runtime, then it’s able to read the configurations. Start the Node.js at local emulator we can find it retrieved the connections, storage account for local. And if we publish our application to azure then it works with WASD and storage service through the configurations for cloud.   Summary In this post I demonstrated how to use Windows Azure SDK for Node.js to interact with storage service, especially the table service. I also demonstrated on how to use WACS service runtime, how to retrieve the configuration settings and the endpoint information. And in order to make the service runtime available to my Node.js application I need to create an entry point element in CSDEF file and set “node.exe” as the entry point. I used five posts to introduce and demonstrate on how to run a Node.js application on Windows platform, how to use Windows Azure Web Site and Windows Azure Cloud Service worker role to host our Node.js application. I also described how to work with other services provided by Windows Azure platform through Windows Azure SDK for Node.js. Node.js is a very new and young network application platform. But since it’s very simple and easy to learn and deploy, as well as, it utilizes single thread non-blocking IO model, Node.js became more and more popular on web application and web service development especially for those IO sensitive projects. And as Node.js is very good at scaling-out, it’s more useful on cloud computing platform. Use Node.js on Windows platform is new, too. The modules for SQL database and Windows Azure SDK are still under development and enhancement. It doesn’t support SQL parameter in “node-sqlserver”. It does support using storage connection string to create the storage client in “azure”. But Microsoft is working on make them easier to use, working on add more features and functionalities.   PS, you can download the source code here. You can download the source code of my “Copy all always” tool here.   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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  • XSLT 1.0 help with recursion logic

    - by DashaLuna
    Hello guys, I'm having troubles with the logic and would apprecite any help/tips. I have <Deposits> elements and <Receipts> elements. However there isn't any identification what receipt was paid toward what deposit. I am trying to update the <Deposits> elements with the following attributes: @DueAmont - the amount that is still due to pay @Status - whether it's paid, outstanding (partly paid) or due @ReceiptDate - the latest receipt's date that was paid towards this deposit Every deposit could be paid with one or more receipts. It also could happen, that 1 receipt could cover one or more deposits. For example. If there are 3 deposits: 500 100 450 That are paid with the following receipts: 200 100 250 I want to get the following info: Deposit 1 is fully paid (status=paid, dueAmount=0, receiptNum=3. Deposit 2 is partly paid (status=outstanding, dueAmount=50, receiptNum=3. Deposit 3 is not paid (status=due, dueAmount=450, receiptNum=NAN. I've added comments in the code explaining what I'm trying to do. I am staring at this code for the 3rd day now non stop - can't see what I'm doing wrong. Please could anyone help me with it? :) Thanks! Set up: $deposits - All the available deposits $receiptsAsc - All the available receipts sorted by their @ActionDate Code: <!-- Accumulate all the deposits with @Status, @DueAmount and @ReceiptDate attributes Provide all deposits, receipts and start with 1st receipt --> <xsl:variable name="depositsClassified"> <xsl:call-template name="classifyDeposits"> <xsl:with-param name="depositsAll" select="$deposits"/> <xsl:with-param name="receiptsAll" select="$receiptsAsc"/> <xsl:with-param name="receiptCount" select="'1'"/> </xsl:call-template> </xsl:variable> <!-- Recursive function to associate deposits' total amounts with overall receipts paid to determine whether a deposit is due, outstanding or paid. Also determine what's the due amount and latest receipt towards the deposit for each deposit --> <xsl:template name="classifyDeposits"> <xsl:param name="depositsAll"/> <xsl:param name="receiptsAll"/> <xsl:param name="receiptCount"/> <!-- If there are deposits to proceed --> <xsl:if test="$depositsAll"> <!-- Get the 1st deposit --> <xsl:variable name="deposit" select="$depositsAll[1]"/> <!-- Calculate the sum of all receipts up to and including currenly considered --> <xsl:variable name="receiptSum"> <xsl:choose> <xsl:when test="$receiptsAll"> <xsl:value-of select="sum($receiptsAll[position() &lt;= $receiptCount]/@ReceiptAmount)"/> </xsl:when> <xsl:otherwise>0</xsl:otherwise> </xsl:choose> </xsl:variable> <!-- Difference between deposit amount and sum of the receipts calculated above --> <xsl:variable name="diff" select="$deposit/@DepositTotalAmount - $receiptSum"/> <xsl:choose> <!-- Deposit isn't paid fully and there are more receipts/payments exist. So consider the same deposit, but take next receipt into calculation as well --> <xsl:when test="($diff &gt; 0) and ($receiptCount &lt; count($receiptsAll))"> <xsl:call-template name="classifyDeposits"> <xsl:with-param name="depositsAll" select="$depositsAll"/> <xsl:with-param name="receiptsAll" select="$receiptsAll"/> <xsl:with-param name="receiptCount" select="$receiptCount + 1"/> </xsl:call-template> </xsl:when> <!-- Deposit is paid or we ran out of receipts --> <xsl:otherwise> <!-- process the deposit. Determine its status and then update corresponding attributes --> <xsl:apply-templates select="$deposit" mode="defineDeposit"> <xsl:with-param name="diff" select="$diff"/> <xsl:with-param name="receiptNum" select="$receiptCount"/> </xsl:apply-templates> <!-- Recursively call the template with the rest of deposits excluding the first. Before hand update the @ReceiptsAmount. For the receipts before current it is now 0, for the current is what left in the $diff, and simply copy over receipts after current one. --> <xsl:variable name="receiptsUpdatedRTF"> <xsl:for-each select="$receiptsAll"> <xsl:choose> <!-- these receipts was fully accounted for the current deposit. Make them 0 --> <xsl:when test="position() &lt; $receiptCount"> <xsl:copy> <xsl:copy-of select="./@*"/> <xsl:attribute name="ReceiptAmount">0</xsl:attribute> </xsl:copy> </xsl:when> <!-- this receipt was partly/fully(in case $diff=0) accounted for the current deposit. Make it whatever is in $diff --> <xsl:when test="position() = $receiptCount"> <xsl:copy> <xsl:copy-of select="./@*"/> <xsl:attribute name="ReceiptAmount"> <xsl:value-of select="format-number($diff, '#.00;#.00')"/> </xsl:attribute> </xsl:copy> </xsl:when> <!-- these receipts weren't yet considered - copy them over --> <xsl:otherwise> <xsl:copy-of select="."/> </xsl:otherwise> </xsl:choose> </xsl:for-each> </xsl:variable> <xsl:variable name="receiptsUpdated" select="msxsl:node-set($receiptsUpdatedRTF)/Receipts"/> <!-- Recursive call for the next deposit. Starting counting receipts from the current one. --> <xsl:call-template name="classifyDeposits"> <xsl:with-param name="depositsAll" select="$deposits[position() != 1]"/> <xsl:with-param name="receiptsAll" select="$receiptsUpdated"/> <xsl:with-param name="receiptCount" select="$receiptCount"/> </xsl:call-template> </xsl:otherwise> </xsl:choose> </xsl:if> </xsl:template> <!-- Determine deposit's status and due amount --> <xsl:template match="MultiDeposits" mode="defineDeposit"> <xsl:param name="diff"/> <xsl:param name="receiptNum"/> <xsl:choose> <xsl:when test="$diff &lt;= 0"> <xsl:apply-templates select="." mode="addAttrs"> <xsl:with-param name="status" select="'paid'"/> <xsl:with-param name="dueAmount" select="'0'"/> <xsl:with-param name="receiptNum" select="$receiptNum"/> </xsl:apply-templates> </xsl:when> <xsl:when test="$diff = ./@DepositTotalAmount"> <xsl:apply-templates select="." mode="addAttrs"> <xsl:with-param name="status" select="'due'"/> <xsl:with-param name="dueAmount" select="$diff"/> </xsl:apply-templates> </xsl:when> <xsl:when test="$diff &lt; ./@DepositTotalAmount"> <xsl:apply-templates select="." mode="addAttrs"> <xsl:with-param name="status" select="'outstanding'"/> <xsl:with-param name="dueAmount" select="$diff"/> <xsl:with-param name="receiptNum" select="$receiptNum"/> </xsl:apply-templates> </xsl:when> <xsl:otherwise/> </xsl:choose> </xsl:template> <!-- Add new attributes (@Status, @DueAmount and @ReceiptDate) to the deposit element --> <xsl:template match="MultiDeposits" mode="addAttrs"> <xsl:param name="status"/> <xsl:param name="dueAmount"/> <xsl:param name="receiptNum" select="''"/> <xsl:copy> <xsl:copy-of select="./@*"/> <xsl:attribute name="Status"><xsl:value-of select="$status"/></xsl:attribute> <xsl:attribute name="DueAmount"><xsl:value-of select="$dueAmount"/></xsl:attribute> <xsl:if test="$receiptNum != ''"> <xsl:attribute name="ReceiptDate"> <xsl:value-of select="$receiptsAsc[position() = $receiptNum]/@ActionDate"/> </xsl:attribute> </xsl:if> <xsl:copy-of select="./*"/> </xsl:copy> </xsl:template>

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  • Debugging HTML & JavaScript with Firebug

    - by MattDiPasquale
    I made a JSONP widget. However, when one of the partner sites put it in their page, (1) it doesn't render at all in IE and (2) in other browsers (Firefox & Google Chrome), the HTML of the widget renders incorrectly: the <aside> closes prematurely, before the Financial Aid Glossary. It's something specific to that page because it works fine on this example college resource center page. To fix these two issues, I tried saving the page source to a local file and messing around with the local file and with Firebug, deleting DOM elements and stuff. I even tried fixing the errors that The W3C Markup Validation Service found. But, I still couldn't get it to render correctly. How should I tell them to change their page so that the widget renders correctly? Or, how should I update the widget script I wrote? They may take their page down since it's not rendering correctly, so here's the source of the page just in case: <!DOCTYPE html> <html> <head id="ctl01_Head1" profile="New Jersey Credit Union League"><title> College Resource Center - New Jersey Credit Union League </title> <link rel='stylesheet' type='text/css' href='http://ajax.googleapis.com/ajax/libs/jqueryui/1.8.6/themes/base/jquery.ui.all.css' /> <link rel='stylesheet' type='text/css' href='/csshandler.ashx?skin=InnerTemplate&amp;s=1&amp;v=2.3.5.8' /> <!--[if IE]> <script defer="defer" src="http://njcul.org/ClientScript/html5.js" type="text/javascript"></script> <![endif]--> <!--[if lt IE 7]> <link rel="stylesheet" href="http://njcul.org/Data/Sites/1/skins/InnerTemplate/IESpecific.css?cb=9d546eec-6752-4067-8f94-9a5b642213e4" type="text/css" id="IE6CSS" /> <![endif]--> <!--[if IE 7]> <link rel="stylesheet" href="http://njcul.org/Data/Sites/1/skins/InnerTemplate/IE7Specific.css?cb=9d546eec-6752-4067-8f94-9a5b642213e4" type="text/css" id="IE7CSS" /> <![endif]--> <meta http-equiv="Content-Type" content="text/html; charset=utf-8" /> <link rel="search" type="application/opensearchdescription+xml" title="New Jersey Credit Union League" href="http://njcul.org/SearchEngineInfo.ashx" /> <!--[if IE]> <meta http-equiv="Page-Enter" content="blendTrans(Duration=0)" /><meta http-equiv="Page-Exit" content="blendTrans(Duration=0)" /> <![endif]--> <meta name="viewport" content="width=670, initial-scale=0.45, minimum-scale=0.45" /> <link rel='shortcut icon' href='http://njcul.org/Data/Sites/1/skins/InnerTemplate/favicon.ico' /> <script src="http://ajax.googleapis.com/ajax/libs/jquery/1.4.4/jquery.min.js" type="text/javascript" ></script> <script src="http://ajax.googleapis.com/ajax/libs/jqueryui/1.8.6/jquery-ui.min.js" type="text/javascript" ></script> <script type="text/javascript"> <!-- function MM_swapImgRestore() { //v3.0 var i,x,a=document.MM_sr; for(i=0;a&&i<a.length&&(x=a[i])&&x.oSrc;i++) x.src=x.oSrc; } function MM_preloadImages() { //v3.0 var d=document; if(d.images){ if(!d.MM_p) d.MM_p=new Array(); var i,j=d.MM_p.length,a=MM_preloadImages.arguments; for(i=0; i<a.length; i++) if (a[i].indexOf("#")!=0){ d.MM_p[j]=new Image; d.MM_p[j++].src=a[i];}} } function MM_findObj(n, d) { //v4.01 var p,i,x; if(!d) d=document; if((p=n.indexOf("?"))>0&&parent.frames.length) { d=parent.frames[n.substring(p+1)].document; n=n.substring(0,p);} if(!(x=d[n])&&d.all) x=d.all[n]; for (i=0;!x&&i<d.forms.length;i++) x=d.forms[i][n]; for(i=0;!x&&d.layers&&i<d.layers.length;i++) x=MM_findObj(n,d.layers[i].document); if(!x && d.getElementById) x=d.getElementById(n); return x; } function MM_swapImage() { //v3.0 var i,j=0,x,a=MM_swapImage.arguments; document.MM_sr=new Array; for(i=0;i<(a.length-2);i+=3) if ((x=MM_findObj(a[i]))!=null){document.MM_sr[j++]=x; if(!x.oSrc) x.oSrc=x.src; x.src=a[i+2];} } //--> </script> <link href="App_Themes/pageskin/theme.css" type="text/css" rel="stylesheet" /> <link rel='canonical' href='http://njcul.org/college-resource-center.aspx' /><style type="text/css"> .ctl01_SiteMenu1_ctl00_0 { background-color:white;visibility:hidden;display:none;position:absolute;left:0px;top:0px; } .ctl01_SiteMenu1_ctl00_1 { text-decoration:none; } .ctl01_SiteMenu1_ctl00_2 { } .ctl01_PageMenu1_ctl01_0 { background-color:white;visibility:hidden;display:none;position:absolute;left:0px;top:0px; } .ctl01_PageMenu1_ctl01_1 { text-decoration:none; } .ctl01_PageMenu1_ctl01_2 { } .ctl01_PageMenu2_ctl01_0 { text-decoration:none; } </style></head> <body class="pagebody" onLoad="MM_preloadImages('ps_menu_down.png')"> <form method="post" action="/college-resource-center.aspx" onsubmit="javascript:return WebForm_OnSubmit();" id="aspnetForm"> <div> <input type="hidden" name="ctl01_ScriptManager1_HiddenField" id="ctl01_ScriptManager1_HiddenField" value="" /> <input type="hidden" name="__EVENTTARGET" id="__EVENTTARGET" value="" /> <input type="hidden" name="__EVENTARGUMENT" 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$("#toolbarbut").fadeOut("slow");} $(document).ready(function(){ $("span.downarr a").click(function() {HideMenuToolbar(); Set_Cookie('openstate', 'closed')}); $("span.showbar a").click(function() {ShowMenuToolbar(); Set_Cookie('openstate', 'open') }); $("span.downarr a, span.showbar a").click(function() { return false; }); var openState = Get_Cookie('openstate'); if(openState != null){ if(openState == 'closed'){HideMenuToolbar();} if(openState == 'open'){ShowMenuToolbar();}} }); </script> <div> <input type="hidden" name="ctl01$ctl06" id="ctl01_ctl06" /> </div> <div> <input type="hidden" name="__EVENTVALIDATION" id="__EVENTVALIDATION" value="/wEWBQKv1e3VCALs75XzDgL+qaz3AwLv26TNCQKS/MC2Dg==" /> </div> <script type="text/javascript">Sys.Application.add_load(function() { var form = Sys.WebForms.PageRequestManager.getInstance()._form; form._initialAction = form.action = window.location.href; }); </script> <script type="text/javascript"> //<![CDATA[ (function() {var fn = function() {$get("ctl01_ScriptManager1_HiddenField").value = '';Sys.Application.remove_init(fn);};Sys.Application.add_init(fn);})(); WebForm_InitCallback();//]]> </script> <script type="text/javascript" > $('div.mojo-accordion').accordion({fx:{opacity:'toggle',duration:'fast'}}); $('div.mojo-accordion-nh').accordion({fx:{opacity:'toggle',duration:'fast'},autoHeight:false}); $('div.mojo-tabs').tabs({fx:{opacity:'toggle',duration:'fast'}}); $('input.jqbutton').button(); </script> <script type="text/javascript">$('#ctl01_spanel1').cycle({fx:'fade',speed:1000,timeout:3000,next:'#ctl01_spanel1'});</script> <script type="text/javascript"> var gaJsHost = (("https:" == document.location.protocol) ? "https://ssl." : "http://www."); document.write(unescape("%3Cscript src='" + gaJsHost + "google-analytics.com/ga.js' type='text/javascript'%3E%3C/script%3E")); </script> <script type="text/javascript"> try{ var mojoPageTracker = _gat._getTracker("UA-19333588-1"); mojoPageTracker._setCustomVar(1, "member-type", "anonymous", 1);mojoPageTracker._trackPageview(); } catch(err) {} </script> <script type="text/javascript"> //<![CDATA[ Sys.Application.initialize(); //]]> </script> </form> </body> </html>

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  • FCKEditor doesn't set Value property on postback!

    - by Shaul
    I'm using FCKEditor on my asp.net web page. It appears beautifully, and the editor looks really good on the front end. Only problem is, the .Value property is not being set on the postback. No matter what changes the user makes to the value of the control on the page, when I click "Submit", the .Value property remains blank. I have Googled for other solutions, and most of them are of the variety where there's some conflict with Ajax, such as this and this. My problem is not solved by these solutions; it's much more fundamental than that. I'm not doing anything to do with Ajax; I'm just a simple asp.net newbie with a simple web form, and the value property is not being set on postback, not in IE and not in FF. It appears that at least one other person has had this problem, but no solution yet. Any ideas? Thanks! New information: I tried this out on a "hello world" test web site - and the test web site works 100%. There is obviously a problem on my page, but I have no idea where to begin tracking this down. Here's the markup of my page, in case anyone can see anything obvious that my newbie eyes can't: <%@ Page Language="vb" AutoEventWireup="false" CodeBehind="EmailTemplateEditForm.aspx.vb" Inherits="EEI_App.EmailTemplateEditForm" %> <%@ Register Assembly="FredCK.FCKeditorV2" Namespace="FredCK.FCKeditorV2" TagPrefix="FCKeditorV2" %> <!DOCTYPE html PUBLIC "-//W3C//DTD XHTML 1.0 Transitional//EN" "http://www.w3.org/TR/xhtml1/DTD/xhtml1-transitional.dtd"> <html xmlns="http://www.w3.org/1999/xhtml"> <head runat="server"> <title>EEI - Email Template</title> <link rel="stylesheet" href="EEI.css"> <script language="javascript" id="jssembleWare" src="sembleWare.js"></script> <style type="text/css"> .style1 { height: 251px; } .style2 { width: 2%; height: 251px; } .style3 { height: 490px; } </style> </head> <body> <form id="form1" runat="server"> <%@ register src="header.ascx" tagname="header" tagprefix="uc1" %> <%@ register src="footer.ascx" tagname="footer" tagprefix="uc1" %> <uc1:header ID="header1" runat="server" /> <!-- main content area --> <div class="content"> <!-- title of the page --> <div class="boxheader"> Email Template </div> <div class="standardbox"> <!-- Start Page Main Contents--> <!-- error messages --> <div class="errorbox"> <asp:Label ID="lblError" CssClass="ErrorControlStyle" runat="server" EnableViewState="False" Width="100%"></asp:Label> </div> <table class="contenttable"> <tr> <td align="left" valign="top" class="style3"> <div class="actionbox"> <div class="navheadertitle"> Navigation</div> <ul> <li> <asp:LinkButton ID="btnSubmit" CssClass="LinkButtonStyle" runat="server">Submit</asp:LinkButton> </li> <li> <asp:LinkButton ID="btnCancel" CssClass="LinkButtonStyle" runat="server" CausesValidation="false">Cancel</asp:LinkButton> </li> </ul> </div> </td> <td align="left" valign="top" class="style3"> <p> </p> <table> <tr class="MCRSFieldRow"> <td class="MCRSFieldLabelCell"> <asp:Label ID="lblEmailTemplate_TemplateName" CssClass="LabelStyle" runat="server" Width="175">Template Name</asp:Label> </td> <td class="MCRSFieldEditCell"> <asp:TextBox ID="txtEmailTemplate_TemplateName" CssClass="TextBoxStyle" runat="server" Width="100%"></asp:TextBox> </td> <td class="MCRSFieldLabelCell"> <asp:Label ID="lblEmailTemplate_TemplateType" CssClass="LabelStyle" runat="server" Width="175">Template Type</asp:Label> </td> <td class="MCRSFieldEditCell"> <asp:RadioButtonList ID="rblEmailTemplate_TemplateType" CssClass="RadioButtonListStyle" runat="server" RepeatColumns="1" RepeatDirection="Horizontal" Width="135px"> <asp:ListItem Value="1">Cover Letter</asp:ListItem> <asp:ListItem Value="2">Email</asp:ListItem> </asp:RadioButtonList> </td> <td class="MCRSRowRightCell"> &nbsp; </td> </tr> <tr class="MCRSFieldRow"> <td class="MCRSFieldLabelCell"> Composition Date </td> <td class="MCRSFieldEditCell"> <asp:Label ID="lblEmailTemplate_CompositionDate" CssClass="ElementLabelStyle" runat="server" Width="175"></asp:Label> </td> <td class="MCRSFieldLabelCell"> Last Used Date </td> <td class="MCRSFieldEditCell"> <asp:Label ID="lblEmailTemplate_LastUsedDate" CssClass="ElementLabelStyle" runat="server" Width="175"></asp:Label> </td> <td class="MCRSRowRightCell"> &nbsp; </td> </tr> <tr class="MCRSFieldRow"> <td class="MCRSFieldLabelCell"> Composed By </td> <td class="MCRSFieldEditCell" colspan="3"> <asp:Label ID="lblPerson_FirstNames" CssClass="ElementLabelStyle" runat="server"></asp:Label> <asp:Label ID="lblPerson_LastName" CssClass="ElementLabelStyle" runat="server"></asp:Label> </td> <td class="MCRSRowRightCell"> &nbsp; </td> </tr> <tr class="MCRSFieldRow"> <td class="MCRSFieldLabelCell"> <asp:Label ID="lblEmailTemplate_Subject" CssClass="LabelStyle" runat="server" Width="175">Subject</asp:Label> </td> <td class="MCRSFieldEditCell" colspan="3"> <asp:TextBox ID="txtEmailTemplate_Subject" CssClass="TextBoxStyle" runat="server" Width="100%"></asp:TextBox> </td> <td class="MCRSRowRightCell"> &nbsp; </td> </tr> <tr class="MCRSFieldRow"> <td class="style1"> <asp:Label ID="lblEmailTemplate_Body" CssClass="LabelStyle" runat="server" Width="175">Body</asp:Label> </td> <td class="style1" colspan="3"> <FCKeditorV2:FCKeditor ID="FCKeditor1" runat="server" Height="500px"> </FCKeditorV2:FCKeditor> </td> <td class="style2"> &nbsp; </td> </tr> </table> </td> </tr> </table> </div> <p> <a class="InputButtonStyle" href="#_swTopOfPage">Top of Page</a> </p> </div> <uc1:footer ID="footer1" runat="server" /> <p> <asp:TextBox ID="txtEmailTemplate_Body" CssClass="TextAreaStyle" Rows="4" runat="server" Width="100%" Height="16px" Visible="False"></asp:TextBox> </p> </form> </body> </html>

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  • css menu <ul><li> dinamically centered or width of buttons that covers the whole page

    - by Tony Stark
    I am building a home page for my minecraft server. Probably in the following 4-6 months I will opend my second and this is why I am in trouble. My first site is 1000 pixel wide, and the second will be 1200. First big difference. My menus are dinamically generated by my php code. It checks on my databases if there is another button or it is over. These buttons can be added or removed directly online. Another big issue is the browser compatibility. In a survey I did on our previous server I had a lot of users using: chrome, internet explorer, safari and firefox. That means that I must find a solution that is compatible with most browsers. What do I have to do? I came up with this CSS, which is touch compatible, it allows menus to be swapped to the left and it is enough to set 1 parameter to fix it for every page width. Sadly it is left aligned. body, nav, ul, li, a {margin: 0; padding: 0;} body {font-family: Verdana,"Helvetica Neue", Helvetica, Arial, sans-serif; } a {text-decoration: none;} .container { max-width: 900px; margin: 0px auto 0px auto; } .toggleMenu { display: none; background: #666; padding: 10px 15px; color: #999999; } .nav { border: 1px solid #424242; background-color: #121212; filter: progid:DXImageTransform.Microsoft.gradient(startColorstr='#686868', endColorstr='#121212'); background-image: -moz-linear-gradient(#686868, #121212); background-image: -webkit-gradient(linear, left top, left bottom, from(#686868), to(#121212)); background-image: -webkit-linear-gradient(#686868, #121212); background-image: -o-linear-gradient(#686868, #121212); background-image: -ms-linear-gradient(#686868, #121212); background-image: linear-gradient(#686868, #121212); -moz-box-shadow: 0 1px 1px #777, 0 1px 0 #666 inset; -webkit-box-shadow: 0 1px 1px #777, 0 1px 0 #666 inset; box-shadow: 0 1px 1px #777, 0 1px 0 #666 inset; list-style: none; *zoom: 1; position: relative; } .nav:before,.nav:after { content: " "; display: table; } .nav:after { clear: both; } .nav ul { list-style: none; width: 11em; z-index: 1; background-color: #121212; -moz-box-shadow: 0 -1px rgba(255,255,255,.3); -webkit-box-shadow: 0 -1px 0 rgba(255,255,255,.3); box-shadow: 0 -1px 0 rgba(255,255,255,.3); } .nav a { padding: 10px 15px; color:#999999; text-transform: uppercase; font: bold 11px Arial, Helvetica; text-decoration: none; text-shadow: 0 1px 0 #000; *zoom: 1; } .nav a:hover{ color:#000000; background-color: #B2B2B2; filter: progid:DXImageTransform.Microsoft.gradient(startColorstr='#D3D3D3', endColorstr='#B2B2B2'); background-image: -moz-linear-gradient(#D3D3D3, #B2B2B2); background-image: -webkit-gradient(linear, left top, left bottom, from(#D3D3D3), to(#B2B2B2)); background-image: -webkit-linear-gradient(#D3D3D3, #B2B2B2); background-image: -o-linear-gradient(#D3D3D3, #B2B2B2); background-image: -ms-linear-gradient(#D3D3D3, #B2B2B2); background-image: linear-gradient(#D3D3D3, #B2B2B2); } /*Delimitazione di ogni tab | HOME | */ .nav li { position: relative; border-right: 1px solid #424242; -moz-box-shadow: 1px 0 0 #686868; -webkit-box-shadow: 1px 0 0 #686868; box-shadow: 1px 0 0 #686868; } .nav > li { float: left; border-top: 1px solid #424242; z-index: 200; } .nav > li > .parent { background-image: url("../downArrow.png"); background-repeat: no-repeat; background-position: center right; } .nav > li li > .parent { background-image: url("../rightArrow.png"); background-repeat: no-repeat; background-position: center right; } .nav > li > a { display: block; } .nav li ul { position: absolute; left: -9999px; z-index: 100; } /* freccetta che indica un sottomenu nell'ultimo tab */ .nav > li:last-child li > .parent{ background-image: url("../leftArrow.png"); background-repeat: no-repeat; background-position: left; } /*flip subsubmenu*/ .nav li.last.hover > ul { left:auto; right: 0; } .nav > li.hover > ul { left: 0; } .nav li li.hover > ul { left: 100%; top: 0; } /* Spostare il 2^ sottomenu a sinistra */ .nav li.last li.hover ul { left:auto; right: 100%; top:0; } .nav li li a { display: block; background-color: #686868; -moz-box-shadow: 0 -1px rgba(255,255,255,.3); -webkit-box-shadow: 0 -1px 0 rgba(255,255,255,.3); box-shadow: 0 -1px 0 rgba(255,255,255,.3); z-index:100; border-top: 1px solid #686868; } .nav li li li a { background-color: #686868; -moz-box-shadow: 0 -1px rgba(255,255,255,.3); -webkit-box-shadow: 0 -1px 0 rgba(255,255,255,.3); box-shadow: 0 -1px 0 rgba(255,255,255,.3); z-index:200; border-top: 1px solid #686868; } .nav li li li li a { display: block; background-color: #686868; -moz-box-shadow: 0 -1px rgba(255,255,255,.3); -webkit-box-shadow: 0 -1px 0 rgba(255,255,255,.3); box-shadow: 0 -1px 0 rgba(255,255,255,.3); z-index:300; border-top: 1px solid #686868; } .nav li li li li a { background-color: #686868; -moz-box-shadow: 0 -1px rgba(255,255,255,.3); -webkit-box-shadow: 0 -1px 0 rgba(255,255,255,.3); box-shadow: 0 -1px 0 rgba(255,255,255,.3); z-index:400; border-top: 1px solid #686868; } @media screen and (max-width: 768px) { .active { display: block; } .nav > li { float: none; } .nav > li > .parent { background-position: 95% 50%; } .nav li li .parent { background-image: url("../downArrow.png"); background-repeat: no-repeat; background-position: 95% 50%; } .nav ul { display: block; width: 100%; } .nav > li.hover > ul , .nav li li.hover ul { position: static; } } My girlfriend (who adapted this code) is really busy for school and cannot help me. Leaving the borders on the whole square (page width), is it possible to make buttons cover the page width dinamically? Or is it possible to center the buttons? Thank you very much!

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  • Zen and the Art of File and Folder Organization

    - by Mark Virtue
    Is your desk a paragon of neatness, or does it look like a paper-bomb has gone off? If you’ve been putting off getting organized because the task is too huge or daunting, or you don’t know where to start, we’ve got 40 tips to get you on the path to zen mastery of your filing system. For all those readers who would like to get their files and folders organized, or, if they’re already organized, better organized—we have compiled a complete guide to getting organized and staying organized, a comprehensive article that will hopefully cover every possible tip you could want. Signs that Your Computer is Poorly Organized If your computer is a mess, you’re probably already aware of it.  But just in case you’re not, here are some tell-tale signs: Your Desktop has over 40 icons on it “My Documents” contains over 300 files and 60 folders, including MP3s and digital photos You use the Windows’ built-in search facility whenever you need to find a file You can’t find programs in the out-of-control list of programs in your Start Menu You save all your Word documents in one folder, all your spreadsheets in a second folder, etc Any given file that you’re looking for may be in any one of four different sets of folders But before we start, here are some quick notes: We’re going to assume you know what files and folders are, and how to create, save, rename, copy and delete them The organization principles described in this article apply equally to all computer systems.  However, the screenshots here will reflect how things look on Windows (usually Windows 7).  We will also mention some useful features of Windows that can help you get organized. Everyone has their own favorite methodology of organizing and filing, and it’s all too easy to get into “My Way is Better than Your Way” arguments.  The reality is that there is no perfect way of getting things organized.  When I wrote this article, I tried to keep a generalist and objective viewpoint.  I consider myself to be unusually well organized (to the point of obsession, truth be told), and I’ve had 25 years experience in collecting and organizing files on computers.  So I’ve got a lot to say on the subject.  But the tips I have described here are only one way of doing it.  Hopefully some of these tips will work for you too, but please don’t read this as any sort of “right” way to do it. At the end of the article we’ll be asking you, the reader, for your own organization tips. Why Bother Organizing At All? For some, the answer to this question is self-evident. And yet, in this era of powerful desktop search software (the search capabilities built into the Windows Vista and Windows 7 Start Menus, and third-party programs like Google Desktop Search), the question does need to be asked, and answered. I have a friend who puts every file he ever creates, receives or downloads into his My Documents folder and doesn’t bother filing them into subfolders at all.  He relies on the search functionality built into his Windows operating system to help him find whatever he’s looking for.  And he always finds it.  He’s a Search Samurai.  For him, filing is a waste of valuable time that could be spent enjoying life! It’s tempting to follow suit.  On the face of it, why would anyone bother to take the time to organize their hard disk when such excellent search software is available?  Well, if all you ever want to do with the files you own is to locate and open them individually (for listening, editing, etc), then there’s no reason to ever bother doing one scrap of organization.  But consider these common tasks that are not achievable with desktop search software: Find files manually.  Often it’s not convenient, speedy or even possible to utilize your desktop search software to find what you want.  It doesn’t work 100% of the time, or you may not even have it installed.  Sometimes its just plain faster to go straight to the file you want, if you know it’s in a particular sub-folder, rather than trawling through hundreds of search results. Find groups of similar files (e.g. all your “work” files, all the photos of your Europe holiday in 2008, all your music videos, all the MP3s from Dark Side of the Moon, all your letters you wrote to your wife, all your tax returns).  Clever naming of the files will only get you so far.  Sometimes it’s the date the file was created that’s important, other times it’s the file format, and other times it’s the purpose of the file.  How do you name a collection of files so that they’re easy to isolate based on any of the above criteria?  Short answer, you can’t. Move files to a new computer.  It’s time to upgrade your computer.  How do you quickly grab all the files that are important to you?  Or you decide to have two computers now – one for home and one for work.  How do you quickly isolate only the work-related files to move them to the work computer? Synchronize files to other computers.  If you have more than one computer, and you need to mirror some of your files onto the other computer (e.g. your music collection), then you need a way to quickly determine which files are to be synced and which are not.  Surely you don’t want to synchronize everything? Choose which files to back up.  If your backup regime calls for multiple backups, or requires speedy backups, then you’ll need to be able to specify which files are to be backed up, and which are not.  This is not possible if they’re all in the same folder. Finally, if you’re simply someone who takes pleasure in being organized, tidy and ordered (me! me!), then you don’t even need a reason.  Being disorganized is simply unthinkable. Tips on Getting Organized Here we present our 40 best tips on how to get organized.  Or, if you’re already organized, to get better organized. Tip #1.  Choose Your Organization System Carefully The reason that most people are not organized is that it takes time.  And the first thing that takes time is deciding upon a system of organization.  This is always a matter of personal preference, and is not something that a geek on a website can tell you.  You should always choose your own system, based on how your own brain is organized (which makes the assumption that your brain is, in fact, organized). We can’t instruct you, but we can make suggestions: You may want to start off with a system based on the users of the computer.  i.e. “My Files”, “My Wife’s Files”, My Son’s Files”, etc.  Inside “My Files”, you might then break it down into “Personal” and “Business”.  You may then realize that there are overlaps.  For example, everyone may want to share access to the music library, or the photos from the school play.  So you may create another folder called “Family”, for the “common” files. You may decide that the highest-level breakdown of your files is based on the “source” of each file.  In other words, who created the files.  You could have “Files created by ME (business or personal)”, “Files created by people I know (family, friends, etc)”, and finally “Files created by the rest of the world (MP3 music files, downloaded or ripped movies or TV shows, software installation files, gorgeous desktop wallpaper images you’ve collected, etc).”  This system happens to be the one I use myself.  See below:  Mark is for files created by meVC is for files created by my company (Virtual Creations)Others is for files created by my friends and familyData is the rest of the worldAlso, Settings is where I store the configuration files and other program data files for my installed software (more on this in tip #34, below). Each folder will present its own particular set of requirements for further sub-organization.  For example, you may decide to organize your music collection into sub-folders based on the artist’s name, while your digital photos might get organized based on the date they were taken.  It can be different for every sub-folder! Another strategy would be based on “currentness”.  Files you have yet to open and look at live in one folder.  Ones that have been looked at but not yet filed live in another place.  Current, active projects live in yet another place.  All other files (your “archive”, if you like) would live in a fourth folder. (And of course, within that last folder you’d need to create a further sub-system based on one of the previous bullet points). Put some thought into this – changing it when it proves incomplete can be a big hassle!  Before you go to the trouble of implementing any system you come up with, examine a wide cross-section of the files you own and see if they will all be able to find a nice logical place to sit within your system. Tip #2.  When You Decide on Your System, Stick to It! There’s nothing more pointless than going to all the trouble of creating a system and filing all your files, and then whenever you create, receive or download a new file, you simply dump it onto your Desktop.  You need to be disciplined – forever!  Every new file you get, spend those extra few seconds to file it where it belongs!  Otherwise, in just a month or two, you’ll be worse off than before – half your files will be organized and half will be disorganized – and you won’t know which is which! Tip #3.  Choose the Root Folder of Your Structure Carefully Every data file (document, photo, music file, etc) that you create, own or is important to you, no matter where it came from, should be found within one single folder, and that one single folder should be located at the root of your C: drive (as a sub-folder of C:\).  In other words, do not base your folder structure in standard folders like “My Documents”.  If you do, then you’re leaving it up to the operating system engineers to decide what folder structure is best for you.  And every operating system has a different system!  In Windows 7 your files are found in C:\Users\YourName, whilst on Windows XP it was C:\Documents and Settings\YourName\My Documents.  In UNIX systems it’s often /home/YourName. These standard default folders tend to fill up with junk files and folders that are not at all important to you.  “My Documents” is the worst offender.  Every second piece of software you install, it seems, likes to create its own folder in the “My Documents” folder.  These folders usually don’t fit within your organizational structure, so don’t use them!  In fact, don’t even use the “My Documents” folder at all.  Allow it to fill up with junk, and then simply ignore it.  It sounds heretical, but: Don’t ever visit your “My Documents” folder!  Remove your icons/links to “My Documents” and replace them with links to the folders you created and you care about! Create your own file system from scratch!  Probably the best place to put it would be on your D: drive – if you have one.  This way, all your files live on one drive, while all the operating system and software component files live on the C: drive – simply and elegantly separated.  The benefits of that are profound.  Not only are there obvious organizational benefits (see tip #10, below), but when it comes to migrate your data to a new computer, you can (sometimes) simply unplug your D: drive and plug it in as the D: drive of your new computer (this implies that the D: drive is actually a separate physical disk, and not a partition on the same disk as C:).  You also get a slight speed improvement (again, only if your C: and D: drives are on separate physical disks). Warning:  From tip #12, below, you will see that it’s actually a good idea to have exactly the same file system structure – including the drive it’s filed on – on all of the computers you own.  So if you decide to use the D: drive as the storage system for your own files, make sure you are able to use the D: drive on all the computers you own.  If you can’t ensure that, then you can still use a clever geeky trick to store your files on the D: drive, but still access them all via the C: drive (see tip #17, below). If you only have one hard disk (C:), then create a dedicated folder that will contain all your files – something like C:\Files.  The name of the folder is not important, but make it a single, brief word. There are several reasons for this: When creating a backup regime, it’s easy to decide what files should be backed up – they’re all in the one folder! If you ever decide to trade in your computer for a new one, you know exactly which files to migrate You will always know where to begin a search for any file If you synchronize files with other computers, it makes your synchronization routines very simple.   It also causes all your shortcuts to continue to work on the other machines (more about this in tip #24, below). Once you’ve decided where your files should go, then put all your files in there – Everything!  Completely disregard the standard, default folders that are created for you by the operating system (“My Music”, “My Pictures”, etc).  In fact, you can actually relocate many of those folders into your own structure (more about that below, in tip #6). The more completely you get all your data files (documents, photos, music, etc) and all your configuration settings into that one folder, then the easier it will be to perform all of the above tasks. Once this has been done, and all your files live in one folder, all the other folders in C:\ can be thought of as “operating system” folders, and therefore of little day-to-day interest for us. Here’s a screenshot of a nicely organized C: drive, where all user files are located within the \Files folder:   Tip #4.  Use Sub-Folders This would be our simplest and most obvious tip.  It almost goes without saying.  Any organizational system you decide upon (see tip #1) will require that you create sub-folders for your files.  Get used to creating folders on a regular basis. Tip #5.  Don’t be Shy About Depth Create as many levels of sub-folders as you need.  Don’t be scared to do so.  Every time you notice an opportunity to group a set of related files into a sub-folder, do so.  Examples might include:  All the MP3s from one music CD, all the photos from one holiday, or all the documents from one client. It’s perfectly okay to put files into a folder called C:\Files\Me\From Others\Services\WestCo Bank\Statements\2009.  That’s only seven levels deep.  Ten levels is not uncommon.  Of course, it’s possible to take this too far.  If you notice yourself creating a sub-folder to hold only one file, then you’ve probably become a little over-zealous.  On the other hand, if you simply create a structure with only two levels (for example C:\Files\Work) then you really haven’t achieved any level of organization at all (unless you own only six files!).  Your “Work” folder will have become a dumping ground, just like your Desktop was, with most likely hundreds of files in it. Tip #6.  Move the Standard User Folders into Your Own Folder Structure Most operating systems, including Windows, create a set of standard folders for each of its users.  These folders then become the default location for files such as documents, music files, digital photos and downloaded Internet files.  In Windows 7, the full list is shown below: Some of these folders you may never use nor care about (for example, the Favorites folder, if you’re not using Internet Explorer as your browser).  Those ones you can leave where they are.  But you may be using some of the other folders to store files that are important to you.  Even if you’re not using them, Windows will still often treat them as the default storage location for many types of files.  When you go to save a standard file type, it can become annoying to be automatically prompted to save it in a folder that’s not part of your own file structure. But there’s a simple solution:  Move the folders you care about into your own folder structure!  If you do, then the next time you go to save a file of the corresponding type, Windows will prompt you to save it in the new, moved location. Moving the folders is easy.  Simply drag-and-drop them to the new location.  Here’s a screenshot of the default My Music folder being moved to my custom personal folder (Mark): Tip #7.  Name Files and Folders Intelligently This is another one that almost goes without saying, but we’ll say it anyway:  Do not allow files to be created that have meaningless names like Document1.doc, or folders called New Folder (2).  Take that extra 20 seconds and come up with a meaningful name for the file/folder – one that accurately divulges its contents without repeating the entire contents in the name. Tip #8.  Watch Out for Long Filenames Another way to tell if you have not yet created enough depth to your folder hierarchy is that your files often require really long names.  If you need to call a file Johnson Sales Figures March 2009.xls (which might happen to live in the same folder as Abercrombie Budget Report 2008.xls), then you might want to create some sub-folders so that the first file could be simply called March.xls, and living in the Clients\Johnson\Sales Figures\2009 folder. A well-placed file needs only a brief filename! Tip #9.  Use Shortcuts!  Everywhere! This is probably the single most useful and important tip we can offer.  A shortcut allows a file to be in two places at once. Why would you want that?  Well, the file and folder structure of every popular operating system on the market today is hierarchical.  This means that all objects (files and folders) always live within exactly one parent folder.  It’s a bit like a tree.  A tree has branches (folders) and leaves (files).  Each leaf, and each branch, is supported by exactly one parent branch, all the way back to the root of the tree (which, incidentally, is exactly why C:\ is called the “root folder” of the C: drive). That hard disks are structured this way may seem obvious and even necessary, but it’s only one way of organizing data.  There are others:  Relational databases, for example, organize structured data entirely differently.  The main limitation of hierarchical filing structures is that a file can only ever be in one branch of the tree – in only one folder – at a time.  Why is this a problem?  Well, there are two main reasons why this limitation is a problem for computer users: The “correct” place for a file, according to our organizational rationale, is very often a very inconvenient place for that file to be located.  Just because it’s correctly filed doesn’t mean it’s easy to get to.  Your file may be “correctly” buried six levels deep in your sub-folder structure, but you may need regular and speedy access to this file every day.  You could always move it to a more convenient location, but that would mean that you would need to re-file back to its “correct” location it every time you’d finished working on it.  Most unsatisfactory. A file may simply “belong” in two or more different locations within your file structure.  For example, say you’re an accountant and you have just completed the 2009 tax return for John Smith.  It might make sense to you to call this file 2009 Tax Return.doc and file it under Clients\John Smith.  But it may also be important to you to have the 2009 tax returns from all your clients together in the one place.  So you might also want to call the file John Smith.doc and file it under Tax Returns\2009.  The problem is, in a purely hierarchical filing system, you can’t put it in both places.  Grrrrr! Fortunately, Windows (and most other operating systems) offers a way for you to do exactly that:  It’s called a “shortcut” (also known as an “alias” on Macs and a “symbolic link” on UNIX systems).  Shortcuts allow a file to exist in one place, and an icon that represents the file to be created and put anywhere else you please.  In fact, you can create a dozen such icons and scatter them all over your hard disk.  Double-clicking on one of these icons/shortcuts opens up the original file, just as if you had double-clicked on the original file itself. Consider the following two icons: The one on the left is the actual Word document, while the one on the right is a shortcut that represents the Word document.  Double-clicking on either icon will open the same file.  There are two main visual differences between the icons: The shortcut will have a small arrow in the lower-left-hand corner (on Windows, anyway) The shortcut is allowed to have a name that does not include the file extension (the “.docx” part, in this case) You can delete the shortcut at any time without losing any actual data.  The original is still intact.  All you lose is the ability to get to that data from wherever the shortcut was. So why are shortcuts so great?  Because they allow us to easily overcome the main limitation of hierarchical file systems, and put a file in two (or more) places at the same time.  You will always have files that don’t play nice with your organizational rationale, and can’t be filed in only one place.  They demand to exist in two places.  Shortcuts allow this!  Furthermore, they allow you to collect your most often-opened files and folders together in one spot for convenient access.  The cool part is that the original files stay where they are, safe forever in their perfectly organized location. So your collection of most often-opened files can – and should – become a collection of shortcuts! If you’re still not convinced of the utility of shortcuts, consider the following well-known areas of a typical Windows computer: The Start Menu (and all the programs that live within it) The Quick Launch bar (or the Superbar in Windows 7) The “Favorite folders” area in the top-left corner of the Windows Explorer window (in Windows Vista or Windows 7) Your Internet Explorer Favorites or Firefox Bookmarks Each item in each of these areas is a shortcut!  Each of those areas exist for one purpose only:  For convenience – to provide you with a collection of the files and folders you access most often. It should be easy to see by now that shortcuts are designed for one single purpose:  To make accessing your files more convenient.  Each time you double-click on a shortcut, you are saved the hassle of locating the file (or folder, or program, or drive, or control panel icon) that it represents. Shortcuts allow us to invent a golden rule of file and folder organization: “Only ever have one copy of a file – never have two copies of the same file.  Use a shortcut instead” (this rule doesn’t apply to copies created for backup purposes, of course!) There are also lesser rules, like “don’t move a file into your work area – create a shortcut there instead”, and “any time you find yourself frustrated with how long it takes to locate a file, create a shortcut to it and place that shortcut in a convenient location.” So how to we create these massively useful shortcuts?  There are two main ways: “Copy” the original file or folder (click on it and type Ctrl-C, or right-click on it and select Copy):  Then right-click in an empty area of the destination folder (the place where you want the shortcut to go) and select Paste shortcut: Right-drag (drag with the right mouse button) the file from the source folder to the destination folder.  When you let go of the mouse button at the destination folder, a menu pops up: Select Create shortcuts here. Note that when shortcuts are created, they are often named something like Shortcut to Budget Detail.doc (windows XP) or Budget Detail – Shortcut.doc (Windows 7).   If you don’t like those extra words, you can easily rename the shortcuts after they’re created, or you can configure Windows to never insert the extra words in the first place (see our article on how to do this). And of course, you can create shortcuts to folders too, not just to files! Bottom line: Whenever you have a file that you’d like to access from somewhere else (whether it’s convenience you’re after, or because the file simply belongs in two places), create a shortcut to the original file in the new location. Tip #10.  Separate Application Files from Data Files Any digital organization guru will drum this rule into you.  Application files are the components of the software you’ve installed (e.g. Microsoft Word, Adobe Photoshop or Internet Explorer).  Data files are the files that you’ve created for yourself using that software (e.g. Word Documents, digital photos, emails or playlists). Software gets installed, uninstalled and upgraded all the time.  Hopefully you always have the original installation media (or downloaded set-up file) kept somewhere safe, and can thus reinstall your software at any time.  This means that the software component files are of little importance.  Whereas the files you have created with that software is, by definition, important.  It’s a good rule to always separate unimportant files from important files. So when your software prompts you to save a file you’ve just created, take a moment and check out where it’s suggesting that you save the file.  If it’s suggesting that you save the file into the same folder as the software itself, then definitely don’t follow that suggestion.  File it in your own folder!  In fact, see if you can find the program’s configuration option that determines where files are saved by default (if it has one), and change it. Tip #11.  Organize Files Based on Purpose, Not on File Type If you have, for example a folder called Work\Clients\Johnson, and within that folder you have two sub-folders, Word Documents and Spreadsheets (in other words, you’re separating “.doc” files from “.xls” files), then chances are that you’re not optimally organized.  It makes little sense to organize your files based on the program that created them.  Instead, create your sub-folders based on the purpose of the file.  For example, it would make more sense to create sub-folders called Correspondence and Financials.  It may well be that all the files in a given sub-folder are of the same file-type, but this should be more of a coincidence and less of a design feature of your organization system. Tip #12.  Maintain the Same Folder Structure on All Your Computers In other words, whatever organizational system you create, apply it to every computer that you can.  There are several benefits to this: There’s less to remember.  No matter where you are, you always know where to look for your files If you copy or synchronize files from one computer to another, then setting up the synchronization job becomes very simple Shortcuts can be copied or moved from one computer to another with ease (assuming the original files are also copied/moved).  There’s no need to find the target of the shortcut all over again on the second computer Ditto for linked files (e.g Word documents that link to data in a separate Excel file), playlists, and any files that reference the exact file locations of other files. This applies even to the drive that your files are stored on.  If your files are stored on C: on one computer, make sure they’re stored on C: on all your computers.  Otherwise all your shortcuts, playlists and linked files will stop working! Tip #13.  Create an “Inbox” Folder Create yourself a folder where you store all files that you’re currently working on, or that you haven’t gotten around to filing yet.  You can think of this folder as your “to-do” list.  You can call it “Inbox” (making it the same metaphor as your email system), or “Work”, or “To-Do”, or “Scratch”, or whatever name makes sense to you.  It doesn’t matter what you call it – just make sure you have one! Once you have finished working on a file, you then move it from the “Inbox” to its correct location within your organizational structure. You may want to use your Desktop as this “Inbox” folder.  Rightly or wrongly, most people do.  It’s not a bad place to put such files, but be careful:  If you do decide that your Desktop represents your “to-do” list, then make sure that no other files find their way there.  In other words, make sure that your “Inbox”, wherever it is, Desktop or otherwise, is kept free of junk – stray files that don’t belong there. So where should you put this folder, which, almost by definition, lives outside the structure of the rest of your filing system?  Well, first and foremost, it has to be somewhere handy.  This will be one of your most-visited folders, so convenience is key.  Putting it on the Desktop is a great option – especially if you don’t have any other folders on your Desktop:  the folder then becomes supremely easy to find in Windows Explorer: You would then create shortcuts to this folder in convenient spots all over your computer (“Favorite Links”, “Quick Launch”, etc). Tip #14.  Ensure You have Only One “Inbox” Folder Once you’ve created your “Inbox” folder, don’t use any other folder location as your “to-do list”.  Throw every incoming or created file into the Inbox folder as you create/receive it.  This keeps the rest of your computer pristine and free of randomly created or downloaded junk.  The last thing you want to be doing is checking multiple folders to see all your current tasks and projects.  Gather them all together into one folder. Here are some tips to help ensure you only have one Inbox: Set the default “save” location of all your programs to this folder. Set the default “download” location for your browser to this folder. If this folder is not your desktop (recommended) then also see if you can make a point of not putting “to-do” files on your desktop.  This keeps your desktop uncluttered and Zen-like: (the Inbox folder is in the bottom-right corner) Tip #15.  Be Vigilant about Clearing Your “Inbox” Folder This is one of the keys to staying organized.  If you let your “Inbox” overflow (i.e. allow there to be more than, say, 30 files or folders in there), then you’re probably going to start feeling like you’re overwhelmed:  You’re not keeping up with your to-do list.  Once your Inbox gets beyond a certain point (around 30 files, studies have shown), then you’ll simply start to avoid it.  You may continue to put files in there, but you’ll be scared to look at it, fearing the “out of control” feeling that all overworked, chaotic or just plain disorganized people regularly feel. So, here’s what you can do: Visit your Inbox/to-do folder regularly (at least five times per day). Scan the folder regularly for files that you have completed working on and are ready for filing.  File them immediately. Make it a source of pride to keep the number of files in this folder as small as possible.  If you value peace of mind, then make the emptiness of this folder one of your highest (computer) priorities If you know that a particular file has been in the folder for more than, say, six weeks, then admit that you’re not actually going to get around to processing it, and move it to its final resting place. Tip #16.  File Everything Immediately, and Use Shortcuts for Your Active Projects As soon as you create, receive or download a new file, store it away in its “correct” folder immediately.  Then, whenever you need to work on it (possibly straight away), create a shortcut to it in your “Inbox” (“to-do”) folder or your desktop.  That way, all your files are always in their “correct” locations, yet you still have immediate, convenient access to your current, active files.  When you finish working on a file, simply delete the shortcut. Ideally, your “Inbox” folder – and your Desktop – should contain no actual files or folders.  They should simply contain shortcuts. Tip #17.  Use Directory Symbolic Links (or Junctions) to Maintain One Unified Folder Structure Using this tip, we can get around a potential hiccup that we can run into when creating our organizational structure – the issue of having more than one drive on our computer (C:, D:, etc).  We might have files we need to store on the D: drive for space reasons, and yet want to base our organized folder structure on the C: drive (or vice-versa). Your chosen organizational structure may dictate that all your files must be accessed from the C: drive (for example, the root folder of all your files may be something like C:\Files).  And yet you may still have a D: drive and wish to take advantage of the hundreds of spare Gigabytes that it offers.  Did you know that it’s actually possible to store your files on the D: drive and yet access them as if they were on the C: drive?  And no, we’re not talking about shortcuts here (although the concept is very similar). By using the shell command mklink, you can essentially take a folder that lives on one drive and create an alias for it on a different drive (you can do lots more than that with mklink – for a full rundown on this programs capabilities, see our dedicated article).  These aliases are called directory symbolic links (and used to be known as junctions).  You can think of them as “virtual” folders.  They function exactly like regular folders, except they’re physically located somewhere else. For example, you may decide that your entire D: drive contains your complete organizational file structure, but that you need to reference all those files as if they were on the C: drive, under C:\Files.  If that was the case you could create C:\Files as a directory symbolic link – a link to D:, as follows: mklink /d c:\files d:\ Or it may be that the only files you wish to store on the D: drive are your movie collection.  You could locate all your movie files in the root of your D: drive, and then link it to C:\Files\Media\Movies, as follows: mklink /d c:\files\media\movies d:\ (Needless to say, you must run these commands from a command prompt – click the Start button, type cmd and press Enter) Tip #18. Customize Your Folder Icons This is not strictly speaking an organizational tip, but having unique icons for each folder does allow you to more quickly visually identify which folder is which, and thus saves you time when you’re finding files.  An example is below (from my folder that contains all files downloaded from the Internet): To learn how to change your folder icons, please refer to our dedicated article on the subject. Tip #19.  Tidy Your Start Menu The Windows Start Menu is usually one of the messiest parts of any Windows computer.  Every program you install seems to adopt a completely different approach to placing icons in this menu.  Some simply put a single program icon.  Others create a folder based on the name of the software.  And others create a folder based on the name of the software manufacturer.  It’s chaos, and can make it hard to find the software you want to run. Thankfully we can avoid this chaos with useful operating system features like Quick Launch, the Superbar or pinned start menu items. Even so, it would make a lot of sense to get into the guts of the Start Menu itself and give it a good once-over.  All you really need to decide is how you’re going to organize your applications.  A structure based on the purpose of the application is an obvious candidate.  Below is an example of one such structure: In this structure, Utilities means software whose job it is to keep the computer itself running smoothly (configuration tools, backup software, Zip programs, etc).  Applications refers to any productivity software that doesn’t fit under the headings Multimedia, Graphics, Internet, etc. In case you’re not aware, every icon in your Start Menu is a shortcut and can be manipulated like any other shortcut (copied, moved, deleted, etc). With the Windows Start Menu (all version of Windows), Microsoft has decided that there be two parallel folder structures to store your Start Menu shortcuts.  One for you (the logged-in user of the computer) and one for all users of the computer.  Having two parallel structures can often be redundant:  If you are the only user of the computer, then having two parallel structures is totally redundant.  Even if you have several users that regularly log into the computer, most of your installed software will need to be made available to all users, and should thus be moved out of the “just you” version of the Start Menu and into the “all users” area. To take control of your Start Menu, so you can start organizing it, you’ll need to know how to access the actual folders and shortcut files that make up the Start Menu (both versions of it).  To find these folders and files, click the Start button and then right-click on the All Programs text (Windows XP users should right-click on the Start button itself): The Open option refers to the “just you” version of the Start Menu, while the Open All Users option refers to the “all users” version.  Click on the one you want to organize. A Windows Explorer window then opens with your chosen version of the Start Menu selected.  From there it’s easy.  Double-click on the Programs folder and you’ll see all your folders and shortcuts.  Now you can delete/rename/move until it’s just the way you want it. Note:  When you’re reorganizing your Start Menu, you may want to have two Explorer windows open at the same time – one showing the “just you” version and one showing the “all users” version.  You can drag-and-drop between the windows. Tip #20.  Keep Your Start Menu Tidy Once you have a perfectly organized Start Menu, try to be a little vigilant about keeping it that way.  Every time you install a new piece of software, the icons that get created will almost certainly violate your organizational structure. So to keep your Start Menu pristine and organized, make sure you do the following whenever you install a new piece of software: Check whether the software was installed into the “just you” area of the Start Menu, or the “all users” area, and then move it to the correct area. Remove all the unnecessary icons (like the “Read me” icon, the “Help” icon (you can always open the help from within the software itself when it’s running), the “Uninstall” icon, the link(s)to the manufacturer’s website, etc) Rename the main icon(s) of the software to something brief that makes sense to you.  For example, you might like to rename Microsoft Office Word 2010 to simply Word Move the icon(s) into the correct folder based on your Start Menu organizational structure And don’t forget:  when you uninstall a piece of software, the software’s uninstall routine is no longer going to be able to remove the software’s icon from the Start Menu (because you moved and/or renamed it), so you’ll need to remove that icon manually. Tip #21.  Tidy C:\ The root of your C: drive (C:\) is a common dumping ground for files and folders – both by the users of your computer and by the software that you install on your computer.  It can become a mess. There’s almost no software these days that requires itself to be installed in C:\.  99% of the time it can and should be installed into C:\Program Files.  And as for your own files, well, it’s clear that they can (and almost always should) be stored somewhere else. In an ideal world, your C:\ folder should look like this (on Windows 7): Note that there are some system files and folders in C:\ that are usually and deliberately “hidden” (such as the Windows virtual memory file pagefile.sys, the boot loader file bootmgr, and the System Volume Information folder).  Hiding these files and folders is a good idea, as they need to stay where they are and are almost never needed to be opened or even seen by you, the user.  Hiding them prevents you from accidentally messing with them, and enhances your sense of order and well-being when you look at your C: drive folder. Tip #22.  Tidy Your Desktop The Desktop is probably the most abused part of a Windows computer (from an organization point of view).  It usually serves as a dumping ground for all incoming files, as well as holding icons to oft-used applications, plus some regularly opened files and folders.  It often ends up becoming an uncontrolled mess.  See if you can avoid this.  Here’s why… Application icons (Word, Internet Explorer, etc) are often found on the Desktop, but it’s unlikely that this is the optimum place for them.  The “Quick Launch” bar (or the Superbar in Windows 7) is always visible and so represents a perfect location to put your icons.  You’ll only be able to see the icons on your Desktop when all your programs are minimized.  It might be time to get your application icons off your desktop… You may have decided that the Inbox/To-do folder on your computer (see tip #13, above) should be your Desktop.  If so, then enough said.  Simply be vigilant about clearing it and preventing it from being polluted by junk files (see tip #15, above).  On the other hand, if your Desktop is not acting as your “Inbox” folder, then there’s no reason for it to have any data files or folders on it at all, except perhaps a couple of shortcuts to often-opened files and folders (either ongoing or current projects).  Everything else should be moved to your “Inbox” folder. In an ideal world, it might look like this: Tip #23.  Move Permanent Items on Your Desktop Away from the Top-Left Corner When files/folders are dragged onto your desktop in a Windows Explorer window, or when shortcuts are created on your Desktop from Internet Explorer, those icons are always placed in the top-left corner – or as close as they can get.  If you have other files, folders or shortcuts that you keep on the Desktop permanently, then it’s a good idea to separate these permanent icons from the transient ones, so that you can quickly identify which ones the transients are.  An easy way to do this is to move all your permanent icons to the right-hand side of your Desktop.  That should keep them separated from incoming items. Tip #24.  Synchronize If you have more than one computer, you’ll almost certainly want to share files between them.  If the computers are permanently attached to the same local network, then there’s no need to store multiple copies of any one file or folder – shortcuts will suffice.  However, if the computers are not always on the same network, then you will at some point need to copy files between them.  For files that need to permanently live on both computers, the ideal way to do this is to synchronize the files, as opposed to simply copying them. We only have room here to write a brief summary of synchronization, not a full article.  In short, there are several different types of synchronization: Where the contents of one folder are accessible anywhere, such as with Dropbox Where the contents of any number of folders are accessible anywhere, such as with Windows Live Mesh Where any files or folders from anywhere on your computer are synchronized with exactly one other computer, such as with the Windows “Briefcase”, Microsoft SyncToy, or (much more powerful, yet still free) SyncBack from 2BrightSparks.  This only works when both computers are on the same local network, at least temporarily. A great advantage of synchronization solutions is that once you’ve got it configured the way you want it, then the sync process happens automatically, every time.  Click a button (or schedule it to happen automatically) and all your files are automagically put where they’re supposed to be. If you maintain the same file and folder structure on both computers, then you can also sync files depend upon the correct location of other files, like shortcuts, playlists and office documents that link to other office documents, and the synchronized files still work on the other computer! Tip #25.  Hide Files You Never Need to See If you have your files well organized, you will often be able to tell if a file is out of place just by glancing at the contents of a folder (for example, it should be pretty obvious if you look in a folder that contains all the MP3s from one music CD and see a Word document in there).  This is a good thing – it allows you to determine if there are files out of place with a quick glance.  Yet sometimes there are files in a folder that seem out of place but actually need to be there, such as the “folder art” JPEGs in music folders, and various files in the root of the C: drive.  If such files never need to be opened by you, then a good idea is to simply hide them.  Then, the next time you glance at the folder, you won’t have to remember whether that file was supposed to be there or not, because you won’t see it at all! To hide a file, simply right-click on it and choose Properties: Then simply tick the Hidden tick-box:   Tip #26.  Keep Every Setup File These days most software is downloaded from the Internet.  Whenever you download a piece of software, keep it.  You’ll never know when you need to reinstall the software. Further, keep with it an Internet shortcut that links back to the website where you originally downloaded it, in case you ever need to check for updates. See tip #33 below for a full description of the excellence of organizing your setup files. Tip #27.  Try to Minimize the Number of Folders that Contain Both Files and Sub-folders Some of the folders in your organizational structure will contain only files.  Others will contain only sub-folders.  And you will also have some folders that contain both files and sub-folders.  You will notice slight improvements in how long it takes you to locate a file if you try to avoid this third type of folder.  It’s not always possible, of course – you’ll always have some of these folders, but see if you can avoid it. One way of doing this is to take all the leftover files that didn’t end up getting stored in a sub-folder and create a special “Miscellaneous” or “Other” folder for them. Tip #28.  Starting a Filename with an Underscore Brings it to the Top of a List Further to the previous tip, if you name that “Miscellaneous” or “Other” folder in such a way that its name begins with an underscore “_”, then it will appear at the top of the list of files/folders. The screenshot below is an example of this.  Each folder in the list contains a set of digital photos.  The folder at the top of the list, _Misc, contains random photos that didn’t deserve their own dedicated folder: Tip #29.  Clean Up those CD-ROMs and (shudder!) Floppy Disks Have you got a pile of CD-ROMs stacked on a shelf of your office?  Old photos, or files you archived off onto CD-ROM (or even worse, floppy disks!) because you didn’t have enough disk space at the time?  In the meantime have you upgraded your computer and now have 500 Gigabytes of space you don’t know what to do with?  If so, isn’t it time you tidied up that stack of disks and filed them into your gorgeous new folder structure? So what are you waiting for?  Bite the bullet, copy them all back onto your computer, file them in their appropriate folders, and then back the whole lot up onto a shiny new 1000Gig external hard drive! Useful Folders to Create This next section suggests some useful folders that you might want to create within your folder structure.  I’ve personally found them to be indispensable. The first three are all about convenience – handy folders to create and then put somewhere that you can always access instantly.  For each one, it’s not so important where the actual folder is located, but it’s very important where you put the shortcut(s) to the folder.  You might want to locate the shortcuts: On your Desktop In your “Quick Launch” area (or pinned to your Windows 7 Superbar) In your Windows Explorer “Favorite Links” area Tip #30.  Create an “Inbox” (“To-Do”) Folder This has already been mentioned in depth (see tip #13), but we wanted to reiterate its importance here.  This folder contains all the recently created, received or downloaded files that you have not yet had a chance to file away properly, and it also may contain files that you have yet to process.  In effect, it becomes a sort of “to-do list”.  It doesn’t have to be called “Inbox” – you can call it whatever you want. Tip #31.  Create a Folder where Your Current Projects are Collected Rather than going hunting for them all the time, or dumping them all on your desktop, create a special folder where you put links (or work folders) for each of the projects you’re currently working on. You can locate this folder in your “Inbox” folder, on your desktop, or anywhere at all – just so long as there’s a way of getting to it quickly, such as putting a link to it in Windows Explorer’s “Favorite Links” area: Tip #32.  Create a Folder for Files and Folders that You Regularly Open You will always have a few files that you open regularly, whether it be a spreadsheet of your current accounts, or a favorite playlist.  These are not necessarily “current projects”, rather they’re simply files that you always find yourself opening.  Typically such files would be located on your desktop (or even better, shortcuts to those files).  Why not collect all such shortcuts together and put them in their own special folder? As with the “Current Projects” folder (above), you would want to locate that folder somewhere convenient.  Below is an example of a folder called “Quick links”, with about seven files (shortcuts) in it, that is accessible through the Windows Quick Launch bar: See tip #37 below for a full explanation of the power of the Quick Launch bar. Tip #33.  Create a “Set-ups” Folder A typical computer has dozens of applications installed on it.  For each piece of software, there are often many different pieces of information you need to keep track of, including: The original installation setup file(s).  This can be anything from a simple 100Kb setup.exe file you downloaded from a website, all the way up to a 4Gig ISO file that you copied from a DVD-ROM that you purchased. The home page of the software manufacturer (in case you need to look up something on their support pages, their forum or their online help) The page containing the download link for your actual file (in case you need to re-download it, or download an upgraded version) The serial number Your proof-of-purchase documentation Any other template files, plug-ins, themes, etc that also need to get installed For each piece of software, it’s a great idea to gather all of these files together and put them in a single folder.  The folder can be the name of the software (plus possibly a very brief description of what it’s for – in case you can’t remember what the software does based in its name).  Then you would gather all of these folders together into one place, and call it something like “Software” or “Setups”. If you have enough of these folders (I have several hundred, being a geek, collected over 20 years), then you may want to further categorize them.  My own categorization structure is based on “platform” (operating system): The last seven folders each represents one platform/operating system, while _Operating Systems contains set-up files for installing the operating systems themselves.  _Hardware contains ROMs for hardware I own, such as routers. Within the Windows folder (above), you can see the beginnings of the vast library of software I’ve compiled over the years: An example of a typical application folder looks like this: Tip #34.  Have a “Settings” Folder We all know that our documents are important.  So are our photos and music files.  We save all of these files into folders, and then locate them afterwards and double-click on them to open them.  But there are many files that are important to us that can’t be saved into folders, and then searched for and double-clicked later on.  These files certainly contain important information that we need, but are often created internally by an application, and saved wherever that application feels is appropriate. A good example of this is the “PST” file that Outlook creates for us and uses to store all our emails, contacts, appointments and so forth.  Another example would be the collection of Bookmarks that Firefox stores on your behalf. And yet another example would be the customized settings and configuration files of our all our software.  Granted, most Windows programs store their configuration in the Registry, but there are still many programs that use configuration files to store their settings. Imagine if you lost all of the above files!  And yet, when people are backing up their computers, they typically only back up the files they know about – those that are stored in the “My Documents” folder, etc.  If they had a hard disk failure or their computer was lost or stolen, their backup files would not include some of the most vital files they owned.  Also, when migrating to a new computer, it’s vital to ensure that these files make the journey. It can be a very useful idea to create yourself a folder to store all your “settings” – files that are important to you but which you never actually search for by name and double-click on to open them.  Otherwise, next time you go to set up a new computer just the way you want it, you’ll need to spend hours recreating the configuration of your previous computer! So how to we get our important files into this folder?  Well, we have a few options: Some programs (such as Outlook and its PST files) allow you to place these files wherever you want.  If you delve into the program’s options, you will find a setting somewhere that controls the location of the important settings files (or “personal storage” – PST – when it comes to Outlook) Some programs do not allow you to change such locations in any easy way, but if you get into the Registry, you can sometimes find a registry key that refers to the location of the file(s).  Simply move the file into your Settings folder and adjust the registry key to refer to the new location. Some programs stubbornly refuse to allow their settings files to be placed anywhere other then where they stipulate.  When faced with programs like these, you have three choices:  (1) You can ignore those files, (2) You can copy the files into your Settings folder (let’s face it – settings don’t change very often), or (3) you can use synchronization software, such as the Windows Briefcase, to make synchronized copies of all your files in your Settings folder.  All you then have to do is to remember to run your sync software periodically (perhaps just before you run your backup software!). There are some other things you may decide to locate inside this new “Settings” folder: Exports of registry keys (from the many applications that store their configurations in the Registry).  This is useful for backup purposes or for migrating to a new computer Notes you’ve made about all the specific customizations you have made to a particular piece of software (so that you’ll know how to do it all again on your next computer) Shortcuts to webpages that detail how to tweak certain aspects of your operating system or applications so they are just the way you like them (such as how to remove the words “Shortcut to” from the beginning of newly created shortcuts).  In other words, you’d want to create shortcuts to half the pages on the How-To Geek website! Here’s an example of a “Settings” folder: Windows Features that Help with Organization This section details some of the features of Microsoft Windows that are a boon to anyone hoping to stay optimally organized. Tip #35.  Use the “Favorite Links” Area to Access Oft-Used Folders Once you’ve created your great new filing system, work out which folders you access most regularly, or which serve as great starting points for locating the rest of the files in your folder structure, and then put links to those folders in your “Favorite Links” area of the left-hand side of the Windows Explorer window (simply called “Favorites” in Windows 7):   Some ideas for folders you might want to add there include: Your “Inbox” folder (or whatever you’ve called it) – most important! The base of your filing structure (e.g. C:\Files) A folder containing shortcuts to often-accessed folders on other computers around the network (shown above as Network Folders) A folder containing shortcuts to your current projects (unless that folder is in your “Inbox” folder) Getting folders into this area is very simple – just locate the folder you’re interested in and drag it there! Tip #36.  Customize the Places Bar in the File/Open and File/Save Boxes Consider the screenshot below: The highlighted icons (collectively known as the “Places Bar”) can be customized to refer to any folder location you want, allowing instant access to any part of your organizational structure. Note:  These File/Open and File/Save boxes have been superseded by new versions that use the Windows Vista/Windows 7 “Favorite Links”, but the older versions (shown above) are still used by a surprisingly large number of applications. The easiest way to customize these icons is to use the Group Policy Editor, but not everyone has access to this program.  If you do, open it up and navigate to: User Configuration > Administrative Templates > Windows Components > Windows Explorer > Common Open File Dialog If you don’t have access to the Group Policy Editor, then you’ll need to get into the Registry.  Navigate to: HKEY_CURRENT_USER \ Software \ Microsoft  \ Windows \ CurrentVersion \ Policies \ comdlg32 \ Placesbar It should then be easy to make the desired changes.  Log off and log on again to allow the changes to take effect. Tip #37.  Use the Quick Launch Bar as a Application and File Launcher That Quick Launch bar (to the right of the Start button) is a lot more useful than people give it credit for.  Most people simply have half a dozen icons in it, and use it to start just those programs.  But it can actually be used to instantly access just about anything in your filing system: For complete instructions on how to set this up, visit our dedicated article on this topic. Tip #38.  Put a Shortcut to Windows Explorer into Your Quick Launch Bar This is only necessary in Windows Vista and Windows XP.  The Microsoft boffins finally got wise and added it to the Windows 7 Superbar by default. Windows Explorer – the program used for managing your files and folders – is one of the most useful programs in Windows.  Anyone who considers themselves serious about being organized needs instant access to this program at any time.  A great place to create a shortcut to this program is in the Windows XP and Windows Vista “Quick Launch” bar: To get it there, locate it in your Start Menu (usually under “Accessories”) and then right-drag it down into your Quick Launch bar (and create a copy). Tip #39.  Customize the Starting Folder for Your Windows 7 Explorer Superbar Icon If you’re on Windows 7, your Superbar will include a Windows Explorer icon.  Clicking on the icon will launch Windows Explorer (of course), and will start you off in your “Libraries” folder.  Libraries may be fine as a starting point, but if you have created yourself an “Inbox” folder, then it would probably make more sense to start off in this folder every time you launch Windows Explorer. To change this default/starting folder location, then first right-click the Explorer icon in the Superbar, and then right-click Properties:Then, in Target field of the Windows Explorer Properties box that appears, type %windir%\explorer.exe followed by the path of the folder you wish to start in.  For example: %windir%\explorer.exe C:\Files If that folder happened to be on the Desktop (and called, say, “Inbox”), then you would use the following cleverness: %windir%\explorer.exe shell:desktop\Inbox Then click OK and test it out. Tip #40.  Ummmmm…. No, that’s it.  I can’t think of another one.  That’s all of the tips I can come up with.  I only created this one because 40 is such a nice round number… Case Study – An Organized PC To finish off the article, I have included a few screenshots of my (main) computer (running Vista).  The aim here is twofold: To give you a sense of what it looks like when the above, sometimes abstract, tips are applied to a real-life computer, and To offer some ideas about folders and structure that you may want to steal to use on your own PC. Let’s start with the C: drive itself.  Very minimal.  All my files are contained within C:\Files.  I’ll confine the rest of the case study to this folder: That folder contains the following: Mark: My personal files VC: My business (Virtual Creations, Australia) Others contains files created by friends and family Data contains files from the rest of the world (can be thought of as “public” files, usually downloaded from the Net) Settings is described above in tip #34 The Data folder contains the following sub-folders: Audio:  Radio plays, audio books, podcasts, etc Development:  Programmer and developer resources, sample source code, etc (see below) Humour:  Jokes, funnies (those emails that we all receive) Movies:  Downloaded and ripped movies (all legal, of course!), their scripts, DVD covers, etc. Music:  (see below) Setups:  Installation files for software (explained in full in tip #33) System:  (see below) TV:  Downloaded TV shows Writings:  Books, instruction manuals, etc (see below) The Music folder contains the following sub-folders: Album covers:  JPEG scans Guitar tabs:  Text files of guitar sheet music Lists:  e.g. “Top 1000 songs of all time” Lyrics:  Text files MIDI:  Electronic music files MP3 (representing 99% of the Music folder):  MP3s, either ripped from CDs or downloaded, sorted by artist/album name Music Video:  Video clips Sheet Music:  usually PDFs The Data\Writings folder contains the following sub-folders: (all pretty self-explanatory) The Data\Development folder contains the following sub-folders: Again, all pretty self-explanatory (if you’re a geek) The Data\System folder contains the following sub-folders: These are usually themes, plug-ins and other downloadable program-specific resources. The Mark folder contains the following sub-folders: From Others:  Usually letters that other people (friends, family, etc) have written to me For Others:  Letters and other things I have created for other people Green Book:  None of your business Playlists:  M3U files that I have compiled of my favorite songs (plus one M3U playlist file for every album I own) Writing:  Fiction, philosophy and other musings of mine Mark Docs:  Shortcut to C:\Users\Mark Settings:  Shortcut to C:\Files\Settings\Mark The Others folder contains the following sub-folders: The VC (Virtual Creations, my business – I develop websites) folder contains the following sub-folders: And again, all of those are pretty self-explanatory. Conclusion These tips have saved my sanity and helped keep me a productive geek, but what about you? What tips and tricks do you have to keep your files organized?  Please share them with us in the comments.  Come on, don’t be shy… Similar Articles Productive Geek Tips Fix For When Windows Explorer in Vista Stops Showing File NamesWhy Did Windows Vista’s Music Folder Icon Turn Yellow?Print or Create a Text File List of the Contents in a Directory the Easy WayCustomize the Windows 7 or Vista Send To MenuAdd Copy To / Move To on Windows 7 or Vista Right-Click Menu 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 Acronis Online Backup DVDFab 6 Revo Uninstaller Pro Registry Mechanic 9 for Windows Track Daily Goals With 42Goals Video Toolbox is a Superb Online Video Editor Fun with 47 charts and graphs Tomorrow is Mother’s Day Check the Average Speed of YouTube Videos You’ve Watched OutlookStatView Scans and Displays General Usage Statistics

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  • Steganography Experiment - Trouble hiding message bits in DCT coefficients

    - by JohnHankinson
    I have an application requiring me to be able to embed loss-less data into an image. As such I've been experimenting with steganography, specifically via modification of DCT coefficients as the method I select, apart from being loss-less must also be relatively resilient against format conversion, scaling/DSP etc. From the research I've done thus far this method seems to be the best candidate. I've seen a number of papers on the subject which all seem to neglect specific details (some neglect to mention modification of 0 coefficients, or modification of AC coefficient etc). After combining the findings and making a few modifications of my own which include: 1) Using a more quantized version of the DCT matrix to ensure we only modify coefficients that would still be present should the image be JPEG'ed further or processed (I'm using this in place of simply following a zig-zag pattern). 2) I'm modifying bit 4 instead of the LSB and then based on what the original bit value was adjusting the lower bits to minimize the difference. 3) I'm only modifying the blue channel as it should be the least visible. This process must modify the actual image and not the DCT values stored in file (like jsteg) as there is no guarantee the file will be a JPEG, it may also be opened and re-saved at a later stage in a different format. For added robustness I've included the message multiple times and use the bits that occur most often, I had considered using a QR code as the message data or simply applying the reed-solomon error correction, but for this simple application and given that the "message" in question is usually going to be between 10-32 bytes I have plenty of room to repeat it which should provide sufficient redundancy to recover the true bits. No matter what I do I don't seem to be able to recover the bits at the decode stage. I've tried including / excluding various checks (even if it degrades image quality for the time being). I've tried using fixed point vs. double arithmetic, moving the bit to encode, I suspect that the message bits are being lost during the IDCT back to image. Any thoughts or suggestions on how to get this working would be hugely appreciated. (PS I am aware that the actual DCT/IDCT could be optimized from it's naive On4 operation using row column algorithm, or an FDCT like AAN, but for now it just needs to work :) ) Reference Papers: http://www.lokminglui.com/dct.pdf http://arxiv.org/ftp/arxiv/papers/1006/1006.1186.pdf Code for the Encode/Decode process in C# below: using System; using System.Collections.Generic; using System.Linq; using System.Text; using System.Drawing.Imaging; using System.Drawing; namespace ImageKey { public class Encoder { public const int HIDE_BIT_POS = 3; // use bit position 4 (1 << 3). public const int HIDE_COUNT = 16; // Number of times to repeat the message to avoid error. // JPEG Standard Quantization Matrix. // (to get higher quality multiply by (100-quality)/50 .. // for lower than 50 multiply by 50/quality. Then round to integers and clip to ensure only positive integers. public static double[] Q = {16,11,10,16,24,40,51,61, 12,12,14,19,26,58,60,55, 14,13,16,24,40,57,69,56, 14,17,22,29,51,87,80,62, 18,22,37,56,68,109,103,77, 24,35,55,64,81,104,113,92, 49,64,78,87,103,121,120,101, 72,92,95,98,112,100,103,99}; // Maximum qauality quantization matrix (if all 1's doesn't modify coefficients at all). public static double[] Q2 = {1,1,1,1,1,1,1,1, 1,1,1,1,1,1,1,1, 1,1,1,1,1,1,1,1, 1,1,1,1,1,1,1,1, 1,1,1,1,1,1,1,1, 1,1,1,1,1,1,1,1, 1,1,1,1,1,1,1,1, 1,1,1,1,1,1,1,1}; public static Bitmap Encode(Bitmap b, string key) { Bitmap response = new Bitmap(b.Width, b.Height, PixelFormat.Format32bppArgb); uint imgWidth = ((uint)b.Width) & ~((uint)7); // Maximum usable X resolution (divisible by 8). uint imgHeight = ((uint)b.Height) & ~((uint)7); // Maximum usable Y resolution (divisible by 8). // Start be transferring the unmodified image portions. // As we'll be using slightly less width/height for the encoding process we'll need the edges to be populated. for (int y = 0; y < b.Height; y++) for (int x = 0; x < b.Width; x++) { if( (x >= imgWidth && x < b.Width) || (y>=imgHeight && y < b.Height)) response.SetPixel(x, y, b.GetPixel(x, y)); } // Setup the counters and byte data for the message to encode. StringBuilder sb = new StringBuilder(); for(int i=0;i<HIDE_COUNT;i++) sb.Append(key); byte[] codeBytes = System.Text.Encoding.ASCII.GetBytes(sb.ToString()); int bitofs = 0; // Current bit position we've encoded too. int totalBits = (codeBytes.Length * 8); // Total number of bits to encode. for (int y = 0; y < imgHeight; y += 8) { for (int x = 0; x < imgWidth; x += 8) { int[] redData = GetRedChannelData(b, x, y); int[] greenData = GetGreenChannelData(b, x, y); int[] blueData = GetBlueChannelData(b, x, y); int[] newRedData; int[] newGreenData; int[] newBlueData; if (bitofs < totalBits) { double[] redDCT = DCT(ref redData); double[] greenDCT = DCT(ref greenData); double[] blueDCT = DCT(ref blueData); int[] redDCTI = Quantize(ref redDCT, ref Q2); int[] greenDCTI = Quantize(ref greenDCT, ref Q2); int[] blueDCTI = Quantize(ref blueDCT, ref Q2); int[] blueDCTC = Quantize(ref blueDCT, ref Q); HideBits(ref blueDCTI, ref blueDCTC, ref bitofs, ref totalBits, ref codeBytes); double[] redDCT2 = DeQuantize(ref redDCTI, ref Q2); double[] greenDCT2 = DeQuantize(ref greenDCTI, ref Q2); double[] blueDCT2 = DeQuantize(ref blueDCTI, ref Q2); newRedData = IDCT(ref redDCT2); newGreenData = IDCT(ref greenDCT2); newBlueData = IDCT(ref blueDCT2); } else { newRedData = redData; newGreenData = greenData; newBlueData = blueData; } MapToRGBRange(ref newRedData); MapToRGBRange(ref newGreenData); MapToRGBRange(ref newBlueData); for(int dy=0;dy<8;dy++) { for(int dx=0;dx<8;dx++) { int col = (0xff<<24) + (newRedData[dx+(dy*8)]<<16) + (newGreenData[dx+(dy*8)]<<8) + (newBlueData[dx+(dy*8)]); response.SetPixel(x+dx,y+dy,Color.FromArgb(col)); } } } } if (bitofs < totalBits) throw new Exception("Failed to encode data - insufficient cover image coefficients"); return (response); } public static void HideBits(ref int[] DCTMatrix, ref int[] CMatrix, ref int bitofs, ref int totalBits, ref byte[] codeBytes) { int tempValue = 0; for (int u = 0; u < 8; u++) { for (int v = 0; v < 8; v++) { if ( (u != 0 || v != 0) && CMatrix[v+(u*8)] != 0 && DCTMatrix[v+(u*8)] != 0) { if (bitofs < totalBits) { tempValue = DCTMatrix[v + (u * 8)]; int bytePos = (bitofs) >> 3; int bitPos = (bitofs) % 8; byte mask = (byte)(1 << bitPos); byte value = (byte)((codeBytes[bytePos] & mask) >> bitPos); // 0 or 1. if (value == 0) { int a = DCTMatrix[v + (u * 8)] & (1 << HIDE_BIT_POS); if (a != 0) DCTMatrix[v + (u * 8)] |= (1 << HIDE_BIT_POS) - 1; DCTMatrix[v + (u * 8)] &= ~(1 << HIDE_BIT_POS); } else if (value == 1) { int a = DCTMatrix[v + (u * 8)] & (1 << HIDE_BIT_POS); if (a == 0) DCTMatrix[v + (u * 8)] &= ~((1 << HIDE_BIT_POS) - 1); DCTMatrix[v + (u * 8)] |= (1 << HIDE_BIT_POS); } if (DCTMatrix[v + (u * 8)] != 0) bitofs++; else DCTMatrix[v + (u * 8)] = tempValue; } } } } } public static void MapToRGBRange(ref int[] data) { for(int i=0;i<data.Length;i++) { data[i] += 128; if(data[i] < 0) data[i] = 0; else if(data[i] > 255) data[i] = 255; } } public static int[] GetRedChannelData(Bitmap b, int sx, int sy) { int[] data = new int[8 * 8]; for (int y = sy; y < (sy + 8); y++) { for (int x = sx; x < (sx + 8); x++) { uint col = (uint)b.GetPixel(x,y).ToArgb(); data[(x - sx) + ((y - sy) * 8)] = (int)((col >> 16) & 0xff) - 128; } } return (data); } public static int[] GetGreenChannelData(Bitmap b, int sx, int sy) { int[] data = new int[8 * 8]; for (int y = sy; y < (sy + 8); y++) { for (int x = sx; x < (sx + 8); x++) { uint col = (uint)b.GetPixel(x, y).ToArgb(); data[(x - sx) + ((y - sy) * 8)] = (int)((col >> 8) & 0xff) - 128; } } return (data); } public static int[] GetBlueChannelData(Bitmap b, int sx, int sy) { int[] data = new int[8 * 8]; for (int y = sy; y < (sy + 8); y++) { for (int x = sx; x < (sx + 8); x++) { uint col = (uint)b.GetPixel(x, y).ToArgb(); data[(x - sx) + ((y - sy) * 8)] = (int)((col >> 0) & 0xff) - 128; } } return (data); } public static int[] Quantize(ref double[] DCTMatrix, ref double[] Q) { int[] DCTMatrixOut = new int[8*8]; for (int u = 0; u < 8; u++) { for (int v = 0; v < 8; v++) { DCTMatrixOut[v + (u * 8)] = (int)Math.Round(DCTMatrix[v + (u * 8)] / Q[v + (u * 8)]); } } return(DCTMatrixOut); } public static double[] DeQuantize(ref int[] DCTMatrix, ref double[] Q) { double[] DCTMatrixOut = new double[8*8]; for (int u = 0; u < 8; u++) { for (int v = 0; v < 8; v++) { DCTMatrixOut[v + (u * 8)] = (double)DCTMatrix[v + (u * 8)] * Q[v + (u * 8)]; } } return(DCTMatrixOut); } public static double[] DCT(ref int[] data) { double[] DCTMatrix = new double[8 * 8]; for (int v = 0; v < 8; v++) { for (int u = 0; u < 8; u++) { double cu = 1; if (u == 0) cu = (1.0 / Math.Sqrt(2.0)); double cv = 1; if (v == 0) cv = (1.0 / Math.Sqrt(2.0)); double sum = 0.0; for (int y = 0; y < 8; y++) { for (int x = 0; x < 8; x++) { double s = data[x + (y * 8)]; double dctVal = Math.Cos((2 * y + 1) * v * Math.PI / 16) * Math.Cos((2 * x + 1) * u * Math.PI / 16); sum += s * dctVal; } } DCTMatrix[u + (v * 8)] = (0.25 * cu * cv * sum); } } return (DCTMatrix); } public static int[] IDCT(ref double[] DCTMatrix) { int[] Matrix = new int[8 * 8]; for (int y = 0; y < 8; y++) { for (int x = 0; x < 8; x++) { double sum = 0; for (int v = 0; v < 8; v++) { for (int u = 0; u < 8; u++) { double cu = 1; if (u == 0) cu = (1.0 / Math.Sqrt(2.0)); double cv = 1; if (v == 0) cv = (1.0 / Math.Sqrt(2.0)); double idctVal = (cu * cv) / 4.0 * Math.Cos((2 * y + 1) * v * Math.PI / 16) * Math.Cos((2 * x + 1) * u * Math.PI / 16); sum += (DCTMatrix[u + (v * 8)] * idctVal); } } Matrix[x + (y * 8)] = (int)Math.Round(sum); } } return (Matrix); } } public class Decoder { public static string Decode(Bitmap b, int expectedLength) { expectedLength *= Encoder.HIDE_COUNT; uint imgWidth = ((uint)b.Width) & ~((uint)7); // Maximum usable X resolution (divisible by 8). uint imgHeight = ((uint)b.Height) & ~((uint)7); // Maximum usable Y resolution (divisible by 8). // Setup the counters and byte data for the message to decode. byte[] codeBytes = new byte[expectedLength]; byte[] outBytes = new byte[expectedLength / Encoder.HIDE_COUNT]; int bitofs = 0; // Current bit position we've decoded too. int totalBits = (codeBytes.Length * 8); // Total number of bits to decode. for (int y = 0; y < imgHeight; y += 8) { for (int x = 0; x < imgWidth; x += 8) { int[] blueData = ImageKey.Encoder.GetBlueChannelData(b, x, y); double[] blueDCT = ImageKey.Encoder.DCT(ref blueData); int[] blueDCTI = ImageKey.Encoder.Quantize(ref blueDCT, ref Encoder.Q2); int[] blueDCTC = ImageKey.Encoder.Quantize(ref blueDCT, ref Encoder.Q); if (bitofs < totalBits) GetBits(ref blueDCTI, ref blueDCTC, ref bitofs, ref totalBits, ref codeBytes); } } bitofs = 0; for (int i = 0; i < (expectedLength / Encoder.HIDE_COUNT) * 8; i++) { int bytePos = (bitofs) >> 3; int bitPos = (bitofs) % 8; byte mask = (byte)(1 << bitPos); List<int> values = new List<int>(); int zeroCount = 0; int oneCount = 0; for (int j = 0; j < Encoder.HIDE_COUNT; j++) { int val = (codeBytes[bytePos + ((expectedLength / Encoder.HIDE_COUNT) * j)] & mask) >> bitPos; values.Add(val); if (val == 0) zeroCount++; else oneCount++; } if (oneCount >= zeroCount) outBytes[bytePos] |= mask; bitofs++; values.Clear(); } return (System.Text.Encoding.ASCII.GetString(outBytes)); } public static void GetBits(ref int[] DCTMatrix, ref int[] CMatrix, ref int bitofs, ref int totalBits, ref byte[] codeBytes) { for (int u = 0; u < 8; u++) { for (int v = 0; v < 8; v++) { if ((u != 0 || v != 0) && CMatrix[v + (u * 8)] != 0 && DCTMatrix[v + (u * 8)] != 0) { if (bitofs < totalBits) { int bytePos = (bitofs) >> 3; int bitPos = (bitofs) % 8; byte mask = (byte)(1 << bitPos); int value = DCTMatrix[v + (u * 8)] & (1 << Encoder.HIDE_BIT_POS); if (value != 0) codeBytes[bytePos] |= mask; bitofs++; } } } } } } } UPDATE: By switching to using a QR Code as the source message and swapping a pair of coefficients in each block instead of bit manipulation I've been able to get the message to survive the transform. However to get the message to come through without corruption I have to adjust both coefficients as well as swap them. For example swapping (3,4) and (4,3) in the DCT matrix and then respectively adding 8 and subtracting 8 as an arbitrary constant seems to work. This survives a re-JPEG'ing of 96 but any form of scaling/cropping destroys the message again. I was hoping that by operating on mid to low frequency values that the message would be preserved even under some light image manipulation.

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  • Failed to Install Xdebug

    - by burnt1ce
    've registered xdebug in php.ini (as per http://xdebug.org/docs/install) but it's not showing up when i run "php -m" or when i get a test page to run "phpinfo()". I've just installed the latest version of XAMPP. I've used both "zend_extention" and "zend_extention_ts" to specify the path of the xdebug dll. I ensured that my apache server restarted and used the latest change of my php.ini by executing "httpd -k restart". Can anyone provide any suggestions in getting xdebug to show up? Here are the contents of my php.ini file. [PHP] ;;;;;;;;;;;;;;;;;;; ; About php.ini ; ;;;;;;;;;;;;;;;;;;; ; PHP's initialization file, generally called php.ini, is responsible for ; configuring many of the aspects of PHP's behavior. ; PHP attempts to find and load this configuration from a number of locations. ; The following is a summary of its search order: ; 1. SAPI module specific location. ; 2. The PHPRC environment variable. (As of PHP 5.2.0) ; 3. A number of predefined registry keys on Windows (As of PHP 5.2.0) ; 4. Current working directory (except CLI) ; 5. The web server's directory (for SAPI modules), or directory of PHP ; (otherwise in Windows) ; 6. The directory from the --with-config-file-path compile time option, or the ; Windows directory (C:\windows or C:\winnt) ; See the PHP docs for more specific information. ; http://php.net/configuration.file ; The syntax of the file is extremely simple. Whitespace and Lines ; beginning with a semicolon are silently ignored (as you probably guessed). ; Section headers (e.g. [Foo]) are also silently ignored, even though ; they might mean something in the future. ; Directives following the section heading [PATH=/www/mysite] only ; apply to PHP files in the /www/mysite directory. Directives ; following the section heading [HOST=www.example.com] only apply to ; PHP files served from www.example.com. Directives set in these ; special sections cannot be overridden by user-defined INI files or ; at runtime. Currently, [PATH=] and [HOST=] sections only work under ; CGI/FastCGI. ; http://php.net/ini.sections ; Directives are specified using the following syntax: ; directive = value ; Directive names are *case sensitive* - foo=bar is different from FOO=bar. ; Directives are variables used to configure PHP or PHP extensions. ; There is no name validation. If PHP can't find an expected ; directive because it is not set or is mistyped, a default value will be used. ; The value can be a string, a number, a PHP constant (e.g. E_ALL or M_PI), one ; of the INI constants (On, Off, True, False, Yes, No and None) or an expression ; (e.g. E_ALL & ~E_NOTICE), a quoted string ("bar"), or a reference to a ; previously set variable or directive (e.g. ${foo}) ; Expressions in the INI file are limited to bitwise operators and parentheses: ; | bitwise OR ; ^ bitwise XOR ; & bitwise AND ; ~ bitwise NOT ; ! boolean NOT ; Boolean flags can be turned on using the values 1, On, True or Yes. ; They can be turned off using the values 0, Off, False or No. ; An empty string can be denoted by simply not writing anything after the equal ; sign, or by using the None keyword: ; foo = ; sets foo to an empty string ; foo = None ; sets foo to an empty string ; foo = "None" ; sets foo to the string 'None' ; If you use constants in your value, and these constants belong to a ; dynamically loaded extension (either a PHP extension or a Zend extension), ; you may only use these constants *after* the line that loads the extension. ;;;;;;;;;;;;;;;;;;; ; About this file ; ;;;;;;;;;;;;;;;;;;; ; PHP comes packaged with two INI files. One that is recommended to be used ; in production environments and one that is recommended to be used in ; development environments. ; php.ini-production contains settings which hold security, performance and ; best practices at its core. But please be aware, these settings may break ; compatibility with older or less security conscience applications. We ; recommending using the production ini in production and testing environments. ; php.ini-development is very similar to its production variant, except it's ; much more verbose when it comes to errors. We recommending using the ; development version only in development environments as errors shown to ; application users can inadvertently leak otherwise secure information. ;;;;;;;;;;;;;;;;;;; ; Quick Reference ; ;;;;;;;;;;;;;;;;;;; ; The following are all the settings which are different in either the production ; or development versions of the INIs with respect to PHP's default behavior. ; Please see the actual settings later in the document for more details as to why ; we recommend these changes in PHP's behavior. ; allow_call_time_pass_reference ; Default Value: On ; Development Value: Off ; Production Value: Off ; display_errors ; Default Value: On ; Development Value: On ; Production Value: Off ; display_startup_errors ; Default Value: Off ; Development Value: On ; Production Value: Off ; error_reporting ; Default Value: E_ALL & ~E_NOTICE ; Development Value: E_ALL | E_STRICT ; Production Value: E_ALL & ~E_DEPRECATED ; html_errors ; Default Value: On ; Development Value: On ; Production value: Off ; log_errors ; Default Value: Off ; Development Value: On ; Production Value: On ; magic_quotes_gpc ; Default Value: On ; Development Value: Off ; Production Value: Off ; max_input_time ; Default Value: -1 (Unlimited) ; Development Value: 60 (60 seconds) ; Production Value: 60 (60 seconds) ; output_buffering ; Default Value: Off ; Development Value: 4096 ; Production Value: 4096 ; register_argc_argv ; Default Value: On ; Development Value: Off ; Production Value: Off ; register_long_arrays ; Default Value: On ; Development Value: Off ; Production Value: Off ; request_order ; Default Value: None ; Development Value: "GP" ; Production Value: "GP" ; session.bug_compat_42 ; Default Value: On ; Development Value: On ; Production Value: Off ; session.bug_compat_warn ; Default Value: On ; Development Value: On ; Production Value: Off ; session.gc_divisor ; Default Value: 100 ; Development Value: 1000 ; Production Value: 1000 ; session.hash_bits_per_character ; Default Value: 4 ; Development Value: 5 ; Production Value: 5 ; short_open_tag ; Default Value: On ; Development Value: Off ; Production Value: Off ; track_errors ; Default Value: Off ; Development Value: On ; Production Value: Off ; url_rewriter.tags ; Default Value: "a=href,area=href,frame=src,form=,fieldset=" ; Development Value: "a=href,area=href,frame=src,input=src,form=fakeentry" ; Production Value: "a=href,area=href,frame=src,input=src,form=fakeentry" ; variables_order ; Default Value: "EGPCS" ; Development Value: "GPCS" ; Production Value: "GPCS" ;;;;;;;;;;;;;;;;;;;; ; php.ini Options ; ;;;;;;;;;;;;;;;;;;;; ; Name for user-defined php.ini (.htaccess) files. Default is ".user.ini" ;user_ini.filename = ".user.ini" ; To disable this feature set this option to empty value ;user_ini.filename = ; TTL for user-defined php.ini files (time-to-live) in seconds. Default is 300 seconds (5 minutes) ;user_ini.cache_ttl = 300 ;;;;;;;;;;;;;;;;;;;; ; Language Options ; ;;;;;;;;;;;;;;;;;;;; ; Enable the PHP scripting language engine under Apache. ; http://php.net/engine engine = On ; This directive determines whether or not PHP will recognize code between ; <? and ?> tags as PHP source which should be processed as such. It's been ; recommended for several years that you not use the short tag "short cut" and ; instead to use the full <?php and ?> tag combination. With the wide spread use ; of XML and use of these tags by other languages, the server can become easily ; confused and end up parsing the wrong code in the wrong context. But because ; this short cut has been a feature for such a long time, it's currently still ; supported for backwards compatibility, but we recommend you don't use them. ; Default Value: On ; Development Value: Off ; Production Value: Off ; http://php.net/short-open-tag short_open_tag = Off ; Allow ASP-style <% %> tags. ; http://php.net/asp-tags asp_tags = Off ; The number of significant digits displayed in floating point numbers. ; http://php.net/precision precision = 14 ; Enforce year 2000 compliance (will cause problems with non-compliant browsers) ; http://php.net/y2k-compliance y2k_compliance = On ; Output buffering is a mechanism for controlling how much output data ; (excluding headers and cookies) PHP should keep internally before pushing that ; data to the client. If your application's output exceeds this setting, PHP ; will send that data in chunks of roughly the size you specify. ; Turning on this setting and managing its maximum buffer size can yield some ; interesting side-effects depending on your application and web server. ; You may be able to send headers and cookies after you've already sent output ; through print or echo. You also may see performance benefits if your server is ; emitting less packets due to buffered output versus PHP streaming the output ; as it gets it. On production servers, 4096 bytes is a good setting for performance ; reasons. ; Note: Output buffering can also be controlled via Output Buffering Control ; functions. ; Possible Values: ; On = Enabled and buffer is unlimited. (Use with caution) ; Off = Disabled ; Integer = Enables the buffer and sets its maximum size in bytes. ; Note: This directive is hardcoded to Off for the CLI SAPI ; Default Value: Off ; Development Value: 4096 ; Production Value: 4096 ; http://php.net/output-buffering output_buffering = Off ; You can redirect all of the output of your scripts to a function. For ; example, if you set output_handler to "mb_output_handler", character ; encoding will be transparently converted to the specified encoding. ; Setting any output handler automatically turns on output buffering. ; Note: People who wrote portable scripts should not depend on this ini ; directive. Instead, explicitly set the output handler using ob_start(). ; Using this ini directive may cause problems unless you know what script ; is doing. ; Note: You cannot use both "mb_output_handler" with "ob_iconv_handler" ; and you cannot use both "ob_gzhandler" and "zlib.output_compression". ; Note: output_handler must be empty if this is set 'On' !!!! ; Instead you must use zlib.output_handler. ; http://php.net/output-handler ;output_handler = ; Transparent output compression using the zlib library ; Valid values for this option are 'off', 'on', or a specific buffer size ; to be used for compression (default is 4KB) ; Note: Resulting chunk size may vary due to nature of compression. PHP ; outputs chunks that are few hundreds bytes each as a result of ; compression. If you prefer a larger chunk size for better ; performance, enable output_buffering in addition. ; Note: You need to use zlib.output_handler instead of the standard ; output_handler, or otherwise the output will be corrupted. ; http://php.net/zlib.output-compression zlib.output_compression = Off ; http://php.net/zlib.output-compression-level ;zlib.output_compression_level = -1 ; You cannot specify additional output handlers if zlib.output_compression ; is activated here. This setting does the same as output_handler but in ; a different order. ; http://php.net/zlib.output-handler ;zlib.output_handler = ; Implicit flush tells PHP to tell the output layer to flush itself ; automatically after every output block. This is equivalent to calling the ; PHP function flush() after each and every call to print() or echo() and each ; and every HTML block. Turning this option on has serious performance ; implications and is generally recommended for debugging purposes only. ; http://php.net/implicit-flush ; Note: This directive is hardcoded to On for the CLI SAPI implicit_flush = Off ; The unserialize callback function will be called (with the undefined class' ; name as parameter), if the unserializer finds an undefined class ; which should be instantiated. A warning appears if the specified function is ; not defined, or if the function doesn't include/implement the missing class. ; So only set this entry, if you really want to implement such a ; callback-function. unserialize_callback_func = ; When floats & doubles are serialized store serialize_precision significant ; digits after the floating point. The default value ensures that when floats ; are decoded with unserialize, the data will remain the same. serialize_precision = 100 ; This directive allows you to enable and disable warnings which PHP will issue ; if you pass a value by reference at function call time. Passing values by ; reference at function call time is a deprecated feature which will be removed ; from PHP at some point in the near future. The acceptable method for passing a ; value by reference to a function is by declaring the reference in the functions ; definition, not at call time. This directive does not disable this feature, it ; only determines whether PHP will warn you about it or not. These warnings ; should enabled in development environments only. ; Default Value: On (Suppress warnings) ; Development Value: Off (Issue warnings) ; Production Value: Off (Issue warnings) ; http://php.net/allow-call-time-pass-reference allow_call_time_pass_reference = On ; Safe Mode ; http://php.net/safe-mode safe_mode = Off ; By default, Safe Mode does a UID compare check when ; opening files. If you want to relax this to a GID compare, ; then turn on safe_mode_gid. ; http://php.net/safe-mode-gid safe_mode_gid = Off ; When safe_mode is on, UID/GID checks are bypassed when ; including files from this directory and its subdirectories. ; (directory must also be in include_path or full path must ; be used when including) ; http://php.net/safe-mode-include-dir safe_mode_include_dir = ; When safe_mode is on, only executables located in the safe_mode_exec_dir ; will be allowed to be executed via the exec family of functions. ; http://php.net/safe-mode-exec-dir safe_mode_exec_dir = ; Setting certain environment variables may be a potential security breach. ; This directive contains a comma-delimited list of prefixes. In Safe Mode, ; the user may only alter environment variables whose names begin with the ; prefixes supplied here. By default, users will only be able to set ; environment variables that begin with PHP_ (e.g. PHP_FOO=BAR). ; Note: If this directive is empty, PHP will let the user modify ANY ; environment variable! ; http://php.net/safe-mode-allowed-env-vars safe_mode_allowed_env_vars = PHP_ ; This directive contains a comma-delimited list of environment variables that ; the end user won't be able to change using putenv(). These variables will be ; protected even if safe_mode_allowed_env_vars is set to allow to change them. ; http://php.net/safe-mode-protected-env-vars safe_mode_protected_env_vars = LD_LIBRARY_PATH ; open_basedir, if set, limits all file operations to the defined directory ; and below. This directive makes most sense if used in a per-directory ; or per-virtualhost web server configuration file. This directive is ; *NOT* affected by whether Safe Mode is turned On or Off. ; http://php.net/open-basedir ;open_basedir = ; This directive allows you to disable certain functions for security reasons. ; It receives a comma-delimited list of function names. This directive is ; *NOT* affected by whether Safe Mode is turned On or Off. ; http://php.net/disable-functions disable_functions = ; This directive allows you to disable certain classes for security reasons. ; It receives a comma-delimited list of class names. This directive is ; *NOT* affected by whether Safe Mode is turned On or Off. ; http://php.net/disable-classes disable_classes = ; Colors for Syntax Highlighting mode. Anything that's acceptable in ; <span style="color: ???????"> would work. ; http://php.net/syntax-highlighting ;highlight.string = #DD0000 ;highlight.comment = #FF9900 ;highlight.keyword = #007700 ;highlight.bg = #FFFFFF ;highlight.default = #0000BB ;highlight.html = #000000 ; If enabled, the request will be allowed to complete even if the user aborts ; the request. Consider enabling it if executing long requests, which may end up ; being interrupted by the user or a browser timing out. PHP's default behavior ; is to disable this feature. ; http://php.net/ignore-user-abort ;ignore_user_abort = On ; Determines the size of the realpath cache to be used by PHP. This value should ; be increased on systems where PHP opens many files to reflect the quantity of ; the file operations performed. ; http://php.net/realpath-cache-size ;realpath_cache_size = 16k ; Duration of time, in seconds for which to cache realpath information for a given ; file or directory. For systems with rarely changing files, consider increasing this ; value. ; http://php.net/realpath-cache-ttl ;realpath_cache_ttl = 120 ;;;;;;;;;;;;;;;;; ; Miscellaneous ; ;;;;;;;;;;;;;;;;; ; Decides whether PHP may expose the fact that it is installed on the server ; (e.g. by adding its signature to the Web server header). It is no security ; threat in any way, but it makes it possible to determine whether you use PHP ; on your server or not. ; http://php.net/expose-php expose_php = On ;;;;;;;;;;;;;;;;;;; ; Resource Limits ; ;;;;;;;;;;;;;;;;;;; ; Maximum execution time of each script, in seconds ; http://php.net/max-execution-time ; Note: This directive is hardcoded to 0 for the CLI SAPI max_execution_time = 60 ; Maximum amount of time each script may spend parsing request data. It's a good ; idea to limit this time on productions servers in order to eliminate unexpectedly ; long running scripts. ; Note: This directive is hardcoded to -1 for the CLI SAPI ; Default Value: -1 (Unlimited) ; Development Value: 60 (60 seconds) ; Production Value: 60 (60 seconds) ; http://php.net/max-input-time max_input_time = 60 ; Maximum input variable nesting level ; http://php.net/max-input-nesting-level ;max_input_nesting_level = 64 ; Maximum amount of memory a script may consume (128MB) ; http://php.net/memory-limit memory_limit = 128M ;;;;;;;;;;;;;;;;;;;;;;;;;;;;;; ; Error handling and logging ; ;;;;;;;;;;;;;;;;;;;;;;;;;;;;;; ; This directive informs PHP of which errors, warnings and notices you would like ; it to take action for. The recommended way of setting values for this ; directive is through the use of the error level constants and bitwise ; operators. The error level constants are below here for convenience as well as ; some common settings and their meanings. ; By default, PHP is set to take action on all errors, notices and warnings EXCEPT ; those related to E_NOTICE and E_STRICT, which together cover best practices and ; recommended coding standards in PHP. For performance reasons, this is the ; recommend error reporting setting. Your production server shouldn't be wasting ; resources complaining about best practices and coding standards. That's what ; development servers and development settings are for. ; Note: The php.ini-development file has this setting as E_ALL | E_STRICT. This ; means it pretty much reports everything which is exactly what you want during ; development and early testing. ; ; Error Level Constants: ; E_ALL - All errors and warnings (includes E_STRICT as of PHP 6.0.0) ; E_ERROR - fatal run-time errors ; E_RECOVERABLE_ERROR - almost fatal run-time errors ; E_WARNING - run-time warnings (non-fatal errors) ; E_PARSE - compile-time parse errors ; E_NOTICE - run-time notices (these are warnings which often result ; from a bug in your code, but it's possible that it was ; intentional (e.g., using an uninitialized variable and ; relying on the fact it's automatically initialized to an ; empty string) ; E_STRICT - run-time notices, enable to have PHP suggest changes ; to your code which will ensure the best interoperability ; and forward compatibility of your code ; E_CORE_ERROR - fatal errors that occur during PHP's initial startup ; E_CORE_WARNING - warnings (non-fatal errors) that occur during PHP's ; initial startup ; E_COMPILE_ERROR - fatal compile-time errors ; E_COMPILE_WARNING - compile-time warnings (non-fatal errors) ; E_USER_ERROR - user-generated error message ; E_USER_WARNING - user-generated warning message ; E_USER_NOTICE - user-generated notice message ; E_DEPRECATED - warn about code that will not work in future versions ; of PHP ; E_USER_DEPRECATED - user-generated deprecation warnings ; ; Common Values: ; E_ALL & ~E_NOTICE (Show all errors, except for notices and coding standards warnings.) ; E_ALL & ~E_NOTICE | E_STRICT (Show all errors, except for notices) ; E_COMPILE_ERROR|E_RECOVERABLE_ERROR|E_ERROR|E_CORE_ERROR (Show only errors) ; E_ALL | E_STRICT (Show all errors, warnings and notices including coding standards.) ; Default Value: E_ALL & ~E_NOTICE ; Development Value: E_ALL | E_STRICT ; Production Value: E_ALL & ~E_DEPRECATED ; http://php.net/error-reporting error_reporting = E_ALL & ~E_NOTICE & ~E_DEPRECATED ; This directive controls whether or not and where PHP will output errors, ; notices and warnings too. Error output is very useful during development, but ; it could be very dangerous in production environments. Depending on the code ; which is triggering the error, sensitive information could potentially leak ; out of your application such as database usernames and passwords or worse. ; It's recommended that errors be logged on production servers rather than ; having the errors sent to STDOUT. ; Possible Values: ; Off = Do not display any errors ; stderr = Display errors to STDERR (affects only CGI/CLI binaries!) ; On or stdout = Display errors to STDOUT ; Default Value: On ; Development Value: On ; Production Value: Off ; http://php.net/display-errors display_errors = On ; The display of errors which occur during PHP's startup sequence are handled ; separately from display_errors. PHP's default behavior is to suppress those ; errors from clients. Turning the display of startup errors on can be useful in ; debugging configuration problems. But, it's strongly recommended that you ; leave this setting off on production servers. ; Default Value: Off ; Development Value: On ; Production Value: Off ; http://php.net/display-startup-errors display_startup_errors = On ; Besides displaying errors, PHP can also log errors to locations such as a ; server-specific log, STDERR, or a location specified by the error_log ; directive found below. While errors should not be displayed on productions ; servers they should still be monitored and logging is a great way to do that. ; Default Value: Off ; Development Value: On ; Production Value: On ; http://php.net/log-errors log_errors = Off ; Set maximum length of log_errors. In error_log information about the source is ; added. The default is 1024 and 0 allows to not apply any maximum length at all. ; http://php.net/log-errors-max-len log_errors_max_len = 1024 ; Do not log repeated messages. Repeated errors must occur in same file on same ; line unless ignore_repeated_source is set true. ; http://php.net/ignore-repeated-errors ignore_repeated_errors = Off ; Ignore source of message when ignoring repeated messages. When this setting ; is On you will not log errors with repeated messages from different files or ; source lines. ; http://php.net/ignore-repeated-source ignore_repeated_source = Off ; If this parameter is set to Off, then memory leaks will not be shown (on ; stdout or in the log). This has only effect in a debug compile, and if ; error reporting includes E_WARNING in the allowed list ; http://php.net/report-memleaks report_memleaks = On ; This setting is on by default. ;report_zend_debug = 0 ; Store the last error/warning message in $php_errormsg (boolean). Setting this value ; to On can assist in debugging and is appropriate for development servers. It should ; however be disabled on production servers. ; Default Value: Off ; Development Value: On ; Production Value: Off ; http://php.net/track-errors track_errors = Off ; Turn off normal error reporting and emit XML-RPC error XML ; http://php.net/xmlrpc-errors ;xmlrpc_errors = 0 ; An XML-RPC faultCode ;xmlrpc_error_number = 0 ; When PHP displays or logs an error, it has the capability of inserting html ; links to documentation related to that error. This directive controls whether ; those HTML links appear in error messages or not. For performance and security ; reasons, it's recommended you disable this on production servers. ; Note: This directive is hardcoded to Off for the CLI SAPI ; Default Value: On ; Development Value: On ; Production value: Off ; http://php.net/html-errors html_errors = On ; If html_errors is set On PHP produces clickable error messages that direct ; to a page describing the error or function causing the error in detail. ; You can download a copy of the PHP manual from http://php.net/docs ; and change docref_root to the base URL of your local copy including the ; leading '/'. You must also specify the file extension being used including ; the dot. PHP's default behavior is to leave these settings empty. ; Note: Never use this feature for production boxes. ; http://php.net/docref-root ; Examples ;docref_root = "/phpmanual/" ; http://php.net/docref-ext ;docref_ext = .html ; String to output before an error message. PHP's default behavior is to leave ; this setting blank. ; http://php.net/error-prepend-string ; Example: ;error_prepend_string = "<font color=#ff0000>" ; String to output after an error message. PHP's default behavior is to leave ; this setting blank. ; http://php.net/error-append-string ; Example: ;error_append_string = "</font>" ; Log errors to specified file. PHP's default behavior is to leave this value ; empty. ; http://php.net/error-log ; Example: ;error_log = php_errors.log ; Log errors to syslog (Event Log on NT, not valid in Windows 95). ;error_log = syslog ;error_log = "C:\xampp\apache\logs\php_error.log" ;;;;;;;;;;;;;;;;; ; Data Handling ; ;;;;;;;;;;;;;;;;; ; Note - track_vars is ALWAYS enabled ; The separator used in PHP generated URLs to separate arguments. ; PHP's default setting is "&". ; http://php.net/arg-separator.output ; Example: arg_separator.output = "&amp;" ; List of separator(s) used by PHP to parse input URLs into variables. ; PHP's default setting is "&

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