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• Migrating R Scripts from Development to Production

  - by Mark Hornick

  Normal 0 false false false EN-US X-NONE X-NONE MicrosoftInternetExplorer4 “How do I move my R scripts stored in one database instance to another? I have my development/test system and want to migrate to production.” Users of Oracle R Enterprise Embedded R Execution will often store their R scripts in the R Script Repository in Oracle Database, especially when using the ORE SQL API. From previous blog posts, you may recall that Embedded R Execution enables running R scripts managed by Oracle Database using both R and SQL interfaces. In ORE 1.3.1., the SQL API requires scripts to be stored in the database and referenced by name in SQL queries. The SQL API enables seamless integration with database-based applications and ease of production deployment. Loading R scripts in the repository Before talking about migration, we’ll first introduce how users store R scripts in Oracle Database. Users can add R scripts to the repository in R using the function ore.scriptCreate, or SQL using the function sys.rqScriptCreate. For the sample R script     id <- 1:10     plot(1:100,rnorm(100),pch=21,bg="red",cex =2)     data.frame(id=id, val=id / 100) users wrap this in a function and store it in the R Script Repository with a name. In R, this looks like ore.scriptCreate("RandomRedDots", function () { line-height: 115%; font-family: "Courier New";">     id <- 1:10     plot(1:100,rnorm(100),pch=21,bg="red",cex =2)     data.frame(id=id, val=id / 100)) }) In SQL, this looks like begin sys.rqScriptCreate('RandomRedDots',  'function(){     id <- 1:10     plot(1:100,rnorm(100),pch=21,bg="red",cex =2)     data.frame(id=id, val=id / 100)   }'); end; / The R function ore.scriptDrop and SQL function sys.rqScriptDrop can be used to drop these scripts as well. Note that the system will give an error if the script name already exists. Accessing R scripts once they’ve been loaded If you’re not using a source code control system, it is possible that your R scripts can be misplaced or files modified, making what is stored in Oracle Database to only or best copy of your R code. If you’ve loaded your R scripts to the database, it is straightforward to access these scripts from the database table SYS.RQ_SCRIPTS. For example, select * from sys.rq_scripts where name='myScriptName'; From R, scripts in the repository can be loaded into the R client engine using a function similar to the following: ore.scriptLoad <- function(name) { query <- paste("select script from sys.rq_scripts where name='",name,"'",sep="") str.f <- OREbase:::.ore.dbGetQuery(query) assign(name,eval(parse(text = str.f)),pos=1) } ore.scriptLoad("myFunctionName") This function is also useful if you want to load an existing R script from the repository into another R script in the repository – think modular coding style. Just include this function in the body of the other function and load the named script. Migrating R scripts from one database instance to another To move a set of functions from one system to another, the following script loads the functions from one R script repository into the client R engine, then connects to the target database and creates the scripts there with the same names. scriptNames <- OREbase:::.ore.dbGetQuery("select name from sys.rq_scripts where name not like 'RQG$%' and name not like 'RQ$%'")$NAME for(s in scriptNames) { cat(s,"\n") ore.scriptLoad(s) } ore.disconnect() ore.connect("rquser","orcl","localhost","rquser") for(s in scriptNames) { cat(s,"\n") ore.scriptDrop(s) ore.scriptCreate(s,get(s)) } Best Practice When naming R scripts, keep in mind that the name can be up to 128 characters. As such, consider organizing scripts in a directory structure manner. For example, if an organization has multiple groups or applications sharing the same database and there are multiple components, use “/” to facilitate the function organization: line-height: 115%;">ore.scriptCreate("/org1/app1/component1/myFuntion1", myFunction1) ore.scriptCreate("/org1/app1/component1/myFuntion2", myFunction2) ore.scriptCreate("/org1/app2/component2/myFuntion2", myFunction2) ore.scriptCreate("/org2/app2/component1/myFuntion3", myFunction3) ore.scriptCreate("/org3/app2/component1/myFuntion4", myFunction4) Users can then query for all functions using the path prefix when looking up functions. /* Style Definitions */ table.MsoNormalTable {mso-style-name:"Table Normal"; mso-tstyle-rowband-size:0; mso-tstyle-colband-size:0; mso-style-noshow:yes; mso-style-priority:99; mso-style-qformat:yes; mso-style-parent:""; mso-padding-alt:0in 5.4pt 0in 5.4pt; mso-para-margin:0in; mso-para-margin-bottom:.0001pt; mso-pagination:widow-orphan; font-size:11.0pt; font-family:"Calibri","sans-serif"; mso-ascii-font-family:Calibri; mso-ascii-theme-font:minor-latin; mso-fareast-font-family:"Times New Roman"; mso-fareast-theme-font:minor-fareast; mso-hansi-font-family:Calibri; mso-hansi-theme-font:minor-latin; mso-bidi-font-family:"Times New Roman"; mso-bidi-theme-font:minor-bidi;}

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• Implementing a Custom Coherence PartitionAssignmentStrategy

  - by jpurdy

  A recent A-Team engagement required the development of a custom PartitionAssignmentStrategy (PAS). By way of background, a PAS is an implementation of a Java interface that controls how a Coherence partitioned cache service assigns partitions (primary and backup copies) across the available set of storage-enabled members. While seemingly straightforward, this is actually a very difficult problem to solve. Traditionally, Coherence used a distributed algorithm spread across the cache servers (and as of Coherence 3.7, this is still the default implementation). With the introduction of the PAS interface, the model of operation was changed so that the logic would run solely in the cache service senior member. Obviously, this makes the development of a custom PAS vastly less complex, and in practice does not introduce a significant single point of failure/bottleneck. Note that Coherence ships with a default PAS implementation but it is not used by default. Further, custom PAS implementations are uncommon (this engagement was the first custom implementation that we know of). The particular implementation mentioned above also faced challenges related to managing multiple backup copies but that won't be discussed here. There were a few challenges that arose during design and implementation: Naive algorithms had an unreasonable upper bound of computational cost. There was significant complexity associated with configurations where the member count varied significantly between physical machines. Most of the complexity of a PAS is related to rebalancing, not initial assignment (which is usually fairly simple). A custom PAS may need to solve several problems simultaneously, such as: Ensuring that each member has a similar number of primary and backup partitions (e.g. each member has the same number of primary and backup partitions) Ensuring that each member carries similar responsibility (e.g. the most heavily loaded member has no more than one partition more than the least loaded). Ensuring that each partition is on the same member as a corresponding local resource (e.g. for applications that use partitioning across message queues, to ensure that each partition is collocated with its corresponding message queue). Ensuring that a given member holds no more than a given number of partitions (e.g. no member has more than 10 partitions) Ensuring that backups are placed far enough away from the primaries (e.g. on a different physical machine or a different blade enclosure) Achieving the above goals while ensuring that partition movement is minimized. These objectives can be even more complicated when the topology of the cluster is irregular. For example, if multiple cluster members may exist on each physical machine, then clearly the possibility exists that at certain points (e.g. following a member failure), the number of members on each machine may vary, in certain cases significantly so. Consider the case where there are three physical machines, with 3, 3 and 9 members each (respectively). This introduces complexity since the backups for the 9 members on the the largest machine must be spread across the other 6 members (to ensure placement on different physical machines), preventing an even distribution. For any given problem like this, there are usually reasonable compromises available, but the key point is that objectives may conflict under extreme (but not at all unlikely) circumstances. The most obvious general purpose partition assignment algorithm (possibly the only general purpose one) is to define a scoring function for a given mapping of partitions to members, and then apply that function to each possible permutation, selecting the most optimal permutation. This would result in N! (factorial) evaluations of the scoring function. This is clearly impractical for all but the smallest values of N (e.g. a partition count in the single digits). It's difficult to prove that more efficient general purpose algorithms don't exist, but the key take away from this is that algorithms will tend to either have exorbitant worst case performance or may fail to find optimal solutions (or both) -- it is very important to be able to show that worst case performance is acceptable. This quickly leads to the conclusion that the problem must be further constrained, perhaps by limiting functionality or by using domain-specific optimizations. Unfortunately, it can be very difficult to design these more focused algorithms. In the specific case mentioned, we constrained the solution space to very small clusters (in terms of machine count) with small partition counts and supported exactly two backup copies, and accepted the fact that partition movement could potentially be significant (preferring to solve that issue through brute force). We then used the out-of-the-box PAS implementation as a fallback, delegating to it for configurations that were not supported by our algorithm. Our experience was that the PAS interface is quite usable, but there are intrinsic challenges to designing PAS implementations that should be very carefully evaluated before committing to that approach.

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• Using XA Transactions in Coherence-based Applications

  - by jpurdy

  While the costs of XA transactions are well known (e.g. increased data contention, higher latency, significant disk I/O for logging, availability challenges, etc.), in many cases they are the most attractive option for coordinating logical transactions across multiple resources. There are a few common approaches when integrating Coherence into applications via the use of an application server's transaction manager: Use of Coherence as a read-only cache, applying transactions to the underlying database (or any system of record) instead of the cache. Use of TransactionMap interface via the included resource adapter. Use of the new ACID transaction framework, introduced in Coherence 3.6.   Each of these may have significant drawbacks for certain workloads. Using Coherence as a read-only cache is the simplest option. In this approach, the application is responsible for managing both the database and the cache (either within the business logic or via application server hooks). This approach also tends to provide limited benefit for many workloads, particularly those workloads that either have queries (given the complexity of maintaining a fully cached data set in Coherence) or are not read-heavy (where the cost of managing the cache may outweigh the benefits of reading from it). All updates are made synchronously to the database, leaving it as both a source of latency as well as a potential bottleneck. This approach also prevents addressing "hot data" problems (when certain objects are updated by many concurrent transactions) since most database servers offer no facilities for explicitly controlling concurrent updates. Finally, this option tends to be a better fit for key-based access (rather than filter-based access such as queries) since this makes it easier to aggressively invalidate cache entries without worrying about when they will be reloaded. The advantage of this approach is that it allows strong data consistency as long as optimistic concurrency control is used to ensure that database updates are applied correctly regardless of whether the cache contains stale (or even dirty) data. Another benefit of this approach is that it avoids the limitations of Coherence's write-through caching implementation. TransactionMap is generally used when Coherence acts as system of record. TransactionMap is not generally compatible with write-through caching, so it will usually be either used to manage a standalone cache or when the cache is backed by a database via write-behind caching. TransactionMap has some restrictions that may limit its utility, the most significant being: The lock-based concurrency model is relatively inefficient and may introduce significant latency and contention. As an example, in a typical configuration, a transaction that updates 20 cache entries will require roughly 40ms just for lock management (assuming all locks are granted immediately, and excluding validation and writing which will require a similar amount of time). This may be partially mitigated by denormalizing (e.g. combining a parent object and its set of child objects into a single cache entry), at the cost of increasing false contention (e.g. transactions will conflict even when updating different child objects). If the client (application server JVM) fails during the commit phase, locks will be released immediately, and the transaction may be partially committed. In practice, this is usually not as bad as it may sound since the commit phase is usually very short (all locks having been previously acquired). Note that this vulnerability does not exist when a single NamedCache is used and all updates are confined to a single partition (generally implying the use of partition affinity). The unconventional TransactionMap API is cumbersome but manageable. Only a few methods are transactional, primarily get(), put() and remove(). The ACID transactions framework (accessed via the Connection class) provides atomicity guarantees by implementing the NamedCache interface, maintaining its own cache data and transaction logs inside a set of private partitioned caches. This feature may be used as either a local transactional resource or as logging XA resource. However, a lack of database integration precludes the use of this functionality for most applications. A side effect of this is that this feature has not seen significant adoption, meaning that any use of this is subject to the usual headaches associated with being an early adopter (greater chance of bugs and greater risk of hitting an unoptimized code path). As a result, for the moment, we generally recommend against using this feature. In summary, it is possible to use Coherence in XA-oriented applications, and several customers are doing this successfully, but it is not a core usage model for the product, so care should be taken before committing to this path. For most applications, the most robust solution is normally to use Coherence as a read-only cache of the underlying data resources, even if this prevents taking advantage of certain product features.

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• How to move a rectangle properly?

  - by bodycountPP

  I recently started to learn OpenGL. Right now I finished the first chapter of the "OpenGL SuperBible". There were two examples. The first had the complete code and showed how to draw a simple triangle. The second example is supposed to show how to move a rectangle using SpecialKeys. The only code provided for this example was the SpecialKeys method. I still tried to implement it but I had two problems. In the previous example I declared and instaciated vVerts in the SetupRC() method. Now as it is also used in the SpecialKeys() method, I moved the declaration and instantiation to the top of the code. Is this proper c++ practice? I copied the part where vertex positions are recalculated from the book, but I had to pick the vertices for the rectangle on my own. So now every time I press a key for the first time the rectangle's upper left vertex is moved to (-0,5:-0.5). This ok because of GLfloat blockX = vVerts[0]; //Upper left X GLfloat blockY = vVerts[7]; // Upper left Y But I also think that this is the reason why my rectangle is shifted in the beginning. After the first time a key was pressed everything works just fine. Here is my complete code I hope you can help me on those two points. GLBatch squareBatch; GLShaderManager shaderManager; //Load up a triangle GLfloat vVerts[] = {-0.5f,0.5f,0.0f, 0.5f,0.5f,0.0f, 0.5f,-0.5f,0.0f, -0.5f,-0.5f,0.0f}; //Window has changed size, or has just been created. //We need to use the window dimensions to set the viewport and the projection matrix. void ChangeSize(int w, int h) { glViewport(0,0,w,h); } //Called to draw the scene. void RenderScene(void) { //Clear the window with the current clearing color glClear(GL_COLOR_BUFFER_BIT|GL_DEPTH_BUFFER_BIT|GL_STENCIL_BUFFER_BIT); GLfloat vRed[] = {1.0f,0.0f,0.0f,1.0f}; shaderManager.UseStockShader(GLT_SHADER_IDENTITY,vRed); squareBatch.Draw(); //perform the buffer swap to display the back buffer glutSwapBuffers(); } //This function does any needed initialization on the rendering context. //This is the first opportunity to do any OpenGL related Tasks. void SetupRC() { //Blue Background glClearColor(0.0f,0.0f,1.0f,1.0f); shaderManager.InitializeStockShaders(); squareBatch.Begin(GL_QUADS,4); squareBatch.CopyVertexData3f(vVerts); squareBatch.End(); } //Respond to arrow keys by moving the camera frame of reference void SpecialKeys(int key,int x,int y) { GLfloat stepSize = 0.025f; GLfloat blockSize = 0.5f; GLfloat blockX = vVerts[0]; //Upper left X GLfloat blockY = vVerts[7]; // Upper left Y if(key == GLUT_KEY_UP) { blockY += stepSize; } if(key == GLUT_KEY_DOWN){blockY -= stepSize;} if(key == GLUT_KEY_LEFT){blockX -= stepSize;} if(key == GLUT_KEY_RIGHT){blockX += stepSize;} //Recalculate vertex positions vVerts[0] = blockX; vVerts[1] = blockY - blockSize*2; vVerts[3] = blockX + blockSize * 2; vVerts[4] = blockY - blockSize *2; vVerts[6] = blockX+blockSize*2; vVerts[7] = blockY; vVerts[9] = blockX; vVerts[10] = blockY; squareBatch.CopyVertexData3f(vVerts); glutPostRedisplay(); } //Main entry point for GLUT based programs int main(int argc, char** argv) { //Sets the working directory. Not really needed gltSetWorkingDirectory(argv[0]); //Passes along the command-line parameters and initializes the GLUT library. glutInit(&argc,argv); //Tells the GLUT library what type of display mode to use, when creating the window. //Double buffered window, RGBA-Color mode,depth-buffer as part of our display, stencil buffer also available glutInitDisplayMode(GLUT_DOUBLE|GLUT_RGBA|GLUT_DEPTH|GLUT_STENCIL); //Window size glutInitWindowSize(800,600); glutCreateWindow("MoveRect"); glutReshapeFunc(ChangeSize); glutDisplayFunc(RenderScene); glutSpecialFunc(SpecialKeys); //initialize GLEW library GLenum err = glewInit(); //Check that nothing goes wrong with the driver initialization before we try and do any rendering. if(GLEW_OK != err) { fprintf(stderr,"Glew Error: %s\n",glewGetErrorString); return 1; } SetupRC(); glutMainLoop(); return 0; }

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• Web optimization

  - by hmloo

  1. CSS Optimization Organize your CSS code Good CSS organization helps with future maintainability of the site, it helps you and your team member understand the CSS more quickly and jump to specific styles. Structure CSS code For small project, you can break your CSS code in separate blocks according to the structure of the page or page content. for example you can break your CSS document according the content of your web page(e.g. Header, Main Content, Footer) Structure CSS file For large project, you may feel having too much CSS code in one place, so it's the best to structure your CSS into more CSS files, and use a master style sheet to import these style sheets. this solution can not only organize style structure, but also reduce server request./*--------------Master style sheet--------------*/ @import "Reset.css"; @import "Structure.css"; @import "Typography.css"; @import "Forms.css"; Create index for your CSS Another important thing is to create index at the beginning of your CSS file, index can help you quickly understand the whole CSS structure./*---------------------------------------- 1. Header 2. Navigation 3. Main Content 4. Sidebar 5. Footer ------------------------------------------*/ Writing efficient CSS selectors keep in mind that browsers match CSS selectors from right to left and the order of efficiency for selectors 1. id (#myid) 2. class (.myclass) 3. tag (div, h1, p) 4. adjacent sibling (h1 + p) 5. child (ul > li) 6. descendent (li a) 7. universal (*) 8. attribute (a[rel="external"]) 9. pseudo-class and pseudo element (a:hover, li:first) the rightmost selector is called "key selector", so when you write your CSS code, you should choose more efficient key selector. Here are some best practice: Don't tag-qualify Never do this:div#myid div.myclass .myclass#myid IDs are unique, classes are more unique than a tag so they don't need a tag. Doing so makes the selector less efficient. Avoid overqualifying selectors for example#nav a is more efficient thanul#nav li a Don't repeat declarationExample: body {font-size:12px;}h1 {font-size:12px;font-weight:bold;} since h1 is already inherited from body, so you don't need to repeate atrribute. Using 0 instead of 0px Always using #selector { margin: 0; } There’s no need to include the px after 0, removing all those superfluous px can reduce the size of your CSS file. Group declaration Example: h1 { font-size: 16pt; } h1 { color: #fff; } h1 { font-family: Arial, sans-serif; } it’s much better to combine them:h1 { font-size: 16pt; color: #fff; font-family: Arial, sans-serif; } Group selectorsExample: h1 { color: #fff; font-family: Arial, sans-serif; } h2 { color: #fff; font-family: Arial, sans-serif; } it would be much better if setup as:h1, h2 { color: #fff; font-family: Arial, sans-serif; } Group attributeExample: h1 { color: #fff; font-family: Arial, sans-serif; } h2 { color: #fff; font-family: Arial, sans-serif; font-size: 16pt; } you can set different rules for specific elements after setting a rule for a grouph1, h2 { color: #fff; font-family: Arial, sans-serif; } h2 { font-size: 16pt; } Using Shorthand PropertiesExample: #selector { margin-top: 8px; margin-right: 4px; margin-bottom: 8px; margin-left: 4px; }Better: #selector { margin: 8px 4px 8px 4px; }Best: #selector { margin: 8px 4px; } a good diagram illustrated how shorthand declarations are interpreted depending on how many values are specified for margin and padding property. instead of using:#selector { background-image: url(”logo.png”); background-position: top left; background-repeat: no-repeat; } is used:#selector { background: url(logo.png) no-repeat top left; } 2. Image Optimization Image Optimizer Image Optimizer is a free Visual Studio2010 extension that optimizes PNG, GIF and JPG file sizes without quality loss. It uses SmushIt and PunyPNG for the optimization. Just right click on any folder or images in Solution Explorer and choose optimize images, then it will automatically optimize all PNG, GIF and JPEG files in that folder. CSS Image Sprites CSS Image Sprites are a way to combine a collection of images to a single image, then use CSS background-position property to shift the visible area to show the required image, many images can take a long time to load and generates multiple server requests, so Image Sprite can reduce the number of server requests and improve site performance. You can use many online tools to generate your image sprite and CSS, and you can also try the Sprite and Image Optimization framework released by The ASP.NET team.

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• Tweaking Hudson memory usage

  - by rovarghe

  Hudson 3.1 has some performance optimizations that greatly reduces its memory footprint. Prior to this Hudson used to always hold the entire data model (all jobs and all builds) in memory which affected scalability. Some installations configured heap sizes in excess of 1GB to counteract this. Hudson 3.1.x maintains an MRU cache and only loads jobs and builds as they are required. Because of the inability to change existing APIs and be backward compatible with plugins, there were limits to how far we could go with this approach. Memory optimizations almost always come with a related cost, in this case its additional I/O that has to be performed to load data on request. On a small site that has frequent traffic, this is usually not noticeable since the MRU cache will usually hold on to all the data. A large site with infrequent traffic might experience some delays when the first request hits the server after a long gap. If you have a large heap and are able to allocate more memory, the cache settings can be adjusted to take advantage of this and even go back to pre-3.1 behavior. All the cache settings can be passed as options to the JVM container (Tomcat or the default Jetty container) using the -D option. There are two caches, independant of each other, one for Jobs and the other for Builds. For the jobs cache: hudson.jobs.cache.evict_in_seconds ( default=60 ) Seconds from last access (could be because of a servlet request or a background cron thread) a job should be purged from the cache. Set this to 0 to never purge based on time. hudson.jobs.cache.initial_capacity ( default=1024 ) Initial number of jobs the cache can accomodate. Setting this to the number of jobs you typically display on your Hudson landing page or home page will speed up consecutive access to that page. If the default is too large you may consider downsizing and using that memory for the Builds cache instead. hudson.jobs.cache.max_entries ( default=1024) Maximum number of jobs in the cache. The default is large enough for most installations, but if you find I/O activity when always accessing the hudson home page you might consider increasing this, but first verify if the I/O is caused by frequent eviction (see above), rather than by the cache not being large enough. For the builds cache: The builds cache is used to store Build objects as they are read from storage. Typically this happens when a user drills down into the details of a particular Job from the hudson hom epage. The cache is shared among builds for different jobs since in most installations all jobs are not accessed with the same frequency, so a per-job builds cache would be a waste of memory. hudson.job.builds.cache.evict_in_seconds ( default=60 ) Same as the equivalent Job cache, applied to Build. hudson.job.builds.cache.initial_capacity" ( default=512 ) Same as equivalent Job cache setting. Note the smaller initial size. If your site stores a large number of builds and has frequent access to more builds you might consider bumping this up. hudson.job.builds.cache.max_entries ( default=10240 ) The default max is large enough for most installations, the builds cache has bigger sized objects, so be careful about increasing the upper limit on this. See section on monitoring below. Sample usage: java -jar hudson-war-3.1.2-SNAPSHOT.war -Dhudson.jobs.cache.evict_in_seconds=300 \ -Dhudson.job.builds.cache.evict_in_seconds=300 Monitoring cache usage The 'jmap' tool that comes with the JDK can be used to monitor cache performance in an indirect way by looking at the number of Job and Build objects in each cache. Find the PID of the hudson instance and run $ jmap -histo:live <pid | grep 'hudson.model.*Lazy.*Key$' Here's a sample output: num #instances #bytes class name 523: 28 896 hudson.model.RunMap$LazyRunValue$Key 1200: 3 96 hudson.model.LazyTopLevelItem$Key These are the keys to the Jobs (LazyTopLevelItem$Key) and Builds (RunMap$LazyRunValue$Key) in the caches, so counting the number of keys is a good indicator of the number of items in the cache at any given moment. The size in bytes can be ignored, they are just the size of the keys, not the actual sizes of the objects they hold. Those sizes can only be obtained with a profiler. With the output above we can conclude that there are 3 jobs and 28 builds in memory. The 28 builds can all be from 1 job or all 3 jobs. Over time on an idle system, these should get evicted and memory cache should be empty. In practice, because of background cron threads and triggers, jobs rarely fall down to zero. Access of a job or a build by a cron thread resets the eviction timer.

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• Dealing with HTTP w00tw00t attacks

  - by Saif Bechan

  I have a server with apache and I recently installed mod_security2 because I get attacked a lot by this: My apache version is apache v2.2.3 and I use mod_security2.c This were the entries from the error log: [Wed Mar 24 02:35:41 2010] [error] [client 88.191.109.38] client sent HTTP/1.1 request without hostname (see RFC2616 section 14.23): /w00tw00t.at.ISC.SANS.DFind:) [Wed Mar 24 02:47:31 2010] [error] [client 202.75.211.90] client sent HTTP/1.1 request without hostname (see RFC2616 section 14.23): /w00tw00t.at.ISC.SANS.DFind:) [Wed Mar 24 02:47:49 2010] [error] [client 95.228.153.177] client sent HTTP/1.1 request without hostname (see RFC2616 section 14.23): /w00tw00t.at.ISC.SANS.DFind:) [Wed Mar 24 02:48:03 2010] [error] [client 88.191.109.38] client sent HTTP/1.1 request without hostname (see RFC2616 section 14.23): /w00tw00t.at.ISC.SANS.DFind:) Here are the errors from the access_log: 202.75.211.90 - - [29/Mar/2010:10:43:15 +0200] "GET /w00tw00t.at.ISC.SANS.DFind:) HTTP/1.1" 400 392 "-" "-" 211.155.228.169 - - [29/Mar/2010:11:40:41 +0200] "GET /w00tw00t.at.ISC.SANS.DFind:) HTTP/1.1" 400 392 "-" "-" 211.155.228.169 - - [29/Mar/2010:12:37:19 +0200] "GET /w00tw00t.at.ISC.SANS.DFind:) HTTP/1.1" 400 392 "-" "-" I tried configuring mod_security2 like this: SecFilterSelective REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind" SecFilterSelective REQUEST_URI "\w00tw00t\.at\.ISC\.SANS" SecFilterSelective REQUEST_URI "w00tw00t\.at\.ISC\.SANS" SecFilterSelective REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind:" SecFilterSelective REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind:\)" The thing in mod_security2 is that SecFilterSelective can not be used, it gives me errors. Instead I use a rule like this: SecRule REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind" SecRule REQUEST_URI "\w00tw00t\.at\.ISC\.SANS" SecRule REQUEST_URI "w00tw00t\.at\.ISC\.SANS" SecRule REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind:" SecRule REQUEST_URI "w00tw00t\.at\.ISC\.SANS\.DFind:\)" Even this does not work. I don't know what to do anymore. Anyone have any advice? Update 1 I see that nobody can solve this problem using mod_security. So far using ip-tables seems like the best option to do this but I think the file will become extremely large because the ip changes serveral times a day. I came up with 2 other solutions, can someone comment on them on being good or not. The first solution that comes to my mind is excluding these attacks from my apache error logs. This will make is easier for me to spot other urgent errors as they occur and don't have to spit trough a long log. The second option is better i think, and that is blocking hosts that are not sent in the correct way. In this example the w00tw00t attack is send without hostname, so i think i can block the hosts that are not in the correct form. Update 2 After going trough the answers I came to the following conclusions. To have custom logging for apache will consume some unnecessary recourses, and if there really is a problem you probably will want to look at the full log without anything missing. It is better to just ignore the hits and concentrate on a better way of analyzing your error logs. Using filters for your logs a good approach for this. Final thoughts on the subject The attack mentioned above will not reach your machine if you at least have an up to date system so there are basically no worries. It can be hard to filter out all the bogus attacks from the real ones after a while, because both the error logs and access logs get extremely large. Preventing this from happening in any way will cost you resources and they it is a good practice not to waste your resources on unimportant stuff. The solution i use now is Linux logwatch. It sends me summaries of the logs and they are filtered and grouped. This way you can easily separate the important from the unimportant. Thank you all for the help, and I hope this post can be helpful to someone else too.

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• Roaming user profile issues on Server 2008

  - by Alicia White

  I thought I cleared a user's profile from 2008, but it keeps coming back. So, I was looking for the best way to clear a roaming profile in Server 2008, but I have been unable to find anything. But, I did see the post here: http://serverfault.com/questions/18724/user-profile-keeps-loading-temp-profile I wanted to add a comment to that post, but it was closed as not being related to sysadmin. But, I think it IS related because I dealt with precisely this same problem on our Wndows 2008 terminal server. Here was the issue: we have a user who was getting an "unable to load your roaming profile" type of error at logon in Windows 2008. Looking at the server, we could see her temp profile listed in the profile list while she was loggged (listed as a "temporary" and not a "roaming" profile). While she was logged on, a folder called C:\Users\Temp.DOMAIN existed in the users folder, but that disappeared as soon as she logged out. When this thing happened in 2003, we would clear the contents of the roaming profile folder & delete the temp folder in C:\Documents and Settings. The thing is, 2008 behaves a bit differently. Server 2008 created a new roaming profile folder in the roaming profile folder share: \SERVER\ProfileShare\UserName.V2 The local profile disappears from the profile list in System Properties, so there is no profile to clear Also the local profile folder, C:\Users\Temp.DOMAIN doesn't stay on the server when the user logs out, so we can't delete that as we would normally do when this sort of thing happens in Windows 2003 Despite all of this, every time the user logs back on, the frickin' Temp profile always comes back. One of my team-mates, who is much more experienced with 2008, said I should check the registry for the user's profile in this key (the users are listed by SID): HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\WindowsNT\CurrentVersion\ProfileList I saw the user's SID listed there, but it ended in .BAK. I checked several other servers where she is having the same profile errors: in all cases, her SID ended with .BAK. For example (xxx replacing the LONG SID): HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Windows NT\CurrentVersion\ProfileList\S-1-5-21-xxxxx-xxxx.bak On the server she was logged on to, there were two keys for her profile in the registry: HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Windows NT\CurrentVersion\ProfileList\S-1-5-21-xxxxx-xxxx and HKEY_LOCAL_MACHINE\SOFTWARE\Microsoft\Windows NT\CurrentVersion\ProfileList\S-1-5-21-xxxxx-xxxx.bak So, here is how I cleared up the issue. I had the user log off. I deleted the apparently bad profiles ending in .BAK from the ProfileList key on each server where it appeared. I made sure her roaming profile folder was empty I made sure that all the TEMP profile folders were gone The user logged back on: no more profile errors! Anyway, I wanted to make a comment on that closed question, but I didn't see any way to re-open the question so I could add it. But, I also would like to know if this is the best practice to clear out a bad roaming profile for Server 2008? I'm having a hard time finding any instructions on line on how best to do this, but this method I used seemed to work. I'd like to find some documentation to give to our Level 1 support staff so they will know how to clear user profiles on 2008 since this seems to be more involved that clearing user profiles in server 2003. Thanks, Alicia

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• How do you handle authentication across domains?

  - by William Ratcliff

  I'm trying to save users of our services from having to have multiple accounts/passwords. I'm in a large organization and there's one group that handles part of user authentication for users who are from outside the facility (primarily for administrative functions). They store a secure cookie to establish a session and communicate only via HTTPS via the browser. Sessions expire either through: 1) explicit logout of the user 2) Inactivity 3) Browser closes My team is trying to write a web application to help users analyze data that they've taken (or are currently taking) while at our facility. We need to determine if a user is 1) authenticated 2) Some identifier for that user so we can store state for them (what analysis they are working on, etc.) So, the problem is how do you authenticate across domains (the authentication server for the other application lives in a border region between public and private--we will live in the public region). We have come up with some scenarios and I'd like advice about what is best practice, or if there is one we haven't considered. Let's start with the case where the user is authenticated with the authentication server. 1) The authentication server leaves a public cookie in the browser with their primary key for a user. If this is deemed sensitive, they encrypt it on their server and we have the key to decrypt it on our server. When the user visits our site, we check for this public cookie. We extract the user_id and use a public api for the authentication server to request if the user is logged in. If they are, they send us a response with: response={ userid :we can then map this to our own user ids. If necessary, we can request additional information such as email-address/display name once (to notify them if long running jobs are done, or to share results with other people, like with google_docs). account_is_active:Make sure that the account is still valid session_is_active: Is their session still active? If we query this for a valid user, this will have a side effect that we will reset the last_time_session_activated value and thus prolong their session with the authentication server last_time_session_activated: let us know how much time they have left ip_address_session_started_from:make sure the person at our site is coming from the same ip as they started the session at } Given this response, we either accept them as authenticated and move on with our app, or redirect them to the login page for the authentication server (question: if we give an encrypted portion of the response (signed by us) with the page to redirect them to, do we open any gaping security holes in the authentication server)? The flaw that we've found with this is that if the user visits evilsite.com and they look at the session cookie and send a query to the public api of the authentication server, they can keep the session alive and if our original user leaves the machine without logging out, then the next user will be able to access their session (this was possible before, but having the session alive eternally makes this worse). 2) The authentication server redirects all requests made to our domain to us and we send responses back through them to the user. Essentially, they act as a proxy. The advantage of this is that we can handshake with the authentication server, so it's safe to be trusted with the email address/name of the user and they don't have to reenter it So, if the user tries to go to: authentication_site/mysite_page1 they are redirected to mysite. Which would you choose, or is there a better way? The goal is to minimize the "Yet Another Password/Yet another username" problem... Thanks!!!!

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• Improving TCP performance over a gigabit network with lots of connections and high traffic of small packets

  - by MinimeDJ

  I’m trying to improve my TCP throughput over a “gigabit network with lots of connections and high traffic of small packets”. My server OS is Ubuntu 11.10 Server 64bit. There are about 50.000 (and growing) clients connected to my server through TCP Sockets (all on the same port). 95% of of my packets have size of 1-150 bytes (TCP header and payload). The rest 5% vary from 150 up to 4096+ bytes. With the config below my server can handle traffic up to 30 Mbps (full duplex). Can you please advice best practice to tune OS for my needs? My /etc/sysctl.cong looks like this: kernel.pid_max = 1000000 net.ipv4.ip_local_port_range = 2500 65000 fs.file-max = 1000000 # net.core.netdev_max_backlog=3000 net.ipv4.tcp_sack=0 # net.core.rmem_max = 16777216 net.core.wmem_max = 16777216 net.core.somaxconn = 2048 # net.ipv4.tcp_rmem = 4096 87380 16777216 net.ipv4.tcp_wmem = 4096 65536 16777216 # net.ipv4.tcp_synack_retries = 2 net.ipv4.tcp_syncookies = 1 net.ipv4.tcp_mem = 50576 64768 98152 # net.core.wmem_default = 65536 net.core.rmem_default = 65536 net.ipv4.tcp_window_scaling=1 # net.ipv4.tcp_mem= 98304 131072 196608 # net.ipv4.tcp_timestamps = 0 net.ipv4.tcp_rfc1337 = 1 net.ipv4.ip_forward = 0 net.ipv4.tcp_congestion_control=cubic net.ipv4.tcp_tw_recycle = 0 net.ipv4.tcp_tw_reuse = 0 # net.ipv4.tcp_orphan_retries = 1 net.ipv4.tcp_fin_timeout = 25 net.ipv4.tcp_max_orphans = 8192 Here are my limits: $ ulimit -a core file size (blocks, -c) 0 data seg size (kbytes, -d) unlimited scheduling priority (-e) 0 file size (blocks, -f) unlimited pending signals (-i) 193045 max locked memory (kbytes, -l) 64 max memory size (kbytes, -m) unlimited open files (-n) 1000000 pipe size (512 bytes, -p) 8 POSIX message queues (bytes, -q) 819200 real-time priority (-r) 0 stack size (kbytes, -s) 8192 cpu time (seconds, -t) unlimited max user processes (-u) 1000000 [ADDED] My NICs are the following: $ dmesg | grep Broad [ 2.473081] Broadcom NetXtreme II 5771x 10Gigabit Ethernet Driver bnx2x 1.62.12-0 (2011/03/20) [ 2.477808] bnx2x 0000:02:00.0: eth0: Broadcom NetXtreme II BCM57711E XGb (A0) PCI-E x4 5GHz (Gen2) found at mem fb000000, IRQ 28, node addr d8:d3:85:bd:23:08 [ 2.482556] bnx2x 0000:02:00.1: eth1: Broadcom NetXtreme II BCM57711E XGb (A0) PCI-E x4 5GHz (Gen2) found at mem fa000000, IRQ 40, node addr d8:d3:85:bd:23:0c [ADDED 2] ethtool -k eth0 Offload parameters for eth0: rx-checksumming: on tx-checksumming: on scatter-gather: on tcp-segmentation-offload: on udp-fragmentation-offload: off generic-segmentation-offload: on generic-receive-offload: on large-receive-offload: on rx-vlan-offload: on tx-vlan-offload: on ntuple-filters: off receive-hashing: off [ADDED 3] sudo ethtool -S eth0|grep -vw 0 NIC statistics: [1]: rx_bytes: 17521104292 [1]: rx_ucast_packets: 118326392 [1]: tx_bytes: 35351475694 [1]: tx_ucast_packets: 191723897 [2]: rx_bytes: 16569945203 [2]: rx_ucast_packets: 114055437 [2]: tx_bytes: 36748975961 [2]: tx_ucast_packets: 194800859 [3]: rx_bytes: 16222309010 [3]: rx_ucast_packets: 109397802 [3]: tx_bytes: 36034786682 [3]: tx_ucast_packets: 198238209 [4]: rx_bytes: 14884911384 [4]: rx_ucast_packets: 104081414 [4]: rx_discards: 5828 [4]: rx_csum_offload_errors: 1 [4]: tx_bytes: 35663361789 [4]: tx_ucast_packets: 194024824 [5]: rx_bytes: 16465075461 [5]: rx_ucast_packets: 110637200 [5]: tx_bytes: 43720432434 [5]: tx_ucast_packets: 202041894 [6]: rx_bytes: 16788706505 [6]: rx_ucast_packets: 113123182 [6]: tx_bytes: 38443961940 [6]: tx_ucast_packets: 202415075 [7]: rx_bytes: 16287423304 [7]: rx_ucast_packets: 110369475 [7]: rx_csum_offload_errors: 1 [7]: tx_bytes: 35104168638 [7]: tx_ucast_packets: 184905201 [8]: rx_bytes: 12689721791 [8]: rx_ucast_packets: 87616037 [8]: rx_discards: 2638 [8]: tx_bytes: 36133395431 [8]: tx_ucast_packets: 196547264 [9]: rx_bytes: 15007548011 [9]: rx_ucast_packets: 98183525 [9]: rx_csum_offload_errors: 1 [9]: tx_bytes: 34871314517 [9]: tx_ucast_packets: 188532637 [9]: tx_mcast_packets: 12 [10]: rx_bytes: 12112044826 [10]: rx_ucast_packets: 84335465 [10]: rx_discards: 2494 [10]: tx_bytes: 36562151913 [10]: tx_ucast_packets: 195658548 [11]: rx_bytes: 12873153712 [11]: rx_ucast_packets: 89305791 [11]: rx_discards: 2990 [11]: tx_bytes: 36348541675 [11]: tx_ucast_packets: 194155226 [12]: rx_bytes: 12768100958 [12]: rx_ucast_packets: 89350917 [12]: rx_discards: 2667 [12]: tx_bytes: 35730240389 [12]: tx_ucast_packets: 192254480 [13]: rx_bytes: 14533227468 [13]: rx_ucast_packets: 98139795 [13]: tx_bytes: 35954232494 [13]: tx_ucast_packets: 194573612 [13]: tx_bcast_packets: 2 [14]: rx_bytes: 13258647069 [14]: rx_ucast_packets: 92856762 [14]: rx_discards: 3509 [14]: rx_csum_offload_errors: 1 [14]: tx_bytes: 35663586641 [14]: tx_ucast_packets: 189661305 rx_bytes: 226125043936 rx_ucast_packets: 1536428109 rx_bcast_packets: 351 rx_discards: 20126 rx_filtered_packets: 8694 rx_csum_offload_errors: 11 tx_bytes: 548442367057 tx_ucast_packets: 2915571846 tx_mcast_packets: 12 tx_bcast_packets: 2 tx_64_byte_packets: 35417154 tx_65_to_127_byte_packets: 2006984660 tx_128_to_255_byte_packets: 373733514 tx_256_to_511_byte_packets: 378121090 tx_512_to_1023_byte_packets: 77643490 tx_1024_to_1522_byte_packets: 43669214 tx_pause_frames: 228 Some info about SACK: When to turn TCP SACK off?

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• Active Directory Time Synchronisation - Time-Service Event ID 50

  - by George

  I have an Active Directory domain with two DCs. The first DC in the forest/domain is Server 2012, the second is 2008 R2. The first DC holds the PDC Emulator role. I sporadically receive a warning from the Time-Service source, event ID 50: The time service detected a time difference of greater than %1 milliseconds for %2 seconds. The time difference might be caused by synchronization with low-accuracy time sources or by suboptimal network conditions. The time service is no longer synchronized and cannot provide the time to other clients or update the system clock. When a valid time stamp is received from a time service provider, the time service will correct itself. Time sync in the domain is configured with the second DC to synchronise using the /syncfromflags:DOMHIER flag. The first DC is configured to sync time using a /syncfromflags:MANUAL /reliable:YES, from a peerlist consisting of a number of UK based stratum 2 servers, such as ntp2d.mcc.ac.uk. I'm confused why I receive this event warning. It implies that my PDC emulator cannot synchronise time with a supposedly reliable external time source, and it quotes a time difference of 5 seconds for 900 seconds. It's worth also mentioning that I used to use a UK pool from ntp.org but I would receive the warning much more often. Since updating to a number of UK based academic time servers, it seems to be more reliable. Can someone with more experience shed some light on this - perhaps it is purely transient? Should I disregard the warning? Is my configuration sound? EDIT: I should add that the DCs are virtual, and installed on two separate VMware ESXi/vSphere physical hosts. I can also confirm that as per MDMarra's comment and best practice, VMware timesync is disabled, since: c:\Program Files\VMware\VMware Tools\VMwareToolboxCmd.exe timesync status returns Disabled. EDIT 2 Some strange new issue has cropped up. I've noticed a pattern. Originally, the event ID 50 warnings would occur at about 1230pm each day. This is interesting since our veeam backup happens at 12 midday. Since I made the changes discussed here, I now receive an event ID 51 instead of 50. The new warning says that: The time sample received from peer server.ac.uk differs from the local time by -40 seconds (Or approximately 40 seconds). This has happened two days in a row. Now I'm even more confused. Obviously the time never updates until I manually intervene. The issue seems to be related to virtualisation and veeam. Something may be occuring when veeam is backing up the PDCe. Any suggestions? UPDATE & SUMMARY msemack's excellent list of resources below (the accepted answer) provided enough information to correctly configure the time service in the domain. This should be the first port of call for any future people looking to verify their configuration. The final "40 second jump" issue I have resolved (there are no more warnings) through adjusting the VMware time sync settings as noted in the veeam knowledge base article here: http://www.veeam.com/kb1202 In any case, should any future reader use ESXi, veeam or not, the resources here are an excellent source of information on the time sync topic and msemack's answer is particularly invaluable.

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• Symantec Protection Suite and System Recovery 2011 Desktop Edition

  - by rihatum

  I am re-posting this as my previous question was being treated as if I am "Shopping or seeking Product Recommendations" even though I was NOT - BTW they have deleted my comments too which were not offensive in nature. anyway - I have re-phrased some parts of my question and I hope SF Admins "Do Not Modify / Edit" this one - will be most grateful for that. I have a lot of respect for the People who visit this SITE and help others ! Just To clarify : Just to go by SF rules - I am not seeking someone to Design this solution, I am simply seeking real world examples, experiences, technical expert opinions / suggestions, any tips or tricks they may have or any problems they may have faced while doing something similar above with these products. I am also not asking for Capacity Planning for Storage, We have done some research and I am seeking Expert Assurance / Suggestions. We (our company) are planning to deploy Symantec Endpoint Protection and Symantec Desktop Recovery 2011 Desktop Edition to our 3000 - 4000 workstations (Windows7 32 and 64) with a few 100s with Windows XP 32/64 Bit. I have read the implementation guide for SEP and have read tech-notes for Desktop Recovery 2011. Our team have planned to deploy this as follows : 1 x dedicated SQL 2008R2 for Symantec Endpoint Protection (Instead of using the Embedded Database) 1 x Dedicated SQL 2008R2 for Symantec Desktop Recovery 2011 (Instead of using the Embedded Database) 1 x Dedicated W2K8 R2 Box for the SEPM (Symantec Endpoint Protection Manager - Mgmt. APP) 1 x Dedicated W2K8 R2 Box for the Symantec Desktop Recovery 2011 Management Application Agent Deployment : As per Symantec Documentation for both of the above, an agent can be pushed via the Mgmt. Application (provided no firewalls are blocking ports required etc. - we have Windows firewall disabled already). Server Hardware : Per SQL Server : 16GB RAM + SAS DISKS + Dual XEON, RAID-10 for the SQL DB or I can always mount a LUN from our existing Hitachi or EMC SAN. SEPM Server : 16GB RAM + SAS DISKS + DUAL XEON System Recovery MGMT SERVER : 16GB RAM + SAS DISKS + DUAL XEON Above is the initial plan we have for 3000 - 4000 client workstation (Windows) Now my Questions :-) a) If we had these users distributed amongst two sites with AD DC / GC in each site, How would I restrict SEPM and Desktop Mgmt. solution to only check for users in their respective site ? b) At present all users are under one building but we are going to move some dept. to a new location (with dedicated connectivity), How would we control which SEPM / MGMT Server is responsible for which site ? c) We have netbackup in our environment backing up other servers, I am planning to protect these 4 (2 x SQL, 1 x SEPM, 1 x System Recovery Mgmt. Server) via netbackup or I can use System recovery 2011 server edition on all 4 of these boxes as well. (License is not an issue as we have the complete symantec portfolio included in our license). d) Now - Saving Desktop backups - What strategies have you implemented ? Any best practice recommendation for a large user base ? I was thinking to either mount a LUN from our Hitachi SAN on the Symantec Recovery Server itself or backup to the users hard drive locally and then copy it over to a network location ? Suggestions welcome :-) If you have anything to add / correct - that will be really helpful before diving into the actual implementation phase. Will be most grateful with your suggestions, recommendations and corrections with above - Many Thanks !

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• after return PartialView() Url.Actionlink("Action", "Controller"), the Controller is lost

  - by Johannes

  Well the Question is related to a problem I posted before (http://stackoverflow.com/questions/2403899/asp-net-mvc-partial-view-does-not-call-my-action). In practice I've a partial view which contains a Form, after submitting the Form the Controller returns the Partial View. Well the Problem is if I reload the page which contains the partial view the function <%= Url.Action("ChangePassword", "Account") %> returns "Account/ChangePassword", if I submit the form and the partial is returned by the controller. Using return PartialView() the function <%= Url.Action("ChangePassword", "Account") %> returns only "ChangePassword". Any Idea because? The View looks like: <form action="<%= Url.Action("ChangePassword", "Account") %>" method="post" id="jform"> <div> <fieldset> <legend>Account Information</legend> <p> <label for="currentPassword">Current password:</label> <%= Html.Password("currentPassword") %> <%= Html.ValidationMessage("currentPassword") %> </p> <p> <label for="newPassword">New password:</label> <%= Html.Password("newPassword") %> <%= Html.ValidationMessage("newPassword") %> </p> <p> <label for="confirmPassword">Confirm new password:</label> <%= Html.Password("confirmPassword") %> <%= Html.ValidationMessage("confirmPassword") %> </p> <p> <input type="submit" value="Change Password" /> </p> </fieldset> </div> </form> </div> <script> $(function() { $('#jform').submit(function() { $('#jform').ajaxSubmit({ target: '#FmChangePassword' }); return false; }); }); </script> Part of the Controller: if (!ValidateChangePassword(currentPassword, newPassword, confirmPassword)) { return PartialView(ViewData); }

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• Very slow compile times on Visual Studio

  - by johnc

  We are getting very slow compile times, which can take upwards of 20+ minutes on dual core 2GHz, 2G Ram machines. A lot of this is due to the size of our solution which has grown to 70+ projects, as well as VSS which is a bottle neck in itself when you have a lot of files. (swapping out VSS is not an option unfortunately, so I don't want this to descend into a VSS bash) We are looking at combing projects (not nice, as we like the separation of concerns, but is a good opportunity to refactor away some dead wood). We are also looking at having multiple solutions to achieve greater separation of concerns and quicker compile times for each element of the application. This I can see will become a dll hell as we try to keep things in synch. I am interested to know how other teams have dealt with this scaling issue, what do you do when your code base reaches a critical mass that you are wasting half the day watching the status bar deliver compile messages UPDATE Apologies, I neglected to mention this is a C# solution. Thanks for all the cpp suggestions, but it's been a few years since I've had to worry about headers. At a distance I say I miss C++, but I'm not sure I want to go back EDIT: Nice suggestions that have helped so far (not saying there aren't other nice suggestions below, just what has helped) New 3GHz laptop - the power of lost utilization works wonders when whinging to management Disable Anti Virus during compile 'Disconnecting' from VSS (actually the network) during compile - I may get us to remove VS-VSS integration altogether and stick to using the VSS UI Still not rip-snorting through a compile, but every bit helps. Orion did mention in a comment that generics may have a play also. From my tests there does appear to be a minimal performance hit, but not high enough to sure - compile times can be inconsistent due to disc activity. Due to time limitations, my tests didn't include as many Generics, or as much code, as would appear in live system, so that may accumulate. I wouldn't avoid using generics where they are supposed to be used, just for compile time performance WORKAROUND We are testing the practice of building new areas of the application in new solutions, importing in the latest dlls as required, them integrating them into the larger solution when we are happy with them. We may also do them same to existing code by creating temporary solutions that just encapsulate the areas we need to work on, and throwing them away after reintegrating the code. We need to weigh up the time it will take to reintegrate this code against the time we gain by not having Rip Van Winkle like experiences with rapid recompiling during development.

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• Weight doesn't work in a custom android component

  - by RCB

  I wanted to create a custom bottom button bar layout, I've created a xml file : <LinearLayout xmlns:android="http://schemas.android.com/apk/res/android" style="@android:style/ButtonBar" android:layout_width="fill_parent" android:layout_height="wrap_content" android:orientation="horizontal" android:padding="0dp" > <Button android:id="@+id/media_menu_button" style="?android:attr/buttonStyleSmall" android:layout_width="0dp" android:layout_height="match_parent" android:layout_margin="0dp" android:layout_weight="1" android:text="@string/media_menu_button" /> <Button android:id="@+id/scenario_menu_button" style="?android:attr/buttonStyleSmall" android:layout_width="0dp" android:layout_height="match_parent" android:layout_margin="0dp" android:layout_weight="1" android:text="@string/scenario_menu_button" /> <Button android:id="@+id/rooms_menu_button" style="?android:attr/buttonStyleSmall" android:layout_width="0dp" android:layout_height="match_parent" android:layout_margin="0dp" android:layout_weight="1" android:text="@string/rooms_menu_button" /> <Button android:id="@+id/shortcut_menu_button" style="?android:attr/buttonStyleSmall" android:layout_width="0dp" android:layout_height="match_parent" android:layout_margin="0dp" android:layout_weight="1" android:text="@string/shortcut_menu_button" /> as you can see I've given all the buttons width 0dp and weight of 1. then, I've created a class that extends the linear layout class : public class BeLightBottomBar extends LinearLayout implements OnClickListener { private LayoutInflater mInflater; private Context contexnt; private Button mShortcutMenuButton; private Button mRoomsMenuButton; private Button mScenarioMenuButton; private Button mMediaMenuButton; public BeLightBottomBar(Context context, AttributeSet attrs) { super(context, attrs); //inflate the view this.contexnt = context; mInflater = (LayoutInflater) context.getSystemService(Context.LAYOUT_INFLATER_SERVICE); LinearLayout barView = (LinearLayout) mInflater.inflate(R.layout.belight_bottombar, null); addView(barView); //get all the instances of the components of the bar mShortcutMenuButton = (Button) barView.findViewById(R.id.shortcut_menu_button); mRoomsMenuButton = (Button) barView.findViewById(R.id.rooms_menu_button); mScenarioMenuButton = (Button) barView.findViewById(R.id.scenario_menu_button); mMediaMenuButton = (Button) barView.findViewById(R.id.media_menu_button); //set this as a click listener mShortcutMenuButton.setOnClickListener(this); mRoomsMenuButton.setOnClickListener(this); mScenarioMenuButton.setOnClickListener(this); mMediaMenuButton.setOnClickListener(this); ... ... ... } the problem is when i add this class to the main activity xml <belight.homecontrol.components.BeLightBottomBar android:id="@+id/button_bar" android:layout_width="match_parent" android:layout_height="wrap_content" android:layout_alignParentBottom="true" android:layout_margin="0dp" android:padding="0dp" /> the weight stops working, all the buttons are different. and I don't know why !? if I just copy paste the bottom's bar xml code to the main xml file it works fine, the problem only occurs when using it as a whole. P.S. Is it a good practice to create a component this way? Or maybe I'm doing something wrong? Thanks, Dan

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• TFS and shared projects in multiple solutions

  - by David Stratton

  Our .NET team works on projects for our company that fall into distinct categories. Some are internal web apps, some are external (publicly facing) web apps, we also have internal Windows applications for our corporate office users, and Windows Forms apps for our retail locations (stores). Of course, because we hate code reuse, we have a ton of code that is shared among the different applications. Currently we're using SVN as our source control, and we've got our repository laid out like this: - = folder, | = Visual Studio Solution -SVN - Internet | Ourcompany.com | Oursecondcompany.com - Intranet | UniformOrdering website | MessageCenter website - Shared | ErrorLoggingModule | RegularExpressionGenerator | Anti-Xss | OrgChartModule etc... So.. The OurCompany.com solution in the Internet folder would have a website project, and it would also include the ErrorLoggingModule, RegularExpressionGenerator, and Anti-Xss projects from the shared directory. Similarly, our UniformOrdering website solution would have each of these projects included in the solution as well. We prefer to have a project reference to a .dll reference because, first of all, if we need to add or fix a function in the ErrorLoggingModule while working on the OurCompany.com website, it's right there. Also, this allows us to build each solution and see if changes to shared code break any other applications. This should work well on a build server as well if I'm correct. In SVN, there is no problem with this. SVN and Visual Studio aren't tied together in the way TFS's source control is. We never figured out how to work this type of structure in TFS when we were using it, because in TFS, the TFS project was always tied to a Visual Studio Solution. The Source Code repository was a child of the TFS Project, so if we wanted to do this, we had to duplicate the Shared code in each TFS project's source code repository. As my co-worker put it, this "breaks every known best practice about code reuse and simplicity". It was enough of a deal breaker for us that we switched to SVN. Now, however, we're faced with truly fixing our development processes, and the Application Lifecycle Management of TFS is pretty close to exactly what we want, and how we want to work. Our one sticking point is the shared code issue. We're evaluating other commercial and open source solutions, but since we're already paying for TFS with our MSDN Subscriptions, and TFS is pretty much exactly what we want, we'd REALLY like to find a way around this issue. Has anybody else faced this and come up with a solution? If you've seen an article or posting on this that you can share with me, that would help as well. As always, I'm open to answers like "You're looking at it all wrong, bonehead, HERE'S the way it SHOULD be done.

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• How to support both HTTP and HTTPS channels in Flex/BlazeDS?

  - by digitalsanctum

  I've been trying to find the right configuration for supporting both http/s requests in a Flex app. I've read all the docs and they allude to doing something like the following: <default-channels> <channel ref="my-secure-amf"> <serialization> <log-property-errors>true</log-property-errors> </serialization> </channel> <channel ref="my-amf"> <serialization> <log-property-errors>true</log-property-errors> </serialization> </channel> This works great when hitting the app via https but get intermittent communication failures when hitting the same app via http. Here's an abbreviated services-config.xml: <channel-definition id="my-amf" class="mx.messaging.channels.AMFChannel"> <endpoint url="http://{server.name}:{server.port}/{context.root}/messagebroker/amf" class="flex.messaging.endpoints.AMFEndpoint"/> <properties> <!-- HTTPS requests don't work on IE when pragma "no-cache" headers are set so you need to set the add-no-cache-headers property to false --> <add-no-cache-headers>false</add-no-cache-headers> <!-- Use to limit the client channel's connect attempt to the specified time interval. --> <connect-timeout-seconds>10</connect-timeout-seconds> </properties> </channel-definition> <channel-definition id="my-secure-amf" class="mx.messaging.channels.SecureAMFChannel"> <!--<endpoint url="https://{server.name}:{server.port}/{context.root}/messagebroker/amfsecure" class="flex.messaging.endpoints.SecureAMFEndpoint"/>--> <endpoint url="https://{server.name}:{server.port}/{context.root}/messagebroker/amfsecure" class="flex.messaging.endpoints.AMFEndpoint"/> <properties> <add-no-cache-headers>false</add-no-cache-headers> <connect-timeout-seconds>10</connect-timeout-seconds> </properties> </channel-definition> I'm running with Tomcat 5.5.17 and Java 5. The BlazeDS docs say this is the best practice. Is there a better way? With this config, there seems to be 2-3 retries associated with each channel defined in the default-channels element so it always takes ~20s before the my-amf channel connects via a http request. Is there a way to override the 2-3 retries to say, 1 retry for each channel? Thanks in advance for answers.

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• JLabel animation in JPanel

  - by Trizicus

  After scratching around I found that it's best to implement a custom image component by extending a JLabel. So far that has worked great as I can add multiple "images" (jlabels without the layout breaking. I just have a question that I hope someone can answer for me. I noticed that in order to animate JLabels across the screen I need to setlayout(null); and setbounds of the component and then to animate eventually setlocation(x,y);. Is this a best practice or a terrible way to animate a component? I plan on eventually making an animation class but I don't want to do so and end up having to chuck it. I have included relevant code for a quick review check. import java.awt.Color; import java.awt.Graphics; import java.awt.Graphics2D; import java.awt.event.ActionEvent; import java.awt.event.ActionListener; import javax.swing.JPanel; import javax.swing.Timer; public class GraphicsPanel extends JPanel { private Timer timer; private long startTime = 0; private int numFrames = 0; private float fps = 0.0f; private int x = 0; GraphicsPanel() { final Entity ent1 = new Entity(); ent1.setBounds(x, 0, ent1.getWidth(), ent1.getHeight()); add(ent1); //ESSENTIAL setLayout(null); //GAMELOOP timer = new Timer(30, new ActionListener() { public void actionPerformed(ActionEvent e) { getFPS(); incX(); ent1.setLocation(x, 0); repaint(); } }); timer.start(); } public void incX() { x++; } @Override public void paintComponent(Graphics g) { super.paintComponent(g); Graphics2D g2 = (Graphics2D) g.create(); g2.setClip(0, 0, getWidth(), getHeight()); g2.setColor(Color.BLACK); g2.drawString("FPS: " + fps, 1, 15); } public void getFPS() { ++numFrames; if (startTime == 0) { startTime = System.currentTimeMillis(); } else { long currentTime = System.currentTimeMillis(); long delta = (currentTime - startTime); if (delta > 1000) { fps = (numFrames * 1000) / delta; numFrames = 0; startTime = currentTime; } } } } Thank you!

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• Pass or Get a value from Parent ViewModel down to Sub-ViewModel?

  - by mkmurray

  I am using the MVVM Light framework as well as Unity for DI. I have some nested Views, each bound to a corresponding ViewModel. The ViewModels are bound to each View's root control DataContext via the ViewModelLocator idea that Laurent Bugnion has put into MVVM Light. This allows for finding ViewModels via a static resource and for controlling the lifetime of ViewModels via a Dependency Injection framework, in this case Unity. It also allows for Expression Blend to see everything in regard to ViewModels and how to bind them. As I stated the Views have a healthy dose of nesting, but the ViewModels don't really know anything about each other. A parent view binds to its corresponding ViewModel via the static resource ViewModelLocator (which uses Unity to control the construction and lifetime of the ViewModel object). That parent view contains a user control in it that is another sub-view, which then goes and gets its corresponding ViewModel via the ViewModelLocator as well. The ViewModels don't have references to each other or know any hierarchy in regard to each other. So here's an example of how the ViewModels do interact via messaging. I've got a parent View that has a ComboBox databound to an ObservableCollection in its ViewModel. The ComboBox's SelectedItem is also bound (two-way) to a property on the ViewModel. When the selection of the ComboBox changes, this is to trigger updates in other Views and sub-Views. Currently I am accomplishing this via the Messaging system that is found in MVVM Light. So I'm wondering what the best practice would be to get information from one ViewModel to another? In this case, what I need to pass down to sub-ViewModels is basically a user Guid representing the currently logged in user. The top-most parent View (well, ViewModel) will know this information, but I'm not sure how to get it down into the sub-ViewModels. Some possible approaches I can think of: Should the sub-ViewModel ask the static resource ViewModelLocator for a reference to the same object the parent View is using and access the property that way? It seems like ViewModels going through each other's properties is not very clean and couples them together unnecessarily. I'm already using messaging to notify the sub-Views that the user selected a new item in the ComboBox and to update accordingly. But the object type that is being selected in the ComboBox is not really directly related to this data value that the sub-Views need.

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• ASP.NET web forms as ASP.NET MVC

  - by lopkiju

  I am sorry for possible misleading about the title, but I have no idea for a proper title. Feel free to edit. Anyway, I am using ASP.NET Web Forms, and maybe this isn't how web forms is intended to be used, but I like to construct and populate HTML elements manually. It gives me more control. I don't use DataBinding and that kind of stuff. I use SqlConnection, SqlCommand and SqlDataReader, set SQL string etc. and read the data from the DataReader. Old school if you like. :) I do create WebControls so that I don't have to copy-paste every time I need some control, but mostly, I need WebControls to render as HTML so I can append that HTML into some other function that renders the final output with the control inside. I know I can render a control with control.RenderControl(writer), but this can only be done in (pre)Render or RenderContents overrides. For example. I have a dal.cs file where is stored all static functions and voids that communicate with the database. Functions mostly return string so that it can be appended into some other function to render the final result. The reason I am doing like this is that I want to separate the coding from the HTML as much as I can so that I don't do <% while (dataReader.Read()) % in HTML and display the data. I moved this into a CodeBehind. I also use this functions to render in the HttpHandler for AJAX response. That works perfectly, but when I want to add a control (ASP.NET Server control (.cs extension, not .ascx)) I don't know how to do that, so I see my self writing the same control as function that returns string or another function inside that control that returns string and replaces a job that would RenderContents do, so that I can call that function when I need control to be appended into a another string. I know this may not be a very good practice. As I see all the tutorials/videos about the ASP.NET MVC, I think it suite my needs as with the MVC you have to construct everything (or most of it) by your self, which I am already doing right now with web forms. After this long intro, I want to ask how can I build my controls so I can use them as I mentioned (return string) or I have to forget about server controls and build the controls as functions and used them that way? Is that even possible with ASP.NET Server Controls (.cs extension) or am I right when I said that I am not using it right. To be clear, I am talking about how to properly use a web forms, but to avoid data binders because I want to construct everything by my self (render HTML in Code Behind). Someone might think that I am appending strings like "some " + "string", which I am not. I am using StringBuilder for that so there's no slowness. Every opinion is welcome.

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• WCF Timeout issue - should there even be a socket connection?

  - by stiank81

  I have a .Net application which is split into a client and server side. The communication between them is handled using WCF. I'm not using the automagic service references, but instead I've built the connection manually like described in the Screencast by Miguel Castro. Summarized this means that I create a console application on the server side that holds ServiceHost objects for the different services: var myServiceHost = new System.ServiceModel.ServiceHost(typeof(MyService), new Uri("net.tcp://localhost:8002")); myServiceHost.Open(); And on the client side I have service proxies creating channels using the ChannelFactory: IMyService proxy = new ChannelFactory<IMyService>("MyServiceEndpoint").CreateChannel(); The client and server side share the service contract defined in the interface IMyService. And another advantage is that I get minimal App.config files - without all the autogenerated stuff created through the Service References. Example from client side: <?xml version="1.0"?> <configuration> <system.serviceModel> <client> <endpoint address="net.tcp://localhost:8002/MyEndpoint" binding="netTcpBinding" contract="IMyService" name="MyServiceEndpoint"/> </client> </system.serviceModel> </configuration> So - to my problem. I create the proxy once, and it holds a channel all the way through the application. However, if I leave the application without use for a few minutes the channel has timed out, and I get the following exception: The socket connection was aborted. This could be caused by an error processing your message or a receive timeout being exceeded by the remote host, or an underlying network resource issue. Local socket timeout was '00:00:59.9979998'. How do I prevent this? I'm assuming I need to specify a higher timeout in my configuration? But I don't want it to ever time out. But on the other hand - I don't want a socket connection! Do I need one? Thought I could go connection less with WCF... What's the permanent solution and best practice on solving this? Set timeout to "never".. Create a new channel for each request? I'm assuming there is some overhead creating the channel?.. Increase the timeout to e.g. 5minutes and create new channel if the connection did timeout? Make it connection less somehow? (Without the overhead of creating channels..) Something else...

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• Is MVVM pointless?

  - by joebeazelman

  Is orthodox MVVM implementation pointless? I am creating a new application and I considered Windows Forms and WPF. I chose WPF because it's future-proof and offer lots of flexibility. There is less code and easier to make significant changes to your UI using XAML. Since the choice for WPF is obvious, I figured that I may as well go all the way by using MVVM as my application architecture since it offers blendability, separation concerns and unit testability. Theoretically, it seems beautiful like the holy grail of UI programming. This brief adventure; however, has turned into a real headache. As expected in practice, I’m finding that I’ve traded one problem for another. I tend to be an obsessive programmer in that I want to do things the right way so that I can get the right results and possibly become a better programmer. The MVVM pattern just flunked my test on productivity and has just turned into a big yucky hack! The clear case in point is adding support for a Modal dialog box. The correct way is to put up a dialog box and tie it to a view model. Getting this to work is difficult. In order to benefit from the MVVM pattern, you have to distribute code in several places throughout the layers of your application. You also have to use esoteric programming constructs like templates and lamba expressions. Stuff that makes you stare at the screen scratching your head. This makes maintenance and debugging a nightmare waiting to happen as I recently discovered. I had an about box working fine until I got an exception the second time I invoked it, saying that it couldn’t show the dialog box again once it is closed. I had to add an event handler for the close functionality to the dialog window, another one in the IDialogView implementation of it and finally another in the IDialogViewModel. I thought MVVM would save us from such extravagant hackery! There are several folks out there with competing solutions to this problem and they are all hacks and don’t provide a clean, easily reusable, elegant solution. Most of the MVVM toolkits gloss over dialogs and when they do address them, they are just alert boxes that don’t require custom interfaces or view models. I’m planning on giving up on the MVVM view pattern, at least its orthodox implementation of it. What do you think? Has it been worth the trouble for you if you had any? Am I just a incompetent programmer or does MVVM not what it's hyped up to be?

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• When should I use Perl's AUTOLOAD?

  - by Robert S. Barnes

  In "Perl Best Practices" the very first line in the section on AUTOLOAD is: Don't use AUTOLOAD However all the cases he describes are dealing with OO or Modules. I have a stand alone script in which some command line switches control which versions of particular functions get defined. Now I know I could just take the conditionals and the evals and stick them naked at the top of my file before everything else, but I find it convenient and cleaner to put them in AUTOLOAD at the end of the file. Is this bad practice / style? If you think so why, and is there a another way to do it? As per brian's request I'm basically using this to do conditional compilation based on command line switches. I don't mind some constructive criticism. sub AUTOLOAD { our $AUTOLOAD; (my $method = $AUTOLOAD) =~ s/.*:://s; # remove package name if ($method eq 'tcpdump' && $tcpdump) { eval q( sub tcpdump { my $msg = shift; warn gf_time()." Thread ".threads->tid().": $msg\n"; } ); } elsif ($method eq 'loginfo' && $debug) { eval q( sub loginfo { my $msg = shift; $msg =~ s/$CRLF/\n/g; print gf_time()." Thread ".threads->tid().": $msg\n"; } ); } elsif ($method eq 'build_get') { if ($pipelining) { eval q( sub build_get { my $url = shift; my $base = shift; $url = "http://".$url unless $url =~ /^http/; return "GET $url HTTP/1.1${CRLF}Host: $base$CRLF$CRLF"; } ); } else { eval q( sub build_get { my $url = shift; my $base = shift; $url = "http://".$url unless $url =~ /^http/; return "GET $url HTTP/1.1${CRLF}Host: $base${CRLF}Connection: close$CRLF$CRLF"; } ); } } elsif ($method eq 'grow') { eval q{ require Convert::Scalar qw(grow); }; if ($@) { eval q( sub grow {} ); } goto &$method; } else { eval "sub $method {}"; return; } die $@ if $@; goto &$method; }

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• Streaming Video with Blackberry Simulator

  - by Jenny

  So, I wrote a quick little app for the iphone that takes in an http URL, and plays the .mp4 video located at that URL (it does more than that, of course, but that's the meat of it). Naturally, I wanted to have it on more than just a single mobile platform, so I decided to target Blackberry next. However, I'm running into a lot of problems with the Blackberry Environment. First of all, I learn that I can only download 256k files! I learn how to set that variable in my MDS simulator (and learn that this is NOT a production solution, because any end users will have to have their BES or MDS admin change the setting there). Then, I find a video less than 2 MB I can practice with. Going to the browser prompts me to save the video (rather than it playing in the browser like I expected). After saving the video, it refuses to play, saying it's the wrong format. So. I can't find a reference to IF Blackberry can stream with HTTP (i"ve heard it CAN use RTSP, though, and heard some rumors that it can't use HTTP, which would really suck). I also can't find a reference to what format blackberry uses (although I can find a million programs that will convert one file to the 'blackberry' format). Surely SOMEONE must have tried to stream video with the blackberry before. How did they go about doing so? Is it just a hopeless pipedream? Will I have to go with RTSP? Sorry for the lack of a concrete question...I'm just really lost, and I hate how so many tutorials or forum posts seem to assume I know the capabilities of the Blackberry... Edit: I finally found out that the .3gp (which I'd never heard of ) format is what Blackberry uses. Still have no idea how to stream videos off the web, though. I found a tutorial: http://www.blackberry.com/knowledgecenterpublic/livelink.exe/fetch/2000/348583/800332/1089414/How%5FTo%5F-%5FPlay%5Fvideo%5Fwithin%5Fa%5FBlackBerry%5Fsmartphone%5Fapplication.html?nodeid=1383173&vernum=0 That seemed to be useful, but the code doesn't work if you give it a URL (even though it claims it does).

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• How to avoid using duplicate savepoint names in nested transactions in nested stored procs?

  - by Gary McGill

  I have a pattern that I almost always follow, where if I need to wrap up an operation in a transaction, I do this: BEGIN TRANSACTION SAVE TRANSACTION TX -- Stuff IF @error <> 0 ROLLBACK TRANSACTION TX COMMIT TRANSACTION That's served me well enough in the past, but after years of using this pattern (and copy-pasting the above code), I've suddenly discovered a flaw which comes as a complete shock. Quite often, I'll have a stored procedure calling other stored procedures, all of which use this same pattern. What I've discovered (to my cost) is that because I'm using the same savepoint name everywhere, I can get into a situation where my outer transaction is partially committed - precisely the opposite of the atomicity that I'm trying to achieve. I've put together an example that exhibits the problem. This is a single batch (no nested stored procs), and so it looks a little odd in that you probably wouldn't use the same savepoint name twice in the same batch, but my real-world scenario would be too confusing to post. CREATE TABLE Test (test INTEGER NOT NULL) BEGIN TRAN SAVE TRAN TX BEGIN TRAN SAVE TRAN TX INSERT INTO Test(test) VALUES (1) COMMIT TRAN TX BEGIN TRAN SAVE TRAN TX INSERT INTO Test(test) VALUES (2) COMMIT TRAN TX DELETE FROM Test ROLLBACK TRAN TX COMMIT TRAN TX SELECT * FROM Test DROP TABLE Test When I execute this, it lists one record, with value "1". In other words, even though I rolled back my outer transaction, a record was added to the table. What's happening is that the ROLLBACK TRANSACTION TX at the outer level is rolling back as far as the last SAVE TRANSACTION TX at the inner level. Now that I write this all out, I can see the logic behind it: the server is looking back through the log file, treating it as a linear stream of transactions; it doesn't understand the nesting/hierarchy implied by either the nesting of the transactions (or, in my real-world scenario, by the calls to other stored procedures). So, clearly, I need to start using unique savepoint names instead of blindly using "TX" everywhere. But - and this is where I finally get to the point - is there a way to do this in a copy-pastable way so that I can still use the same code everywhere? Can I auto-generate the savepoint name on the fly somehow? Is there a convention or best-practice for doing this sort of thing? It's not exactly hard to come up with a unique name every time you start a transaction (could base it off the SP name, or somesuch), but I do worry that eventually there would be a conflict - and you wouldn't know about it because rather than causing an error it just silently destroys your data... :-(

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