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  • Oracle Traffic Director – download and check out new cool features in 11.1.1.7.0 by Frances Zhao

    - by JuergenKress
    As Oracle's strategic layer-7 software load balancer product, Oracle Traffic Direct is fast, reliable, secure, easy-to-use and scalable; that you can deploy as the reliable entry point for all TCP, HTTP and HTTPS traffic to application servers and web servers in your network. The latest release Oracle Traffic Director 11.1.1.7.0 is available for ExaLogic and Database Appliance! For download and details please visit the Traffic Director OTN website. It this release, we have introduced some major new functionality and improvements. Web application firewall. Oracle Traffic Director supports web application firewalls. A web application firewall (WAF) is a filter or server plugin that applies a set of rules, called rule sets, to an HTTP request. Using a web application firewall, users can inspect traffic and deny requests to protect back-end applications from CSRF vulnerabilities and common attacks such as cross-site scripting. WebSocket Connections. Oracle Traffic Director handles WebSocket connections by default. WebSocket connections are long-lived and allow support for live content, games in real-time, video chatting, and so on. Support for LDAP/T3 Load Balancing. Oracle Traffic Director now supports basic LDAP/T3 load balancing at layer 7, where requests are handled as generic TCP connections for traffic tunneling. It works in full-NAT mode. Please download and try it out. For more information, check out the data sheet and the documentation. For regular information become a member in the WebLogic Partner Community please visit: http://www.oracle.com/partners/goto/wls-emea ( OPN account required). If you need support with your account please contact the Oracle Partner Business Center. Blog Twitter LinkedIn Mix Forum Wiki Technorati Tags: traffic director,WebLogic Community,Oracle,OPN,Jürgen Kress

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  • Check parameters annotated with @Nonnull for null?

    - by David Harkness
    We've begun using FindBugs with and annotating our parameters with @Nonnull appropriately, and it works great to point out bugs early in the cycle. So far we have continued checking these arguments for null using Guava's checkNotNull, but I would prefer to check for null only at the edges--places where the value can come in without having been checked for null, e.g., a SOAP request. // service layer accessible from outside public Person createPerson(@CheckForNull String name) { return new Person(Preconditions.checkNotNull(name)); } ... // internal constructor accessed only by the service layer public Person(@Nonnull String name) { this.name = Preconditions.checkNotNull(name); // remove this check? } I understand that @Nonnull does not block null values itself. However, given that FindBugs will point out anywhere a value is transferred from an unmarked field to one marked @Nonnull, can't we depend on it to catch these cases (which it does) without having to check these values for null everywhere they get passed around in the system? Am I naive to want to trust the tool and avoid these verbose checks? Bottom line: While it seems safe to remove the second null check below, is it bad practice? This question is perhaps too similar to Should one check for null if he does not expect null, but I'm asking specifically in relation to the @Nonnull annotation.

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  • Designing a flexible tile-based engine

    - by Vee
    I'm trying to create a flexible tile-based game engine to make all sorts of non-realtime puzzle games, just as Bejeweled, Civilization, Sokoban, and so on. The first approach I had was to have a 2D array of Tile objects, and then have classes inheriting from Tile that represented the game objects. Unfortunately that way I couldn't stack more game elements on the same Tile without having a 3D array. Then I did something different: I still had the 2D array of Tile objects, but every Tile object contained a List where I put and different entities. This worked fine until 20 minutes ago, when I realized that it's too expensive to do many things, look at this example: I have a Wall entity. Every update I have to check the 8 adjacent Tiles, then check all of the entities in the Tile's List, check if any of those entities is a Wall, then finally draw the correct sprite. (This is done to draw walls that are next to each other seamlessly) The only solution I see now is having a 3D array, with many layers, that could suit every situation. But that way I can't stack two entities that share the same layer on the same tile. Whenever I want to do that I have to create a new layer. Is there a better solution? What would you do?

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  • Proper updating of GeoClipMaps

    - by thr
    I have been working on an implementation of gpu-based geo clip maps, but there is a section of the GPU Gems 2 article that i just can't seem to understand, specifically this paragraph and more precisely the bolded part: The choice of grid size n = 2k-1 has the further advantage that the finer level is never exactly centered with respect to its parent next-coarser level. In other words, it is always offset by 1 grid unit either left or right, as well as either top or bottom (see Figure 2-4), depending on the position of the viewpoint. In fact, it is necessary to allow a finer level to shift while its next-coarser level stays fixed, and therefore the finer level must sometimes be off-center with respect to the next-coarser level. An alternative choice of grid size, such as n = 2k-3, would provide the possibility for exact centering Let's take an example image from the article: My "understanding" of the way the clip maps were update was that you floor the position of the viewpoint to an int, and such get the center vertex point if this is not the same as the previous center point, you update the entire map. Now, this obviously is not the case - but what I am failing to understand is this: If you look at the image above, if the viewpoint was to move one unit to the right, then the inner ring (the one just around the view point + white center square) would end up getting a 1 unit space on both the left and right side of itself. But there is nothing in the paper that deals with this, what i mean is that it would end up looking like this (excuse my crummy cut-and-paste editing of the above image): This is obviously not a valid state of the. So, would the solution be that a clip ring (layer) can only move in increments of the ring/layer it's contained within? Wouldn't this end up being very restrictive? I feel like I am missing some crucial understanding of parts of the algorithm, but I have been over both this paper and the original paper from 2004 and I just can't see what I am not getting.

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  • ADF Essentials - Available for free and certified on GlassFish!

    - by delabassee
    If you are an Oracle customer, you are probably familiar with Oracle ADF (Application Development Framework). If you are not, ADF is, in a nutshell, a Java EE based framework that simplifies the development of enterprise applications. It is the development framework that was used, among other things, to build Oracle Fusion Applications. Oracle has just released ADF Essentials, a free to develop and deploy version of Oracle ADF's core technologies. As a good news never come alone, GlassFish 3.1.2 is now a certified container for ADF Essentials! ADF Essentials leverage core ADF features and includes: Oracle ADF Faces - a set of more than 150 JSF 2.0 rich components that simplify the creation of rich Web user interfaces (charting, data vizualization, advanced tables, drag and drop, touch gesture support, extensive windowing capabilities, etc.) Oracle ADF Controller - an extension of the JSF controller that helps build reusable process flows and provides the ability to create dynamic regions within Web pages. Oracle ADF Binding - an XML-based, meta-data abstraction layer to connect user interfaces to business services. Oracle ADF Business Components – a declaratively-configured layer that simplifies developing business services against relational databases by providing reusable components that implement common design patterns. ADF is a highly declarative framework, it has always had a very good tooling support. Visual development for Oracle ADF Essentials is provided in Oracle JDeveloper 11.1.2.3. Eclispe support is planned for a later OEPE (Oracle Enterprise Pack for Eclipse) release. Here are some relevant links to quickly learn on how to use ADF Essentials on GlassFish: Video : Oracle ADF Essentials Overview and Demo Deploying Oracle ADF Essentials Applications to Glassfish OTN : Oracle ADF Essentials Ressources

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  • What stops HTML5 and JS apps to perform as good as native apps?

    - by Amogh Talpallikar
    From what I understand, HTML is a mark-up language, so is the content of XAML, XIB and whatever Android uses and other native UI development frameworks. JavaScript is a programming language used along with it to handle client side scripting which will include things like event handling, client side validations and anything else C#,Java,Objective-C or C++ do in various such frameworks. There are MVC/MVVM patterns available in form frameworks like Sencha's, Angular etc. We have localStorage in form of both sqlite and key-value store as other frameworks have and you have API specification for almost everything that it missing. Whenever a native UI frameworks has to render UI , it has to parse a similar the markup and render the UI. Question break-down What stops from doing the same in HTML and JS itself ? Instead of having a web-control or browser as a layer in between why can't HTML(along with CSS) and JS be made to perform the same way ? Even if there is a layer,so does .net runtime and JVM are in other cases where C++,C are not being used. So Lets take the case of Android, like Dalvik, why Can't Chromium be another option(along with dalvik and NDK) where HTML does what android markup does and JavaScript is used to do what Java does ? So the Question is, Even if current implementations aren't as good, but theoretically is it possible to get HTML5 based applications to work as other native apps specially on mobile ?

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  • Need advice on framework design: how to make extending easy

    - by beginner_
    I'm creating a framework/library for a rather specific use-case (data type). It uses diverse spring components, including spring-data. The library has a set of entity classes properly set up and according service and dao layers. The main work or main benefit of the framework lies in the dao and service layer. Developers using the framework should be able to extend my entity classes to add additional fields they require. Therefore I made dao and service layer generic so it can be used by such extended entity classes. I now face an issue in the IO part of the framework. It must be able to import the according "special data type" into the database. In this part I need to create a new entity instance and hence need the actual class used. My current solution is to configure in spring a bean of the actual class used. The problem with this is that an application using the framework could only use 1 implementation of the entity (the original one from me or exactly 1 subclass but not 2 different classes of the same hierarchy. I'm looking for suggestions / desgins for solving this issue. Any ideas?

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  • Navigation in Win8 Metro Style applications

    - by Dennis Vroegop
    In Windows 8, Touch is, as they say, a first class citizen. Now, to be honest: they also said that in Windows 7. However in Win8 this is actually true. Applications are meant to be used by touch. Yes, you can still use mouse, keyboard and pen and your apps should take that into account but touch is where you should focus on initially. Will all users have touch enabled devices? No, not in the first place. I don’t think touchscreens will be on every device sold next year. But in 5 years? Who knows? Don’t forget: if your app is successful it will be around for a long time and by that time touchscreens will be everywhere. Another reason to embrace touch is that it’s easier to develop a touch-oriented app and then to make sure that keyboard, nouse and pen work as doing it the other way around. Porting a mouse-based application to a touch based application almost never works. The reverse gives you much more chances for success. That being said, there are some things that you need to think about. Most people have more than one finger, while most users only use one mouse at the time. Still, most touch-developers translate their mouse-knowledge to the touch and think they did a good job. Martin Tirion from Microsoft said that since Touch is a new language people face the same challenges they do when learning a new real spoken language. The first thing people try when learning a new language is simply replace the words in their native language to the newly learned words. At first they don’t care about grammar. To a native speaker of that other language this sounds all wrong but they still will be able to understand what the intention was. If you don’t believe me: try Google translate to translate something for you from your language to another and then back and see what happens. The same thing happens with Touch. Most developers translate a mouse-click into a tap-event and think they’re done. Well matey, you’re not done. Not by far. There are things you can do with a mouse that you cannot do with touch. Think hover. A mouse has the ability to ‘slide’ over UI elements. Touch doesn’t (I know: with Pen you can do this but I’m talking about actual fingers here). A touch is either there or it isn’t. And right-click? Forget about it. A click is a click.  Yes, you have more than one finger but the machine doesn’t know which finger you use… The other way around is also true. Like I said: most users only have one mouse but they are likely to have more than one finger. So how do we take that into account? Thinking about this is really worth the time: you might come up with some surprisingly good ideas! Still: don’t forget that not every user has touch-enabled hardware so make sure your app is useable for both groups. Keep this in mind: we’re going to need it later on! Now. Apps should be easy to use. You don’t want your user to read through pages and pages of documentation before they can use the app. Imagine that spotter next to an airfield suddenly seeing a prototype of a Concorde 2 landing on the nearby runway. He probably wants to enter that information in our app NOW and not after he’s taken a 3 day course. Even if he still has to download the app, install it for the first time and then run it he should be on his way immediately. At least, fast enough to note down the details of that unique, rare and possibly exciting sighting he just did. So.. How do we do this? Well, I am not talking about games here. Games are in a league of their own. They fall outside the scope of the apps I am describing. But all the others can roughly be characterized as being one of two flavors: the navigation is either flat or hierarchical. That’s it. And if it’s hierarchical it’s no more than three levels deep. Not more. Your users will get lost otherwise and we don’t want that. Flat is simple. Just imagine we have one screen that is as high as our physical screen is and as wide as you need it to be. Don’t worry if it doesn’t fit on the screen: people can scroll to the right and left. Don’t combine up/down and left/right scrolling: it’s confusing. Next to that, since most users will hold their device in landscape mode it’s very natural to scroll horizontal. So let’s use that when we have a flat model. The same applies to the hierarchical model. Try to have at most three levels. If you need more space, find a way to group the items in such a way that you can fit it in three, very wide lanes. At the highest level we have the so called hub level. This is the entry point of the app and as such it should give the user an immediate feeling of what the app is all about. If your app has categories if items then you might show these categories here. And while you’re at it: also show 2 or 3 of the items itself here to give the user a taste of what lies beneath. If the user selects a category you go to the section part. Here you show several sections (again, go as wide as you need) with again some detail examples. After that: the details layer shows each item. By giving some samples of the underlaying layer you achieve several things: you make the layer attractive by showing several different things, you show some highlights so the user sees actual content and you provide a shortcut to the layers underneath. The image below is borrowed from the http://design.windows.com website which has tons and tons of examples: For our app we’ll use this layout. So what will we show? Well, let’s see what sorts of features our app has to offer. I’ll repeat them here: Note planes Add pictures of that plane Notify friends of new spots Share new spots on social media Write down arrival times Write down departure times Write down the runway they take I am sure you can think of some more items but for now we'll use these. In the hub we’ll show something that represents “Spots”, “Friends”, “Social”. Apparently we have an inner list of spotter-friends that are in the app, while we also have to whole world in social. In the layer below we show something else, depending on what the user choose. When they choose “Spots” we’ll display the last spots, last spots by our friends (so we can actually jump from this category to the one next to it) and so on. When they choose a “spot” (or press the + icon in the App bar, which I’ll talk about next time) they go to the lowest and final level that shows details about that spot, including a picture, date and time and the notes belonging to that entry. You’d be amazed at how easy it is to organize your app this way. If you don’t have enough room in these three layers you probably could easily get away with grouping items. Take a look at our hub: we have three completely different things in one place. If you still can’t fit it all in in a logical and consistent way, chances are you are trying to do too much in this app. Go back to your mission statement, determine if it is specific enough and if your feature list helps that statement or makes it unclear. Go ahead. Give it a go! Next time we’ll talk about the look and feel, the charms and the app-bar….

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  • Centralizing a resource file among multiple projects in one solution (C#/WPF)

    - by MarkPearl
    One of the challenges one faces when doing multi language support in WPF is when one has several projects in one solution (i.e. a business layer & ui layer) and you want multi language support. Typically each solution would have a resource file – meaning if you have 3 projects in a solution you will have 3 resource files.   For me this isn’t an ideal solution, as you normally want to send the resource files to a translator and the more resource files you have, the more fragmented the dictionary will be and the more complicated it will be for the translator. This can easily be overcome by creating a single project that just holds your translation resources and then exposing it to the other projects as a reference as explained in the following steps. Step 1 Step 1 -  Add a class library to your solution that will contain just the resource files. Your solution will now have an additional project as illustrated below. Step 2 Reference this project to the other projects. Step 3 Move all the resources from the other resource files to the translation projects resource file. Step 4 Set the translations projects resource files access modifier to public. Step 5 Reference all other projects to use the translation resource file instead of their local resource file. To do this in xaml you would need to expose the project as a namespace at the top of the xaml file… note that the example below is for a project called MaxCutLanguages – you need to put the correct project name in its place.   xmlns:MaxCutLanguages="clr-namespace:MaxCutLanguages;assembly=MaxCutLanguages"   And then in the actual xaml you need to replace any text with a reference to the resource file. <TextBlock Text="{x:Static MaxCutLanguages:Properties.Resources.HelloWorld}"/> End Result You can now delete all the resource files in the other projects as you now have one centralized resource file.

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  • How is determined an impact of a requirement change on the existing code?

    - by MainMa
    Hi, How companies working on large projects evaluate an impact of a single modification on an existing code? Since my question is probably not very clear, here's an example: Let's take a sample business application which deals with tasks. In the database, each task has a state, 0 being "Pending", ... 5 - "Finished". A new requirement adds a new state, between 2nd and 3rd one. It means that: A constraint on the values 1 - 5 in the database must be changed, Business layer and code contracts must be changed to add a new state, Data access layer must be changed to take in account that, for example the state StateReady is now 6 instead of 5, etc. The application must implement a new state visually, add new controls for it, new localized strings for tool-tips, etc. When an application is written recently by one developer, it's more or less easy to predict every change to do. On the other hand, when an application was written for years by many people, no single person can anticipate every change immediately, without any investigation. So since this situation (such changes in requirements) is very frequent, I imagine there are already some clever techniques and ways to predict the impact. Is there any? Do you know any books which deal about this subject? Note: my question is not related to How do you deal with changing requirements? question. In fact, I'm not interested in evaluating the cost of a change, but rather the way to predict the parts of an application which will be concerned by the change. What will be those changes and how difficult they are really doesn't matter in my question.

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  • Layers - Logical seperation vs physical

    - by P.Brian.Mackey
    Some programmers recommend logical seperation of layers over physical. For example, given a DL, this means we create a DL namespace not a DL assembly. Benefits include: faster compilation time simpler deployment Faster startup time for your program Less assemblies to reference Im on a small team of 5 devs. We have over 50 assemblies to maintain. IMO this ratio is far from ideal. I prefer an extreme programming approach. Where if 100 assemblies are easier to maintain than 10,000...then 1 assembly must be easier than 100. Given technical limits, we should strive for < 5 assemblies. New assemblies are created out of technical need not layer requirements. Developers are worried for a few reasons. A. People like to work in their own environment so they dont step on eachothers toes. B. Microsoft tends to create new assemblies. E.G. Asp.net has its own DLL, so does winforms. Etc. C. Devs view this drive for a common assembly as a threat. Some team members Have a tendency to change the common layer without regard for how it will impact dependencies. My personal view: I view A. as silos, aka cowboy programming and suggest we implement branching to create isolation. C. First, that is a human problem and we shouldnt create technical work arounds for human behavior. Second, my goal is not to put everything in common. Rather, I want partitions to be made in namespaces not assemblies. Having a shared assembly doesnt make everything common. I want the community to chime in and tell me if Ive gone off my rocker. Is a drive for a single assembly or my viewpoint illogical or otherwise a bad idea?

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  • Encapsulate standard C functions?

    - by Jack Stout
    While studying the C programming language and learning safe practices, I'm inclined to write a layer of functionality over several parts of the standard library. This would serve two purposes: I could use standard parts of the language in ways that feel more familiar or rational to me, and I could easily replace that functionality with my own, if I needed to. I could benefit from this, but should I do it? As an example, we can consider memory management. If I've written malloc() into the constructors of each of my objects, then decide that I need to handle memory allocation on my own, I have to edit the constructor associated with every object. By referencing my own function, I can change the contents of that function without writing a new constructors. It seems obvious that I should do this, but I'm used to Python. I'm extremely comfortable in that environment and have no problem linking to any part of the standard library from any part of my program because I know I will almost certainly leave that relationship untouched for the life of the project. The situation I'm running into with C feels like I'm trying to hide the language from myself. Will writing a layer of functionality over the C standard library help me in learning the language and developing a codebase, or will it stifle my understanding going forward?

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  • Avoiding bloated Domain Objects

    - by djcredo
    We're trying to move data from our bloated Service layer into our Domain layer using a DDD approach. We currently have a lot of business logic in our services, which is spread out all over the place and doesn't benefit from inheritance. We have a central Domain class which is the focus of most of our work - a Trade. The Trade object will know how to price itself, how to estimate risk, validate itself, etc. We can then replace conditionals with polymorphism. Eg: SimpleTrade will price itself one way, but ComplexTrade will price itself another. However, we are worried that this will bloat the Trade class(s). It really should be in charge of its own processing but the class size is going to increase exponentially as more features are added. So we have choices: Put processing logic in Trade class. Processing logic is now polymorphic based on the type of the trade, but Trade class is now has multiple responsibilites (pricing, risk, etc) and is large Put processing logic into other class such as TradePricingService. No longer polymorphic with the Trade inheritance tree, but classes are smaller and easier to test. What would be the suggested approach?

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  • decouple software components via nameconvention

    - by csteinmueller
    I'm currently evaluating alternatives to refactor a drivermanagement. In my multitier architecture I have Baseclass DAL.Device //my entity Interfaces BL.IDriver //handles the dataprocessing between application and device BL.IDriverCreator //creates an IDriver from a Device BL.IDriverFactory //handles the driver creation requests Every specialization of Device has a corresponding IDriver implementation and a corresponding IDriverCreator implementation. At the moment the mapping is fix via a type check within the business layer / DriverFactory. That means every new driver needs a) changing code within the DriverFactory and b) referencing the new IDriver implementation / assembly. On a customers point of view that means, every new driver, used or not, needs a complex revalidation of their hardware environment, because it's a critical process. My first inspiration was to use a caliburn micro like nameconvention see Caliburn.Micro: Xaml Made Easy BL.RestDriver BL.RestDriverCreator DAL.RestDevice After receiving the RestDevicewithin the IDriverFactory I can load all driver dlls via reflection and do a namesplitting/comparing (extracting the xx from xxDriverCreator and xxDevice) Another idea would be a custom attribute (which also leads to comparing strings). My question: is that a good approach above layer borders? If not, what would be a good approach?

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  • Am I just not understanding TDD unit testing (Asp.Net MVC project)?

    - by KallDrexx
    I am trying to figure out how to correctly and efficiently unit test my Asp.net MVC project. When I started on this project I bought the Pro ASP.Net MVC, and with that book I learned about TDD and unit testing. After seeing the examples, and the fact that I work as a software engineer in QA in my current company, I was amazed at how awesome TDD seemed to be. So I started working on my project and went gun-ho writing unit tests for my database layer, business layer, and controllers. Everything got a unit test prior to implementation. At first I thought it was awesome, but then things started to go downhill. Here are the issues I started encountering: I ended up writing application code in order to make it possible for unit tests to be performed. I don't mean this in a good way as in my code was broken and I had to fix it so the unit test pass. I mean that abstracting out the database to a mock database is impossible due to the use of linq for data retrieval (using the generic repository pattern). The reason is that with linq-sql or linq-entities you can do joins just by doing: var objs = select p from _container.Projects select p.Objects; However, if you mock the database layer out, in order to have that linq pass the unit test you must change the linq to be var objs = select p from _container.Projects join o in _container.Objects on o.ProjectId equals p.Id select o; Not only does this mean you are changing your application logic just so you can unit test it, but you are making your code less efficient for the sole purpose of testability, and getting rid of a lot of advantages using an ORM has in the first place. Furthermore, since a lot of the IDs for my models are database generated, I proved to have to write additional code to handle the non-database tests since IDs were never generated and I had to still handle those cases for the unit tests to pass, yet they would never occur in real scenarios. Thus I ended up throwing out my database unit testing. Writing unit tests for controllers was easy as long as I was returning views. However, the major part of my application (and the one that would benefit most from unit testing) is a complicated ajax web application. For various reasons I decided to change the app from returning views to returning JSON with the data I needed. After this occurred my unit tests became extremely painful to write, as I have not found any good way to write unit tests for non-trivial json. After pounding my head and wasting a ton of time trying to find a good way to unit test the JSON, I gave up and deleted all of my controller unit tests (all controller actions are focused on this part of the app so far). So finally I was left with testing the Service layer (BLL). Right now I am using EF4, however I had this issue with linq-sql as well. I chose to do the EF4 model-first approach because to me, it makes sense to do it that way (define my business objects and let the framework figure out how to translate it into the sql backend). This was fine at the beginning but now it is becoming cumbersome due to relationships. For example say I have Project, User, and Object entities. One Object must be associated to a project, and a project must be associated to a user. This is not only a database specific rule, these are my business rules as well. However, say I want to do a unit test that I am able to save an object (for a simple example). I now have to do the following code just to make sure the save worked: User usr = new User { Name = "Me" }; _userService.SaveUser(usr); Project prj = new Project { Name = "Test Project", Owner = usr }; _projectService.SaveProject(prj); Object obj = new Object { Name = "Test Object" }; _objectService.SaveObject(obj); // Perform verifications There are many issues with having to do all this just to perform one unit test. There are several issues with this. For starters, if I add a new dependency, such as all projects must belong to a category, I must go into EVERY single unit test that references a project, add code to save the category then add code to add the category to the project. This can be a HUGE effort down the road for a very simple business logic change, and yet almost none of the unit tests I will be modifying for this requirement are actually meant to test that feature/requirement. If I then add verifications to my SaveProject method, so that projects cannot be saved unless they have a name with at least 5 characters, I then have to go through every Object and Project unit test to make sure that the new requirement doesn't make any unrelated unit tests fail. If there is an issue in the UserService.SaveUser() method it will cause all project, and object unit tests to fail and it the cause won't be immediately noticeable without having to dig through the exceptions. Thus I have removed all service layer unit tests from my project. I could go on and on, but so far I have not seen any way for unit testing to actually help me and not get in my way. I can see specific cases where I can, and probably will, implement unit tests, such as making sure my data verification methods work correctly, but those cases are few and far between. Some of my issues can probably be mitigated but not without adding extra layers to my application, and thus making more points of failure just so I can unit test. Thus I have no unit tests left in my code. Luckily I heavily use source control so I can get them back if I need but I just don't see the point. Everywhere on the internet I see people talking about how great TDD unit tests are, and I'm not just talking about the fanatical people. The few people who dismiss TDD/Unit tests give bad arguments claiming they are more efficient debugging by hand through the IDE, or that their coding skills are amazing that they don't need it. I recognize that both of those arguments are utter bullocks, especially for a project that needs to be maintainable by multiple developers, but any valid rebuttals to TDD seem to be few and far between. So the point of this post is to ask, am I just not understanding how to use TDD and automatic unit tests?

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  • Problems with real-valued input deep belief networks (of RBMs)

    - by Junier
    I am trying to recreate the results reported in Reducing the dimensionality of data with neural networks of autoencoding the olivetti face dataset with an adapted version of the MNIST digits matlab code, but am having some difficulty. It seems that no matter how much tweaking I do on the number of epochs, rates, or momentum the stacked RBMs are entering the fine-tuning stage with a large amount of error and consequently fail to improve much at the fine-tuning stage. I am also experiencing a similar problem on another real-valued dataset. For the first layer I am using a RBM with a smaller learning rate (as described in the paper) and with negdata = poshidstates*vishid' + repmat(visbiases,numcases,1); I'm fairly confident I am following the instructions found in the supporting material but I cannot achieve the correct errors. Is there something I am missing? See the code I'm using for real-valued visible unit RBMs below, and for the whole deep training. The rest of the code can be found here. rbmvislinear.m: epsilonw = 0.001; % Learning rate for weights epsilonvb = 0.001; % Learning rate for biases of visible units epsilonhb = 0.001; % Learning rate for biases of hidden units weightcost = 0.0002; initialmomentum = 0.5; finalmomentum = 0.9; [numcases numdims numbatches]=size(batchdata); if restart ==1, restart=0; epoch=1; % Initializing symmetric weights and biases. vishid = 0.1*randn(numdims, numhid); hidbiases = zeros(1,numhid); visbiases = zeros(1,numdims); poshidprobs = zeros(numcases,numhid); neghidprobs = zeros(numcases,numhid); posprods = zeros(numdims,numhid); negprods = zeros(numdims,numhid); vishidinc = zeros(numdims,numhid); hidbiasinc = zeros(1,numhid); visbiasinc = zeros(1,numdims); sigmainc = zeros(1,numhid); batchposhidprobs=zeros(numcases,numhid,numbatches); end for epoch = epoch:maxepoch, fprintf(1,'epoch %d\r',epoch); errsum=0; for batch = 1:numbatches, if (mod(batch,100)==0) fprintf(1,' %d ',batch); end %%%%%%%%% START POSITIVE PHASE %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% data = batchdata(:,:,batch); poshidprobs = 1./(1 + exp(-data*vishid - repmat(hidbiases,numcases,1))); batchposhidprobs(:,:,batch)=poshidprobs; posprods = data' * poshidprobs; poshidact = sum(poshidprobs); posvisact = sum(data); %%%%%%%%% END OF POSITIVE PHASE %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% poshidstates = poshidprobs > rand(numcases,numhid); %%%%%%%%% START NEGATIVE PHASE %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% negdata = poshidstates*vishid' + repmat(visbiases,numcases,1);% + randn(numcases,numdims) if not using mean neghidprobs = 1./(1 + exp(-negdata*vishid - repmat(hidbiases,numcases,1))); negprods = negdata'*neghidprobs; neghidact = sum(neghidprobs); negvisact = sum(negdata); %%%%%%%%% END OF NEGATIVE PHASE %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% err= sum(sum( (data-negdata).^2 )); errsum = err + errsum; if epoch>5, momentum=finalmomentum; else momentum=initialmomentum; end; %%%%%%%%% UPDATE WEIGHTS AND BIASES %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% vishidinc = momentum*vishidinc + ... epsilonw*( (posprods-negprods)/numcases - weightcost*vishid); visbiasinc = momentum*visbiasinc + (epsilonvb/numcases)*(posvisact-negvisact); hidbiasinc = momentum*hidbiasinc + (epsilonhb/numcases)*(poshidact-neghidact); vishid = vishid + vishidinc; visbiases = visbiases + visbiasinc; hidbiases = hidbiases + hidbiasinc; %%%%%%%%%%%%%%%% END OF UPDATES %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% end fprintf(1, '\nepoch %4i error %f \n', epoch, errsum); end dofacedeepauto.m: clear all close all maxepoch=200; %In the Science paper we use maxepoch=50, but it works just fine. numhid=2000; numpen=1000; numpen2=500; numopen=30; fprintf(1,'Pretraining a deep autoencoder. \n'); fprintf(1,'The Science paper used 50 epochs. This uses %3i \n', maxepoch); load fdata %makeFaceData; [numcases numdims numbatches]=size(batchdata); fprintf(1,'Pretraining Layer 1 with RBM: %d-%d \n',numdims,numhid); restart=1; rbmvislinear; hidrecbiases=hidbiases; save mnistvh vishid hidrecbiases visbiases; maxepoch=50; fprintf(1,'\nPretraining Layer 2 with RBM: %d-%d \n',numhid,numpen); batchdata=batchposhidprobs; numhid=numpen; restart=1; rbm; hidpen=vishid; penrecbiases=hidbiases; hidgenbiases=visbiases; save mnisthp hidpen penrecbiases hidgenbiases; fprintf(1,'\nPretraining Layer 3 with RBM: %d-%d \n',numpen,numpen2); batchdata=batchposhidprobs; numhid=numpen2; restart=1; rbm; hidpen2=vishid; penrecbiases2=hidbiases; hidgenbiases2=visbiases; save mnisthp2 hidpen2 penrecbiases2 hidgenbiases2; fprintf(1,'\nPretraining Layer 4 with RBM: %d-%d \n',numpen2,numopen); batchdata=batchposhidprobs; numhid=numopen; restart=1; rbmhidlinear; hidtop=vishid; toprecbiases=hidbiases; topgenbiases=visbiases; save mnistpo hidtop toprecbiases topgenbiases; backpropface; Thanks for your time

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  • Problems with real-valued deep belief networks (of RBMs)

    - by Junier
    I am trying to recreate the results reported in Reducing the dimensionality of data with neural networks of autoencoding the olivetti face dataset with an adapted version of the MNIST digits matlab code, but am having some difficulty. It seems that no matter how much tweaking I do on the number of epochs, rates, or momentum the stacked RBMs are entering the fine-tuning stage with a large amount of error and consequently fail to improve much at the fine-tuning stage. I am also experiencing a similar problem on another real-valued dataset. For the first layer I am using a RBM with a smaller learning rate (as described in the paper) and with negdata = poshidstates*vishid' + repmat(visbiases,numcases,1); I'm fairly confident I am following the instructions found in the supporting material but I cannot achieve the correct errors. Is there something I am missing? See the code I'm using for real-valued visible unit RBMs below, and for the whole deep training. The rest of the code can be found here. rbmvislinear.m: epsilonw = 0.001; % Learning rate for weights epsilonvb = 0.001; % Learning rate for biases of visible units epsilonhb = 0.001; % Learning rate for biases of hidden units weightcost = 0.0002; initialmomentum = 0.5; finalmomentum = 0.9; [numcases numdims numbatches]=size(batchdata); if restart ==1, restart=0; epoch=1; % Initializing symmetric weights and biases. vishid = 0.1*randn(numdims, numhid); hidbiases = zeros(1,numhid); visbiases = zeros(1,numdims); poshidprobs = zeros(numcases,numhid); neghidprobs = zeros(numcases,numhid); posprods = zeros(numdims,numhid); negprods = zeros(numdims,numhid); vishidinc = zeros(numdims,numhid); hidbiasinc = zeros(1,numhid); visbiasinc = zeros(1,numdims); sigmainc = zeros(1,numhid); batchposhidprobs=zeros(numcases,numhid,numbatches); end for epoch = epoch:maxepoch, fprintf(1,'epoch %d\r',epoch); errsum=0; for batch = 1:numbatches, if (mod(batch,100)==0) fprintf(1,' %d ',batch); end %%%%%%%%% START POSITIVE PHASE %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% data = batchdata(:,:,batch); poshidprobs = 1./(1 + exp(-data*vishid - repmat(hidbiases,numcases,1))); batchposhidprobs(:,:,batch)=poshidprobs; posprods = data' * poshidprobs; poshidact = sum(poshidprobs); posvisact = sum(data); %%%%%%%%% END OF POSITIVE PHASE %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% poshidstates = poshidprobs > rand(numcases,numhid); %%%%%%%%% START NEGATIVE PHASE %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% negdata = poshidstates*vishid' + repmat(visbiases,numcases,1);% + randn(numcases,numdims) if not using mean neghidprobs = 1./(1 + exp(-negdata*vishid - repmat(hidbiases,numcases,1))); negprods = negdata'*neghidprobs; neghidact = sum(neghidprobs); negvisact = sum(negdata); %%%%%%%%% END OF NEGATIVE PHASE %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% err= sum(sum( (data-negdata).^2 )); errsum = err + errsum; if epoch>5, momentum=finalmomentum; else momentum=initialmomentum; end; %%%%%%%%% UPDATE WEIGHTS AND BIASES %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% vishidinc = momentum*vishidinc + ... epsilonw*( (posprods-negprods)/numcases - weightcost*vishid); visbiasinc = momentum*visbiasinc + (epsilonvb/numcases)*(posvisact-negvisact); hidbiasinc = momentum*hidbiasinc + (epsilonhb/numcases)*(poshidact-neghidact); vishid = vishid + vishidinc; visbiases = visbiases + visbiasinc; hidbiases = hidbiases + hidbiasinc; %%%%%%%%%%%%%%%% END OF UPDATES %%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%%% end fprintf(1, '\nepoch %4i error %f \n', epoch, errsum); end dofacedeepauto.m: clear all close all maxepoch=200; %In the Science paper we use maxepoch=50, but it works just fine. numhid=2000; numpen=1000; numpen2=500; numopen=30; fprintf(1,'Pretraining a deep autoencoder. \n'); fprintf(1,'The Science paper used 50 epochs. This uses %3i \n', maxepoch); load fdata %makeFaceData; [numcases numdims numbatches]=size(batchdata); fprintf(1,'Pretraining Layer 1 with RBM: %d-%d \n',numdims,numhid); restart=1; rbmvislinear; hidrecbiases=hidbiases; save mnistvh vishid hidrecbiases visbiases; maxepoch=50; fprintf(1,'\nPretraining Layer 2 with RBM: %d-%d \n',numhid,numpen); batchdata=batchposhidprobs; numhid=numpen; restart=1; rbm; hidpen=vishid; penrecbiases=hidbiases; hidgenbiases=visbiases; save mnisthp hidpen penrecbiases hidgenbiases; fprintf(1,'\nPretraining Layer 3 with RBM: %d-%d \n',numpen,numpen2); batchdata=batchposhidprobs; numhid=numpen2; restart=1; rbm; hidpen2=vishid; penrecbiases2=hidbiases; hidgenbiases2=visbiases; save mnisthp2 hidpen2 penrecbiases2 hidgenbiases2; fprintf(1,'\nPretraining Layer 4 with RBM: %d-%d \n',numpen2,numopen); batchdata=batchposhidprobs; numhid=numopen; restart=1; rbmhidlinear; hidtop=vishid; toprecbiases=hidbiases; topgenbiases=visbiases; save mnistpo hidtop toprecbiases topgenbiases; backpropface; Thanks for your time

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  • Exchange Server 2007 Setup

    - by AlamedaDad
    Hi, I'm working on a upgrade to Exchange 2007 and I wanted to get some advise on hardware choices. We currently have an Exchange 2003 STD server with 400 users split between 6 AD Sites, that is housed on a single server. We need to move to a redundant, fault tolerant system to support our users. I'm planning on installing 2 Dell 1950 servers with W2k8-std to act as CAS and Hub servers, with NLB to allow abstraction of the actual server name to the users. There won't be an edge system since we have a Barracuda box already that will handle in/out spam/virus filtering. Backend I'm planning on 2 mailbox servers which will be Dell 2950s with 16GB RAM, 2 either dual-core or quad-core CPUs and 6 300GB SAS drives in some RAID config. These systems will be clustered using W2k8 Ent clustering and running CCR in Exchange. My questions are as follows: Is 16GB enough RAM for serving that many mailboxes along with the windows clustering and ccr? I'm trying to figure out disk layouts and I'm unsure of whether to use all local disk or some local and some SAN, via an OpenFiler iSCSI server. The SAN would be a Dell 2850 with 6 - 300GB SCSI drives and a PERC controller to slice as I want, with 8GB RAM. Option 1: 2 drives, RAID 1 - OS 2 drives, RAID 1 - Logs 2 drives, RAID 1 - Mail stores Option 2: 2 drives, RAID 1 - OS and logs 4 drives, RAID 5 - Mail Stores and scratch space for eseutil. Option 3: 2 drives, RAID 1 - OS 2 drives, RAID 1 - Logs 2 drives, RAID 0 - scratch space ~300GB iSCSI volume for mail stores Option 4: 2 drives, RAID 1 - OS 4 drives, RAID 5 - scratch space ~300GB iSCSI volume for mail stores ~300GB iSCSI volume for logs I have 2 sockets for CPUs and need to chose between dual and quad cores. The dual core have faster clocks but less cache and I'm thinking older architecture. Am I better off with more cores and cache while sacraficing clock speed? I am planning on adding the new E2K7 cluster to the E2K3 server and then move each mailbox over, all at once, then remove the old server. This seems more complicated than simply getting rid of the 2003 server and then adding the 2007 cluster and restoring the mailboxes using PowerControls or exmerge. The migration option lets me do this on my time, where a cutover means it all needs to work at once. If I go with the cutover method, how can I prebuild the servers and add them to the domain right after removing the 2003 server, or can't I? I think the answer is no and the migration is my only real option if I want to prebuild. I need to also migrate about 30GB of Public Folders. Is there anything special about this, other than specifying in the E2K7 install that I want older Outlook clients and PF's setup? I guess I could even keep the E2K3 server to host just the PFs? Lastly, if I have a mix of Outlook 200, 2003 and 2007 what do I need to do to make sure they all have access to the GAL and OAB? At time of cutover, we'll be at like 90% 2007, but we will have some older stuff around. My plan is to use Outlook Anywhere on laptops that are used outside the physical network. Are there any gotchas involved in that? I'm even thinking about using is for all Outlook clients, does anyone do that? The reason I'm considering it is that our WAN is really VPN tunnels over internet connections, so not a fully messhed, stable WAN. Thank you all very much for the assistance in advance and I look forward to discussion of these points! Regards...Michael

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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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  • May 20th Links: ASP.NET MVC, ASP.NET, .NET 4, VS 2010, Silverlight

    - by ScottGu
    Here is the latest in my link-listing series.  Also check out my VS 2010 and .NET 4 series and ASP.NET MVC 2 series for other on-going blog series I’m working on. [In addition to blogging, I am also now using Twitter for quick updates and to share links. Follow me at: twitter.com/scottgu] ASP.NET MVC How to Localize an ASP.NET MVC Application: Michael Ceranski has a good blog post that describes how to localize ASP.NET MVC 2 applications. ASP.NET MVC with jTemplates Part 1 and Part 2: Steve Gentile has a nice two-part set of blog posts that demonstrate how to use the jTemplate and DataTable jQuery libraries to implement client-side data binding with ASP.NET MVC. CascadingDropDown jQuery Plugin for ASP.NET MVC: Raj Kaimal has a nice blog post that demonstrates how to implement a dynamically constructed cascading dropdownlist on the client using jQuery and ASP.NET MVC. How to Configure VS 2010 Code Coverage for ASP.NET MVC Unit Tests: Visual Studio enables you to calculate the “code coverage” of your unit tests.  This measures the percentage of code within your application that is exercised by your tests – and can give you a sense of how much test coverage you have.  Gunnar Peipman demonstrates how to configure this for ASP.NET MVC projects. Shrinkr URL Shortening Service Sample: A nice open source application and code sample built by Kazi Manzur that demonstrates how to implement a URL Shortening Services (like bit.ly) using ASP.NET MVC 2 and EF4.  More details here. Creating RSS Feeds in ASP.NET MVC: Damien Guard has a nice post that describes a cool new “FeedResult” class he created that makes it easy to publish and expose RSS feeds from within ASP.NET MVC sites. NoSQL with MongoDB, NoRM and ASP.NET MVC Part 1 and Part 2: Nice two-part blog series by Shiju Varghese on how to use MongoDB (a document database) with ASP.NET MVC.  If you are interested in document databases also make sure to check out the Raven DB project from Ayende. Using the FCKEditor with ASP.NET MVC: Quick blog post that describes how to use FCKEditor – an open source HTML Text Editor – with ASP.NET MVC. ASP.NET Replace Html.Encode Calls with the New HTML Encoding Syntax: Phil Haack has a good blog post that describes a useful way to quickly update your ASP.NET pages and ASP.NET MVC views to use the new <%: %> encoding syntax in ASP.NET 4.  I blogged about the new <%: %> syntax – it provides an easy and concise way to HTML encode content. Integrating Twitter into an ASP.NET Website using OAuth: Scott Mitchell has a nice article that describes how to take advantage of Twiter within an ASP.NET Website using the OAuth protocol – which is a simple, secure protocol for granting API access. Creating an ASP.NET report using VS 2010 Part 1, Part 2, and Part 3: Raj Kaimal has a nice three part set of blog posts that detail how to use SQL Server Reporting Services, ASP.NET 4 and VS 2010 to create a dynamic reporting solution. Three Hidden Extensibility Gems in ASP.NET 4: Phil Haack blogs about three obscure but useful extensibility points enabled with ASP.NET 4. .NET 4 Entity Framework 4 Video Series: Julie Lerman has a nice, free, 7-part video series on MSDN that walks through how to use the new EF4 capabilities with VS 2010 and .NET 4.  I’ll be covering EF4 in a blog series that I’m going to start shortly as well. Getting Lazy with System.Lazy: System.Lazy and System.Lazy<T> are new features in .NET 4 that provide a way to create objects that may need to perform time consuming operations and defer the execution of the operation until it is needed.  Derik Whittaker has a nice write-up that describes how to use it. LINQ to Twitter: Nifty open source library on Codeplex that enables you to use LINQ syntax to query Twitter. Visual Studio 2010 Using Intellitrace in VS 2010: Chris Koenig has a nice 10 minute video that demonstrates how to use the new Intellitrace features of VS 2010 to enable DVR playback of your debug sessions. Make the VS 2010 IDE Colors look like VS 2008: Scott Hanselman has a nice blog post that covers the Visual Studio Color Theme Editor extension – which allows you to customize the VS 2010 IDE however you want. How to understand your code using Dependency Graphs, Sequence Diagrams, and the Architecture Explorer: Jennifer Marsman has a nice blog post describes how to take advantage of some of the new architecture features within VS 2010 to quickly analyze applications and legacy code-bases. How to maintain control of your code using Layer Diagrams: Another great blog post by Jennifer Marsman that demonstrates how to setup a “layer diagram” within VS 2010 to enforce clean layering within your applications.  This enables you to enforce a compiler error if someone inadvertently violates a layer design rule. Collapse Selection in Solution Explorer Extension: Useful VS 2010 extension that enables you to quickly collapse “child nodes” within the Visual Studio Solution Explorer.  If you have deeply nested project structures this extension is useful. Silverlight and Windows Phone 7 Building a Simple Windows Phone 7 Application: A nice tutorial blog post that demonstrates how to take advantage of Expression Blend to create an animated Windows Phone 7 application. If you haven’t checked out my Windows Phone 7 Twitter Tutorial I also recommend reading that. Hope this helps, Scott P.S. If you haven’t already, check out this month’s "Find a Hoster” page on the www.asp.net website to learn about great (and very inexpensive) ASP.NET hosting offers.

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  • SQL SERVER – Enumerations in Relational Database – Best Practice

    - by pinaldave
    Marko Parkkola This article has been submitted by Marko Parkkola, Data systems designer at Saarionen Oy, Finland. Marko is excellent developer and always thinking at next level. You can read his earlier comment which created very interesting discussion here: SQL SERVER- IF EXISTS(Select null from table) vs IF EXISTS(Select 1 from table). I must express my special thanks to Marko for sending this best practice for Enumerations in Relational Database. He has really wrote excellent piece here and welcome comments here. Enumerations in Relational Database This is a subject which is very basic thing in relational databases but often not very well understood and sometimes badly implemented. There are of course many ways to do this but I concentrate only two cases, one which is “the right way” and one which is definitely wrong way. The concept Let’s say we have table Person in our database. Person has properties/fields like Firstname, Lastname, Birthday and so on. Then there’s a field that tells person’s marital status and let’s name it the same way; MaritalStatus. Now MaritalStatus is an enumeration. In C# I would definitely make it an enumeration with values likes Single, InRelationship, Married, Divorced. Now here comes the problem, SQL doesn’t have enumerations. The wrong way This is, in my opinion, absolutely the wrong way to do this. It has one upside though; you’ll see the enumeration’s description instantly when you do simple SELECT query and you don’t have to deal with mysterious values. There’s plenty of downsides too and one would be database fragmentation. Consider this (I’ve left all indexes and constraints out of the query on purpose). CREATE TABLE [dbo].[Person] ( [Firstname] NVARCHAR(100), [Lastname] NVARCHAR(100), [Birthday] datetime, [MaritalStatus] NVARCHAR(10) ) You have nvarchar(20) field in the table that tells the marital status. Obvious problem with this is that what if you create a new value which doesn’t fit into 20 characters? You’ll have to come and alter the table. There are other problems also but I’ll leave those for the reader to think about. The correct way Here’s how I’ve done this in many projects. This model still has one problem but it can be alleviated in the application layer or with CHECK constraints if you like. First I will create a namespace table which tells the name of the enumeration. I will add one row to it too. I’ll write all the indexes and constraints here too. CREATE TABLE [CodeNamespace] ( [Id] INT IDENTITY(1, 1), [Name] NVARCHAR(100) NOT NULL, CONSTRAINT [PK_CodeNamespace] PRIMARY KEY ([Id]), CONSTRAINT [IXQ_CodeNamespace_Name] UNIQUE NONCLUSTERED ([Name]) ) GO INSERT INTO [CodeNamespace] SELECT 'MaritalStatus' GO Then I create a table that holds the actual values and which reference to namespace table in order to group the values under different namespaces. I’ll add couple of rows here too. CREATE TABLE [CodeValue] ( [CodeNamespaceId] INT NOT NULL, [Value] INT NOT NULL, [Description] NVARCHAR(100) NOT NULL, [OrderBy] INT, CONSTRAINT [PK_CodeValue] PRIMARY KEY CLUSTERED ([CodeNamespaceId], [Value]), CONSTRAINT [FK_CodeValue_CodeNamespace] FOREIGN KEY ([CodeNamespaceId]) REFERENCES [CodeNamespace] ([Id]) ) GO -- 1 is the 'MaritalStatus' namespace INSERT INTO [CodeValue] SELECT 1, 1, 'Single', 1 INSERT INTO [CodeValue] SELECT 1, 2, 'In relationship', 2 INSERT INTO [CodeValue] SELECT 1, 3, 'Married', 3 INSERT INTO [CodeValue] SELECT 1, 4, 'Divorced', 4 GO Now there’s four columns in CodeValue table. CodeNamespaceId tells under which namespace values belongs to. Value tells the enumeration value which is used in Person table (I’ll show how this is done below). Description tells what the value means. You can use this, for example, column in UI’s combo box. OrderBy tells if the values needs to be ordered in some way when displayed in the UI. And here’s the Person table again now with correct columns. I’ll add one row here to show how enumerations are to be used. CREATE TABLE [dbo].[Person] ( [Firstname] NVARCHAR(100), [Lastname] NVARCHAR(100), [Birthday] datetime, [MaritalStatus] INT ) GO INSERT INTO [Person] SELECT 'Marko', 'Parkkola', '1977-03-04', 3 GO Now I said earlier that there is one problem with this. MaritalStatus column doesn’t have any database enforced relationship to the CodeValue table so you can enter any value you like into this field. I’ve solved this problem in the application layer by selecting all the values from the CodeValue table and put them into a combobox / dropdownlist (with Value field as value and Description as text) so the end user can’t enter any illegal values; and of course I’ll check the entered value in data access layer also. I said in the “The wrong way” section that there is one benefit to it. In fact, you can have the same benefit here by using a simple view, which I schema bound so you can even index it if you like. CREATE VIEW [dbo].[Person_v] WITH SCHEMABINDING AS SELECT p.[Firstname], p.[Lastname], p.[BirthDay], c.[Description] MaritalStatus FROM [dbo].[Person] p JOIN [dbo].[CodeValue] c ON p.[MaritalStatus] = c.[Value] JOIN [dbo].[CodeNamespace] n ON n.[Id] = c.[CodeNamespaceId] AND n.[Name] = 'MaritalStatus' GO -- Select from View SELECT * FROM [dbo].[Person_v] GO This is excellent write up byMarko Parkkola. Do you have this kind of design setup at your organization? Let us know your opinion. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Best Practices, Database, DBA, Readers Contribution, Software Development, SQL, SQL Authority, SQL Documentation, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • Adding DTrace Probes to PHP Extensions

    - by cj
    The powerful DTrace tracing facility has some PHP-specific probes that can be enabled with --enable-dtrace. DTrace for Linux is being created by Oracle and is currently in tech preview. Currently it doesn't support userspace tracing so, in the meantime, Systemtap can be used to monitor the probes implemented in PHP. This was recently outlined in David Soria Parra's post Probing PHP with Systemtap on Linux. My post shows how DTrace probes can be added to PHP extensions and traced on Linux. I was using Oracle Linux 6.3. Not all Linux kernels are built with Systemtap, since this can impact stability. Check whether your running kernel (or others installed) have Systemtap enabled, and reboot with such a kernel: # grep CONFIG_UTRACE /boot/config-`uname -r` # grep CONFIG_UTRACE /boot/config-* When you install Systemtap itself, the package systemtap-sdt-devel is needed since it provides the sdt.h header file: # yum install systemtap-sdt-devel You can now install and build PHP as shown in David's article. Basically the build is with: $ cd ~/php-src $ ./configure --disable-all --enable-dtrace $ make (For me, running 'make' a second time failed with an error. The workaround is to do 'git checkout Zend/zend_dtrace.d' and then rerun 'make'. See PHP Bug 63704) David's article shows how to trace the probes already implemented in PHP. You can also use Systemtap to trace things like userspace PHP function calls. For example, create test.php: <?php $c = oci_connect('hr', 'welcome', 'localhost/orcl'); $s = oci_parse($c, "select dbms_xmlgen.getxml('select * from dual') xml from dual"); $r = oci_execute($s); $row = oci_fetch_array($s, OCI_NUM); $x = $row[0]->load(); $row[0]->free(); echo $x; ?> The normal output of this file is the XML form of Oracle's DUAL table: $ ./sapi/cli/php ~/test.php <?xml version="1.0"?> <ROWSET> <ROW> <DUMMY>X</DUMMY> </ROW> </ROWSET> To trace the PHP function calls, create the tracing file functrace.stp: probe process("sapi/cli/php").function("zif_*") { printf("Started function %s\n", probefunc()); } probe process("sapi/cli/php").function("zif_*").return { printf("Ended function %s\n", probefunc()); } This makes use of the way PHP userspace functions (not builtins) like oci_connect() map to C functions with a "zif_" prefix. Login as root, and run System tap on the PHP script: # cd ~cjones/php-src # stap -c 'sapi/cli/php ~cjones/test.php' ~cjones/functrace.stp Started function zif_oci_connect Ended function zif_oci_connect Started function zif_oci_parse Ended function zif_oci_parse Started function zif_oci_execute Ended function zif_oci_execute Started function zif_oci_fetch_array Ended function zif_oci_fetch_array Started function zif_oci_lob_load <?xml version="1.0"?> <ROWSET> <ROW> <DUMMY>X</DUMMY> </ROW> </ROWSET> Ended function zif_oci_lob_load Started function zif_oci_free_descriptor Ended function zif_oci_free_descriptor Each call and return is logged. The Systemtap scripting language allows complex scripts to be built. There are many examples on the web. To augment this generic capability and the PHP probes in PHP, other extensions can have probes too. Below are the steps I used to add probes to OCI8: I created a provider file ext/oci8/oci8_dtrace.d, enabling three probes. The first one will accept a parameter that runtime tracing can later display: provider php { probe oci8__connect(char *username); probe oci8__nls_start(); probe oci8__nls_done(); }; I updated ext/oci8/config.m4 with the PHP_INIT_DTRACE macro. The patch is at the end of config.m4. The macro takes the provider prototype file, a name of the header file that 'dtrace' will generate, and a list of sources files with probes. When --enable-dtrace is used during PHP configuration, then the outer $PHP_DTRACE check is true and my new probes will be enabled. I've chosen to define an OCI8 specific macro, HAVE_OCI8_DTRACE, which can be used in the OCI8 source code: diff --git a/ext/oci8/config.m4 b/ext/oci8/config.m4 index 34ae76c..f3e583d 100644 --- a/ext/oci8/config.m4 +++ b/ext/oci8/config.m4 @@ -341,4 +341,17 @@ if test "$PHP_OCI8" != "no"; then PHP_SUBST_OLD(OCI8_ORACLE_VERSION) fi + + if test "$PHP_DTRACE" = "yes"; then + AC_CHECK_HEADERS([sys/sdt.h], [ + PHP_INIT_DTRACE([ext/oci8/oci8_dtrace.d], + [ext/oci8/oci8_dtrace_gen.h],[ext/oci8/oci8.c]) + AC_DEFINE(HAVE_OCI8_DTRACE,1, + [Whether to enable DTrace support for OCI8 ]) + ], [ + AC_MSG_ERROR( + [Cannot find sys/sdt.h which is required for DTrace support]) + ]) + fi + fi In ext/oci8/oci8.c, I added the probes at, for this example, semi-arbitrary places: diff --git a/ext/oci8/oci8.c b/ext/oci8/oci8.c index e2241cf..ffa0168 100644 --- a/ext/oci8/oci8.c +++ b/ext/oci8/oci8.c @@ -1811,6 +1811,12 @@ php_oci_connection *php_oci_do_connect_ex(char *username, int username_len, char } } +#ifdef HAVE_OCI8_DTRACE + if (DTRACE_OCI8_CONNECT_ENABLED()) { + DTRACE_OCI8_CONNECT(username); + } +#endif + /* Initialize global handles if they weren't initialized before */ if (OCI_G(env) == NULL) { php_oci_init_global_handles(TSRMLS_C); @@ -1870,11 +1876,22 @@ php_oci_connection *php_oci_do_connect_ex(char *username, int username_len, char size_t rsize = 0; sword result; +#ifdef HAVE_OCI8_DTRACE + if (DTRACE_OCI8_NLS_START_ENABLED()) { + DTRACE_OCI8_NLS_START(); + } +#endif PHP_OCI_CALL_RETURN(result, OCINlsEnvironmentVariableGet, (&charsetid_nls_lang, 0, OCI_NLS_CHARSET_ID, 0, &rsize)); if (result != OCI_SUCCESS) { charsetid_nls_lang = 0; } smart_str_append_unsigned_ex(&hashed_details, charsetid_nls_lang, 0); + +#ifdef HAVE_OCI8_DTRACE + if (DTRACE_OCI8_NLS_DONE_ENABLED()) { + DTRACE_OCI8_NLS_DONE(); + } +#endif } timestamp = time(NULL); The oci_connect(), oci_pconnect() and oci_new_connect() calls all use php_oci_do_connect_ex() internally. The first probe simply records that the PHP application made a connection call. I already showed a way to do this without needing a probe, but adding a specific probe lets me record the username. The other two probes can be used to time how long the globalization initialization takes. The relationships between the oci8_dtrace.d names like oci8__connect, the probe guards like DTRACE_OCI8_CONNECT_ENABLED() and probe names like DTRACE_OCI8_CONNECT() are obvious after seeing the pattern of all three probes. I included the new header that will be automatically created by the dtrace tool when PHP is built. I did this in ext/oci8/php_oci8_int.h: diff --git a/ext/oci8/php_oci8_int.h b/ext/oci8/php_oci8_int.h index b0d6516..c81fc5a 100644 --- a/ext/oci8/php_oci8_int.h +++ b/ext/oci8/php_oci8_int.h @@ -44,6 +44,10 @@ # endif # endif /* osf alpha */ +#ifdef HAVE_OCI8_DTRACE +#include "oci8_dtrace_gen.h" +#endif + #if defined(min) #undef min #endif Now PHP can be rebuilt: $ cd ~/php-src $ rm configure && ./buildconf --force $ ./configure --disable-all --enable-dtrace \ --with-oci8=instantclient,/home/cjones/instantclient $ make If 'make' fails, do the 'git checkout Zend/zend_dtrace.d' trick I mentioned. The new probes can be seen by logging in as root and running: # stap -l 'process.provider("php").mark("oci8*")' -c 'sapi/cli/php -i' process("sapi/cli/php").provider("php").mark("oci8__connect") process("sapi/cli/php").provider("php").mark("oci8__nls_done") process("sapi/cli/php").provider("php").mark("oci8__nls_start") To test them out, create a new trace file, oci.stp: global numconnects; global start; global numcharlookups = 0; global tottime = 0; probe process.provider("php").mark("oci8-connect") { printf("Connected as %s\n", user_string($arg1)); numconnects += 1; } probe process.provider("php").mark("oci8-nls_start") { start = gettimeofday_us(); numcharlookups++; } probe process.provider("php").mark("oci8-nls_done") { tottime += gettimeofday_us() - start; } probe end { printf("Connects: %d, Charset lookups: %ld\n", numconnects, numcharlookups); printf("Total NLS charset initalization time: %ld usecs/connect\n", (numcharlookups 0 ? tottime/numcharlookups : 0)); } This calculates the average time that the NLS character set lookup takes. It also prints out the username of each connection, as an example of using parameters. Login as root and run Systemtap over the PHP script: # cd ~cjones/php-src # stap -c 'sapi/cli/php ~cjones/test.php' ~cjones/oci.stp Connected as cj <?xml version="1.0"?> <ROWSET> <ROW> <DUMMY>X</DUMMY> </ROW> </ROWSET> Connects: 1, Charset lookups: 1 Total NLS charset initalization time: 164 usecs/connect This shows the time penalty of making OCI8 look up the default character set. This time would be zero if a character set had been passed as the fourth argument to oci_connect() in test.php.

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  • Multiple Monitors

    - by mroberts
    At my workplace .Net developers get pretty much the same equipment. A decent Dell workstation / Desktop, mine is a Dell Precision 390. One dual core 2.40 GHz. Eight GB RAM. Windows 7 Enterprise 64-bit. Two Dell 20.1 Monitors. I'm happy with this.  The machine is about 3 years old but still runs with some decent speed. New developers are getting a Dell workstation with dual quad processors. I just put in a request for myself and three other developers for an upgraded video card and two additional monitors, for a total of four monitors per person.  We suggested this card, BTW, mainly for the cost.  The move from one monitor to two was fantastic (one might even say life (or work) changing) and truly did increase productivity. Now what about going from 2 monitors to 4?  I'm sure the change is not as dramatic as one to two, but I can't help but to think four monitors is better than two.  But if four is better than two, should we have asked for six?!? Also what about mixing monitor types?  Right now my monitors are the older square type vs. wide-screen.  It's been rumored that we will be getting monitors out of current stock and they will be 22 inch wide-screens.  I understand this, recession and all.  2-20 inch square monitors with 2-22 inch wide-screen monitors...hmmmmm.  I'm thinking I'd rather get 2 additional 17 inch square monitors to put on each side of my 20's. Also, a question was raised about the layout of four monitors. By default, my thought was I'll just put them all on my desk, kinda in a line. I've heard others say they want to stack them in a 2 x 2 square. BTW, loving multi monitor support in Visual studio 2010! I’d love some comments on your experience with one, two, four, or however many monitors from a developers perspective.

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  • Configure 27" 2560x1440 for a monitor with corrupt EDID

    - by Aras
    I am trying to get a monitor work with my Ubuntu laptop. The monitor is this cheap 27" Korean monitors which has a 2560x1440 resolution -- and nothing else. Here are some specifications of this monitor: 2560x1440 @60Hz Only one dual link DVI-D input -- no other input port (no HDMI or display port) no OSD no scalar reports corrupt EDID does 2560x1440 @60Hz, did I say that already? Anyways, the monitor works beautifully with my Ubuntu desktop which has an nVidia card with DVI output. However, I am having problem using this monitor with my laptop. After some searching around I found a few posts suggesting to use an active adaptor for mini display port, so I went and bought a mini display to dual link DVI-D adaptor.. When using this adaptor the monitor is recognized by nvidia-settings tool but with incorrect resolution information. As you can see the monitor is incorrectly recognized and there are no other resolution available to set. This post on ubuntu forums and this other post on overclock both suggest that the monitor is reporting corrupt EDID file. I have tried following their instructions, but so far I have not been able to display any image on the monitor from my laptop. The laptop I am using is an ASUS G75VW with a 1920x1080 screen. It has a VGA, an HDMI 1.4a, and a mini display port. The graphic card is an nvidia gforce gtx 660M with 2GB dedicated memory. I am running Ubuntu 12.10 on here which I upgrade from 12.04 a few weeks ago. As I said I have tried several suggestions, including specifying Modeline in xorg.conf and also linking to EDID files I found from those forum posts above. However, I am not sure if the EDID files I found are suitable for my monitor. I think the solution to my problem consist of obtaining the EDID file of my monitor and then fixing it and modifying xorg.conf to force nvidia driver to load the correct resolution. However, I am not sure what steps I need to take to do this. Here is the part of sudo xrandr --prop output that is related to this monitor: DP-1 connected 800x600+1920+0 (normal left inverted right x axis y axis) 0mm x 0mm SignalFormat: DisplayPort supported: DisplayPort ConnectorType: DisplayPort ConnectorNumber: 3 (0x00000003) _ConnectorLocation: 3 (0x00000003) 800x600 60.3*+ I was expecting to see the EDID file in this output as was mentioned in this post, but it is not there. After several hours of tweaking X configurations, I decided it was time to ask for help here. I would really appreciate if someone with experience regarding EDID and X configuration could give me a hand to solve this issue.

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  • Apple iPad 2 In April, iPhone 5 in June With New Hardware[Rumours]

    - by Gopinath
    Blogs and news sites are buzzing with the rumours of Apple’s next generation iPad and iPhone devices. These rumours interests the bloggers, geeks and end users of Apple devices as Apple maintains very tight lip on the new features of their upcoming products. The gadget blog Engadget has some very interesting rumours on the release of iPad 2 & iPhone 5 as well the new hardware they are going to have. Lets get into the details if you love to read the rumours of high profile blogs iPad 2 Release Date and Specs Apple seems to be all set to release iPad 2 in April, that is almost an year after the release of first iPad. It’s common for Apple to enjoy an one year long time to release a new version of their products. So if at all the rumours are to be believed, I can place an order of iPad 2 in April. Just like many of you out there, I’m also holding my iPad buying instinct and waiting for iPad 2 as it’s going to have at the minimum retina display,  Facetime features and few game changing features in Apple’s style. The report claims, iPad 2 will have a front and back cameras retina display SD Card slot (seems to be no USB) a dual GSM / CDMA chipset, that lets you use it with both GSM(AT &T, Airte) and CDMA(Verizon, Reliance) telecom providers iPhone 5 Release Date and Specs When it comes to iPhone 5 information, the rumour claims that the new iPhone is a completed redesigned device and it’s slated to release in summer of United States(i.e. June 2011). The device is also being tested by senior Apple executives right inside the campus and strictly not allowed to carry it outside. This restriction is to make sure that iPhone 5 will not land land up in a bar and then in the hands of geek blogs like how it happened with iPhone 4 last year. When it comes to the hardware of iPhone 5 Apple’s new A5 CPU (a Cortex A9-based, multi-core chip) a dual GSM / CDMA chipset, that lets you use it with both GSM(AT &T, Airte) and CDMA(Verizon, Reliance) telecom providers via Engadget and cc image credit flickr/mr-blixt This article titled,Apple iPad 2 In April, iPhone 5 in June With New Hardware[Rumours], was originally published at Tech Dreams. Grab our rss feed or fan us on Facebook to get updates from us.

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