Search Results

Search found 23274 results on 931 pages for 'on call'.

Page 416/931 | < Previous Page | 412 413 414 415 416 417 418 419 420 421 422 423  | Next Page >

  • Integrating Coherence & Java EE 6 Applications using ActiveCache

    - by Ricardo Ferreira
    OK, so you are a developer and are starting a new Java EE 6 application using the most wonderful features of the Java EE platform like Enterprise JavaBeans, JavaServer Faces, CDI, JPA e another cool stuff technologies. And your architecture need to hold piece of data into distributed caches to improve application's performance, scalability and reliability? If this is your current facing scenario, maybe you should look closely in the solutions provided by Oracle WebLogic Server. Oracle had integrated WebLogic Server and its champion data caching technology called Oracle Coherence. This seamless integration between this two products provides a comprehensive environment to develop applications without the complexity of extra Java code to manage cache as a dependency, since Oracle provides an DI ("Dependency Injection") mechanism for Coherence, the same DI mechanism available in standard Java EE applications. This feature is called ActiveCache. In this article, I will show you how to configure ActiveCache in WebLogic and at your Java EE application. Configuring WebLogic to manage Coherence Before you start changing your application to use Coherence, you need to configure your Coherence distributed cache. The good news is, you can manage all this stuff without writing a single line of code of XML or even Java. This configuration can be done entirely in the WebLogic administration console. The first thing to do is the setup of a Coherence cluster. A Coherence cluster is a set of Coherence JVMs configured to form one single view of the cache. This means that you can insert or remove members of the cluster without the client application (the application that generates or consume data from the cache) knows about the changes. This concept allows your solution to scale-out without changing the application server JVMs. You can growth your application only in the data grid layer. To start the configuration, you need to configure an machine that points to the server in which you want to execute the Coherence JVMs. WebLogic Server allows you to do this very easily using the Administration Console. In this example, I will call the machine as "coherence-server". Remember that in order to the machine concept works, you need to ensure that the NodeManager are being executed in the target server that the machine points to. The NodeManager executable can be found in <WLS_HOME>/server/bin/startNodeManager.sh. The next thing to do is to configure a Coherence cluster. In the WebLogic administration console, go to Environment > Coherence Clusters and click in "New". Call this Coherence cluster of "my-coherence-cluster". Click in next. Specify a valid cluster address and port. The Coherence members will communicate with each other through this address and port. Our Coherence cluster are now configured. Now it is time to configure the Coherence members and add them to this cluster. In the WebLogic administration console, go to Environment > Coherence Servers and click in "New". In the field "Name" set to "coh-server-1". In the field "Machine", associate this Coherence server to the machine "coherence-server". In the field "Cluster", associate this Coherence server to the cluster named "my-coherence-cluster". Click in "Finish". Start the Coherence server using the "Control" tab of WebLogic administration console. This will instruct WebLogic to start a new JVM of Coherence in the target machine that should join the pre-defined Coherence cluster. Configuring your Java EE Application to Access Coherence Now lets pass to the funny part of the configuration. The first thing to do is to inform your Java EE application which Coherence cluster to join. Oracle had updated WebLogic server deployment descriptors so you will not have to change your code or the containers deployment descriptors like application.xml, ejb-jar.xml or web.xml. In this example, I will show you how to enable DI ("Dependency Injection") to a Coherence cache from a Servlet 3.0 component. In the WEB-INF/weblogic.xml deployment descriptor, put the following metadata information: <?xml version="1.0" encoding="UTF-8"?> <wls:weblogic-web-app xmlns:wls="http://xmlns.oracle.com/weblogic/weblogic-web-app" xmlns:xsi="http://www.w3.org/2001/XMLSchema-instance" xsi:schemaLocation="http://java.sun.com/xml/ns/javaee http://java.sun.com/xml/ns/javaee/web-app_2_5.xsd http://xmlns.oracle.com/weblogic/weblogic-web-app http://xmlns.oracle.com/weblogic/weblogic-web-app/1.4/weblogic-web-app.xsd"> <wls:context-root>myWebApp</wls:context-root> <wls:coherence-cluster-ref> <wls:coherence-cluster-name>my-coherence-cluster</wls:coherence-cluster-name> </wls:coherence-cluster-ref> </wls:weblogic-web-app> As you can see, using the "coherence-cluster-name" tag, we are informing our Java EE application that it should join the "my-coherence-cluster" when it loads in the web container. Without this information, the application will not be able to access the predefined Coherence cluster. It will form its own Coherence cluster without any members. So never forget to put this information. Now put the coherence.jar and active-cache-1.0.jar dependencies at your WEB-INF/lib application classpath. You need to deploy this dependencies so ActiveCache can automatically take care of the Coherence cluster join phase. This dependencies can be found in the following locations: - <WLS_HOME>/common/deployable-libraries/active-cache-1.0.jar - <COHERENCE_HOME>/lib/coherence.jar Finally, you need to write down the access code to the Coherence cache at your Servlet. In the following example, we have a Servlet 3.0 component that access a Coherence cache named "transactions" and prints into the browser output the content (the ammount property) of one specific transaction. package com.oracle.coherence.demo.activecache; import java.io.IOException; import javax.annotation.Resource; import javax.servlet.ServletException; import javax.servlet.annotation.WebServlet; import javax.servlet.http.HttpServlet; import javax.servlet.http.HttpServletRequest; import javax.servlet.http.HttpServletResponse; import com.tangosol.net.NamedCache; @WebServlet("/demo/specificTransaction") public class TransactionServletExample extends HttpServlet { @Resource(mappedName = "transactions") NamedCache transactions; protected void doGet(HttpServletRequest request, HttpServletResponse response) throws ServletException, IOException { int transId = Integer.parseInt(request.getParameter("transId")); Transaction transaction = (Transaction) transactions.get(transId); response.getWriter().println("<center>" + transaction.getAmmount() + "</center>"); } } Thats it! No more configuration is necessary and you have all set to start producing and getting data to/from Coherence. As you can see in the example code, the Coherence cache are treated as a normal dependency in the Java EE container. The magic happens behind the scenes when the ActiveCache allows your application to join the defined Coherence cluster. The most interesting thing about this approach is, no matter which type of Coherence cache your are using (Distributed, Partitioned, Replicated, WAN-Remote) for the client application, it is just a simple attribute member of com.tangosol.net.NamedCache type. And its all managed by the Java EE container as an dependency. This means that if you inject the same dependency (the Coherence cache named "transactions") in another Java EE component (JSF managed-bean, Stateless EJB) the cache will be the same. Cool isn't it? Thanks to the CDI technology, we can extend the same support for non-Java EE standards components like simple POJOs. This means that you are not forced to only use Servlets, EJBs or JSF in order to inject Coherence caches. You can do the same approach for regular POJOs created for you and managed by lightweight containers like Spring or Seam.

    Read the article

  • The Shift: how Orchard painlessly shifted to document storage, and how it’ll affect you

    - by Bertrand Le Roy
    We’ve known it all along. The storage for Orchard content items would be much more efficient using a document database than a relational one. Orchard content items are composed of parts that serialize naturally into infoset kinds of documents. Storing them as relational data like we’ve done so far was unnatural and requires the data for a single item to span multiple tables, related through 1-1 relationships. This means lots of joins in queries, and a great potential for Select N+1 problems. Document databases, unfortunately, are still a tough sell in many places that prefer the more familiar relational model. Being able to x-copy Orchard to hosters has also been a basic constraint in the design of Orchard. Combine those with the necessity at the time to run in medium trust, and with license compatibility issues, and you’ll find yourself with very few reasonable choices. So we went, a little reluctantly, for relational SQL stores, with the dream of one day transitioning to document storage. We have played for a while with the idea of building our own document storage on top of SQL databases, and Sébastien implemented something more than decent along those lines, but we had a better way all along that we didn’t notice until recently… In Orchard, there are fields, which are named properties that you can add dynamically to a content part. Because they are so dynamic, we have been storing them as XML into a column on the main content item table. This infoset storage and its associated API are fairly generic, but were only used for fields. The breakthrough was when Sébastien realized how this existing storage could give us the advantages of document storage with minimal changes, while continuing to use relational databases as the substrate. public bool CommercialPrices { get { return this.Retrieve(p => p.CommercialPrices); } set { this.Store(p => p.CommercialPrices, value); } } This code is very compact and efficient because the API can infer from the expression what the type and name of the property are. It is then able to do the proper conversions for you. For this code to work in a content part, there is no need for a record at all. This is particularly nice for site settings: one query on one table and you get everything you need. This shows how the existing infoset solves the data storage problem, but you still need to query. Well, for those properties that need to be filtered and sorted on, you can still use the current record-based relational system. This of course continues to work. We do however provide APIs that make it trivial to store into both record properties and the infoset storage in one operation: public double Price { get { return Retrieve(r => r.Price); } set { Store(r => r.Price, value); } } This code looks strikingly similar to the non-record case above. The difference is that it will manage both the infoset and the record-based storages. The call to the Store method will send the data in both places, keeping them in sync. The call to the Retrieve method does something even cooler: if the property you’re looking for exists in the infoset, it will return it, but if it doesn’t, it will automatically look into the record for it. And if that wasn’t cool enough, it will take that value from the record and store it into the infoset for the next time it’s required. This means that your data will start automagically migrating to infoset storage just by virtue of using the code above instead of the usual: public double Price { get { return Record.Price; } set { Record.Price = value; } } As your users browse the site, it will get faster and faster as Select N+1 issues will optimize themselves away. If you preferred, you could still have explicit migration code, but it really shouldn’t be necessary most of the time. If you do already have code using QueryHints to mitigate Select N+1 issues, you might want to reconsider those, as with the new system, you’ll want to avoid joins that you don’t need for filtering or sorting, further optimizing your queries. There are some rare cases where the storage of the property must be handled differently. Check out this string[] property on SearchSettingsPart for example: public string[] SearchedFields { get { return (Retrieve<string>("SearchedFields") ?? "") .Split(new[] {',', ' '}, StringSplitOptions.RemoveEmptyEntries); } set { Store("SearchedFields", String.Join(", ", value)); } } The array of strings is transformed by the property accessors into and from a comma-separated list stored in a string. The Retrieve and Store overloads used in this case are lower-level versions that explicitly specify the type and name of the attribute to retrieve or store. You may be wondering what this means for code or operations that look directly at the database tables instead of going through the new infoset APIs. Even if there is a record, the infoset version of the property will win if it exists, so it is necessary to keep the infoset up-to-date. It’s not very complicated, but definitely something to keep in mind. Here is what a product record looks like in Nwazet.Commerce for example: And here is the same data in the infoset: The infoset is stored in Orchard_Framework_ContentItemRecord or Orchard_Framework_ContentItemVersionRecord, depending on whether the content type is versionable or not. A good way to find what you’re looking for is to inspect the record table first, as it’s usually easier to read, and then get the item record of the same id. Here is the detailed XML document for this product: <Data> <ProductPart Inventory="40" Price="18" Sku="pi-camera-box" OutOfStockMessage="" AllowBackOrder="false" Weight="0.2" Size="" ShippingCost="null" IsDigital="false" /> <ProductAttributesPart Attributes="" /> <AutoroutePart DisplayAlias="camera-box" /> <TitlePart Title="Nwazet Pi Camera Box" /> <BodyPart Text="[...]" /> <CommonPart CreatedUtc="2013-09-10T00:39:00Z" PublishedUtc="2013-09-14T01:07:47Z" /> </Data> The data is neatly organized under each part. It is easy to see how that document is all you need to know about that content item, all in one table. If you want to modify that data directly in the database, you should be careful to do it in both the record table and the infoset in the content item record. In this configuration, the record is now nothing more than an index, and will only be used for sorting and filtering. Of course, it’s perfectly fine to mix record-backed properties and record-less properties on the same part. It really depends what you think must be sorted and filtered on. In turn, this potentially simplifies migrations considerably. So here it is, the great shift of Orchard to document storage, something that Orchard has been designed for all along, and that we were able to implement with a satisfying and surprising economy of resources. Expect this code to make its way into the 1.8 version of Orchard when that’s available.

    Read the article

  • Slick2d/Nifty-gui input

    - by eerongal
    I'm trying to get input from slick2d into nifty gui. Ive searched online, and I've seen a few examples, but I can't seem to get it working right. i've tried the example on here but I can't seem to get everything working. I'm not entirely sure what I'm doing wrong. I've also looked at examples using the JMonkeyEngine to help point me in the right direction, but still having issues with input. I can get everything else working like i need. Here's the code for my element controller: package gui; import java.util.Properties; import de.lessvoid.nifty.Nifty; import de.lessvoid.nifty.controls.Controller; import de.lessvoid.nifty.elements.Element; import de.lessvoid.nifty.input.NiftyInputEvent; import de.lessvoid.nifty.screen.Screen; import de.lessvoid.xml.xpp3.Attributes; public class BaseElementController implements Controller { private Element element; public void bind(Nifty arg0, Screen arg1, Element arg2, Properties arg3, Attributes arg4) { this.element = element; } public void init(Properties arg0, Attributes arg1) { // TODO Auto-generated method stub } public boolean inputEvent(NiftyInputEvent arg0) { // TODO Auto-generated method stub return false; } public void onFocus(boolean arg0) { // TODO Auto-generated method stub } public void onStartScreen() { // TODO Auto-generated method stub } public void test() { System.out.println("test"); } public void bam() { System.out.println("bam"); } } Here's my XML file: <?xml version="1.0" encoding="UTF-8" standalone="no"?> <nifty> <useStyles filename="nifty-default-styles.xml"/> <useControls filename="nifty-default-controls.xml"/> <screen id="screen2" controller="gui.BaseScreenController"> <layer backgroundColor="#fff0" childLayout="absolute" id="layer4" controller="gui.BaseElementController"> <panel childLayout="center" height="30%" id="panel1" style="nifty-panel-simple" width="50%" x="282" y="334" controller="gui.BaseElementController"> <control id="checkbox1" name="checkbox"/> <control childLayout="center" id="button2" label="button2" name="button" x="381" y="224" visibleToMouse="true" controller="gui.BaseElementController"> <interact onClick="bam()"/> </control> </panel> <text text="${CALL.getPlayerName()}" style="nifty-label" width="100%" height="100%" x="0" y="10" /> </layer> </screen> </nifty> Here's how I'm trying to bind the controller: public void init(GameContainer gc) throws SlickException { Input input = gc.getInput(); inputSystem = new PlainSlickInputSystem(); inputSystem.setInput(input); gui = new Gui(); gui.init(gc, inputSystem, "gui/tset.xml", "screen2"); input.removeListener(this); input.removeListener(inputSystem); input.addListener(inputSystem); } Essentially, all that happens right now is the screen loads up and displays, and it grabs the variable correctly in the label, but none of the input seems to be getting forwarded to Nifty from slick. I assume there's something I'm missing, but I can't seem to figure out what that is. In so far as what I have tried, I attempted to define a custom input listener to pick up events and assign that to my game in order to pick up input, which did not work, so i dropped that implementation, at current i'm trying to take the default inputs and bind then with a PlainSlickInputSystem and assigning that to the input (as shown in the first example link). On code execution, all the code is hit, and i've put several system.out.println's to get ouput of what is happening (the code above has been cleaned for presentation), and i even see the elements getting bound to the controller, yet it doesn't pick up controller events. As far as EXACTLY what's wrong, that I don't know, because I've followed all implementations i can find of this, and none of them seem to do anything it's like the input is just getting thrown out. None of the objects from niftyGui appear to be recognizing any input. Here is the binding from my objects at run time: ******INITIALIZED SCREEN: de.lessvoid.nifty.screen.Screen@4a1ab1c1 ******INITIALIZED ELEMENT: button2 (de.lessvoid.nifty.elements.Element@1e8c1be9) ******INITIALIZED ELEMENT: focusable => true, width => 100px {nifty-button#panel}, backgroundImage => button/button.png {nifty-button#panel}, label => button2, paddingLeft => 7px {nifty-button#panel}, imageMode => sprite-resize:100,23,0,2,96,2,2,2,96,2,19,2,96,2,2 {nifty-button#panel}, paddingRight => 7px {nifty-button#panel}, id => button2, visibleToMouse => true, height => 23px {nifty-button#panel}, style => nifty-button, name => button, inputMapping => de.lessvoid.nifty.input.mapping.MenuInputMapping, childLayout => center, controller => gui.BaseElementController, y => 224, x => 381 ******INITIALIZED SCREEN: de.lessvoid.nifty.screen.Screen@4a1ab1c1 ******INITIALIZED ELEMENT: panel1 (de.lessvoid.nifty.elements.Element@373ec894) ******INITIALIZED ELEMENT: id => panel1, height => 30%, style => nifty-panel-simple, width => 50%, backgroundImage => panel/nifty-panel-simple.png {nifty-panel-simple}, controller => gui.BaseElementController, childLayout => center, padding => 5px {nifty-panel-simple}, imageMode => resize:9,2,9,9,9,2,9,2,9,2,9,9 {nifty-panel-simple}, y => 334, x => 282 ******INITIALIZED SCREEN: de.lessvoid.nifty.screen.Screen@4a1ab1c1 ******INITIALIZED ELEMENT: layer4 (de.lessvoid.nifty.elements.Element@6427d489) ******INITIALIZED ELEMENT: id => layer4, backgroundColor => #fff0, controller => gui.BaseElementController, childLayout => absolute the button2 object is getting bound to my BaseElementController, but i can't seem to get it into the defined "onClick" call.

    Read the article

  • Sliding collision response

    - by dbostream
    I have been reading plenty of tutorials about sliding collision responses yet I am not able to implement it properly in my project. What I want to do is make a puck slide along the rounded corner boards of a hockey rink. In my latest attempt the puck does slide along the boards but there are some strange velocity behaviors. First of all the puck slows down a lot pretty much right away and then it slides for awhile and stops before exiting the corner. Even if I double the speed I get a similar behavior and the puck does not make it out of the corner. I used some ideas from this document http://www.peroxide.dk/papers/collision/collision.pdf. This is what I have: Update method called from the game loop when it is time to update the puck (I removed some irrelevant parts). I use two states (current, previous) which are used to interpolate the position during rendering. public override void Update(double fixedTimeStep) { /* Acceleration is set to 0 for now. */ Acceleration.Zero(); PreviousState = CurrentState; _collisionRecursionDepth = 0; CurrentState.Position = SlidingCollision(CurrentState.Position, CurrentState.Velocity * fixedTimeStep + 0.5 * Acceleration * fixedTimeStep * fixedTimeStep); /* Should not this be affected by a sliding collision? and not only the position. */ CurrentState.Velocity = CurrentState.Velocity + Acceleration * fixedTimeStep; Heading = Vector2.NormalizeRet(CurrentState.Velocity); } private Vector2 SlidingCollision(Vector2 position, Vector2 velocity) { if(_collisionRecursionDepth > 5) return position; bool collisionFound = false; Vector2 futurePosition = position + velocity; Vector2 intersectionPoint = new Vector2(); Vector2 intersectionPointNormal = new Vector2(); /* I did not include the collision detection code, if a collision is detected the intersection point and normal in that point is returned. */ if(!collisionFound) return futurePosition; /* If no collision was detected it is safe to move to the future position. */ /* It is not exactly the intersection point, but slightly before. */ Vector2 newPosition = intersectionPoint; /* oldVelocity is set to the distance from the newPosition(intersection point) to the position it had moved to had it not collided. */ Vector2 oldVelocity = futurePosition - newPosition; /* Project the distance left to move along the intersection normal. */ Vector2 newVelocity = oldVelocity - intersectionPointNormal * oldVelocity.DotProduct(intersectionPointNormal); if(newVelocity.LengthSq() < 0.001) return newPosition; /* If almost no speed, no need to continue. */ _collisionRecursionDepth++; return SlidingCollision(newPosition, newVelocity); } What am I doing wrong with the velocity? I have been staring at this for very long so I have gone blind. I have tried different values of recursion depth but it does not seem to make it better. Let me know if you need more information. I appreciate any help. EDIT: A combination of Patrick Hughes' and teodron's answers solved the velocity problem (I think), thanks a lot! This is the new code: I decided to use a separate recursion method now too since I don't want to recalculate the acceleration in each recursion. public override void Update(double fixedTimeStep) { Acceleration.Zero();// = CalculateAcceleration(fixedTimeStep); PreviousState = new MovingEntityState(CurrentState.Position, CurrentState.Velocity); CurrentState = SlidingCollision(CurrentState, fixedTimeStep); Heading = Vector2.NormalizeRet(CurrentState.Velocity); } private MovingEntityState SlidingCollision(MovingEntityState state, double timeStep) { bool collisionFound = false; /* Calculate the next position given no detected collision. */ Vector2 futurePosition = state.Position + state.Velocity * timeStep; Vector2 intersectionPoint = new Vector2(); Vector2 intersectionPointNormal = new Vector2(); /* I did not include the collision detection code, if a collision is detected the intersection point and normal in that point is returned. */ /* If no collision was detected it is safe to move to the future position. */ if (!collisionFound) return new MovingEntityState(futurePosition, state.Velocity); /* Set new position to the intersection point (slightly before). */ Vector2 newPosition = intersectionPoint; /* Project the new velocity along the intersection normal. */ Vector2 newVelocity = state.Velocity - 1.90 * intersectionPointNormal * state.Velocity.DotProduct(intersectionPointNormal); /* Calculate the time of collision. */ double timeOfCollision = Math.Sqrt((newPosition - state.Position).LengthSq() / (futurePosition - state.Position).LengthSq()); /* Calculate new time step, remaining time of full step after the collision * current time step. */ double newTimeStep = timeStep * (1 - timeOfCollision); return SlidingCollision(new MovingEntityState(newPosition, newVelocity), newTimeStep); } Even though the code above seems to slide the puck correctly please have a look at it. I have a few questions, if I don't multiply by 1.90 in the newVelocity calculation it doesn't work (I get a stack overflow when the puck enters the corner because the timeStep decreases very slowly - a collision is found early in every recursion), why is that? what does 1.90 really do and why 1.90? Also I have a new problem, the puck does not move parallell to the short side after exiting the curve; to be more exact it moves outside the rink (I am not checking for any collisions with the short side at the moment). When I perform the collision detection I first check that the puck is in the correct quadrant. For example bottom-right corner is quadrant four i.e. circleCenter.X < puck.X && circleCenter.Y puck.Y is this a problem? or should the short side of the rink be the one to make the puck go parallell to it and not the last collision in the corner? EDIT2: This is the code I use for collision detection, maybe it has something to do with the fact that I can't make the puck slide (-1.0) but only reflect (-2.0): /* Point is the current position (not the predicted one) and quadrant is 4 for the bottom-right corner for example. */ if (GeometryHelper.PointInCircleQuadrant(circleCenter, circleRadius, state.Position, quadrant)) { /* The line is: from = state.Position, to = futurePosition. So a collision is detected when from is inside the circle and to is outside. */ if (GeometryHelper.LineCircleIntersection2d(state.Position, futurePosition, circleCenter, circleRadius, intersectionPoint, quadrant)) { collisionFound = true; /* Set the intersection point to slightly before the real intersection point (I read somewhere this was good to do because of floting point precision, not sure exactly how much though). */ intersectionPoint = intersectionPoint - Vector2.NormalizeRet(state.Velocity) * 0.001; /* Normal at the intersection point. */ intersectionPointNormal = Vector2.NormalizeRet(circleCenter - intersectionPoint) } } When I set the intersection point, if I for example use 0.1 instead of 0.001 the puck travels further before it gets stuck, but for all values I have tried (including 0 - the real intersection point) it gets stuck somewhere (but I necessarily not get a stack overflow). Can something in this part be the cause of my problem? I can see why I get the stack overflow when using -1.0 when calculating the new velocity vector; but not how to solve it. I traced the time steps used in the recursion (initial time step is always 1/60 ~ 0.01666): Recursion depth Time step next recursive call [Start recursion, time step ~ 0.016666] 0 0,000985806527246773 [No collision, stop recursion] [Start recursion, time step ~ 0.016666] 0 0,0149596704364629 1 0,0144883449376379 2 0,0143155612984837 3 0,014224925727213 4 0,0141673917461608 5 0,0141265435314026 6 0,0140953966184117 7 0,0140704653746625 ...and so on. As you can see the collision is detected early in every recursive call which means the next time step decreases very slowly thus the recursion depth gets very big - stack overflow.

    Read the article

  • Cost Comparison Hard Disk Drive to Solid State Drive on Price per Gigabyte - dispelling a myth!

    - by tonyrogerson
    It is often said that Hard Disk Drive storage is significantly cheaper per GiByte than Solid State Devices – this is wholly inaccurate within the database space. People need to look at the cost of the complete solution and not just a single component part in isolation to what is really required to meet the business requirement. Buying a single Hitachi Ultrastar 600GB 3.5” SAS 15Krpm hard disk drive will cost approximately £239.60 (http://scan.co.uk, 22nd March 2012) compared to an OCZ 600GB Z-Drive R4 CM84 PCIe costing £2,316.54 (http://scan.co.uk, 22nd March 2012); I’ve not included FusionIO ioDrive because there is no public pricing available for it – something I never understand and personally when companies do this I immediately think what are they hiding, luckily in FusionIO’s case the product is proven though is expensive compared to OCZ enterprise offerings. On the face of it the single 15Krpm hard disk has a price per GB of £0.39, the SSD £3.86; this is what you will see in the press and this is what sales people will use in comparing the two technologies – do not be fooled by this bullshit people! What is the requirement? The requirement is the database will have a static size of 400GB kept static through archiving so growth and trim will balance the database size, the client requires resilience, there will be several hundred call centre staff querying the database where queries will read a small amount of data but there will be no hot spot in the data so the randomness will come across the entire 400GB of the database, estimates predict that the IOps required will be approximately 4,000IOps at peak times, because it’s a call centre system the IO latency is important and must remain below 5ms per IO. The balance between read and write is 70% read, 30% write. The requirement is now defined and we have three of the most important pieces of the puzzle – space required, estimated IOps and maximum latency per IO. Something to consider with regard SQL Server; write activity requires synchronous IO to the storage media specifically the transaction log; that means the write thread will wait until the IO is completed and hardened off until the thread can continue execution, the requirement has stated that 30% of the system activity will be write so we can expect a high amount of synchronous activity. The hardware solution needs to be defined; two possible solutions: hard disk or solid state based; the real question now is how many hard disks are required to achieve the IO throughput, the latency and resilience, ditto for the solid state. Hard Drive solution On a test on an HP DL380, P410i controller using IOMeter against a single 15Krpm 146GB SAS drive, the throughput given on a transfer size of 8KiB against a 40GiB file on a freshly formatted disk where the partition is the only partition on the disk thus the 40GiB file is on the outer edge of the drive so more sectors can be read before head movement is required: For 100% sequential IO at a queue depth of 16 with 8 worker threads 43,537 IOps at an average latency of 2.93ms (340 MiB/s), for 100% random IO at the same queue depth and worker threads 3,733 IOps at an average latency of 34.06ms (34 MiB/s). The same test was done on the same disk but the test file was 130GiB: For 100% sequential IO at a queue depth of 16 with 8 worker threads 43,537 IOps at an average latency of 2.93ms (340 MiB/s), for 100% random IO at the same queue depth and worker threads 528 IOps at an average latency of 217.49ms (4 MiB/s). From the result it is clear random performance gets worse as the disk fills up – I’m currently writing an article on short stroking which will cover this in detail. Given the work load is random in nature looking at the random performance of the single drive when only 40 GiB of the 146 GB is used gives near the IOps required but the latency is way out. Luckily I have tested 6 x 15Krpm 146GB SAS 15Krpm drives in a RAID 0 using the same test methodology, for the same test above on a 130 GiB for each drive added the performance boost is near linear, for each drive added throughput goes up by 5 MiB/sec, IOps by 700 IOps and latency reducing nearly 50% per drive added (172 ms, 94 ms, 65 ms, 47 ms, 37 ms, 30 ms). This is because the same 130GiB is spread out more as you add drives 130 / 1, 130 / 2, 130 / 3 etc. so implicit short stroking is occurring because there is less file on each drive so less head movement required. The best latency is still 30 ms but we have the IOps required now, but that’s on a 130GiB file and not the 400GiB we need. Some reality check here: a) the drive randomness is more likely to be 50/50 and not a full 100% but the above has highlighted the effect randomness has on the drive and the more a drive fills with data the worse the effect. For argument sake let us assume that for the given workload we need 8 disks to do the job, for resilience reasons we will need 16 because we need to RAID 1+0 them in order to get the throughput and the resilience, RAID 5 would degrade performance. Cost for hard drives: 16 x £239.60 = £3,833.60 For the hard drives we will need disk controllers and a separate external disk array because the likelihood is that the server itself won’t take the drives, a quick spec off DELL for a PowerVault MD1220 which gives the dual pathing with 16 disks 146GB 15Krpm 2.5” disks is priced at £7,438.00, note its probably more once we had two controller cards to sit in the server in, racking etc. Minimum cost taking the DELL quote as an example is therefore: {Cost of Hardware} / {Storage Required} £7,438.60 / 400 = £18.595 per GB £18.59 per GiB is a far cry from the £0.39 we had been told by the salesman and the myth. Yes, the storage array is composed of 16 x 146 disks in RAID 10 (therefore 8 usable) giving an effective usable storage availability of 1168GB but the actual storage requirement is only 400 and the extra disks have had to be purchased to get the  IOps up. Solid State Drive solution A single card significantly exceeds the IOps and latency required, for resilience two will be required. ( £2,316.54 * 2 ) / 400 = £11.58 per GB With the SSD solution only two PCIe sockets are required, no external disk units, no additional controllers, no redundant controllers etc. Conclusion I hope by showing you an example that the myth that hard disk drives are cheaper per GiB than Solid State has now been dispelled - £11.58 per GB for SSD compared to £18.59 for Hard Disk. I’ve not even touched on the running costs, compare the costs of running 18 hard disks, that’s a lot of heat and power compared to two PCIe cards!Just a quick note: I've left a fair amount of information out due to this being a blog! If in doubt, email me :)I'll also deal with the myth that SSD's wear out at a later date as well - that's just way over done still, yes, 5 years ago, but now - no.

    Read the article

  • Nashorn, the rhino in the room

    - by costlow
    Nashorn is a new runtime within JDK 8 that allows developers to run code written in JavaScript and call back and forth with Java. One advantage to the Nashorn scripting engine is that is allows for quick prototyping of functionality or basic shell scripts that use Java libraries. The previous JavaScript runtime, named Rhino, was introduced in JDK 6 (released 2006, end of public updates Feb 2013). Keeping tradition amongst the global developer community, "Nashorn" is the German word for rhino. The Java platform and runtime is an intentional home to many languages beyond the Java language itself. OpenJDK’s Da Vinci Machine helps coordinate work amongst language developers and tool designers and has helped different languages by introducing the Invoke Dynamic instruction in Java 7 (2011), which resulted in two major benefits: speeding up execution of dynamic code, and providing the groundwork for Java 8’s lambda executions. Many of these improvements are discussed at the JVM Language Summit, where language and tool designers get together to discuss experiences and issues related to building these complex components. There are a number of benefits to running JavaScript applications on JDK 8’s Nashorn technology beyond writing scripts quickly: Interoperability with Java and JavaScript libraries. Scripts do not need to be compiled. Fast execution and multi-threading of JavaScript running in Java’s JRE. The ability to remotely debug applications using an IDE like NetBeans, Eclipse, or IntelliJ (instructions on the Nashorn blog). Automatic integration with Java monitoring tools, such as performance, health, and SIEM. In the remainder of this blog post, I will explain how to use Nashorn and the benefit from those features. Nashorn execution environment The Nashorn scripting engine is included in all versions of Java SE 8, both the JDK and the JRE. Unlike Java code, scripts written in nashorn are interpreted and do not need to be compiled before execution. Developers and users can access it in two ways: Users running JavaScript applications can call the binary directly:jre8/bin/jjs This mechanism can also be used in shell scripts by specifying a shebang like #!/usr/bin/jjs Developers can use the API and obtain a ScriptEngine through:ScriptEngine engine = new ScriptEngineManager().getEngineByName("nashorn"); When using a ScriptEngine, please understand that they execute code. Avoid running untrusted scripts or passing in untrusted/unvalidated inputs. During compilation, consider isolating access to the ScriptEngine and using Type Annotations to only allow @Untainted String arguments. One noteworthy difference between JavaScript executed in or outside of a web browser is that certain objects will not be available. For example when run outside a browser, there is no access to a document object or DOM tree. Other than that, all syntax, semantics, and capabilities are present. Examples of Java and JavaScript The Nashorn script engine allows developers of all experience levels the ability to write and run code that takes advantage of both languages. The specific dialect is ECMAScript 5.1 as identified by the User Guide and its standards definition through ECMA international. In addition to the example below, Benjamin Winterberg has a very well written Java 8 Nashorn Tutorial that provides a large number of code samples in both languages. Basic Operations A basic Hello World application written to run on Nashorn would look like this: #!/usr/bin/jjs print("Hello World"); The first line is a standard script indication, so that Linux or Unix systems can run the script through Nashorn. On Windows where scripts are not as common, you would run the script like: jjs helloWorld.js. Receiving Arguments In order to receive program arguments your jjs invocation needs to use the -scripting flag and a double-dash to separate which arguments are for jjs and which are for the script itself:jjs -scripting print.js -- "This will print" #!/usr/bin/jjs var whatYouSaid = $ARG.length==0 ? "You did not say anything" : $ARG[0] print(whatYouSaid); Interoperability with Java libraries (including 3rd party dependencies) Another goal of Nashorn was to allow for quick scriptable prototypes, allowing access into Java types and any libraries. Resources operate in the context of the script (either in-line with the script or as separate threads) so if you open network sockets and your script terminates, those sockets will be released and available for your next run. Your code can access Java types the same as regular Java classes. The “import statements” are written somewhat differently to accommodate for language. There is a choice of two styles: For standard classes, just name the class: var ServerSocket = java.net.ServerSocket For arrays or other items, use Java.type: var ByteArray = Java.type("byte[]")You could technically do this for all. The same technique will allow your script to use Java types from any library or 3rd party component and quickly prototype items. Building a user interface One major difference between JavaScript inside and outside of a web browser is the availability of a DOM object for rendering views. When run outside of the browser, JavaScript has full control to construct the entire user interface with pre-fabricated UI controls, charts, or components. The example below is a variation from the Nashorn and JavaFX guide to show how items work together. Nashorn has a -fx flag to make the user interface components available. With the example script below, just specify: jjs -fx -scripting fx.js -- "My title" #!/usr/bin/jjs -fx var Button = javafx.scene.control.Button; var StackPane = javafx.scene.layout.StackPane; var Scene = javafx.scene.Scene; var clickCounter=0; $STAGE.title = $ARG.length>0 ? $ARG[0] : "You didn't provide a title"; var button = new Button(); button.text = "Say 'Hello World'"; button.onAction = myFunctionForButtonClicking; var root = new StackPane(); root.children.add(button); $STAGE.scene = new Scene(root, 300, 250); $STAGE.show(); function myFunctionForButtonClicking(){   var text = "Click Counter: " + clickCounter;   button.setText(text);   clickCounter++;   print(text); } For a more advanced post on using Nashorn to build a high-performing UI, see JavaFX with Nashorn Canvas example. Interoperable with frameworks like Node, Backbone, or Facebook React The major benefit of any language is the interoperability gained by people and systems that can read, write, and use it for interactions. Because Nashorn is built for the ECMAScript specification, developers familiar with JavaScript frameworks can write their code and then have system administrators deploy and monitor the applications the same as any other Java application. A number of projects are also running Node applications on Nashorn through Project Avatar and the supported modules. In addition to the previously mentioned Nashorn tutorial, Benjamin has also written a post about Using Backbone.js with Nashorn. To show the multi-language power of the Java Runtime, there is another interesting example that unites Facebook React and Clojure on JDK 8’s Nashorn. Summary Nashorn provides a simple and fast way of executing JavaScript applications and bridging between the best of each language. By making the full range of Java libraries to JavaScript applications, and the quick prototyping style of JavaScript to Java applications, developers are free to work as they see fit. Software Architects and System Administrators can take advantage of one runtime and leverage any work that they have done to tune, monitor, and certify their systems. Additional information is available within: The Nashorn Users’ Guide Java Magazine’s article "Next Generation JavaScript Engine for the JVM." The Nashorn team’s primary blog or a very helpful collection of Nashorn links.

    Read the article

  • tile_static, tile_barrier, and tiled matrix multiplication with C++ AMP

    - by Daniel Moth
    We ended the previous post with a mechanical transformation of the C++ AMP matrix multiplication example to the tiled model and in the process introduced tiled_index and tiled_grid. This is part 2. tile_static memory You all know that in regular CPU code, static variables have the same value regardless of which thread accesses the static variable. This is in contrast with non-static local variables, where each thread has its own copy. Back to C++ AMP, the same rules apply and each thread has its own value for local variables in your lambda, whereas all threads see the same global memory, which is the data they have access to via the array and array_view. In addition, on an accelerator like the GPU, there is a programmable cache, a third kind of memory type if you'd like to think of it that way (some call it shared memory, others call it scratchpad memory). Variables stored in that memory share the same value for every thread in the same tile. So, when you use the tiled model, you can have variables where each thread in the same tile sees the same value for that variable, that threads from other tiles do not. The new storage class for local variables introduced for this purpose is called tile_static. You can only use tile_static in restrict(direct3d) functions, and only when explicitly using the tiled model. What this looks like in code should be no surprise, but here is a snippet to confirm your mental image, using a good old regular C array // each tile of threads has its own copy of locA, // shared among the threads of the tile tile_static float locA[16][16]; Note that tile_static variables are scoped and have the lifetime of the tile, and they cannot have constructors or destructors. tile_barrier In amp.h one of the types introduced is tile_barrier. You cannot construct this object yourself (although if you had one, you could use a copy constructor to create another one). So how do you get one of these? You get it, from a tiled_index object. Beyond the 4 properties returning index objects, tiled_index has another property, barrier, that returns a tile_barrier object. The tile_barrier class exposes a single member, the method wait. 15: // Given a tiled_index object named t_idx 16: t_idx.barrier.wait(); 17: // more code …in the code above, all threads in the tile will reach line 16 before a single one progresses to line 17. Note that all threads must be able to reach the barrier, i.e. if you had branchy code in such a way which meant that there is a chance that not all threads could reach line 16, then the code above would be illegal. Tiled Matrix Multiplication Example – part 2 So now that we added to our understanding the concepts of tile_static and tile_barrier, let me obfuscate rewrite the matrix multiplication code so that it takes advantage of tiling. Before you start reading this, I suggest you get a cup of your favorite non-alcoholic beverage to enjoy while you try to fully understand the code. 01: void MatrixMultiplyTiled(vector<float>& vC, const vector<float>& vA, const vector<float>& vB, int M, int N, int W) 02: { 03: static const int TS = 16; 04: array_view<const float,2> a(M, W, vA); 05: array_view<const float,2> b(W, N, vB); 06: array_view<writeonly<float>,2> c(M,N,vC); 07: parallel_for_each(c.grid.tile< TS, TS >(), 08: [=] (tiled_index< TS, TS> t_idx) restrict(direct3d) 09: { 10: int row = t_idx.local[0]; int col = t_idx.local[1]; 11: float sum = 0.0f; 12: for (int i = 0; i < W; i += TS) { 13: tile_static float locA[TS][TS], locB[TS][TS]; 14: locA[row][col] = a(t_idx.global[0], col + i); 15: locB[row][col] = b(row + i, t_idx.global[1]); 16: t_idx.barrier.wait(); 17: for (int k = 0; k < TS; k++) 18: sum += locA[row][k] * locB[k][col]; 19: t_idx.barrier.wait(); 20: } 21: c[t_idx.global] = sum; 22: }); 23: } Notice that all the code up to line 9 is the same as per the changes we made in part 1 of tiling introduction. If you squint, the body of the lambda itself preserves the original algorithm on lines 10, 11, and 17, 18, and 21. The difference being that those lines use new indexing and the tile_static arrays; the tile_static arrays are declared and initialized on the brand new lines 13-15. On those lines we copy from the global memory represented by the array_view objects (a and b), to the tile_static vanilla arrays (locA and locB) – we are copying enough to fit a tile. Because in the code that follows on line 18 we expect the data for this tile to be in the tile_static storage, we need to synchronize the threads within each tile with a barrier, which we do on line 16 (to avoid accessing uninitialized memory on line 18). We also need to synchronize the threads within a tile on line 19, again to avoid the race between lines 14, 15 (retrieving the next set of data for each tile and overwriting the previous set) and line 18 (not being done processing the previous set of data). Luckily, as part of the awesome C++ AMP debugger in Visual Studio there is an option that helps you find such races, but that is a story for another blog post another time. May I suggest reading the next section, and then coming back to re-read and walk through this code with pen and paper to really grok what is going on, if you haven't already? Cool. Why would I introduce this tiling complexity into my code? Funny you should ask that, I was just about to tell you. There is only one reason we tiled our extent, had to deal with finding a good tile size, ensure the number of threads we schedule are correctly divisible with the tile size, had to use a tiled_index instead of a normal index, and had to understand tile_barrier and to figure out where we need to use it, and double the size of our lambda in terms of lines of code: the reason is to be able to use tile_static memory. Why do we want to use tile_static memory? Because accessing tile_static memory is around 10 times faster than accessing the global memory on an accelerator like the GPU, e.g. in the code above, if you can get 150GB/second accessing data from the array_view a, you can get 1500GB/second accessing the tile_static array locA. And since by definition you are dealing with really large data sets, the savings really pay off. We have seen tiled implementations being twice as fast as their non-tiled counterparts. Now, some algorithms will not have performance benefits from tiling (and in fact may deteriorate), e.g. algorithms that require you to go only once to global memory will not benefit from tiling, since with tiling you already have to fetch the data once from global memory! Other algorithms may benefit, but you may decide that you are happy with your code being 150 times faster than the serial-version you had, and you do not need to invest to make it 250 times faster. Also algorithms with more than 3 dimensions, which C++ AMP supports in the non-tiled model, cannot be tiled. Also note that in future releases, we may invest in making the non-tiled model, which already uses tiling under the covers, go the extra step and use tile_static memory on your behalf, but it is obviously way to early to commit to anything like that, and we certainly don't do any of that today. Comments about this post by Daniel Moth welcome at the original blog.

    Read the article

  • Grow Your Business with Security

    - by Darin Pendergraft
    Author: Kevin Moulton Kevin Moulton has been in the security space for more than 25 years, and with Oracle for 7 years. He manages the East EnterpriseSecurity Sales Consulting Team. He is also a Distinguished Toastmaster. Follow Kevin on Twitter at twitter.com/kevin_moulton, where he sometimes tweets about security, but might also tweet about running, beer, food, baseball, football, good books, or whatever else grabs his attention. Kevin will be a regular contributor to this blog so stay tuned for more posts from him. It happened again! There I was, reading something interesting online, and realizing that a friend might find it interesting too. I clicked on the little email link, thinking that I could easily forward this to my friend, but no! Instead, a new screen popped up where I was asked to create an account. I was expected to create a User ID and password, not to mention providing some personally identifiable information, just for the privilege of helping that website spread their word. Of course, I didn’t want to have to remember a new account and password, I didn’t want to provide the requisite information, and I didn’t want to waste my time. I gave up, closed the web page, and moved on to something else. I was left with a bad taste in my mouth, and my friend might never find her way to this interesting website. If you were this content provider, would this be the outcome you were looking for? A few days later, I had a similar experience, but this one went a little differently. I was surfing the web, when I happened upon some little chotcke that I just had to have. I added it to my cart. When I went to buy the item, I was again brought to a page to create account. Groan! But wait! On this page, I also had the option to sign in with my OpenID account, my Facebook account, my Yahoo account, or my Google Account. I have all of those! No new account to create, no new password to remember, and no personally identifiable information to be given to someone else (I’ve already given it all to those other guys, after all). In this case, the vendor was easy to deal with, and I happily completed the transaction. That pleasant experience will bring me back again. This is where security can grow your business. It’s a differentiator. You’ve got to have a presence on the web, and that presence has to take into account all the smart phones everyone’s carrying, and the tablets that took over cyber Monday this year. If you are a company that a customer can deal with securely, and do so easily, then you are a company customers will come back to again and again. I recently had a need to open a new bank account. Every bank has a web presence now, but they are certainly not all the same. I wanted one that I could deal with easily using my laptop, but I also wanted 2-factor authentication in case I had to login from a shared machine, and I wanted an app for my iPad. I found a bank with all three, and that’s who I am doing business with. Let’s say, for example, that I’m in a regular Texas Hold-em game on Friday nights, so I move a couple of hundred bucks from checking to savings on Friday afternoons. I move a similar amount each week and I do it from the same machine. The bank trusts me, and they trust my machine. Most importantly, they trust my behavior. This is adaptive authentication. There should be no reason for my bank to make this transaction difficult for me. Now let's say that I login from a Starbucks in Uzbekistan, and I transfer $2,500. What should my bank do now? Should they stop the transaction? Should they call my home number? (My former bank did exactly this once when I was taking money out of an ATM on a business trip, when I had provided my cell phone number as my primary contact. When I asked them why they called my home number rather than my cell, they told me that their “policy” is to call the home number. If I'm on the road, what exactly is the use of trying to reach me at home to verify my transaction?) But, back to Uzbekistan… Should my bank assume that I am happily at home in New Jersey, and someone is trying to hack into my account? Perhaps they think they are protecting me, but I wouldn’t be very happy if I happened to be traveling on business in Central Asia. What if my bank were to automatically analyze my behavior and calculate a risk score? Clearly, this scenario would be outside of my typical behavior, so my risk score would necessitate something more than a simple login and password. Perhaps, in this case, a one-time password to my cell phone would prove that this is not just some hacker half way around the world. But, what if you're not a bank? Do you need this level of security? If you want to be a business that is easy to deal with while also protecting your customers, then of course you do. You want your customers to trust you, but you also want them to enjoy doing business with you. Make it easy for them to do business with you, and they’ll come back, and perhaps even Tweet about it, or Like you, and then their friends will follow. How can Oracle help? Oracle has the technology and expertise to help you to grown your business with security. Oracle Adaptive Access Manager will help you to prevent fraud while making it easier for your customers to do business with you by providing the risk analysis I discussed above, step-up authentication, and much more. Oracle Mobile and Social Access Service will help you to secure mobile access to applications by expanding on your existing back-end identity management infrastructure, and allowing your customers to transact business with you using the social media accounts they already know. You also have device fingerprinting and metrics to help you to grow your business securely. Security is not just a cost anymore. It’s a way to set your business apart. With Oracle’s help, you can be the business that everyone’s tweeting about. Image courtesy of Flickr user shareski

    Read the article

  • Big Visible Charts

    - by Robert May
    An important part of Agile is the concept of transparency and visibility. In proper functioning teams, stakeholders can look at any team at any time in the iteration or release and see how that team is doing by simply looking at what we call Big Visible Charts. If you’ve done Scrum, you’ve seen these charts. However, interpreting these charts can often be an art form. There are several different charts that can be useful. In this newsletter, I’ll focus on the Iteration Burndown and Cumulative Flow charts. I’ve included a copy of the spreadsheet that I used to create the charts, and if you don’t have a tool that creates them for you, you can use this spreadsheet to do so. Our preferred tool for managing Scrum projects is Rally. Rally creates all of these charts for you, saving you quite a bit of time. The Iteration Burndown and Cumulative Flow Charts This is the main chart that teams use. Although less useful to stakeholders, this chart is critical to the team and provides quite a bit of information to the team about how their iteration is going. Most charts are a combination of the charts below, so you may need to combine aspects of each section to understand what is happening in your iterations. Ideal Ah, isn’t that a pretty picture? Unfortunately, it’s also very unrealistic. I’ve seen iterations that come close to ideal, but never that match perfectly. If your iteration matches perfectly, chances are, someone is playing with the numbers. Reality is just too difficult to have a burndown chart that matches this exactly. Late Planning Iteration started, but the team didn’t. You can tell this by the fact that the real number of estimated hours didn’t appear until day two. In the cumulative flow, you can also see that nothing was defined in Day one and two. You want to avoid situations like this. You’ll note that the team had to burn faster than is ideal to meet the iteration because of the late planning. This often results in long weeks and days. Testing Starved Determining whether or not testing is starved is difficult without the cumulative flow. The pattern in the burndown could be nothing more that developers not completing stories early enough or could be caused by stories being too big. With the cumulative flow, however, you see that only small bites are in progress and stories were completed early, but testing didn’t start testing until the end of the iteration, and didn’t complete testing all stories in the iteration. When this happens, question whether or not your testing resources are sufficient for your team and whether or not acceptance is adequately defined. No Testing With this one, both graphs show the same thing; the team needs testers and testing! Without testing, what was completed cannot be verified to make sure that it is acceptable to the business. If you find yourself in this situation, review your testing practices and acceptance testing process and make changes today. Late Development With this situation, both graphs tell a story. In the top graph, you can see that the hours failed to burn down as quickly as the team expected. This could be caused by the team not correctly estimating their hours or the team could have had illness or some other issue that affected them. Often, when teams are tackling something that is more unknown, they’ll run into technical barriers that cause the burn down to happen slower than expected. In the cumulative flow graph, you can see that not much was completed in the first few days. This could be because of illness or technical barriers or simply poor estimation. Testing was able to keep up with everything that was completed, however. No Tool Updating When you see graphs that look like this, you can be assured that it’s because the team is not updating the tool that generates the graphs. Review your policy for when they are to update. On the teams that I run, I require that each team member updates the tool at least once daily. You should also check to see how well the team is breaking down stories into tasks. If they’re creating few large tasks, graphs can look similar to this. As a general rule, I never allow tasks, other than Unit Testing and Uncertainty, to be greater than eight hours in duration. Scope Increase I always encourage team members to enter in however much time they think they have left on a task, even if that means increasing the total amount of time left to do. You get a much better and more realistic picture this way. Increasing time remaining could explain the burndown graph, but by looking at the cumulative flow graph, we can see that stories were added to the iteration and scope was increased. Since planning should consume all of the hours in the iteration, this is almost always a bad thing. If the scope change happened late in the iteration and the hours remaining were well below the ideal burn, then increasing scope is probably o.k., but estimation needs to get better. However, with the charts above, that’s clearly not what happened and the team was required to do extra work to make the iteration. If you find this happening, your product owner and ScrumMasters need training. The team also needs to learn to say no. Scope Decrease Scope decreases are just as bad as scope increases. Usually, graphs above show that the team did a poor job of estimating their stories and part way through had to reduce scope to change the iteration. This will happen once in a while, but if you find it’s a pattern on your team, you need to re-evaluate planning. Some teams are hopelessly optimistic. In those cases, I’ll introduce a task I call “Uncertainty.” With Uncertainty, the team estimates how many hours they might need if things don’t go well with the tasks they’ve defined. They try to estimate things that could go poorly and increase the time appropriately. Having an Uncertainty task allows them to have a low and high estimate. Uncertainty should not just be an arbitrary buffer. It must correlate to real uncertainty in the tasks that have been defined. Stories are too Big Often, we see graphs like the ones above. Note that the burndown looks fairly good, other than the chunky acceptance of stories. However, when you look at cumulative flow, you can see that at one point, everything is in progress. This is a bad thing. When you see graphs like this, you’re in one of two states. You may just have a very small team and can only handle one or two stories in your iteration. If you have more than one or two people, then the most likely problem is that your stories are far too big. To combat this, break large high hour stories into smaller pieces that can be completed independently and accepted independently. If you don’t, you’ll likely be requiring your testers to do heroic things to complete testing on the last day of the iteration and you’re much more likely to have the entire iteration fail, because of the limited amount of things that can be completed. Summary There are other charts that can be useful when doing scrum. If you don’t have any big visible charts, you really need to evaluate your process and change. These charts can provide the team a wealth of information and help you write better software. If you have any questions about charts that you’re seeing on your team, contact me with a screen capture of the charts and I’ll tell you what I’m seeing in those charts. I always want this information to be useful, so please let me know if you have other questions. Technorati Tags: Agile

    Read the article

  • BRE (Business Rules Engine) Data Services is out...!!!

    - by Vishal
    A few months ago we at Tellago had open sourced the BizTalk Data Services. We were meanwhile working on other artifacts which comes along with BizTalk Server like the “Business Rules Engine”.  We are happy to announce the first version of BRE Data Services. BRE Data Services is a same concept which we covered through BTS Data Services, providing a RESTFul OData – based API to interact with the Business Rules Engine via HTTP using ATOM Publishing Protocol or JSON as the encoding mechanism.   In the first version release, we mainly focused on the browsing, querying and searching BRE artifacts via a RESTFul interface. Also along with that we provide the functionality to execute Business Rules by inserting the Facts for policies via the IUpdatable implementation of WCF Data Services.   The BRE Data Services API provides a lightweight interface for managing Business Rules Engine artifacts such as Policies, Rules, Vocabularies, Conditions, Actions, Facts etc. The following are some examples which details some of the available features in the current version of the API.   Basic Querying: Querying BRE Policies http://localhost/BREDataServices/BREMananagementService.svc/Policies Querying BRE Rules http://localhost/BREDataServices/BREMananagementService.svc/Rules Querying BRE Vocabularies http://localhost/BREDataServices/BREMananagementService.svc/Vocabularies   Navigation: The BRE Data Services API also leverages WCF Data Services to enable navigation across related different BRE objects. Querying a specific Policy http://localhost/BREDataServices/BREMananagementService.svc/Policies(‘PolicyName’) Querying a specific Rule http://localhost/BREDataServices/BREMananagementService.svc/Rules(‘RuleName’) Querying all Rules under a Policy http://localhost/BREDataServices/BREMananagementService.svc/Policies('PolicyName')/Rules Querying all Facts under a Policy http://localhost/BREDataServices/BREMananagementService.svc/Policies('PolicyName')/Facts Querying all Actions for a specific Rule http://localhost/BREDataServices/BREMananagementService.svc/Rules('RuleName')/Actions Querying all Conditions for a specific Rule http://localhost/BREDataServices/BREMananagementService.svc/Rules('RuleName')/Actions Querying a specific Vocabulary: http://localhost/BREDataServices/BREMananagementService.svc/Vocabularies('VocabName')   Implementation: With the BRE Data Services, we also provide the functionality of executing a particular policy via HTTP. There are couple of ways you can do that though the API.   Ø First is though Service Operations feature of WCF Data Services in which you can execute the Facts by passing them in the URL itself. This is a very simple implementations of the executing the policies due to the limitations & restrictions (only primitive types of input parameters which can be passed) currently of the Service Operations of the WCF Data Services. Below is a code sample.                Below is a traced Request/Response message.                                 Ø Second is through the IUpdatable Interface of WCF Data Services. In this method, you can first query the rule which you want to execute and then inserts Facts for that particular Rules and finally when you perform the SaveChanges() call for the IUpdatable Interface API, it executes the policy with the facts which you inserted at runtime. Below is a sample of client side code. Due to the limitations of current version of WCF Data Services where there is no way you can return back the updates happening on the service side back to the client via the SaveChanges() method. Here we are executing the rule passing a serialized XML as Facts and there is no changes made to any data where we can query back to fetch the changes. This is overcome though the first way to executing the policies which is by executing it as a Service Operation call.     This actually generates a AtomPub message shown as below:   POST /Tellago.BRE.REST.ServiceHost/BREMananagementService.svc/$batch HTTP/1.1 User-Agent: Microsoft ADO.NET Data Services DataServiceVersion: 1.0;NetFx MaxDataServiceVersion: 2.0;NetFx Accept: application/atom+xml,application/xml Accept-Charset: UTF-8 Content-Type: multipart/mixed; boundary=batch_6b9a5ced-5ecb-4585-940a-9d5e704c28c7 Host: localhost:8080 Content-Length: 1481 Expect: 100-continue   --batch_6b9a5ced-5ecb-4585-940a-9d5e704c28c7 Content-Type: multipart/mixed; boundary=changeset_184a8c59-a714-4ba9-bb3d-889a88fe24bf   --changeset_184a8c59-a714-4ba9-bb3d-889a88fe24bf Content-Type: application/http Content-Transfer-Encoding: binary   MERGE http://localhost:8080/Tellago.BRE.REST.ServiceHost/BREMananagementService.svc/Facts('TestPolicy') HTTP/1.1 Content-ID: 4 Content-Type: application/atom+xml;type=entry Content-Length: 927   <?xml version="1.0" encoding="utf-8" standalone="yes"?> <entry xmlns:d="http://schemas.microsoft.com/ado/2007/08/dataservices" xmlns:m="http://schemas.microsoft.com/ado/2007/08/dataservices/metadata" font-size: x-small"http://www.w3.org/2005/Atom">   <category scheme="http://schemas.microsoft.com/ado/2007/08/dataservices/scheme" term="Tellago.BRE.REST.Resources.Fact" />   <title />   <author>     <name />   </author>   <updated>2011-01-31T20:09:15.0023982Z</updated>   <id>http://localhost:8080/Tellago.BRE.REST.ServiceHost/BREMananagementService.svc/Facts('TestPolicy')</id>   <content type="application/xml">     <m:properties>       <d:FactInstance>&lt;ns0:LoanStatus xmlns:ns0="http://tellago.com"&gt;&lt;Age&gt;10&lt;/Age&gt;&lt;Status&gt;true&lt;/Status&gt;&lt;/ns0:LoanStatus&gt;</d:FactInstance>       <d:FactType>TestSchema</d:FactType>       <d:ID>TestPolicy</d:ID>     </m:properties>   </content> </entry> --changeset_184a8c59-a714-4ba9-bb3d-889a88fe24bf-- --batch_6b9a5ced-5ecb-4585-940a-9d5e704c28c7—     Installation: The installation of the BRE Data Services is pretty straight forward. ·         Create a new IIS website say BREDataServices. ·         Download the SourceCode from TellagoCodeplex and copy the content from Tellago.BRE.REST.ServiceHost to the physical location of the above created website.     ·         The appPool account running the website should have admin access to the BizTalkRuleEngineDb database. ·         TheRight click the BREManagementService.svc in the IIS ContentView for the website and wala..     Conclusion: The BRE Data Services API is an experiment intended to bring the capabilities of RESTful/OData based services to the Traditional BTS/BRE Solutions. The future releases will target on technologies like BAM, ESB Toolkit. This version has been tested with various version of BizTalk Server and we have uploaded the source code to our Tellago's DevLabs workspace at Codeplex. I hope you guys enjoy this release. Keep an eye on our new releases @ Tellago Codeplex. We are working on various other Biztalk Artifacts like BAM, ESB Toolkit.     Till than happy BizzRuling…!!!     Thanks,   Vishal Mody

    Read the article

  • The Sensemaking Spectrum for Business Analytics: Translating from Data to Business Through Analysis

    - by Joe Lamantia
    One of the most compelling outcomes of our strategic research efforts over the past several years is a growing vocabulary that articulates our cumulative understanding of the deep structure of the domains of discovery and business analytics. Modes are one example of the deep structure we’ve found.  After looking at discovery activities across a very wide range of industries, question types, business needs, and problem solving approaches, we've identified distinct and recurring kinds of sensemaking activity, independent of context.  We label these activities Modes: Explore, compare, and comprehend are three of the nine recognizable modes.  Modes describe *how* people go about realizing insights.  (Read more about the programmatic research and formal academic grounding and discussion of the modes here: https://www.researchgate.net/publication/235971352_A_Taxonomy_of_Enterprise_Search_and_Discovery) By analogy to languages, modes are the 'verbs' of discovery activity.  When applied to the practical questions of product strategy and development, the modes of discovery allow one to identify what kinds of analytical activity a product, platform, or solution needs to support across a spread of usage scenarios, and then make concrete and well-informed decisions about every aspect of the solution, from high-level capabilities, to which specific types of information visualizations better enable these scenarios for the types of data users will analyze. The modes are a powerful generative tool for product making, but if you've spent time with young children, or had a really bad hangover (or both at the same time...), you understand the difficult of communicating using only verbs.  So I'm happy to share that we've found traction on another facet of the deep structure of discovery and business analytics.  Continuing the language analogy, we've identified some of the ‘nouns’ in the language of discovery: specifically, the consistently recurring aspects of a business that people are looking for insight into.  We call these discovery Subjects, since they identify *what* people focus on during discovery efforts, rather than *how* they go about discovery as with the Modes. Defining the collection of Subjects people repeatedly focus on allows us to understand and articulate sense making needs and activity in more specific, consistent, and complete fashion.  In combination with the Modes, we can use Subjects to concretely identify and define scenarios that describe people’s analytical needs and goals.  For example, a scenario such as ‘Explore [a Mode] the attrition rates [a Measure, one type of Subject] of our largest customers [Entities, another type of Subject] clearly captures the nature of the activity — exploration of trends vs. deep analysis of underlying factors — and the central focus — attrition rates for customers above a certain set of size criteria — from which follow many of the specifics needed to address this scenario in terms of data, analytical tools, and methods. We can also use Subjects to translate effectively between the different perspectives that shape discovery efforts, reducing ambiguity and increasing impact on both sides the perspective divide.  For example, from the language of business, which often motivates analytical work by asking questions in business terms, to the perspective of analysis.  The question posed to a Data Scientist or analyst may be something like “Why are sales of our new kinds of potato chips to our largest customers fluctuating unexpectedly this year?” or “Where can innovate, by expanding our product portfolio to meet unmet needs?”.  Analysts translate questions and beliefs like these into one or more empirical discovery efforts that more formally and granularly indicate the plan, methods, tools, and desired outcomes of analysis.  From the perspective of analysis this second question might become, “Which customer needs of type ‘A', identified and measured in terms of ‘B’, that are not directly or indirectly addressed by any of our current products, offer 'X' potential for ‘Y' positive return on the investment ‘Z' required to launch a new offering, in time frame ‘W’?  And how do these compare to each other?”.  Translation also happens from the perspective of analysis to the perspective of data; in terms of availability, quality, completeness, format, volume, etc. By implication, we are proposing that most working organizations — small and large, for profit and non-profit, domestic and international, and in the majority of industries — can be described for analytical purposes using this collection of Subjects.  This is a bold claim, but simplified articulation of complexity is one of the primary goals of sensemaking frameworks such as this one.  (And, yes, this is in fact a framework for making sense of sensemaking as a category of activity - but we’re not considering the recursive aspects of this exercise at the moment.) Compellingly, we can place the collection of subjects on a single continuum — we call it the Sensemaking Spectrum — that simply and coherently illustrates some of the most important relationships between the different types of Subjects, and also illuminates several of the fundamental dynamics shaping business analytics as a domain.  As a corollary, the Sensemaking Spectrum also suggests innovation opportunities for products and services related to business analytics. The first illustration below shows Subjects arrayed along the Sensemaking Spectrum; the second illustration presents examples of each kind of Subject.  Subjects appear in colors ranging from blue to reddish-orange, reflecting their place along the Spectrum, which indicates whether a Subject addresses more the viewpoint of systems and data (Data centric and blue), or people (User centric and orange).  This axis is shown explicitly above the Spectrum.  Annotations suggest how Subjects align with the three significant perspectives of Data, Analysis, and Business that shape business analytics activity.  This rendering makes explicit the translation and bridging function of Analysts as a role, and analysis as an activity. Subjects are best understood as fuzzy categories [http://georgelakoff.files.wordpress.com/2011/01/hedges-a-study-in-meaning-criteria-and-the-logic-of-fuzzy-concepts-journal-of-philosophical-logic-2-lakoff-19731.pdf], rather than tightly defined buckets.  For each Subject, we suggest some of the most common examples: Entities may be physical things such as named products, or locations (a building, or a city); they could be Concepts, such as satisfaction; or they could be Relationships between entities, such as the variety of possible connections that define linkage in social networks.  Likewise, Events may indicate a time and place in the dictionary sense; or they may be Transactions involving named entities; or take the form of Signals, such as ‘some Measure had some value at some time’ - what many enterprises understand as alerts.   The central story of the Spectrum is that though consumers of analytical insights (represented here by the Business perspective) need to work in terms of Subjects that are directly meaningful to their perspective — such as Themes, Plans, and Goals — the working realities of data (condition, structure, availability, completeness, cost) and the changing nature of most discovery efforts make direct engagement with source data in this fashion impossible.  Accordingly, business analytics as a domain is structured around the fundamental assumption that sense making depends on analytical transformation of data.  Analytical activity incrementally synthesizes more complex and larger scope Subjects from data in its starting condition, accumulating insight (and value) by moving through a progression of stages in which increasingly meaningful Subjects are iteratively synthesized from the data, and recombined with other Subjects.  The end goal of  ‘laddering’ successive transformations is to enable sense making from the business perspective, rather than the analytical perspective.Synthesis through laddering is typically accomplished by specialized Analysts using dedicated tools and methods. Beginning with some motivating question such as seeking opportunities to increase the efficiency (a Theme) of fulfillment processes to reach some level of profitability by the end of the year (Plan), Analysts will iteratively wrangle and transform source data Records, Values and Attributes into recognizable Entities, such as Products, that can be combined with Measures or other data into the Events (shipment of orders) that indicate the workings of the business.  More complex Subjects (to the right of the Spectrum) are composed of or make reference to less complex Subjects: a business Process such as Fulfillment will include Activities such as confirming, packing, and then shipping orders.  These Activities occur within or are conducted by organizational units such as teams of staff or partner firms (Networks), composed of Entities which are structured via Relationships, such as supplier and buyer.  The fulfillment process will involve other types of Entities, such as the products or services the business provides.  The success of the fulfillment process overall may be judged according to a sophisticated operating efficiency Model, which includes tiered Measures of business activity and health for the transactions and activities included.  All of this may be interpreted through an understanding of the operational domain of the businesses supply chain (a Domain).   We'll discuss the Spectrum in more depth in succeeding posts.

    Read the article

  • Pathfinding results in false path costs that are too high

    - by user2144536
    I'm trying to implement pathfinding in a game I'm programming using this method. I'm implementing it with recursion but some of the values after the immediate circle of tiles around the player are way off. For some reason I cannot find the problem with it. This is a screen cap of the problem: The pathfinding values are displayed in the center of every tile. Clipped blocks are displayed with the value of 'c' because the values were too high and were covering up the next value. The red circle is the first value that is incorrect. The code below is the recursive method. //tileX is the coordinates of the current tile, val is the current pathfinding value, used[][] is a boolean //array to keep track of which tiles' values have already been assigned public void pathFind(int tileX, int tileY, int val, boolean[][] used) { //increment pathfinding value int curVal = val + 1; //set current tile to true if it hasn't been already used[tileX][tileY] = true; //booleans to know which tiles the recursive call needs to be used on boolean topLeftUsed = false, topUsed = false, topRightUsed = false, leftUsed = false, rightUsed = false, botomLeftUsed = false, botomUsed = false, botomRightUsed = false; //set value of top left tile if necessary if(tileX - 1 >= 0 && tileY - 1 >= 0) { //isClipped(int x, int y) returns true if the coordinates givin are in a tile that can't be walked through (IE walls) //occupied[][] is an array that keeps track of which tiles have an enemy in them // //if the tile is not clipped and not occupied set the pathfinding value if(isClipped((tileX - 1) * 50 + 25, (tileY - 1) * 50 + 25) == false && occupied[tileX - 1][tileY - 1] == false && !(used[tileX - 1][tileY - 1])) { pathFindingValues[tileX - 1][tileY - 1] = curVal; topLeftUsed = true; used[tileX - 1][tileY - 1] = true; } //if it is occupied set it to an arbitrary high number so enemies find alternate routes if the best is clogged if(occupied[tileX - 1][tileY - 1] == true) pathFindingValues[tileX - 1][tileY - 1] = 1000000000; //if it is clipped set it to an arbitrary higher number so enemies don't travel through walls if(isClipped((tileX - 1) * 50 + 25, (tileY - 1) * 50 + 25) == true) pathFindingValues[tileX - 1][tileY - 1] = 2000000000; } //top middle if(tileY - 1 >= 0 ) { if(isClipped(tileX * 50 + 25, (tileY - 1) * 50 + 25) == false && occupied[tileX][tileY - 1] == false && !(used[tileX][tileY - 1])) { pathFindingValues[tileX][tileY - 1] = curVal; topUsed = true; used[tileX][tileY - 1] = true; } if(occupied[tileX][tileY - 1] == true) pathFindingValues[tileX][tileY - 1] = 1000000000; if(isClipped(tileX * 50 + 25, (tileY - 1) * 50 + 25) == true) pathFindingValues[tileX][tileY - 1] = 2000000000; } //top right if(tileX + 1 <= used.length && tileY - 1 >= 0) { if(isClipped((tileX + 1) * 50 + 25, (tileY - 1) * 50 + 25) == false && occupied[tileX + 1][tileY - 1] == false && !(used[tileX + 1][tileY - 1])) { pathFindingValues[tileX + 1][tileY - 1] = curVal; topRightUsed = true; used[tileX + 1][tileY - 1] = true; } if(occupied[tileX + 1][tileY - 1] == true) pathFindingValues[tileX + 1][tileY - 1] = 1000000000; if(isClipped((tileX + 1) * 50 + 25, (tileY - 1) * 50 + 25) == true) pathFindingValues[tileX + 1][tileY - 1] = 2000000000; } //left if(tileX - 1 >= 0) { if(isClipped((tileX - 1) * 50 + 25, (tileY) * 50 + 25) == false && occupied[tileX - 1][tileY] == false && !(used[tileX - 1][tileY])) { pathFindingValues[tileX - 1][tileY] = curVal; leftUsed = true; used[tileX - 1][tileY] = true; } if(occupied[tileX - 1][tileY] == true) pathFindingValues[tileX - 1][tileY] = 1000000000; if(isClipped((tileX - 1) * 50 + 25, (tileY) * 50 + 25) == true) pathFindingValues[tileX - 1][tileY] = 2000000000; } //right if(tileX + 1 <= used.length) { if(isClipped((tileX + 1) * 50 + 25, (tileY) * 50 + 25) == false && occupied[tileX + 1][tileY] == false && !(used[tileX + 1][tileY])) { pathFindingValues[tileX + 1][tileY] = curVal; rightUsed = true; used[tileX + 1][tileY] = true; } if(occupied[tileX + 1][tileY] == true) pathFindingValues[tileX + 1][tileY] = 1000000000; if(isClipped((tileX + 1) * 50 + 25, (tileY) * 50 + 25) == true) pathFindingValues[tileX + 1][tileY] = 2000000000; } //botom left if(tileX - 1 >= 0 && tileY + 1 <= used[0].length) { if(isClipped((tileX - 1) * 50 + 25, (tileY + 1) * 50 + 25) == false && occupied[tileX - 1][tileY + 1] == false && !(used[tileX - 1][tileY + 1])) { pathFindingValues[tileX - 1][tileY + 1] = curVal; botomLeftUsed = true; used[tileX - 1][tileY + 1] = true; } if(occupied[tileX - 1][tileY + 1] == true) pathFindingValues[tileX - 1][tileY + 1] = 1000000000; if(isClipped((tileX - 1) * 50 + 25, (tileY + 1) * 50 + 25) == true) pathFindingValues[tileX - 1][tileY + 1] = 2000000000; } //botom middle if(tileY + 1 <= used[0].length) { if(isClipped((tileX) * 50 + 25, (tileY + 1) * 50 + 25) == false && occupied[tileX][tileY + 1] == false && !(used[tileX][tileY + 1])) { pathFindingValues[tileX][tileY + 1] = curVal; botomUsed = true; used[tileX][tileY + 1] = true; } if(occupied[tileX][tileY + 1] == true) pathFindingValues[tileX][tileY + 1] = 1000000000; if(isClipped((tileX) * 50 + 25, (tileY + 1) * 50 + 25) == true) pathFindingValues[tileX][tileY + 1] = 2000000000; } //botom right if(tileX + 1 <= used.length && tileY + 1 <= used[0].length) { if(isClipped((tileX + 1) * 50 + 25, (tileY + 1) * 50 + 25) == false && occupied[tileX + 1][tileY + 1] == false && !(used[tileX + 1][tileY + 1])) { pathFindingValues[tileX + 1][tileY + 1] = curVal; botomRightUsed = true; used[tileX + 1][tileY + 1] = true; } if(occupied[tileX + 1][tileY + 1] == true) pathFindingValues[tileX + 1][tileY + 1] = 1000000000; if(isClipped((tileX + 1) * 50 + 25, (tileY + 1) * 50 + 25) == true) pathFindingValues[tileX + 1][tileY + 1] = 2000000000; } //call the method on the tiles that need it if(tileX - 1 >= 0 && tileY - 1 >= 0 && topLeftUsed) pathFind(tileX - 1, tileY - 1, curVal, used); if(tileY - 1 >= 0 && topUsed) pathFind(tileX , tileY - 1, curVal, used); if(tileX + 1 <= used.length && tileY - 1 >= 0 && topRightUsed) pathFind(tileX + 1, tileY - 1, curVal, used); if(tileX - 1 >= 0 && leftUsed) pathFind(tileX - 1, tileY, curVal, used); if(tileX + 1 <= used.length && rightUsed) pathFind(tileX + 1, tileY, curVal, used); if(tileX - 1 >= 0 && tileY + 1 <= used[0].length && botomLeftUsed) pathFind(tileX - 1, tileY + 1, curVal, used); if(tileY + 1 <= used[0].length && botomUsed) pathFind(tileX, tileY + 1, curVal, used); if(tileX + 1 <= used.length && tileY + 1 <= used[0].length && botomRightUsed) pathFind(tileX + 1, tileY + 1, curVal, used); }

    Read the article

  • Combining Shared Secret and Username Token – Azure Service Bus

    - by Michael Stephenson
    As discussed in the introduction article this walkthrough will explain how you can implement WCF security with the Windows Azure Service Bus to ensure that you can protect your endpoint in the cloud with a shared secret but also flow through a username token so that in your listening WCF service you will be able to identify who sent the message. This could either be in the form of an application or a user depending on how you want to use your token. Prerequisites Before going into the walk through I want to explain a few assumptions about the scenario we are implementing but to keep the article shorter I am not going to walk through all of the steps in how to setup some of this. In the solution we have a simple console application which will represent the client application. There is also the services WCF application which contains the WCF service we will expose via the Windows Azure Service Bus. The WCF Service application in this example was hosted in IIS 7 on Windows 2008 R2 with AppFabric Server installed and configured to auto-start the WCF listening services. I am not going to go through significant detail around the IIS setup because it should not matter in relation to this article however if you want to understand more about how to configure WCF and IIS for such a scenario please refer to the following paper which goes into a lot of detail about how to configure this. The link is: http://tinyurl.com/8s5nwrz   The Service Component To begin with let's look at the service component and how it can be configured to listen to the service bus using a shared secret but to also accept a username token from the client. In the sample the service component is called Acme.Azure.ServiceBus.Poc.UN.Services. It has a single service which is the Visual Studio template for a WCF service when you add a new WCF Service Application so we have a service called Service1 with its Echo method. Nothing special so far!.... The next step is to look at the web.config file to see how we have configured the WCF service. In the services section of the WCF configuration you can see I have created my service and I have created a local endpoint which I simply used to do a little bit of diagnostics and to check it was working, but more importantly there is the Windows Azure endpoint which is using the ws2007HttpRelayBinding (note that this should also work just the same if your using netTcpRelayBinding). The key points to note on the above picture are the service behavior called MyServiceBehaviour and the service bus endpoints behavior called MyEndpointBehaviour. We will go into these in more detail later.   The Relay Binding The relay binding for the service has been configured to use the TransportWithMessageCredential security mode. This is the important bit where the transport security really relates to the interaction between the service and listening to the Azure Service Bus and the message credential is where we will use our username token like we have specified in the message/clientCrentialType attribute. Note also that we have left the relayClientAuthenticationType set to RelayAccessToken. This means that authentication will be made against ACS for accessing the service bus and messages will not be accepted from any sender who has not been authenticated by ACS.   The Endpoint Behaviour In the below picture you can see the endpoint behavior which is configured to use the shared secret client credential for accessing the service bus and also for diagnostic purposes I have included the service registry element. Hopefully if you are familiar with using Windows Azure Service Bus relay feature the above is very familiar to you and this is a very common setup for this section. There is nothing specific to the username token implementation here. The Service Behaviour Now we come to the bit with most of the username token bits in it. When you configure the service behavior I have included the serviceCredentials element and then setup to use userNameAuthentication and you can see that I have created my own custom username token validator.   This setup means that WCF will hand off to my class for validating the username token details. I have also added the serviceSecurityAudit element to give me a simple auditing of access capability. My UsernamePassword Validator The below picture shows you the details of the username password validator class I have implemented. WCF will hand off to this class when validating the token and give me a nice way to check the token credentials against an on-premise store. You have all of the validation features with a non-service bus WCF implementation available such as validating the username password against active directory or ASP.net membership features or as in my case above something much simpler.   The Client Now let's take a look at the client side of this solution and how we can configure the client to authenticate against ACS but also send a username token over to the service component so it can implement additional security checks on-premise. I have a console application and in the program class I want to use the proxy generated with Add Service Reference to send a message via the Azure Service Bus. You can see in my WCF client configuration below I have setup my details for the azure service bus url and am using the ws2007HttpRelayBinding. Next is my configuration for the relay binding. You can see below I have configured security to use TransportWithMessageCredential so we will flow the username token with the message and also the RelayAccessToken relayClientAuthenticationType which means the component will validate against ACS before being allowed to access the relay endpoint to send a message.     After the binding we need to configure the endpoint behavior like in the below picture. This is the normal configuration to use a shared secret for accessing a Service Bus endpoint.   Finally below we have the code of the client in the console application which will call the service bus. You can see that we have created our proxy and then made a normal call to a WCF service but this time we have also set the ClientCredentials to use the appropriate username and password which will be flown through the service bus and to our service which will validate them.     Conclusion As you can see from the above walkthrough it is not too difficult to configure a service to use both a shared secret and username token at the same time. This gives you the power and protection offered by the access control service in the cloud but also the ability to flow additional tokens to the on-premise component for additional security features to be implemented. Sample The sample used in this post is available at the following location: https://s3.amazonaws.com/CSCBlogSamples/Acme.Azure.ServiceBus.Poc.UN.zip

    Read the article

  • Checking who is connected to your server, with PowerShell.

    - by Fatherjack
    There are many occasions when, as a DBA, you want to see who is connected to your SQL Server, along with how they are connecting and what sort of activities they are carrying out. I’m going to look at a couple of ways of getting this information and compare the effort required and the results achieved of each. SQL Server comes with a couple of stored procedures to help with this sort of task – sp_who and its undocumented counterpart sp_who2. There is also the pumped up version of these called sp_whoisactive, written by Adam Machanic which does way more than these procedures. I wholly recommend you try it out if you don’t already know how it works. When it comes to serious interrogation of your SQL Server activity then it is absolutely indispensable. Anyway, back to the point of this blog, we are going to look at getting the information from sp_who2 for a remote server. I wrote this Powershell script a week or so ago and was quietly happy with it for a while. I’m relatively new to Powershell so forgive both my rather low threshold for entertainment and the fact that something so simple is a moderate achievement for me. $Server = 'SERVERNAME' $SMOServer = New-Object Microsoft.SQLServer.Management.SMO.Server $Server # connection and query stuff         $ConnectionStr = "Server=$Server;Database=Master;Integrated Security=True" $Query = "EXEC sp_who2" $Connection = new-object system.Data.SQLClient.SQLConnection $Table = new-object "System.Data.DataTable" $Connection.connectionstring = $ConnectionStr try{ $Connection.open() $Command = $Connection.CreateCommand() $Command.commandtext = $Query $result = $Command.ExecuteReader() $Table.Load($result) } catch{ # Show error $error[0] | format-list -Force } $Title = "Data access processes (" + $Table.Rows.Count + ")" $Table | Out-GridView -Title $Title $Connection.close() So this is pretty straightforward, create an SMO object that represents our chosen server, define a connection to the database and a table object for the results when we get them, execute our query over the connection, load the results into our table object and then, if everything is error free display these results to the PowerShell grid viewer. The query simply gets the results of ‘EXEC sp_who2′ for us. Depending on how many connections there are will influence how long the query runs. The grid viewer lets me sort and search the results so it can be a pretty handy way to locate troublesome connections. Like I say, I was quite pleased with this, it seems a pretty simple script and was working well for me, I have added a few parameters to control the output and give me more specific details but then I see a script that uses the $SMOServer object itself to provide the process information and saves having to define the connection object and query specifications. $Server = 'SERVERNAME' $SMOServer = New-Object Microsoft.SQLServer.Management.SMO.Server $Server $Processes = $SMOServer.EnumProcesses() $Title = "SMO processes (" + $Processes.Rows.Count + ")" $Processes | Out-GridView -Title $Title Create the SMO object of our server and then call the EnumProcesses method to get all the process information from the server. Staggeringly simple! The results are a little different though. Some columns are the same and we can see the same basic information so my first thought was to which runs faster – so that I can get my results more quickly and also so that I place less stress on my server(s). PowerShell comes with a great way of testing this – the Measure-Command function. All you have to do is wrap your piece of code in Measure-Command {[your code here]} and it will spit out the time taken to execute the code. So, I placed both of the above methods of getting SQL Server process connections in two Measure-Command wrappers and pressed F5! The Powershell console goes blank for a while as the code is executed internally when Measure-Command is used but the grid viewer windows appear and the console shows this. You can take the output from Measure-Command and format it for easier reading but in a simple comparison like this we can simply cross refer the TotalMilliseconds values from the two result sets to see how the two methods performed. The query execution method (running EXEC sp_who2 ) is the first set of timings and the SMO EnumProcesses is the second. I have run these on a variety of servers and while the results vary from execution to execution I have never seen the SMO version slower than the other. The difference has varied and the time for both has ranged from sub-second as we see above to almost 5 seconds on other systems. This difference, I would suggest is partly due to the cost overhead of having to construct the data connection and so on where as the SMO EnumProcesses method has the connection to the server already in place and just needs to call back the process information. There is also the difference in the data sets to consider. Let’s take a look at what we get and where the two methods differ Query execution method (sp_who2) SMO EnumProcesses Description - Urn What looks like an XML or JSON representation of the server name and the process ID SPID Spid The process ID Status Status The status of the process Login Login The login name of the user executing the command HostName Host The name of the computer where the  process originated BlkBy BlockingSpid The SPID of a process that is blocking this one DBName Database The database that this process is connected to Command Command The type of command that is executing CPUTime Cpu The CPU activity related to this process DiskIO - The Disk IO activity related to this process LastBatch - The time the last batch was executed from this process. ProgramName Program The application that is facilitating the process connection to the SQL Server. SPID1 - In my experience this is always the same value as SPID. REQUESTID - In my experience this is always 0 - Name In my experience this is always the same value as SPID and so could be seen as analogous to SPID1 from sp_who2 - MemUsage An indication of the memory used by this process but I don’t know what it is measured in (bytes, Kb, Mb…) - IsSystem True or False depending on whether the process is internal to the SQL Server instance or has been created by an external connection requesting data. - ExecutionContextID In my experience this is always 0 so could be analogous to REQUESTID from sp_who2. Please note, these are my own very brief descriptions of these columns, detail can be found from MSDN for columns in the sp_who results here http://msdn.microsoft.com/en-GB/library/ms174313.aspx. Where the columns are common then I would use that description, in other cases then the information returned is purely for interpretation by the reader. Rather annoyingly both result sets have useful information that the other doesn’t. sp_who2 returns Disk IO and LastBatch information which is really useful but the SMO processes method give you IsSystem and MemUsage which have their place in fault diagnosis methods too. So which is better? On reflection I think I prefer to use the sp_who2 method primarily but knowing that the SMO Enumprocesses method is there when I need it is really useful and I’m sure I’ll use it regularly. I’m OK with the fact that it is the slower method because Measure-Command has shown me how close it is to the other option and that it really isn’t a large enough margin to matter.

    Read the article

  • MapRedux - PowerShell and Big Data

    - by Dittenhafer Solutions
    MapRedux – #PowerShell and #Big Data Have you been hearing about “big data”, “map reduce” and other large scale computing terms over the past couple of years and been curious to dig into more detail? Have you read some of the Apache Hadoop online documentation and unfortunately concluded that it wasn't feasible to setup a “test” hadoop environment on your machine? More recently, I have read about some of Microsoft’s work to enable Hadoop on the Azure cloud. Being a "Microsoft"-leaning technologist, I am more inclinded to be successful with experimentation when on the Windows platform. Of course, it is not that I am "religious" about one set of technologies other another, but rather more experienced. Anyway, within the past couple of weeks I have been thinking about PowerShell a bit more as the 2012 PowerShell Scripting Games approach and it occured to me that PowerShell's support for Windows Remote Management (WinRM), and some other inherent features of PowerShell might lend themselves particularly well to a simple implementation of the MapReduce framework. I fired up my PowerShell ISE and started writing just to see where it would take me. Quite simply, the ScriptBlock feature combined with the ability of Invoke-Command to create remote jobs on networked servers provides much of the plumbing of a distributed computing environment. There are some limiting factors of course. Microsoft provided some default settings which prevent PowerShell from taking over a network without administrative approval first. But even with just one adjustment, a given Windows-based machine can become a node in a MapReduce-style distributed computing environment. Ok, so enough introduction. Let's talk about the code. First, any machine that will participate as a remote "node" will need WinRM enabled for remote access, as shown below. This is not exactly practical for hundreds of intended nodes, but for one (or five) machines in a test environment it does just fine. C:> winrm quickconfig WinRM is not set up to receive requests on this machine. The following changes must be made: Set the WinRM service type to auto start. Start the WinRM service. Make these changes [y/n]? y Alternatively, you could take the approach described in the Remotely enable PSRemoting post from the TechNet forum and use PowerShell to create remote scheduled tasks that will call Enable-PSRemoting on each intended node. Invoke-MapRedux Moving on, now that you have one or more remote "nodes" enabled, you can consider the actual Map and Reduce algorithms. Consider the following snippet: $MyMrResults = Invoke-MapRedux -MapReduceItem $Mr -ComputerName $MyNodes -DataSet $dataset -Verbose Invoke-MapRedux takes an instance of a MapReduceItem which references the Map and Reduce scriptblocks, an array of computer names which are the remote nodes, and the initial data set to be processed. As simple as that, you can start working with concepts of big data and the MapReduce paradigm. Now, how did we get there? I have published the initial version of my PsMapRedux PowerShell Module on GitHub. The PsMapRedux module provides the Invoke-MapRedux function described above. Feel free to browse the underlying code and even contribute to the project! In a later post, I plan to show some of the inner workings of the module, but for now let's move on to how the Map and Reduce functions are defined. Map Both the Map and Reduce functions need to follow a prescribed prototype. The prototype for a Map function in the MapRedux module is as follows. A simple scriptblock that takes one PsObject parameter and returns a hashtable. It is important to note that the PsObject $dataset parameter is a MapRedux custom object that has a "Data" property which offers an array of data to be processed by the Map function. $aMap = { Param ( [PsObject] $dataset ) # Indicate the job is running on the remote node. Write-Host ($env:computername + "::Map"); # The hashtable to return $list = @{}; # ... Perform the mapping work and prepare the $list hashtable result with your custom PSObject... # ... The $dataset has a single 'Data' property which contains an array of data rows # which is a subset of the originally submitted data set. # Return the hashtable (Key, PSObject) Write-Output $list; } Reduce Likewise, with the Reduce function a simple prototype must be followed which takes a $key and a result $dataset from the MapRedux's partitioning function (which joins the Map results by key). Again, the $dataset is a MapRedux custom object that has a "Data" property as described in the Map section. $aReduce = { Param ( [object] $key, [PSObject] $dataset ) Write-Host ($env:computername + "::Reduce - Count: " + $dataset.Data.Count) # The hashtable to return $redux = @{}; # Return Write-Output $redux; } All Together Now When everything is put together in a short example script, you implement your Map and Reduce functions, query for some starting data, build the MapReduxItem via New-MapReduxItem and call Invoke-MapRedux to get the process started: # Import the MapRedux and SQL Server providers Import-Module "MapRedux" Import-Module “sqlps” -DisableNameChecking # Query the database for a dataset Set-Location SQLSERVER:\sql\dbserver1\default\databases\myDb $query = "SELECT MyKey, Date, Value1 FROM BigData ORDER BY MyKey"; Write-Host "Query: $query" $dataset = Invoke-SqlCmd -query $query # Build the Map function $MyMap = { Param ( [PsObject] $dataset ) Write-Host ($env:computername + "::Map"); $list = @{}; foreach($row in $dataset.Data) { # Write-Host ("Key: " + $row.MyKey.ToString()); if($list.ContainsKey($row.MyKey) -eq $true) { $s = $list.Item($row.MyKey); $s.Sum += $row.Value1; $s.Count++; } else { $s = New-Object PSObject; $s | Add-Member -Type NoteProperty -Name MyKey -Value $row.MyKey; $s | Add-Member -type NoteProperty -Name Sum -Value $row.Value1; $list.Add($row.MyKey, $s); } } Write-Output $list; } $MyReduce = { Param ( [object] $key, [PSObject] $dataset ) Write-Host ($env:computername + "::Reduce - Count: " + $dataset.Data.Count) $redux = @{}; $count = 0; foreach($s in $dataset.Data) { $sum += $s.Sum; $count += 1; } # Reduce $redux.Add($s.MyKey, $sum / $count); # Return Write-Output $redux; } # Create the item data $Mr = New-MapReduxItem "My Test MapReduce Job" $MyMap $MyReduce # Array of processing nodes... $MyNodes = ("node1", "node2", "node3", "node4", "localhost") # Run the Map Reduce routine... $MyMrResults = Invoke-MapRedux -MapReduceItem $Mr -ComputerName $MyNodes -DataSet $dataset -Verbose # Show the results Set-Location C:\ $MyMrResults | Out-GridView Conclusion I hope you have seen through this article that PowerShell has a significant infrastructure available for distributed computing. While it does take some code to expose a MapReduce-style framework, much of the work is already done and PowerShell could prove to be the the easiest platform to develop and run big data jobs in your corporate data center, potentially in the Azure cloud, or certainly as an academic excerise at home or school. Follow me on Twitter to stay up to date on the continuing progress of my Powershell MapRedux module, and thanks for reading! Daniel

    Read the article

  • Google Analytics on Android

    - by pjv
    There is a specific and official analytics SDK for native Android apps (note that I'm not talking about webpages in apps on a phone). This library basically sends pages and events to Google Analytics and you can view your analytics in exactly the same dashboard as for websites. Since my background is apps rather than websites, and since a lot of the Google Analytics terminology seems particularly inapplicable to a native app, I need some pointers. Please discuss my remarks, provide some clarification where you think I'm off-track, and above all share good experiences! 1. Page Views Pages mostly can match different Activities (and Dialogs) being displayed. Activities can be visible behind non-full-screen Activities however, though only the top-level Activity can be interacted. This sort-off clashes with a "(page) view". You'd also want at least one page view for each visit and therefore put one page view tracker in the Application class. However this does not constitute a window or sorts. Usually an Activity will open at the same time, so the time spent on that page will have been 0. This will influence your "time spent" statistics. How are these counted anyway? Moreover, there is a loose coupling between the Activities, by means of Intents. A user can, much like on any website, step in at any Activity, although usually this then concerns resuming the application where he left off. This makes that the hierarchy of Activities usually is very flat. And since there are no url's involved. What meaning would using slashes in page titles have, such as "/Home"? All pages would appear on an equal level in the reports, so no content drilldown. Non-unique page views seem to be counted as some kind of indicator of successfulness: how often does the visitor revisit the page. When the user rotates the screen however usually an Activity resumes again, thus making it a new page view. This happens a lot. Maybe a well-thought-through placement of the call might solve this, or placing several, I'm not sure. How to deal with Page Views? 2. Events I'd say there are two sorts: A user event Something that happened, usually as an indirect consequence of the above. The latter particularly is giving me headaches. First of all, many events aren't written in code any more, but pieced logically together by means of Intents. This means that there is no place to put the analytics call. You'd either have to give up this advantage and start doing it the old-fashioned way in favor of good analytics, or, just be missing some events. Secondly, as a developer you're not so much interested in when a user clicks a button, but if the action that should have been performed really was performed and what the result was. There seems to be no clear way to get resulting data into Google Analytics (what's up with the integers? I want to put in Strings!). The same that applies to the flat pages hierarchy, also goes for the event categories. You could do "vertical" categories (topically, that is), but some code is shared "horizontally" and the tracking will be equally shared. Just as with the Intents mechanism, inheritance makes it hard for you to put the tracking in the right places at all times. And I can't really imagine "horizontal" categories. Unless you start making really small categories, such as all the items form the same menu in one category, I have a hard time grasping the concept. Finally, how do you deal with cancelling? Usually you both have an explicit cancel mechanism by ways of a button, as well as the implicit cancel when the "back"-button is pressed to leave the activity and there were no changes. The latter also applies to "saves", when the back button is pressed and there ARE changes. How are you consequently going to catch all these if not by doing all the "back"-button work yourself? How to deal with events? 3. Goals For goal types I have choice of: URL Destination, Time on Site, and Pages/Visit. Most apps don't have a funnel that leads the user to some "registration done" or "order placed" page. Apps have either already been bought (in which case you want to stimulate the user to love your app, so that he might bring on new buyers) or are paid for by in-app ads. So URL Destination is not a very important goal. Time on Site also seems troublesome. First, I have some doubt on how this would be measured. Second, I don't necessarily want my user to spend a lot of time in my already paid app, just be active and content. Equivalently, why not mention how frequent a user uses your app? Regarding Pages/Visit I already mentioned how screen orientation changes blow up the page view numbers. In an app I'd be most interested in events/visit to measure the user's involvement/activity. If he's intensively using the app then he must be loving it right? Furthermore, I also have some small funnels (that do not lead to conversion though) that I want to see streamlined. In my mind those funnels would end in events rather than page views but that seems not to be possible. I could also measure clickthroughs on in-app ads, but then I'd need to track those as Page Views rather than Events, in view of "URL Destination". What are smart goals for apps and how can you fit them on top of Analytics? 4. Optimisation Is there a smart way to manually do what "Website Optimiser" does for websites? Most importantly, how would I track different landing page designs? 5. Traffic Sources Referrals deal with installation time referrals, if you're smart enough to get them included. But perhaps I'd also want to get some data which third-party app sends users to my app to perform some actions (this app interoperability is possible via Intents). Many of the terminologies related to "Traffic Sources" seem totally meaningless and there is no possibility of connecting in AdSense. What are smart uses of this data? 6. Visitors Of the "Browser capabilities", "Network Properties" and "Mobile" tabs, many things are pointless as they have no influence on / relation with my mostly offline app that won't use flash anyway. Only if you drill down far enough, can you get to OS versions, which do matter a lot. I even forgot where you could check what exact Android devices visited. What are smart uses of this data? How can you make the relevant info more prominent? 7. Other No in-page analytics. I have to register my app as a web-url (What!?)?

    Read the article

  • I Know What I Did This Summer: Put Down Trex Decking

    - by thatjeffsmith
    If you’re wondering why I would bore everyone with my pictures and frequent status updates/tweets from the past week – it’s so I could document the process of refurbishing my deck, or what some would call a porch. When we go to take a vacation, buy a car, do anything – we also read personal blogs to get the real story. So, if you’re curious about what it takes to tackle this sort of project, read on. Skills/Equipment/Manpower We Possessed I took the old decking out by myself. I’m about 230 lbs, more than 6′ tall, and I’m pretty healthy. This took about 8 hours over two afternoons. Three of us put the deck back together. My wife has two engineering degrees. Her father also has two engineering degrees. Lots of brainpower available here. Also, her dad ran the public works department for a country for more than 20 years – so lots and lots of practical experience on hand. We had a compound mitre saw, a skilsaw, 2-3 crowbars, a framing hammer, 3 cordless drills, a corded drill, lots of sawhorses, a power sander, an angle grinder, a 10×10 Coleman canopy tent, a Ford F-150 pickup truck, outdoor speakers and lots of iTunes playlists, plenty of water and cold beer. Why We Did This Our deck was relatively young – it was built in 2005. However, the pressure treated boards must not have been adequately maintained before we bought the house. I had powerwashed the deck every other year and had it stained a few times. The boards just rotted. We’re going to be in the house for a long time, and we wanted something that would look nice and require little maintenance. More bad deck boards The deck boards were in bad shape Things We Learned The two most important things: The hidden fasteners have to be put in JUST right. Wedge them into the grooved board, then bend down the bit that is screwed down. We didn’t do this on the first board and couldn’t get the second board to fit nearly close enough. Watching the official TREX YouTube video helped immensely, and we should have watched that first. When pre-drilling holes for the boards that need screwed down – DO NOT pre-drill through the underlying framing wood. ONLY pre-drill through the TREX itself. The screw won’t seat in the board properly. Instead of sitting down flush with the board, it will stop at the top of the board and just spin. I had to call the the place that sold me the screws to find this out. So about a third of our screws look like crap. If it doesn’t look or feel right – stop everything and pick up your computer or your phone. It’s not right, and it will be much easier to stop and find out why. We didn’t do this, and now I’m going to see every screw that’s not flush with the boards and get upset. Oh well. The Process How much time did it take? Well I spent about 8 hours taking the deck apart. And then the 3 of use spent 8 hours the first day, 10 hours the second day, 8 hours the third, and another 6 hours on the fourth day. That’s like 104 man-hours. We supposedly saved four or five thousand dollars in labor, but don’t do the math here or you might get a bit upset. The main thing is that we got what we wanted, and there won’t be any surprises later. Now for some pictures… This 6”+ pry bar made the destruction of the old deck much easier Most of the joists, once exposed, were OK. This joist wasn’t sitting on ANYTHING before. We think a lazy gas person cut the board to sneak a gas line in. Awesome… These monster lag bolts had to be accounted for when putting in the additional framing The border pattern Sheri wanted to put in required a lot more framing. These were the first boards to go down – we screwed them in as there was no way to attach clips I sat, kicked in the boards, and then drilled these clips in – but my wife was able to go MUCH faster by using her hands to lock the boards in and drill on her knees. I liked locking the board in with my feet when they needed to be ‘encouraged’ to go straight. The first board took FOREVER to go in, but then when we got rolling, we were able to put in a 20′ board in less than 10 minutes. This was end of construction day #2 – we got much further than we thought we would. Ah, the dreaded last 10% – what to do here? Remember those ‘floating’ stringers? Yeah, we fixed that up a bit, too. My wife used a website (and her brain) to calculate exactly how to cut the stringers to give us the rise/run we needed with the proper clearance and all that jazz. The stairs with stringers and toe kicks – this was worth the effort It started raining on us as I screwed down the steps – this we managed to get our shade tent up on the deck to protect us from the rain too The stairs, finished Finished, mostly Good corner shot The top of the stairs Stairs, looking down Celebratory beer In Summary There are a few things we’re not happy with. I think we can fix them up – but later. I have a few things left to finish, rewire the lighting, get the gas grille put back in, and rehang some screen doors. I was expecting this to be a lot worse than it was. If I didn’t have the help, I would have never done it myself. But I’m glad that I did have that help and did do that project. It’s not often you get to spend that kind of qualify time with family and building cool stuff.

    Read the article

  • D2K to OA Framework Transition

    - by PRajkumar
    What is the difference between D2K form and OA Framework? It is a very innocent but important question for someone that desires to make transition from D2K to OA Framework. I hope you have already read and implemented OA Framework Getting Started. I will re-visit my own experience of implementing HelloWorld program in "OA Framework". When I implemented HelloWorld a year ago, I had no clue as to what I was doing & why I was doing those steps. I merely copied the steps from Oracle Tutorial without understanding them. Hence in this blog, I will try to explain in simple manner the meaning of OA Framework HelloWorld Program and compare the steps to D2K form [where possible]. To keep things simple, only basics will be discussed. Following key Steps were needed for HelloWorld Step 1 Create a new Workspace and a new Project as dictated by Oracle's tutorial. When defining project, you will specify a default package, which in this case was oracle.apps.ak.hello This means the following: - ak is the short name of the Application in Oracle           [means fnd_applications.short_name] hello is the name of your project Step 2 Next, you will create a OA Page within hello project Think OA Page as the fmx file itself in D2K. I am saying so because this page gets attached to the form function. This page will be created within hello project, hence the package name oracle.apps.ak.hello.webui Note the webui, it is a convention to have page in webui, means this page represents the Web User Interface You will assign the default AM [OAApplicationModule]. Think of AM "Connection Manager" and "Transaction State Manager" for your page          I can't co-relate this to anything in D2k, as there is no concept of Connection Pooling and that D2k is not stateless. Reason being that as soon as you kick off a D2K Form, it connects to a single session of Oracle and sticks to that single Oracle database session. So is not the case in OAF, hence AM is needed. Step 3 You create Region within the Page. ·         Region is what will store your fields. Text input fields will be of type messageTextInput. Think of Canvas in D2K. You can have nested regions. Stacked Canvas in D2K comes the closest to this component of OA Framework Step 4 Add a button to one of the nested regions The itemStyle should be submitButton, in case you want the page to be submitted when this button is clicked There is no WHEN-BUTTON-PRESSED trigger in OAF. In Framework, you will add a controller java code to handle events like Form Submit button clicks. JDeveloper generates the default code for you. Primarily two functions [should I call methods] will be created processRequest [for UI Rendering Handling] and processFormRequest          Think of processRequest as WHEN-NEW-FORM-INSTANCE, though processRequest is very restrictive. Note What is the difference between processRequest and processFormRequest? These two methods are available in the Default Controller class that gets created. processFormRequest This method is commonly used to react/respond to the event that has taken place, for example click of a button. Some examples are if(oapagecontext.getParameter("Cancel") != null) (Do your processing for Cancellation/ Rollback) if(oapagecontext.getParameter("Submit") != null) (Do your validations and commit here) if(oapagecontext.getParameter("Update") != null) (Do your validations and commit here) In the above three examples, you could be calling oapagecontext.forwardImmediately to re-direct the page navigation to some other page if needed. processRequest In this method, usually page rendering related code is written. Effectively, each GUI component is a bean that gets initialised during processRequest. Those who are familiar with D2K forms, something like pre-query may be written in this method. Step 5 In the controller to access the value in field "HelloName" the command is String userContent = pageContext.getParameter("HelloName"); In D2k, we used :block.field. In OAFramework, at submission of page, all the field values get passed into to OAPageContext object. Use getParameter to access the field value To set the value of the field, use OAMessageTextInputBean field HelloName = (OAMessageTextInputBean)webBean.findChildRecursive("HelloName"); fieldHelloName.setText(pageContext,"Setting the default value" ); Note when setting field value in controller: Note 1. Do not set the value in processFormRequest Note 2. If the field comes from View Object, then do not use setText in controller Note 3. For control fields [that are not based on View Objects], you can use setText to assign values in processRequest method Lets take some notes to expand beyond the HelloWorld Project Note 1 In D2K-forms we sort of created a Window, attached to Canvas, and then fields within that Canvas. However in OA Framework, think of Page being fmx/Window, think of Region being a Canvas, and fields being within Regions. This is not a formal/accurate understanding of analogy between D2k and Framework, but is close to being logical. Note 2 In D2k, your Forms fmb file was compiled to fmx. It was fmx file that was deployed on mid-tier. In case of OAF, your OA Page is nothing but a XML file. We call this MDS [meta data]. Whatever name you give to "Page" in OAF, an XML file of the same name gets created. This xml file must then be loaded into database by using XML Importer command. Note 3 Apart from MDS XML file, almost everything else is merely deployed to your mid-tier. Usually this is underneath $JAVA_TOP/oracle/apps/../.. All java files will go underneath java top/oracle/apps/../.. etc. Note 4 When building tutorial, ignore the steps for setting "Attribute Sets". These are not mandatory. Oracle might just have developed their tutorials without including these. Think of these like Visual Attributes of D2K forms Note 5 Controller is where you will write any java code in OA Framework. You can create a Controller per Page or have a different Controller for each of the Regions with the same Page. Note 6 In the method processFormRequest of the Controller, you can access the values of the page by using notation pageContext.getParameter("<fieldname here>"). This method processFormRequest is executed when the OAF Screen/Page is submitted by click of a button. Note 7 Inside the controller, all the Database Related interactions for example interaction with View Objects happen via Application Module. But why so? Because Application Module Manages the transaction state of the Application. OAApplicationModuleImpl oaapplicationmoduleimpl = OAApplicationModuleImpl)oapagecontext.getApplicationModule(oawebbean); OADBTransaction oadbtransaction = OADBTransaction)oaapplicationmoduleimpl.getDBTransaction(); Note 8 In D2K, we have control block or a block based on database view. Similarly, in OA Framework, if the field does not have view Object attached, then it is like a control field. Hence in HelloWorld example, field HelloName is a control field [in D2K terminology]. A view Object can either be based on a view/table, synonym or on a SQL statement. Note 9 I wish to access the fields in multi record block that is based on view Object. Can I do this in Controller? Sure you can. To traverse through those records, do the below ·         Get the reference to the View Object using (OAViewObject)oapagecontext.getApplicationModule(oawebbean).findViewObject("VO Name Here") ·         Loop through the records in View Objects using count returned from oaviewobject.getFetchedRowCount() ·         For each record, fetch the value of the fields within the loop as oracle.jbo.Row row = oaviewobject.getRowAtRangeIndex(loop index here); (String)row.getAttribute("Column name of VO here ");

    Read the article

  • Recursion in the form of a Recursive Func&lt;T, T&gt;

    - by ToStringTheory
    I gotta admit, I am kind of surprised that I didn’t realize I could do this sooner.  I recently had a problem which required a recursive function call to come up with the answer.  After some time messing around with a recursive method, and creating an API that I was not happy with, I was able to create an API that I enjoy, and seems intuitive. Introduction To bring it to a simple example, consider the summation to n: A mathematically identical formula is: In a .NET function, this can be represented by a function: Func<int, int> summation = x => x*(x+1)/2 Calling summation with an input integer will yield the summation to that number: var sum10 = summation(4); //sum10 would be equal to 10 But what if I wanted to get a second level summation…  First some to n, and then use that argument as the input to the same function, to find the second level summation: So as an easy example, calculate the summation to 3, which yields 6.  Then calculate the summation to 6 which yields 21. Represented as a mathematical formula - So what if I wanted to represent this as .NET functions.  I can always do: //using the summation formula from above var sum3 = summation(3); //sets sum3 to 6 var sum3_2 = summation(sum3); //sets sum3 to 21 I could always create a while loop to perform the calculations too: Func<int, int> summation = x => x*(x+1)/2; //for the interests of a smaller example, using shorthand int sumResultTo = 3; int level = 2; while(level-- > 0) { sumResultTo = summation(sumResultTo); } //sumResultTo is equal to 21 now. Or express it as a for-loop, method calls, etc…  I really didn’t like any of the options that I tried.  Then it dawned on me – since I was using a Func<T, T> anyways, why not use the Func’s output from one call as the input as another directly. Some Code So, I decided that I wanted a recursion class.  Something that I would be generic and reusable in case I ever wanted to do something like this again. It is limited to only the Func<T1, T2> level of Func, and T1 must be the same as T2. The first thing in this class is a private field for the function: private readonly Func<T, T> _functionToRecurse; So, I since I want the function to be unchangeable, I have defined it as readonly.  Therefore my constructor looks like: public Recursion(Func<T, T> functionToRecurse) { if (functionToRecurse == null) { throw new ArgumentNullException("functionToRecurse", "The function to recurse can not be null"); } _functionToRecurse = functionToRecurse; } Simple enough.  If you have any questions, feel free to post them in the comments, and I will be sure to answer them. Next, I want enough. If be able to get the result of a function dependent on how many levels of recursion: private Func<T, T> GetXLevel(int level) { if (level < 1) { throw new ArgumentOutOfRangeException("level", level, "The level of recursion must be greater than 0"); } if (level == 1) return _functionToRecurse; return _GetXLevel(level - 1, _functionToRecurse); } So, if you pass in 1 for the level, you get just the Func<T,T> back.  If you say that you want to go deeper down the rabbit hole, it calls a method which accepts the level it is at, and the function which it needs to use to recurse further: private Func<T, T> _GetXLevel(int level, Func<T, T> prevFunc) { if (level == 1) return y => prevFunc(_functionToRecurse(y)); return _GetXLevel(level - 1, y => prevFunc(_functionToRecurse(y))); } That is really all that is needed for this class. If I exposed the GetXLevel function publicly, I could use that to get the function for a level, and pass in the argument..  But I wanted something better.  So, I used the ‘this’ array operator for the class: public Func<T,T> this[int level] { get { if (level < 1) { throw new ArgumentOutOfRangeException("level", level, "The level of recursion must be greater than 0"); } return this.GetXLevel(level); } } So, using the same example above of finding the second recursion of the summation of 3: var summator = new Recursion<int>(x => (x * (x + 1)) / 2); var sum_3_level2 = summator[2](3); //yields 21 You can even find just store the delegate to the second level summation, and use it multiple times: var summator = new Recursion<int>(x => (x * (x + 1)) / 2); var sum_level2 = summator[2]; var sum_3_level2 = sum_level2(3); //yields 21 var sum_4_level2 = sum_level2(4); //yields 55 var sum_5_level2 = sum_level2(5); //yields 120 Full Code Don’t think I was just going to hold off on the full file together and make you do the hard work…  Copy this into a new class file: public class Recursion<T> { private readonly Func<T, T> _functionToRecurse; public Recursion(Func<T, T> functionToRecurse) { if (functionToRecurse == null) { throw new ArgumentNullException("functionToRecurse", "The function to recurse can not be null"); } _functionToRecurse = functionToRecurse; } public Func<T,T> this[int level] { get { if (level < 1) { throw new ArgumentOutOfRangeException("level", level, "The level of recursion must be greater than 0"); } return this.GetXLevel(level); } } private Func<T, T> GetXLevel(int level) { if (level < 1) { throw new ArgumentOutOfRangeException("level", level, "The level of recursion must be greater than 0"); } if (level == 1) return _functionToRecurse; return _GetXLevel(level - 1, _functionToRecurse); } private Func<T, T> _GetXLevel(int level, Func<T, T> prevFunc) { if (level == 1) return y => prevFunc(_functionToRecurse(y)); return _GetXLevel(level - 1, y => prevFunc(_functionToRecurse(y))); } } Conclusion The great thing about this class, is that it can be used with any function with same input/output parameters.  I strived to find an implementation that I found clean and useful, and I finally settled on this.  If you have feedback – good or bad, I would love to hear it!

    Read the article

  • SOA Suite Integration: Part 3: Loading files

    - by Anthony Shorten
    One of the most common scenarios in SOA Integration is the loading of a file into the product from an external source. In Oracle SOA Suite there is a File Adapter that can process many file types into your BPEL process. For this example I will use the File Adapter to load a file of user and emails to update the user object within the Oracle Utilities Application Framework. Remember you can repeat this process with other objects and other file types. Again I am illustrating the ease of integration. The first thing is to create an empty BPEL process that will hold our flow. In Oracle JDeveloper this can be achieved by specifying the Define Service Later template (as other templates have predefined inputs and outputs and in this case we want to specify those). So I will create simpleFileLoad process to house our process. You will start with an empty canvas so you need to first specify the load part of the process using the File Adapter. Select the File Adapter from the Component Palette under BPEL Services and drag and drop it to the left side Partner Links (left is input). You name the Service. In this case I chose LoadFile. Press Next. We will define the interface as part of the wizard so select Define from operation and schema (specified later). Press Next. We are going to choose Read File to denote that we will read the file and specify the default Operation Name as Read. Press Next. The next step is to tell the Adapter the location of the files, how to process them and what to do with them after they have been processed. I am using hardcoded locations in this example but you can have logical locations as well. Press Next. I am now going to tell the adapter how to recognize the files I want to load. In my case I am using CSV files and more importantly I am tell the adapter to run the process for each record in the file it encounters. Press Next. Now, I tell the adapter how often I want to poll for the files. I have taken the defaults. Press Next. At this stage I have no explanation of the format of the input. So I am going to invoke the Native Format Wizard which will guide me through the process of creating the file input format. Clicking the purple cog icon will start the wizard. After an introduction screen (not shown), you specify the format of the input file. The File Adapter supports multiple format types. For this example, I will use Delimited as I am going to load a CSV file. Press Next. The best way for the wizard to work is with a sample. I have a sample file and the wizard will ask how much of the file to use as a template. I will use the defaults. Note: If you are using a language that has other languages other than US-ASCII, it is at this point you specify the character set to use.  Press Next. The sample contains multiple instances of a single record type. The wizard supports complex types as well. We will use the appropriate setting for our file. Press Next. You have to specify the file element and the record element. This will be used by the input wizard to translate the CSV data into an XML structure (this will make sense later). I am using LoadUsers as my file delimiter (root element) and User Record as my record root element. Press Next. As the file is CSV the delimiter is "," so I will also specify that the End Of Line (EOL) indicator indicates the end of a record. Press Next. Up until this point your have not given the columns their names. In my case my sample includes the column names in the first record. This is not always the case but you can specify the names and formats of columns in this dialog (not shown). Press Next. The wizard now generates the schema for the input file. You can specify a name for the schema. I have used userupdate.xsd. We want to verify the schema so press Test. You can test the schema by specifying an input sample. and pressing the green play button. You will see the delimiters you specified earlier for the file and the records. Press Ok to continue. A confirmation screen will be displayed showing you the location of the schema in your project. Press Finish to return to the File Adapter configuration. You will now see the schema and elements prepopulated from the wizard. Press Next. The File Adapter configuration is now complete. Press Finish. Now you need to receive the input from the LoadFile component so we need to place a Receive node in the BPEL process by drag and dropping the Receive component from the Component Palette under BPEL Constructs onto the BPEL process. We link the receive process with the LoadFile component by dragging the left most connect node of the Receive node to the LoadFile component. Once the link is established you need to name the Receive node appropriately and as in the post of the last part of this series you need to generate input variables for the BPEL process to hold the input records in. You need to now add the product Web Service. The process is the same as described in the post of the last part of this series. You drop the Web Service BPEL Service onto the right side of the process and fill in the details of the WSDL URL . You also have to add an Invoke node to call the service and generate the input and outputs variables for the call in the Invoke node. Now, to get the inputs from File to the service. You have to use a Transform (you can use an Assign action but a Transform action is more flexible). You drag and drop the Transform component from the Component Palette under Oracle Extensions and place it between the Receive and Invoke nodes. We name the Transform Node, Mapper File and associate the source of the mapping the schema from the Receive node and the output will be the input variable from the Invoke node. We now build the transform. We first map the user and email attributes by drag and drop the elements from the left to the right. The reason we needed to use the transform is that we will be telling the AS-User service that we want to issue an update action. Remember when we registered the service we actually used Read as the default. If we do not otherwise inform the service to use the Update action it will use the Read action instead (which is not desired). To specify the update action you need to click on the transactionType node on the right and select Set Text to set the action. You need to specify the transactionType of UPD (for update). The mapping is now complete. The final BPEL process is ready for deployment. You then deploy the BPEL process to the server and to test the service by simply dropping a file, in the same pattern/name as you specified, in the directory you specified in the File Adapter. You will see each record as a separate instance entry in the Fusion Middleware Control console. You can now load files into the product. You can repeat this process for each type of file to process. While this was a simple example it illustrates the method of loading data can be achieved using SOA Suite in conjunction with our products.

    Read the article

  • Making Those PanelBoxes Behave

    - by Duncan Mills
    I have a little problem to solve earlier this week - misbehaving <af:panelBox> components... What do I mean by that? Well here's the scenario, I have a page fragment containing a set of panelBoxes arranged vertically. As it happens, they are stamped out in a loop but that does not really matter. What I want to be able to do is to provide the user with a simple UI to close and open all of the panelBoxes in concert. This could also apply to showDetailHeader and similar items with a disclosed attrubute, but in this case it's good old panelBoxes.  Ok, so the basic solution to this should be self evident. I can set up a suitable scoped managed bean that the panelBoxes all refer to for their disclosed attribute state. Then the open all / close commandButtons in the UI can simply set the state of that bean for all the panelBoxes to pick up via EL on their disclosed attribute. Sound OK? Well that works basically without a hitch, but turns out that there is a slight problem and this is where the framework is attempting to be a little too helpful. The issue is that is the user manually discloses or hides a panelBox then that will override the value that the EL is setting. So for example. I start the page with all panelBoxes collapsed, all set by the EL state I'm storing on the session I manually disclose panelBox no 1. I press the Expand All button - all works as you would hope and all the panelBoxes are now disclosed, including of course panelBox 1 which I just expanded manually. Finally I press the Collapse All button and everything collapses except that first panelBox that I manually disclosed.  The problem is that the component remembers this manual disclosure and that overrides the value provided by the expression. If I change the viewId (navigate away and back) then the panelBox will start to behave again, until of course I touch it again! Now, the more astute amoungst you would think (as I did) Ah, sound like the MDS personalizaton stuff is getting in the way and the solution should simply be to set the dontPersist attribute to disclosed | ALL. Alas this does not fix the issue.  After a little noodling on the best way to approach this I came up with a solution that works well, although if you think of an alternative way do let me know. The principle is simple. In the disclosureListener for the panelBox I take a note of the clientID of the panelBox component that has been touched by the user along with the state. This all gets stored in a Map of Booleans in ViewScope which is keyed by clientID and stores the current disclosed state in the Boolean value.  The listener looks like this (it's held in a request scope backing bean for the page): public void handlePBDisclosureEvent(DisclosureEvent disclosureEvent) { String clientId = disclosureEvent.getComponent().getClientId(FacesContext.getCurrentInstance()); boolean state = disclosureEvent.isExpanded(); pbState.addTouchedPanelBox(clientId, state); } The pbState variable referenced here is a reference to the bean which will hold the state of the panelBoxes that lives in viewScope (recall that everything is re-set when the viewid is changed so keeping this in viewScope is just fine and cleans things up automatically). The addTouchedPanelBox() method looks like this: public void addTouchedPanelBox(String clientId, boolean state) { //create the cache if needed this is just a Map<String,Boolean> if (_touchedPanelBoxState == null) { _touchedPanelBoxState = new HashMap<String, Boolean>(); } // Simply put / replace _touchedPanelBoxState.put(clientId, state); } So that's the first part, we now have a record of every panelBox that the user has touched. So what do we do when the Collapse All or Expand All buttons are pressed? Here we do some JavaScript magic. Basically for each clientID that we have stored away, we issue a client side disclosure event from JavaScript - just as if the user had gone back and changed it manually. So here's the Collapse All button action: public String CloseAllAction() { submitDiscloseOverride(pbState.getTouchedClientIds(true), false); _uiManager.closeAllBoxes(); return null; }  The _uiManager.closeAllBoxes() method is just manipulating the master-state that all of the panelBoxes are bound to using EL. The interesting bit though is the line:  submitDiscloseOverride(pbState.getTouchedClientIds(true), false); To break that down, the first part is a call to that viewScoped state holder to ask for a list of clientIDs that need to be "tweaked": public String getTouchedClientIds(boolean targetState) { StringBuilder sb = new StringBuilder(); if (_touchedPanelBoxState != null && _touchedPanelBoxState.size() > 0) { for (Map.Entry<String, Boolean> entry : _touchedPanelBoxState.entrySet()) { if (entry.getValue() == targetState) { if (sb.length() > 0) { sb.append(','); } sb.append(entry.getKey()); } } } return sb.toString(); } You'll notice that this method only processes those panelBoxes that will be in the wrong state and returns those as a comma separated list. This is then processed by the submitDiscloseOverride() method: private void submitDiscloseOverride(String clientIdList, boolean targetDisclosureState) { if (clientIdList != null && clientIdList.length() > 0) { FacesContext fctx = FacesContext.getCurrentInstance(); StringBuilder script = new StringBuilder(); script.append("overrideDiscloseHandler('"); script.append(clientIdList); script.append("',"); script.append(targetDisclosureState); script.append(");"); Service.getRenderKitService(fctx, ExtendedRenderKitService.class).addScript(fctx, script.toString()); } } This method constructs a JavaScript command to call a routine called overrideDiscloseHandler() in a script attached to the page (using the standard <af:resource> tag). That method parses out the list of clientIDs and sends the correct message to each one: function overrideDiscloseHandler(clientIdList, newState) { AdfLogger.LOGGER.logMessage(AdfLogger.INFO, "Disclosure Hander newState " + newState + " Called with: " + clientIdList); //Parse out the list of clientIds var clientIdArray = clientIdList.split(','); for (var i = 0; i < clientIdArray.length; i++){ var panelBox = flipPanel = AdfPage.PAGE.findComponentByAbsoluteId(clientIdArray[i]); if (panelBox.getComponentType() == "oracle.adf.RichPanelBox"){ panelBox.broadcast(new AdfDisclosureEvent(panelBox, newState)); } }  }  So there you go. You can see how, with a few tweaks the same code could be used for other components with disclosure that might suffer from the same problem, although I'd point out that the behavior I'm working around here us usually desirable. You can download the running example (11.1.2.2) from here. 

    Read the article

  • Azure Task Scheduling Options

    - by charlie.mott
    Currently, the Azure PaaS does not offer a distributed\resilient task scheduling service.  If you do want to host a task scheduling product\solution off-premise (and ideally use Azure), what are your options? PaaS Option 1: Worker Roles Use a worker role to schedule and execute actions at specific time periods.  There are a few frameworks available to assist with this: http://azuretoolkit.codeplex.com https://github.com/Lokad/lokad-cloud/wiki/TaskScheduler http://blog.smarx.com/posts/building-a-task-scheduler-in-windows-azure - This addresses a slightly different set of requirements. It’s a more dynamic approach for queuing up tasks, but not repeatable tasks (e.g. daily). I found the Azure Toolkit option the most simple to implement.  Step 1 : Create a domain entity implementing IJob for each job to schedule.  In this sample, I asynchronously call a WCF service method. 1: namespace Acme.WorkerRole.Jobs 2: { 3: using AzureToolkit; 4: using ScheduledTasksService; 5: 6: public class UploadEmployeesJob : IJob 7: { 8: public void Run() 9: { 10: // Call Tasks Service 11: var client = new ScheduledTasksServiceClient("BasicHttpBinding_IScheduledTasksService"); 12: client.UploadEmployees(); 13: client.Close(); 14: } 15: } 16: } Step 2 : In the worker role run method, add the jobs to the toolkit engine. 1: namespace Acme.WorkerRole 2: { 3: using AzureToolkit.Engine; 4: using Jobs; 5:   6: public class WorkerRole : WorkerRoleEntryPoint 7: { 8: public override void Run() 9: { 10: var engine = new CloudEngine(); 11:   12: // Add Scheduled Jobs (using CronJob syntax - see http://www.adminschoice.com/crontab-quick-reference). 13:   14: // 1. Upload Employee job - 8.00 PM every weekday (Mon-Fri) 15: engine.WithJobScheduler().ScheduleJob<UploadEmployeesJob>(c => { c.CronSchedule = "0 20 * * 1-5"; }); 16: // 2. Purge Data job - 10 AM every Saturday 17: engine.WithJobScheduler().ScheduleJob<PurgeDataJob>(c => { c.CronSchedule = "0 10 * * 6"; }); 18: // 3. Process Exceptions job - Every 5 minutes 19: engine.WithJobScheduler().ScheduleJob<ProcessExceptionsJob>(c => { c.CronSchedule = "*/5 * * * *"; }); 20:   21: engine.Run(); 22: base.Run(); 23: } 24: } 25: } Pros Cons Azure Toolkit option is simple to implement. For the AzureToolkit option, you are limited to a single worker role.  Otherwise, the jobs will be executed multiple times, once for each worker role instance.   Paying for a continuously running worker role, even if it just processes a single job once a week.  If you only have a few scheduled tasks to run calling asynchronous services hosted in different web roles, an extra small worker role likely to be sufficient.  However, for an extra small worker role this still costs $14.40/month (03/09/2012). Option 2: Use Scheduled Task on Azure Web Role calling a console app Setup a Windows Scheduled Task on the Azure Web Role. This calls a console application that calls the WCF service methods that run the task actions. This design is described here: http://www.ronaldwidha.net/2011/02/23/cron-job-on-azure-using-scheduled-task-on-a-web-role-to-replace-azure-worker-role-for-background-job/ http://www.voiceoftech.com/swhitley/index.php/2011/07/windows-azure-task-scheduler/ http://devlicio.us/blogs/vinull/archive/2011/10/23/moving-to-azure-worker-roles-for-nothing-and-tasks-for-free.aspx Pros Cons Fairly easy to implement. Supportability - I RDC’ed onto the Azure server and stopped the scheduled task. I then rebooted the machine and the task was re-started. I also tried deleting the task and rebooting, the same thing occurred. The only way to permanently guarantee that a task is disabled is to do a fresh deployment. I think this is a major supportability concern.   Saleability - multiple instances would trigger multiple tasks. You can only have one instance for the scheduled task web role. The guidance implements setup of the scheduled task as part of a web role instance. But if you have more than one instance in a web role, the task will be triggered multiple times for each scheduled action (once per machine). Workaround: If we wanted to use scheduled tasks for another client with a saleable WCF service, then we could include the console & tasks scripts in a separate web role (e.g. a empty WCF service with no real purpose to it). SaaS Option 3: Azure Marketplace I thought that someone might be offering this type of service via the Azure marketplace. At the point of writing this blog post, I did not find anyone doing so. https://datamarket.azure.com/ Pros Cons   Nobody currently offers this on the Azure Marketplace. Option 4: Online Job Scheduling Service Provider There are plenty of online providers that offer this type of service on a pay-as-you-go approach.  Some of these are free for small usage.   Many of these providers are listed here: http://en.wikipedia.org/wiki/Webcron Pros Cons No bespoke development for scheduler. Reliance on third party. IaaS Option 5: Setup Scheduling Software on Azure IaaS VM’s One of job scheduling software offerings could be installed and configured on Azure VM’s.  A list of software options is listed here: http://en.wikipedia.org/wiki/List_of_job_scheduler_software Pros Cons Enterprise distributed\resilient task scheduling service VM Setup and maintenance   Software Licence Costs Option 6: VM Gallery A the time of writing this blog post, I did not spot a VM in the gallery that included pre-installation of any of the above software options. Pros Cons   No current VM template. Summary For my current project that had a small handful of tasks to schedule with a limited project budget I chose option 1 (a worker role using the Azure Toolkit to schedule tasks).  If I was building an enterprise scale solution for the future, options 4 and 5 are currently worthy of consideration. Hopefully, Microsoft will include tasks scheduling in the future as part of their PaaS offerings.

    Read the article

  • Combining Shared Secret and Certificates

    - by Michael Stephenson
    As discussed in the introduction article this walkthrough will explain how you can implement WCF security with the Windows Azure Service Bus to ensure that you can protect your endpoint in the cloud with a shared secret but also combine this with certificates so that you can identify the sender of the message.   Prerequisites As in the previous article before going into the walk through I want to explain a few assumptions about the scenario we are implementing but to keep the article shorter I am not going to walk through all of the steps in how to setup some of this. In the solution we have a simple console application which will represent the client application. There is also the services WCF application which contains the WCF service we will expose via the Windows Azure Service Bus. The WCF Service application in this example was hosted in IIS 7 on Windows 2008 R2 with AppFabric Server installed and configured to auto-start the WCF listening services. I am not going to go through significant detail around the IIS setup because it should not matter in relation to this article however if you want to understand more about how to configure WCF and IIS for such a scenario please refer to the following paper which goes into a lot of detail about how to configure this. The link is: http://tinyurl.com/8s5nwrz   Setting up the Certificates To keep the post and sample simple I am going to use the local computer store for all certificates but this bit is really just the same as setting up certificates for an example where you are using WCF without using Windows Azure Service Bus. In the sample I have included two batch files which you can use to create the sample certificates or remove them. Basically you will end up with: A certificate called PocServerCert in the personal store for the local computer which will be used by the WCF Service component A certificate called PocClientCert in the personal store for the local computer which will be used by the client application A root certificate in the Root store called PocRootCA with its associated revocation list which is the root from which the client and server certificates were created   For the sample Im just using development certificates like you would normally, and you can see exactly how these are configured and placed in the stores from the batch files in the solution using makecert and certmgr.   The Service Component To begin with let's look at the service component and how it can be configured to listen to the service bus using a shared secret but to also accept a username token from the client. In the sample the service component is called Acme.Azure.ServiceBus.Poc.Cert.Services. It has a single service which is the Visual Studio template for a WCF service when you add a new WCF Service Application so we have a service called Service1 with its Echo method. Nothing special so far!.... The next step is to look at the web.config file to see how we have configured the WCF service. In the services section of the WCF configuration you can see I have created my service and I have created a local endpoint which I simply used to do a little bit of diagnostics and to check it was working, but more importantly there is the Windows Azure endpoint which is using the ws2007HttpRelayBinding (note that this should also work just the same if your using netTcpRelayBinding). The key points to note on the above picture are the service behavior called MyServiceBehaviour and the service bus endpoints behavior called MyEndpointBehaviour. We will go into these in more detail later.   The Relay Binding The relay binding for the service has been configured to use the TransportWithMessageCredential security mode. This is the important bit where the transport security really relates to the interaction between the service and listening to the Azure Service Bus and the message credential is where we will use our certificate like we have specified in the message/clientCrentialType attribute. Note also that we have left the relayClientAuthenticationType set to RelayAccessToken. This means that authentication will be made against ACS for accessing the service bus and messages will not be accepted from any sender who has not been authenticated by ACS.   The Endpoint Behaviour In the below picture you can see the endpoint behavior which is configured to use the shared secret client credential for accessing the service bus and also for diagnostic purposes I have included the service registry element.     Hopefully if you are familiar with using Windows Azure Service Bus relay feature the above is very familiar to you and this is a very common setup for this section. There is nothing specific to the username token implementation here. The Service Behaviour Now we come to the bit with most of the certificate stuff in it. When you configure the service behavior I have included the serviceCredentials element and then setup to use the clientCertificate check and also specifying the serviceCertificate with information on how to find the servers certificate in the store.     I have also added a serviceAuthorization section where I will implement my own authorization component to perform additional security checks after the service has validated that the message was signed with a good certificate. I also have the same serviceSecurityAudit configuration to log access to my service. My Authorization Manager The below picture shows you implementation of my authorization manager. WCF will eventually hand off the message to my authorization component before it calls the service code. This is where I can perform some logic to check if the identity is allowed to access resources. In this case I am simple rejecting messages from anyone except the PocClientCertificate.     The Client Now let's take a look at the client side of this solution and how we can configure the client to authenticate against ACS but also send a certificate over to the service component so it can implement additional security checks on-premise. I have a console application and in the program class I want to use the proxy generated with Add Service Reference to send a message via the Azure Service Bus. You can see in my WCF client configuration below I have setup my details for the azure service bus url and am using the ws2007HttpRelayBinding.   Next is my configuration for the relay binding. You can see below I have configured security to use TransportWithMessageCredential so we will flow the token from a certificate with the message and also the RelayAccessToken relayClientAuthenticationType which means the component will validate against ACS before being allowed to access the relay endpoint to send a message.     After the binding we need to configure the endpoint behavior like in the below picture. This contains the normal transportClientEndpointBehaviour to setup the ACS shared secret configuration but we have also configured the clientCertificate to look for the PocClientCert.     Finally below we have the code of the client in the console application which will call the service bus. You can see that we have created our proxy and then made a normal call to a WCF in exactly the normal way but the configuration will jump in and ensure that a token is passed representing the client certificate.     Conclusion As you can see from the above walkthrough it is not too difficult to configure a service to use both a shared secret and certificate based token at the same time. This gives you the power and protection offered by the access control service in the cloud but also the ability to flow additional tokens to the on-premise component for additional security features to be implemented. Sample The sample used in this post is available at the following location: https://s3.amazonaws.com/CSCBlogSamples/Acme.Azure.ServiceBus.Poc.Cert.zip

    Read the article

  • OpenGL loading functions error [on hold]

    - by Ghilliedrone
    I'm new to OpenGL, and I bought a book on it for beginners. I finished writing the sample code for making a context/window. I get an error on this line at the part PFNWGLCREATECONTEXTATTRIBSARBPROC, saying "Error: expected a ')'": typedef HGLRC(APIENTRYP PFNWGLCREATECONTEXTATTRIBSARBPROC)(HDC, HGLRC, const int*); Replacing it or adding a ")" makes it error, but the error disappears when I use the OpenGL headers included in the books CD, which are OpenGL 3.0. I would like a way to make this work with the newest gl.h/wglext.h and without libraries. Here's the rest of the class if it's needed: #include <ctime> #include <windows.h> #include <iostream> #include <gl\GL.h> #include <gl\wglext.h> #include "Example.h" #include "GLWindow.h" typedef HGLRC(APIENTRYP PFNWGLCREATECONTEXTATTRIBSARBPROC)(HDC, HGLRC, const int*); PFNWGLCREATECONTEXTATTRIBSARBPROC wglCreateContextAttribsARB = NULL; bool GLWindow::create(int width, int height, int bpp, bool fullscreen) { DWORD dwExStyle; //Window Extended Style DWORD dwStyle; //Window Style m_isFullscreen = fullscreen;//Store the fullscreen flag m_windowRect.left = 0L; m_windowRect.right = (long)width; m_windowRect.top = 0L; m_windowRect.bottom = (long)height;//Set bottom to height // fill out the window class structure m_windowClass.cbSize = sizeof(WNDCLASSEX); m_windowClass.style = CS_HREDRAW | CS_VREDRAW; m_windowClass.lpfnWndProc = GLWindow::StaticWndProc; //We set our static method as the event handler m_windowClass.cbClsExtra = 0; m_windowClass.cbWndExtra = 0; m_windowClass.hInstance = m_hinstance; m_windowClass.hIcon = LoadIcon(NULL, IDI_APPLICATION); // default icon m_windowClass.hCursor = LoadCursor(NULL, IDC_ARROW); // default arrow m_windowClass.hbrBackground = NULL; // don't need background m_windowClass.lpszMenuName = NULL; // no menu m_windowClass.lpszClassName = (LPCWSTR)"GLClass"; m_windowClass.hIconSm = LoadIcon(NULL, IDI_WINLOGO); // windows logo small icon if (!RegisterClassEx(&m_windowClass)) { MessageBox(NULL, (LPCWSTR)"Failed to register window class", NULL, MB_OK); return false; } if (m_isFullscreen)//If we are fullscreen, we need to change the display { DEVMODE dmScreenSettings; //Device mode memset(&dmScreenSettings, 0, sizeof(dmScreenSettings)); dmScreenSettings.dmSize = sizeof(dmScreenSettings); dmScreenSettings.dmPelsWidth = width; //Screen width dmScreenSettings.dmPelsHeight = height; //Screen height dmScreenSettings.dmBitsPerPel = bpp; //Bits per pixel dmScreenSettings.dmFields = DM_BITSPERPEL | DM_PELSWIDTH | DM_PELSHEIGHT; if (ChangeDisplaySettings(&dmScreenSettings, CDS_FULLSCREEN) != DISP_CHANGE_SUCCESSFUL) { MessageBox(NULL, (LPCWSTR)"Display mode failed", NULL, MB_OK); m_isFullscreen = false; } } if (m_isFullscreen) //Is it fullscreen? { dwExStyle = WS_EX_APPWINDOW; //Window Extended Style dwStyle = WS_POPUP; //Windows Style ShowCursor(false); //Hide mouse pointer } else { dwExStyle = WS_EX_APPWINDOW | WS_EX_WINDOWEDGE; //Window Exteneded Style dwStyle = WS_OVERLAPPEDWINDOW; //Windows Style } AdjustWindowRectEx(&m_windowRect, dwStyle, false, dwExStyle); //Adjust window to true requested size //Class registered, so now create window m_hwnd = CreateWindowEx(NULL, //Extended Style (LPCWSTR)"GLClass", //Class name (LPCWSTR)"Chapter 2", //App name dwStyle | WS_CLIPCHILDREN | WS_CLIPSIBLINGS, 0, 0, //x, y coordinates m_windowRect.right - m_windowRect.left, m_windowRect.bottom - m_windowRect.top, //Width and height NULL, //Handle to parent NULL, //Handle to menu m_hinstance, //Application instance this); //Pass a pointer to the GLWindow here //Check if window creation failed, hwnd would equal NULL if (!m_hwnd) { return 0; } m_hdc = GetDC(m_hwnd); ShowWindow(m_hwnd, SW_SHOW); UpdateWindow(m_hwnd); m_lastTime = GetTickCount() / 1000.0f; return true; } LRESULT CALLBACK GLWindow::StaticWndProc(HWND hWnd, UINT uMsg, WPARAM wParam, LPARAM lParam) { GLWindow* window = nullptr; //If this is the create message if (uMsg == WM_CREATE) { //Get the pointer we stored during create window = (GLWindow*)((LPCREATESTRUCT)lParam)->lpCreateParams; //Associate the window pointer with the hwnd for the other events to access SetWindowLongPtr(hWnd, GWL_USERDATA, (LONG_PTR)window); } else { //If this is not a creation event, then we should have stored a pointer to the window window = (GLWindow*)GetWindowLongPtr(hWnd, GWL_USERDATA); if (!window) { //Do the default event handling return DefWindowProc(hWnd, uMsg, wParam, lParam); } } //Call our window's member WndProc(allows us to access member variables) return window->WndProc(hWnd, uMsg, wParam, lParam); } LRESULT GLWindow::WndProc(HWND hWnd, UINT uMsg, WPARAM wParam, LPARAM lParam) { switch (uMsg) { case WM_CREATE: { m_hdc = GetDC(hWnd); setupPixelFormat(); //Set the version that we want, in this case 3.0 int attribs[] = { WGL_CONTEXT_MAJOR_VERSION_ARB, 3, WGL_CONTEXT_MINOR_VERSION_ARB, 0, 0}; //Create temporary context so we can get a pointer to the function HGLRC tmpContext = wglCreateContext(m_hdc); //Make the context current wglMakeCurrent(m_hdc, tmpContext); //Get the function pointer wglCreateContextAttribsARB = (PFNWGLCREATECONTEXTATTRIBSARBPROC)wglGetProcAddress("wglCreateContextAttribsARB"); //If this is NULL then OpenGl 3.0 is not supported if (!wglCreateContextAttribsARB) { MessageBox(NULL, (LPCWSTR)"OpenGL 3.0 is not supported", (LPCWSTR)"An error occured", MB_ICONERROR | MB_OK); DestroyWindow(hWnd); return 0; } //Create an OpenGL 3.0 context using the new function m_hglrc = wglCreateContextAttribsARB(m_hdc, 0, attribs); //Delete the temporary context wglDeleteContext(tmpContext); //Make the GL3 context current wglMakeCurrent(m_hdc, m_hglrc); m_isRunning = true; } break; case WM_DESTROY: //Window destroy case WM_CLOSE: //Windows is closing wglMakeCurrent(m_hdc, NULL); wglDeleteContext(m_hglrc); m_isRunning = false; //Stop the main loop PostQuitMessage(0); break; case WM_SIZE: { int height = HIWORD(lParam); //Get height and width int width = LOWORD(lParam); getAttachedExample()->onResize(width, height); //Call the example's resize method } break; case WM_KEYDOWN: if (wParam == VK_ESCAPE) //If the escape key was pressed { DestroyWindow(m_hwnd); } break; default: break; } return DefWindowProc(hWnd, uMsg, wParam, lParam); } void GLWindow::processEvents() { MSG msg; //While there are messages in the queue, store them in msg while (PeekMessage(&msg, NULL, 0, 0, PM_REMOVE)) { //Process the messages TranslateMessage(&msg); DispatchMessage(&msg); } } Here is the header: #pragma once #include <ctime> #include <windows.h> class Example;//Declare our example class class GLWindow { public: GLWindow(HINSTANCE hInstance); //default constructor bool create(int width, int height, int bpp, bool fullscreen); void destroy(); void processEvents(); void attachExample(Example* example); bool isRunning(); //Is the window running? void swapBuffers() { SwapBuffers(m_hdc); } static LRESULT CALLBACK StaticWndProc(HWND wnd, UINT msg, WPARAM wParam, LPARAM lParam); LRESULT CALLBACK WndProc(HWND wnd, UINT msg, WPARAM wParam, LPARAM lParam); float getElapsedSeconds(); private: Example* m_example; //A link to the example program bool m_isRunning; //Is the window still running? bool m_isFullscreen; HWND m_hwnd; //Window handle HGLRC m_hglrc; //Rendering context HDC m_hdc; //Device context RECT m_windowRect; //Window bounds HINSTANCE m_hinstance; //Application instance WNDCLASSEX m_windowClass; void setupPixelFormat(void); Example* getAttachedExample() { return m_example; } float m_lastTime; };

    Read the article

  • SharePoint logging to a list

    - by Norgean
    I recently worked in an environment with several servers. Locating the correct SharePoint log file for error messages, or development trace calls, is cumbersome. And once the solution hit the cloud, it got even worse, as we had no access to the log files at all. Obviously we are not the only ones with this problem, and the current trend seems to be to log to a list. This had become an off-hour project, so rather than do the sensible thing and find a ready-made solution, I decided to do it the hard way. So! Fire up Visual Studio, create yet another empty SharePoint solution, and start to think of some requirements. Easy on/offI want to be able to turn list-logging on and off.Easy loggingFor me, this means being able to use string.Format.Easy filteringLet's have the possibility to add some filtering columns; category and severity, where severity can be "verbose", "warning" or "error". Easy on/off Well, that's easy. Create a new web feature. Add an event receiver, and create the list on activation of the feature. Tear the list down on de-activation. I chose not to create a new content type; I did not feel that it would give me anything extra. I based the list on the generic list - I think a better choice would have been the announcement type. Approximately: public void CreateLog(SPWeb web)         {             var list = web.Lists.TryGetList(LogListName);             if (list == null)             {                 var listGuid = web.Lists.Add(LogListName, "Logging for the masses", SPListTemplateType.GenericList);                 list = web.Lists[listGuid];                 list.Title = LogListTitle;                 list.Update();                 list.Fields.Add(Category, SPFieldType.Text, false);                 var stringColl = new StringCollection();                 stringColl.AddRange(new[]{Error, Information, Verbose});                 list.Fields.Add(Severity, SPFieldType.Choice, true, false, stringColl);                 ModifyDefaultView(list);             }         }Should be self explanatory, but: only create the list if it does not already exist (d'oh). Best practice: create it with a Url-friendly name, and, if necessary, give it a better title. ...because otherwise you'll have to look for a list with a name like "Simple_x0020_Log". I've added a couple of fields; a field for category, and a 'severity'. Both to make it easier to find relevant log messages. Notice that I don't have to call list.Update() after adding the fields - this would cause a nasty error (something along the lines of "List locked by another user"). The function for deleting the log is exactly as onerous as you'd expect:         public void DeleteLog(SPWeb web)         {             var list = web.Lists.TryGetList(LogListTitle);             if (list != null)             {                 list.Delete();             }         } So! "All" that remains is to log. Also known as adding items to a list. Lots of different methods with different signatures end up calling the same function. For example, LogVerbose(web, message) calls LogVerbose(web, null, message) which again calls another method which calls: private static void Log(SPWeb web, string category, string severity, string textformat, params object[] texts)         {             if (web != null)             {                 var list = web.Lists.TryGetList(LogListTitle);                 if (list != null)                 {                     var item = list.AddItem(); // NOTE! NOT list.Items.Add… just don't, mkay?                     var text = string.Format(textformat, texts);                     if (text.Length > 255) // because the title field only holds so many chars. Sigh.                         text = text.Substring(0, 254);                     item[SPBuiltInFieldId.Title] = text;                     item[Degree] = severity;                     item[Category] = category;                     item.Update();                 }             } // omitted: Also log to SharePoint log.         } By adding a params parameter I can call it as if I was doing a Console.WriteLine: LogVerbose(web, "demo", "{0} {1}{2}", "hello", "world", '!'); Ok, that was a silly example, a better one might be: LogError(web, LogCategory, "Exception caught when updating {0}. exception: {1}", listItem.Title, ex); For performance reasons I use list.AddItem rather than list.Items.Add. For completeness' sake, let us include the "ModifyDefaultView" function that I deliberately skipped earlier.         private void ModifyDefaultView(SPList list)         {             // Add fields to default view             var defaultView = list.DefaultView;             var exists = defaultView.ViewFields.Cast<string>().Any(field => String.CompareOrdinal(field, Severity) == 0);               if (!exists)             {                 var field = list.Fields.GetFieldByInternalName(Severity);                 if (field != null)                     defaultView.ViewFields.Add(field);                 field = list.Fields.GetFieldByInternalName(Category);                 if (field != null)                     defaultView.ViewFields.Add(field);                 defaultView.Update();                   var sortDoc = new XmlDocument();                 sortDoc.LoadXml(string.Format("<Query>{0}</Query>", defaultView.Query));                 var orderBy = (XmlElement) sortDoc.SelectSingleNode("//OrderBy");                 if (orderBy != null && sortDoc.DocumentElement != null)                     sortDoc.DocumentElement.RemoveChild(orderBy);                 orderBy = sortDoc.CreateElement("OrderBy");                 sortDoc.DocumentElement.AppendChild(orderBy);                 field = list.Fields[SPBuiltInFieldId.Modified];                 var fieldRef = sortDoc.CreateElement("FieldRef");                 fieldRef.SetAttribute("Name", field.InternalName);                 fieldRef.SetAttribute("Ascending", "FALSE");                 orderBy.AppendChild(fieldRef);                   fieldRef = sortDoc.CreateElement("FieldRef");                 field = list.Fields[SPBuiltInFieldId.ID];                 fieldRef.SetAttribute("Name", field.InternalName);                 fieldRef.SetAttribute("Ascending", "FALSE");                 orderBy.AppendChild(fieldRef);                 defaultView.Query = sortDoc.DocumentElement.InnerXml;                 //defaultView.Query = "<OrderBy><FieldRef Name='Modified' Ascending='FALSE' /><FieldRef Name='ID' Ascending='FALSE' /></OrderBy>";                 defaultView.Update();             }         } First two lines are easy - see if the default view includes the "Severity" column. If it does - quit; our job here is done.Adding "severity" and "Category" to the view is not exactly rocket science. But then? Then we build the sort order query. Through XML. The lines are numerous, but boring. All to achieve the CAML query which is commented out. The major benefit of using the dom to build XML, is that you may get compile time errors for spelling mistakes. I say 'may', because although the compiler will not let you forget to close a tag, it will cheerfully let you spell "Name" as "Naem". Whichever you prefer, at the end of the day the view will sort by modified date and ID, both descending. I added the ID as there may be several items with the same time stamp. So! Simple logging to a list, with sensible a view, and with normal functionality for creating your own filterings. I should probably have added some more views in code, ready filtered for "only errors", "errors and warnings" etc. And it would be nice to block verbose logging completely, but I'm not happy with the alternatives. (yetanotherfeature or an admin page seem like overkill - perhaps just removing it as one of the choices, and not log if it isn't there?) Before you comment - yes, try-catches have been removed for clarity. There is nothing worse than having a logging function that breaks your site!

    Read the article

< Previous Page | 412 413 414 415 416 417 418 419 420 421 422 423  | Next Page >