Search Results

Search found 24037 results on 962 pages for 'every'.

Page 167/962 | < Previous Page | 163 164 165 166 167 168 169 170 171 172 173 174  | Next Page >

  • Issuing Current Time Increments in StreamInsight (A Practical Example)

    The issuing of a Current Time Increment, Cti, in StreamInsight is very definitely one of the most important concepts to learn if you want your Streams to be responsive. A full discussion of how to issue Ctis is beyond the scope of this article but a very good explanation in addition to Books Online can be found in these three articles by a member of the StreamInsight team at Microsoft, Ciprian Gerea. Time in StreamInsight Series http://blogs.msdn.com/b/streaminsight/archive/2010/07/23/time-in-streaminsight-i.aspx http://blogs.msdn.com/b/streaminsight/archive/2010/07/30/time-in-streaminsight-ii.aspx http://blogs.msdn.com/b/streaminsight/archive/2010/08/03/time-in-streaminsight-iii.aspx A lot of the problems I see with unresponsive or stuck streams on the MSDN Forums are to do with how Ctis are enqueued or in a lot of cases not enqueued. If you enqueue events and never enqueue a Cti then StreamInsight will be perfectly happy. You, on the other hand, will never see data on the output as you have not told StreamInsight to flush the stream. This article deals with a specific implementation problem I had recently whilst working on a StreamInsight project. I look at some possible options and discuss why they would not work before showing the way I solved the problem. The stream of data I was dealing with on this project was very bursty that is to say when events were flowing they came through very quickly and in large numbers (1000 events/sec), but when the stream calmed down it could be a few seconds between each event. When enqueuing events into the StreamInsight engne it is best practice to do so with a StartTime that is given to you by the system producing the event . StreamInsight processes events and it doesn't matter whether those events are being pushed into the engine by a source system or the events are being read from something like a flat file in a directory somewhere. You can apply the same logic and temporal algebra to both situations. Reading from a file is an excellent example of where the time of the event on the source itself is very important. We could be reading that file a long time after it was written. Being able to read the StartTime from the events allows us to define windows that will hold the correct sets of events. I was able to do this with my stream but this is where my problems started. Below is a very simple script to create a SQL Server table and populate it with sample data that will show exactly the problem I had. CREATE TABLE [dbo].[t] ( [c1] [int] PRIMARY KEY, [c2] [datetime] NULL ) INSERT t VALUES (1,'20100810'),(2,'20100810'),(3,'20100810') Column c2 defines the StartTime of the event on the source and as you can see the values in all 3 rows of data is the same. If we read Ciprian’s articles we know that we can define how Ctis get injected into the stream in 3 different places The Stream Definition The Input Factory The Input Adapter I personally have always been a fan of enqueing Ctis through the factory. Below is code typical of what I would use to do this On the class itself I do some inheriting public class SimpleInputFactory : ITypedInputAdapterFactory<SimpleInputConfig>, ITypedDeclareAdvanceTimeProperties<SimpleInputConfig> And then I implement the following function public AdapterAdvanceTimeSettings DeclareAdvanceTimeProperties<TPayload>(SimpleInputConfig configInfo, EventShape eventShape) { return new AdapterAdvanceTimeSettings( new AdvanceTimeGenerationSettings(configInfo.CtiFrequency, TimeSpan.FromTicks(-1)), AdvanceTimePolicy.Adjust); } The configInfo .CtiFrequency property is a value I pass through to define after how many events I want a Cti to be injected and this in turn will flush through the stream of data. I usually pass a value of 1 for this setting. The second parameter determines the CTI timestamp in terms of a delay relative to the events. -1 ticks in the past results in 1 tick in the future, i.e., ahead of the event. The problem with this method though is that if consecutive events have the same StartTime then only one of those events will be enqueued. In this example I use the following to define how I assign the StartTime of my events currEvent.StartTime = (DateTimeOffset)dt.c2; If I go ahead and run my StreamInsight process with this configuration i can see on the output adapter that two events have been removed To see this in a little more depth I can use the StreamInsight Debugger and see what happens internally. What is happening here is that the first event arrives and a Cti is injected with a time of 1 tick after the StartTime of that event (Also the EndTime of the event). The second event arrives and it has a StartTime of before the Cti and even though we specified AdvanceTimePolicy.Adjust on the factory we know that a point event can never be adjusted like this and the event is dropped. The same happens for the third event as well (The second and third events get trumped by the Cti). For a more detailed discussion of why this happens look here http://www.sqlis.com/sqlis/post/AdvanceTimePolicy-and-Point-Event-Streams-In-StreamInsight.aspx We end up with a single event being pushed into the output adapter and our result now makes sense. The next way I tried to solve this problem by changing the value of the second parameter to TimeSpan.Zero Here is how my factory code now looks public AdapterAdvanceTimeSettings DeclareAdvanceTimeProperties<TPayload>(SimpleInputConfig configInfo, EventShape eventShape) { return new AdapterAdvanceTimeSettings( new AdvanceTimeGenerationSettings(configInfo.CtiFrequency, TimeSpan.Zero), AdvanceTimePolicy.Adjust); } What I am doing here is declaring a policy that says inject a Cti together with every event and stamp it with a StartTime that is equal to the start time of the event itself (TimeSpan.Zero). This method has plus points as well as a downside. The upside is that no events will be lost by having the same StartTime as previous events. The Downside is that because the Cti is declared with the StartTime of the event itself then it does not actually flush that particular event because in the StreamInsight algebra, a Cti commits only those events that occurred strictly before them. To flush the events we need a Cti to be enqueued with a greater StartTime than the events themselves. Here is what happened when I ran this configuration As you can see all we got through was the Cti and none of the events. The debugger output shows the stamps on the Cti and the events themselves. Because the Cti issued has the same timestamp (StartTime) as the events then none of the events get flushed. I was nearly there but not quite. Because my stream was bursty it was possible that the next event would not come along for a few seconds and this was far too long for an event to be enqueued and not be flushed to the output adapter. I needed another solution. Two possible solutions crossed my mind although only one of them made sense when I explored it some more. Where multiple events have the same StartTime I could add 1 tick to the first event, two to the second, three to third etc thereby giving them unique StartTime values. Add a timer to manually inject Ctis The problem with the first implementation is that I would be giving the events a new StartTime. This would cause me the following problems If I want to define windows over the stream then some events may not be captured in the right windows and therefore any calculations on those windows I did would be wrong What would happen if we had 10,000 events with the same StartTime? I would enqueue them with StartTime + n ticks. Along comes a genuine event with a StartTime of the very first event + 1 tick. It is now too far in the past as far as my stream is concerned and it would be dropped. Not what I would want to do at all. I decided then to look at the Timer based solution I created a timer on my input adapter that elapsed every 200ms. private Timer tmr; public SimpleInputAdapter(SimpleInputConfig configInfo) { ctx = new SimpleTimeExtractDataContext(configInfo.ConnectionString); this.configInfo = configInfo; tmr = new Timer(200); tmr.Elapsed += new ElapsedEventHandler(t_Elapsed); tmr.Enabled = true; } void t_Elapsed(object sender, ElapsedEventArgs e) { ts = DateTime.Now - dtCtiIssued; if (ts.TotalMilliseconds >= 200 && TimerIssuedCti == false) { EnqueueCtiEvent(System.DateTime.Now.AddTicks(-100)); TimerIssuedCti = true; } }   In the t_Elapsed event handler I find out the difference in time between now and when the last event was processed (dtCtiIssued). I then check to see if that is greater than or equal to 200ms and if the last issuing of a Cti was done by the timer or by a genuine event (TimerIssuedCti). If I didn’t do this check then I would enqueue a Cti every time the timer elapsed which is not something I wanted. If the difference between the two times is greater than or equal to 500ms and the last event enqueued was by a real event then I issue a Cti through the timer to flush the event Queue, otherwise I do nothing. When I enqueue the Ctis into my stream in my ProduceEvents method I also set the values of dtCtiIssued and TimerIssuedCti   currEvent = CreateInsertEvent(); currEvent.StartTime = (DateTimeOffset)dt.c2; TimerIssuedCti = false; dtCtiIssued = currEvent.StartTime; If I go ahead and run this configuration I see the following in my output. As we can see the first Cti gets enqueued as before but then another is enqueued by the timer and because this has a later timestamp it flushes the enqueued events through the engine. Conclusion Hopefully this has shown how the enqueuing of Ctis can have a dramatic effect on the responsiveness of your output in StreamInsight. Understanding the temporal nature of the product is for me one of the most important things you can learn. I have attached my solution for the demos. It is all in one project and testing each variation is a simple matter of commenting and un-commenting the parts in the code we have been dealing with here.

    Read the article

  • Project Navigation and File Nesting in ASP.NET MVC Projects

    - by Rick Strahl
    More and more I’m finding myself getting lost in the files in some of my larger Web projects. There’s so much freaking content to deal with – HTML Views, several derived CSS pages, page level CSS, script libraries, application wide scripts and page specific script files etc. etc. Thankfully I use Resharper and the Ctrl-T Go to Anything which autocompletes you to any file, type, member rapidly. Awesome except when I forget – or when I’m not quite sure of the name of what I’m looking for. Project navigation is still important. Sometimes while working on a project I seem to have 30 or more files open and trying to locate another new file to open in the solution often ends up being a mental exercise – “where did I put that thing?” It’s those little hesitations that tend to get in the way of workflow frequently. To make things worse most NuGet packages for client side frameworks and scripts, dump stuff into folders that I generally don’t use. I’ve never been a fan of the ‘Content’ folder in MVC which is just an empty layer that doesn’t serve much of a purpose. It’s usually the first thing I nuke in every MVC project. To me the project root is where the actual content for a site goes – is there really a need to add another folder to force another path into every resource you use? It’s ugly and also inefficient as it adds additional bytes to every resource link you embed into a page. Alternatives I’ve been playing around with different folder layouts recently and found that moving my cheese around has actually made project navigation much easier. In this post I show a couple of things I’ve found useful and maybe you find some of these useful as well or at least get some ideas what can be changed to provide better project flow. The first thing I’ve been doing is add a root Code folder and putting all server code into that. I’m a big fan of treating the Web project root folder as my Web root folder so all content comes from the root without unneeded nesting like the Content folder. By moving all server code out of the root tree (except for Code) the root tree becomes a lot cleaner immediately as you remove Controllers, App_Start, Models etc. and move them underneath Code. Yes this adds another folder level for server code, but it leaves only code related things in one place that’s easier to jump back and forth in. Additionally I find myself doing a lot less with server side code these days, more with client side code so I want the server code separated from that. The root folder itself then serves as the root content folder. Specifically I have the Views folder below it, as well as the Css and Scripts folders which serve to hold only common libraries and global CSS and Scripts code. These days of building SPA style application, I also tend to have an App folder there where I keep my application specific JavaScript files, as well as HTML View templates for client SPA apps like Angular. Here’s an example of what this looks like in a relatively small project: The goal is to keep things that are related together, so I don’t end up jumping around so much in the solution to get to specific project items. The Code folder may irk some of you and hark back to the days of the App_Code folder in non Web-Application projects, but these days I find myself messing with a lot less server side code and much more with client side files – HTML, CSS and JavaScript. Generally I work on a single controller at a time – once that’s open it’s open that’s typically the only server code I work with regularily. Business logic lives in another project altogether, so other than the controller and maybe ViewModels there’s not a lot of code being accessed in the Code folder. So throwing that off the root and isolating seems like an easy win. Nesting Page specific content In a lot of my existing applications that are pure server side MVC application perhaps with some JavaScript associated with them , I tend to have page level javascript and css files. For these types of pages I actually prefer the local files stored in the same folder as the parent view. So typically I have a .css and .js files with the same name as the view in the same folder. This looks something like this: In order for this to work you have to also make a configuration change inside of the /Views/web.config file, as the Views folder is blocked with the BlockViewHandler that prohibits access to content from that folder. It’s easy to fix by changing the path from * to *.cshtml or *.vbhtml so that view retrieval is blocked:<system.webServer> <handlers> <remove name="BlockViewHandler"/> <add name="BlockViewHandler" path="*.cshtml" verb="*" preCondition="integratedMode" type="System.Web.HttpNotFoundHandler" /> </handlers> </system.webServer> With this in place, from inside of your Views you can then reference those same resources like this:<link href="~/Views/Admin/QuizPrognosisItems.css" rel="stylesheet" /> and<script src="~/Views/Admin/QuizPrognosisItems.js"></script> which works fine. JavaScript and CSS files in the Views folder deploy just like the .cshtml files do and can be referenced from this folder as well. Making this happen is not really as straightforward as it should be with just Visual Studio unfortunately, as there’s no easy way to get the file nesting from the VS IDE directly (you have to modify the .csproj file). However, Mads Kristensen has a nice Visual Studio Add-in that provides file nesting via a short cut menu option. Using this you can select each of the ‘child’ files and then nest them under a parent file. In the case above I select the .js and .css files and nest them underneath the .cshtml view. I was even toying with the idea of throwing the controller.cs files into the Views folder, but that’s maybe going a little too far :-) It would work however as Visual Studio doesn’t publish .cs files and the compiler doesn’t care where the files live. There are lots of options and if you think that would make life easier it’s another option to help group related things together. Are there any downside to this? Possibly – if you’re using automated minification/packaging tools like ASP.NET Bundling or Grunt/Gulp with Uglify, it becomes a little harder to group script and css files for minification as you may end up looking in multiple folders instead of a single folder. But – again that’s a one time configuration step that’s easily handled and much less intrusive then constantly having to search for files in your project. Client Side Folders The particular project shown above in the screen shots above is a traditional server side ASP.NET MVC application with most content rendered into server side Razor pages. There’s a fair amount of client side stuff happening on these pages as well – specifically several of these pages are self contained single page Angular applications that deal with 1 or maybe 2 separate views and the layout I’ve shown above really focuses on the server side aspect where there are Razor views with related script and css resources. For applications that are more client centric and have a lot more script and HTML template based content I tend to use the same layout for the server components, but the client side code can often be broken out differently. In SPA type applications I tend to follow the App folder approach where all the application pieces that make the SPA applications end up below the App folder. Here’s what that looks like for me – here this is an AngularJs project: In this case the App folder holds both the application specific js files, and the partial HTML views that get loaded into this single SPA page application. In this particular Angular SPA application that has controllers linked to particular partial views, I prefer to keep the script files that are associated with the views – Angular Js Controllers in this case – with the actual partials. Again I like the proximity of the view with the main code associated with the view, because 90% of the UI application code that gets written is handled between these two files. This approach works well, but only if controllers are fairly closely aligned with the partials. If you have many smaller sub-controllers or lots of directives where the alignment between views and code is more segmented this approach starts falling apart and you’ll probably be better off with separate folders in js folder. Following Angular conventions you’d have controllers/directives/services etc. folders. Please note that I’m not saying any of these ways are right or wrong  – this is just what has worked for me and why! Skipping Project Navigation altogether with Resharper I’ve talked a bit about project navigation in the project tree, which is a common way to navigate and which we all use at least some of the time, but if you use a tool like Resharper – which has Ctrl-T to jump to anything, you can quickly navigate with a shortcut key and autocomplete search. Here’s what Resharper’s jump to anything looks like: Resharper’s Goto Anything box lets you type and quick search over files, classes and members of the entire solution which is a very fast and powerful way to find what you’re looking for in your project, by passing the solution explorer altogether. As long as you remember to use (which I sometimes don’t) and you know what you’re looking for it’s by far the quickest way to find things in a project. It’s a shame that this sort of a simple search interface isn’t part of the native Visual Studio IDE. Work how you like to work Ultimately it all comes down to workflow and how you like to work, and what makes *you* more productive. Following pre-defined patterns is great for consistency, as long as they don’t get in the way you work. A lot of the default folder structures in Visual Studio for ASP.NET MVC were defined when things were done differently. These days we’re dealing with a lot more diverse project content than when ASP.NET MVC was originally introduced and project organization definitely is something that can get in the way if it doesn’t fit your workflow. So take a look and see what works well and what might benefit from organizing files differently. As so many things with ASP.NET, as things evolve and tend to get more complex I’ve found that I end up fighting some of the conventions. The good news is that you don’t have to follow the conventions and you have the freedom to do just about anything that works for you. Even though what I’ve shown here diverges from conventions, I don’t think anybody would stumble over these relatively minor changes and not immediately figure out where things live, even in larger projects. But nevertheless think long and hard before breaking those conventions – if there isn’t a good reason to break them or the changes don’t provide improved workflow then it’s not worth it. Break the rules, but only if there’s a quantifiable benefit. You may not agree with how I’ve chosen to divert from the standard project structures in this article, but maybe it gives you some ideas of how you can mix things up to make your existing project flow a little nicer and make it easier to navigate for your environment. © Rick Strahl, West Wind Technologies, 2005-2014Posted in ASP.NET  MVC   Tweet !function(d,s,id){var js,fjs=d.getElementsByTagName(s)[0];if(!d.getElementById(id)){js=d.createElement(s);js.id=id;js.src="//platform.twitter.com/widgets.js";fjs.parentNode.insertBefore(js,fjs);}}(document,"script","twitter-wjs"); (function() { var po = document.createElement('script'); po.type = 'text/javascript'; po.async = true; po.src = 'https://apis.google.com/js/plusone.js'; var s = document.getElementsByTagName('script')[0]; s.parentNode.insertBefore(po, s); })();

    Read the article

  • Lightning talk: Coderetreat

    - by Michael Williamson
    In the spirit of trying to encourage more deliberate practice amongst coders in Red Gate, Lauri Pesonen had the idea of running a coderetreat in Red Gate. Lauri and I ran the first one a few weeks ago: given that neither of us hadn’t even been to a coderetreat before, let alone run one, I think it turned out quite well. The participants gave positive feedback, saying that they enjoyed the day, wrote some thought-provoking code and would do it again. Sam Blackburn was one of the attendees, and gave a lightning talk to the other developers in one of our regular lightning talk sessions: In case you can’t watch the video, I’ve transcribed the talk below, although I’d recommend watching the video if you can — I didn’t have much time to do the transcribing! So, what is a coderetreat? So it’s not just something in Red Gate, there’s a website and everything, although it’s not a very big website. It calls itself a community network. The basic ideas behind coderetreat are: you’ve got one day, and you split it into one hour sections. You spend three quarters of that coding, and do a little retrospective at the end. You’re supposed to start fresh each, we were told to delete our code after every session. We were in pairs, swapping after each session, and we did the same task every time. In fact, Conway’s Game of Life is the only task mentioned anywhere that I find for coderetreat. So I don’t know what we’ll do next time, or if we’re meant to do the same thing again. There are some guiding principles which felt to us like restrictions, that you have to code in crazy ways to encourage better code. Final thing is that it’s supposed to be free for outsiders to join. It’s meant to be a kind of networking thing, where you link up with people from other companies. We had a pilot day with Michael and Lauri. Since it was basically the first time any of us had done anything like this, everybody was from Red Gate. We didn’t chat to anybody else for the initial one. The task was Conway’s Game of Life, which most of you have probably heard of it, all but one of us knew about it when did the coderetreat. I won’t got into the details of what it is, but it felt like the right size of task, basically one or two groups actually produced something working by the end of the day, and of course that doesn’t mean it’s necessarily a day’s work to produce that because we were starting again every hour. The task really drives you more than trying to create good code, I found. It was really tempting to try and get it working rather than stick to the rules. But it’s really good to stop and try again because there are so many what-ifs when you’ve finished writing something, “what if I’d done it this way?”. You can answer all those questions at a coderetreat because it’s not about getting a product out the door, it’s about learning and playing with ideas. So we had all these different practices we were trying. I’ll try and go through most of these. Single responsibility is this idea that everything should do just one thing. It was the very first session, we were still trying to figure out how do you go about the Game of Life? So by the end of forty-five minutes hadn’t produced very much for that first session. We were still thinking, “Do we start with a board, how do we represent all these squares? It can be infinitely big, help, this is getting really difficult!”. So, most of us didn’t really get anywhere on the first one. Although it was interesting that some people started with the board, one group started with the FateDecider class that decides whether things live or die. A sort of god class, but in a good way. They managed to implement all of the rules without even defining how the squares were arranged or anything like that. Another thing we tried was TDD (test-driven development). I’m sure most of you know what TDD is: Watch a test, watch it fail for the right reason Write code to pass the test, watch it pass Refactor, check the test still passes Repeat! It basically worked, we were able to produce code, but we often found the tests defined the direction that code went, which is obviously the idea of TDD. But you tend to find that by the time you’ve even written your first assertion, which is supposed to be the very first thing you write, because you write your tests backwards from the assertions back to the initial conditions, you’ve already constrained the logic of the code in some way by the time you’ve done that. You then get to this situation of, “Well, we actually want to go in a slightly different direction. Can we do this?”. Can we write tests that don’t constrain the architecture? Wrapping up all primitives: it’s kind of turtles all the way down. We had a Size, which has a Width and Height, which both derive from Dimension. You’ve got pages of code before you’ve even done anything. No getters and setters (use tell don’t ask instead): mocks and stubs for tests are required if you want to assert that your results are what you think they should be. You can’t just check the internal state of the code. And people found that really challenging and it made them think in a different way which I think is really good. Not having mutable state: that was kind of confusing because we weren’t quite sure what fitted within that rule and what didn’t, and I think we were trying too hard to follow the rule rather than the guideline. No if-statements: supposed to use polymorphism instead, but polymorphism still requires a factory with conditional behaviour. We did something really crazy to get around this: public T If(bool condition, Func<T> left, Func<T> right) { var dict = new Dictionary<bool, Func<T>> {{true, left}, {false, right}}; return dict[condition].Invoke(); } That is not really polymorphism, is it? For-loops: you can always replace a for-loop with recursion, but it doesn’t tend to make it any more readable unless it’s the kind of task that really lends itself to that. So it was interesting, it was good practice, but it wouldn’t make it easier it’s the kind of tree-structure algorithm where that would help. Having a limit on the number of levels of indentation: again, I think it does produce very nice, clean code, but it wasn’t actually a challenge because you just extract methods. That’s quite a useful thing because you can apply that to real code and say, “Okay, should this method really be going crazy like this?” No talking: we hated that. It’s like there’s two of you at a computer, and one of you is doing the typing, what does the other guy do if they’re not allowed to talk. The answer is TDD ping-pong – one person writes the tests, and then the other person writes the code to pass the test. And that creates communication without actually having to have discussion about things which is kind of cool. No code comments: just makes no difference to anything. It’s a forty-five minute exercise, so what are you going to put comments in code for? Finally, this is my fault. I discovered an entertaining way of doing the calculation that was kind of cool (using convolutions over the state of the board). Unfortunately, it turns out to be really hard to implement in C#, so didn’t even manage to work out how to do that convolution in C#. It’s trivial in some high-level languages, but you need something matrix-orientated for it to really work. That’s most of it, really. The thoughts that people went away with: we put down our answers to questions like “What have you learnt?” and “What surprised you?”, “How are you going to do things differently?”, and most people said redoing the problem is really, really good for understanding it properly. People hate having a massive legacy codebase that they can’t change, so being able to attack something three different ways in an environment where the end-product isn’t important: that’s something people really enjoyed. Pair-programming: also people said that they wanted to do more of that, especially with TDD ping-pong, where you write the test and somebody else writes the code. Various people thought different things about immutables, but most people thought they were good, they promote functional programming. And TDD people found really hard. “Tell, don’t ask” people found really, really hard and really, really, really hard to do well. And the recursion just made things trickier to debug. But most people agreed that coderetreats are really cool, and we should do more of them.

    Read the article

  • Ingredient Substitutes while Baking

    - by Rekha
    In our normal cooking, we substitute the vegetables for the gravies we prepare. When we start baking, we look for a good recipe. At least one or two ingredient will be missing. We do not know where to substitute what to bring same output. So we finally drop the plan of baking. Again after a month, we get the interest in baking. Again one or two lack of ingredient and that’s it. We keep on doing this for months. When I was going through the cooking blogs, I came across a site with the Ingredient Substitutes for Baking: (*) is to indicate that this substitution is ideal from personal experience. Flour Substitutes ( For 1 cup of Flour) All Purpose Flour 1/2 cup white cake flour plus 1/2 cup whole wheat flour 1 cup self-rising flour (omit using salt and baking powder if the recipe calls for it since self raising flour has it already) 1 cup plus 2 tablespoons cake flour 1/2 cup (75 grams) whole wheat flour 7/8 cup (130 grams) rice flour (starch) (do not replace all of the flour with the rice flour) 7/8 cup whole wheat Bread Flour 1 cup all purpose flour 1 cup all purpose flour plus 1 teaspoon wheat gluten (*) Cake Flour Place 2 tbsp cornstarch in 1 cup and fill the rest up with All Purpose flour (*) 1 cup all purpose flour minus 2 tablespoons Pastry flour Place 2 tbsp cornstarch in 1 cup and fill the rest up with All Purpose flour Equal parts of All purpose flour plus cake flour (*) Self-rising Flour 1½ teaspoons of baking powder plus ½ teaspoon of salt plus 1 cup of all-purpose flour. Cornstarch (1 tbsp) 2 tablespoons all-purpose flour 1 tablespoon arrowroot 4 teaspoons quick-cooking tapioca 1 tablespoon potato starch or rice starch or flour Tapioca (1 tbsp) 1 – 1/2 tablespoons all-purpose flour Cornmeal (stone ground) polenta OR corn flour (gives baked goods a lighter texture) if using cornmeal for breading,crush corn chips in a blender until they have the consistency of cornmeal. maize meal Corn grits Sweeteners ( for Every 1 cup ) * * (HV) denotes Healthy Version for low fat or fat free substitution in Baking Light Brown Sugar 2 tablespoons molasses plus 1 cup of white sugar Dark Brown Sugar 3 tablespoons molasses plus 1 cup of white sugar Confectioner’s/Powdered Sugar Process 1 cup sugar plus 1 tablespoon cornstarch Corn Syrup 1 cup sugar plus 1/4 cup water 1 cup Golden Syrup 1 cup honey (may be little sweeter) 1 cup molasses Golden Syrup Combine two parts light corn syrup plus one part molasses 1/2 cup honey plus 1/2 cup corn syrup 1 cup maple syrup 1 cup corn syrup Honey 1- 1/4 cups sugar plus 1/4 cup water 3/4 cup maple syrup plus 1/2 cup granulated sugar 3/4 cup corn syrup plus 1/2 cup granulated sugar 3/4 cup light molasses plus 1/2 cup granulated white sugar 1 1/4 cups granulated white or brown sugar plus 1/4 cup additional liquid in recipe plus 1/2 teaspoon cream of tartar Maple Syrup 1 cup honey,thinned with water or fruit juice like apple 3/4 cup corn syrup plus 1/4 cup butter 1 cup Brown Rice Syrup 1 cup Brown sugar (in case of cereals) 1 cup light molasses (on pancakes, cereals etc) 1 cup granulated sugar for every 3/4 cup of maple syrup and increase liquid in the recipe by 3 tbsp for every cup of sugar.If baking soda is used, decrease the amount by 1/4 teaspoon per cup of sugar substituted, since sugar is less acidic than maple syrup Molasses 1 cup honey 1 cup dark corn syrup 1 cup maple syrup 3/4 cup brown sugar warmed and dissolved in 1/4 cup of liquid ( use this if taste of molasses is important in the baked good) Cocoa Powder (Natural, Unsweetened) 3 tablespoons (20 grams) Dutch-processed cocoa plus 1/8 teaspoon cream of tartar, lemon juice or white vinegar 1 ounce (30 grams) unsweetened chocolate (reduce fat in recipe by 1 tablespoon) 3 tablespoons (20 grams) carob powder Semisweet baking chocolate (1 oz) 1 oz unsweetened baking chocolate plus 1 Tbsp sugar Unsweetened baking chocolate (1 oz ) 3 Tbsp baking cocoa plus 1 Tbsp vegetable oil or melted shortening or margarine Semisweet chocolate chips (1 cup) 6 oz semisweet baking chocolate, chopped (Alternatively) For 1 cup of Semi sweet chocolate chips you can use : 6 tablespoons unsweetened cocoa powder, 7 tablespoons sugar ,1/4 cup fat (butter or oil) Leaveners and Diary * * (HV) denotes Healthy Version for low fat or fat free substitution in Baking Compressed Yeast (1 cake) 1 envelope or 2 teaspoons active dry yeast 1 packet (1/4 ounce) Active Dry yeast 1 cake fresh compressed yeast 1 tablespoon fast-rising active yeast Baking Powder (1 tsp) 1/3 teaspoon baking soda plus 1/2 teaspoon cream of tartar 1/2 teaspoon baking soda plus 1/2 cup buttermilk or plain yogurt 1/4 teaspoon baking soda plus 1/3 cup molasses. When using the substitutions that include liquid, reduce other liquid in recipe accordingly Baking Soda(1 tsp) 3 tsp Baking Powder ( and reduce the acidic ingredients in the recipe. Ex Instead of buttermilk add milk) 1 tsp potassium bicarbonate Ideal substitution – 2 tsp Baking powder and omit salt in recipe Cream of tartar (1 tsp) 1 teaspoon white vinegar 1 tsp lemon juice Notes from What’s Cooking America – If cream of tartar is used along with baking soda in a cake or cookie recipe, omit both and use baking powder instead. If it calls for baking soda and cream of tarter, just use baking powder.Normally, when cream of tartar is used in a cookie, it is used together with baking soda. The two of them combined work like double-acting baking powder. When substituting for cream of tartar, you must also substitute for the baking soda. If your recipe calls for baking soda and cream of tarter, just use baking powder. One teaspoon baking powder is equivalent to 1/4 teaspoon baking soda plus 5/8 teaspoon cream of tartar. If there is additional baking soda that does not fit into the equation, simply add it to the batter. Buttermilk (1 cup) 1 tablespoon lemon juice or vinegar (white or cider) plus enough milk to make 1 cup (let stand 5-10 minutes) 1 cup plain or low fat yogurt 1 cup sour cream 1 cup water plus 1/4 cup buttermilk powder 1 cup milk plus 1 1/2 – 1 3/4 teaspoons cream of tartar Plain Yogurt (1 cup) 1 cup sour cream 1 cup buttermilk 1 cup crème fraiche 1 cup heavy whipping cream (35% butterfat) plus 1 tablespoon freshly squeezed lemon juice Whole Milk (1 cup) 1 cup fat free milk plus 1 tbsp unsaturated Oil like canola (HV) 1 cup low fat milk (HV) Heavy Cream (1 cup) 3/4 cup milk plus 1/3 cup melted butter.(whipping wont work) Sour Cream (1 cup) (pls refer also Substitutes for Fats in Baking below) 7/8 cup buttermilk or sour milk plus 3 tablespoons butter. 1 cup thickened yogurt plus 1 teaspoon baking soda. 3/4 cup sour milk plus 1/3 cup butter. 3/4 cup buttermilk plus 1/3 cup butter. Cooked sauces: 1 cup yogurt plus 1 tablespoon flour plus 2 teaspoons water. Cooked sauces: 1 cup evaporated milk plus 1 tablespoon vinegar or lemon juice. Let stand 5 minutes to thicken. Dips: 1 cup yogurt (drain through a cheesecloth-lined sieve for 30 minutes in the refrigerator for a thicker texture). Dips: 1 cup cottage cheese plus 1/4 cup yogurt or buttermilk, briefly whirled in a blender. Dips: 6 ounces cream cheese plus 3 tablespoons milk,briefly whirled in a blender. Lower fat: 1 cup low-fat cottage cheese plus 1 tablespoon lemon juice plus 2 tablespoons skim milk, whipped until smooth in a blender. Lower fat: 1 can chilled evaporated milk whipped with 1 teaspoon lemon juice. 1 cup plain yogurt plus 1 tablespoon cornstarch 1 cup plain nonfat yogurt Substitutes for Fats in Baking * * (HV) denoted Healthy Version for low fat or fat free substitution in Baking Butter (1 cup) 1 cup trans-free vegetable shortening 3/4 cups of vegetable oil (example. Canola oil) Fruit purees (example- applesauce, pureed prunes, baby-food fruits). Add it along with some vegetable oil and reduce any other sweeteners needed in the recipe since fruit purees are already sweet. 1 cup polyunsaturated margarine (HV) 3/4 cup polyunsaturated oil like safflower oil (HV) 1 cup mild olive oil (not extra virgin)(HV) Note: Butter creates the flakiness and the richness which an oil/purees cant provide. If you don’t want to compromise that much to taste, replace half the butter with the substitutions. Shortening(1 cup) 1 cup polyunsaturated margarine like Earth Balance or Smart Balance(HV) 1 cup + 2tbsp Butter ( better tasting than shortening but more expensive and has cholesterol and a higher level of saturated fat; makes cookies less crunchy, bread crusts more crispy) 1 cup + 2 tbsp Margarine (better tasting than shortening but more expensive; makes cookies less crunchy, bread crusts tougher) 1 Cup – 2tbsp Lard (Has cholesterol and a higher level of saturated fat) Oil equal amount of apple sauce stiffly beaten egg whites into batter equal parts mashed banana equal parts yogurt prune puree grated raw zucchini or seeds removed if cooked. Works well in quick breads/muffins/coffee cakes and does not alter taste pumpkin puree (if the recipe can handle the taste change) Low fat cottage cheese (use only half of the required fat in the recipe). Can give rubbery texture to the end result Silken Tofu – (use only half of the required fat in the recipe). Can give rubbery texture to the end result Equal parts of fruit juice Note: Fruit purees can alter the taste of the final product is used in large quantities. Cream Cheese (1 cup) 4 tbsps. margarine plus 1 cup low-fat cottage cheese – blended. Add few teaspoons of fat-free milk if needed (HV) Heavy Cream (1 cup) 1 cup evaporated skim milk (or full fat milk) 1/2 cup low fat Yogurt plus 1/2 low fat Cottage Cheese (HV) 1/2 cup Yogurt plus 1/2 Cottage Cheese Sour Cream (1 cup) 1 cup plain yogurt (HV) 3/4 cup buttermilk or plain yogurt plus 1/3 cup melted butter 1 cup crème fraiche 1 tablespoon lemon juice or vinegar plus enough whole milk to fill 1 cup (let stand 5-10 minutes) 1/2 cup low-fat cottage cheese plus 1/2 cup low-fat or nonfat yogurt (HV) 1 cup fat-free sour cream (HV) Note: How to Make Maple Syrup Substitute at home For 1 Cup Maple Syrup 1/2 cup granulated sugar 1 cup brown sugar, firmly packed 1 cup boiling water 1 teaspoon butter 1 teaspoon maple extract or vanilla extract Method In a heavy saucepan, place the granulated sugar and keep stirring until it melts and turns slightly brown. Alternatively in another pan, place brown sugar and water and bring to a boil without stirring. Now mix both the sugars and simmer in low heat until they come together as one thick syrup. Remove from heat, add butter and the extract. Use this in place of maple syrup. Store it in a fridge in an air tight container. Even though this was posted in their site long back, I found it helpful. So posting it for you. via chefinyou . cc image credit: flickr/zetrules

    Read the article

  • Dotfuscator Deep Dive with WP7

    - by Bil Simser
    I thought I would share some experience with code obfuscation (specifically the Dotfuscator product) and Windows Phone 7 apps. These days twitter is a buzz with black hat and white operations coming out about how the marketplace is insecure and Microsoft failed, blah, blah, blah. So it’s that much more important to protect your intellectual property. You should protect it no matter what when releasing apps into the wild but more so when someone is paying for them. You want to protect the time and effort that went into your code and have some comfort that the casual hacker isn’t going to usurp your next best thing. Enter code obfuscation. Code obfuscation is one tool that can help protect your IP. Basically it goes into your compiled assemblies, rewrites things at an IL level (like renaming methods and classes and hiding logic flow) and rewrites it back so that the assembly or executable is still fully functional but prying eyes using a tool like ILDASM or Reflector can’t see what’s going on.  You can read more about code obfuscation here on Wikipedia. A word to the wise. Code obfuscation isn’t 100% secure. More so on the WP7 platform where the OS expects certain things to be as they were meant to be. So don’t expect 100% obfuscation of every class and every method and every property. It’s just not going to happen. What this does do is give you some level of protection but don’t put all your eggs in one basket and call it done. Like I said, this is just one step in the process. There are a few tools out there that provide code obfuscation and support the Windows Phone 7 platform (see links to other tools at the end of this post). One such tool is Dotfuscator from PreEmptive solutions. The thing about Dotfuscator is that they’ve struck a deal with Microsoft to provide a *free* copy of their commercial product for Windows Phone 7. The only drawback is that it only runs until March 31, 2010. However it’s a good place to start and the focus of this article. Getting Started When you fire up Dotfuscator you’re presented with a dialog to start a new project or load a previous one. We’ll start with a new project. You’re then looking at a somewhat blank screen that shows an Input tab (among others) and you’re probably wondering what to do? Click on the folder icon (first one) and browse to where your xap file is. At this point you can save the project and click on the arrow to start the process. Bam! You’re done. Right? Think again. The program did indeed run and create a new version of your xap (doing it’s thing and rewriting back your *obfuscated* assemblies) but let’s take a look at the assembly in Reflector to see the end result. Remember a xap file is really just a glorified zip file (or cab file if you prefer). When you ran Dotfuscator for the first time with the default settings you’ll see it created a new version of your xap in a folder under “My Documents” called “Dotfuscated” (you can configure the output directory in settings). Here’s the new xap file. Since a xap is just a zip, rename it to .cab or .zip or something and open it with your favorite unarchive program (I use WinRar but it doesn’t matter as long as it can unzip files). If you already have the xap file associated with your unarchive tool the rename isn’t needed. Once renamed extract the contents of the xap to your hard drive: Now you’ll have a folder with the contents of the xap file extracted: Double click or load up your assembly (WindowsPhoneDataBoundApplication1.dll in the example) in Reflector and let’s see the results: Hmm. That doesn’t look right. I can see all the methods and the code is all there for my LoadData method I wanted to protect. Product failure. Let’s return it for a refund. Hold your horses. We need to check out the settings in the program first. Remember when we loaded up our xap file. It started us on the Input tab but there was a settings tab before that. Wonder what it does? Here’s the default settings: Renaming Taking a closer look, all of the settings in Feature are disabled. WTF? Yeah, it leaves me scratching my head why an obfuscator by default doesn’t obfuscate. However it’s a simple fix to change these settings. Let’s enable Renaming as it sounds like a good start. Renaming obscures code by renaming methods and fields to names that are not understandable. Great. Run the tool again and go through the process of unzipping the updated xap and let’s take a look in Reflector again at our project. This looks a lot better. Lots of methods named a, b, c, d, etc. That’ll help slow hackers down a bit. What about our logic that we spent days weeks on? Let’s take a look at the LoadData method: What gives? We have renaming enabled but all of our code is still there. If you look through all your methods you’ll find it’s still sitting there out in the open. Control Flow Back to the settings page again. Let’s enable Control Flow now. Control Flow obfuscation synthesizes branching, conditional, and iterative constructs (such as if, for, and while) that produce valid executable logic, but yield non-deterministic semantic results when decompilation is attempted. In other words, the code runs as before, but decompilers cannot reproduce the original code. Do the dance again and let’s see the results in Reflector. Ahh, that’s better. Methods renamed *and* nobody can look at our LoadData method now. Life is good. More than Minimum This is the bare minimum to obfuscate your xap to at least a somewhat comfortable level. However I did find that while this worked in my Hello World demo, it didn’t work on one of my real world apps. I had to do some extra tweaking with that. Below are the screens that I used on one app that worked. I’m not sure what it was about the app that the approach above didn’t work with (maybe the extra assembly?) but it works and I’m happy with it. YMMV. Remember to test your obfuscated app on your device first before submitting to ensure you haven’t obfuscated the obfuscator. settings tab: rename tab: string encryption tab: premark tab: A few final notes Play with the settings and keep bumping up the bar to try to get as much obfuscation as you can. The more the better but remember you can overdo it. Always (always, always, always) deploy your obfuscated xap to your device and test it before submitting to the marketplace. I didn’t and got rejected because I had gone overboard with the obfuscation so the app wouldn’t launch at all. Not everything is going to be obfuscated. Specifically I don’t see a way to obfuscate auto properties and a few other language features. Again, if you crank the settings up you might hide these but I haven’t spent a lot of time optimizing the process. Some people might say to obfuscate your xaml using string encryption but again, test, test, test. Xaml is picky so too much obfuscation (or any) might disable your app or produce odd rendering effets. Remember, obfuscation is not 100% secure! Don’t rely on it as a sole way of protecting your assets. Other Tools Dotfuscator is one just product and isn’t the end-all be-all to obfuscation so check out others below. For example, Crypto can make it so Reflector doesn’t even recognize the app as a .NET one and won’t open it. Others can encrypt resources and Xaml markup files. Here are some other obfuscators that support the Windows Phone 7 platform. Feel free to give them a try and let people know your experience with them! Dotfuscator Windows Phone Edition Crypto Obfuscator for .NET DeepSea Obfuscation

    Read the article

  • Metro: Promises

    - by Stephen.Walther
    The goal of this blog entry is to describe the Promise class in the WinJS library. You can use promises whenever you need to perform an asynchronous operation such as retrieving data from a remote website or a file from the file system. Promises are used extensively in the WinJS library. Asynchronous Programming Some code executes immediately, some code requires time to complete or might never complete at all. For example, retrieving the value of a local variable is an immediate operation. Retrieving data from a remote website takes longer or might not complete at all. When an operation might take a long time to complete, you should write your code so that it executes asynchronously. Instead of waiting for an operation to complete, you should start the operation and then do something else until you receive a signal that the operation is complete. An analogy. Some telephone customer service lines require you to wait on hold – listening to really bad music – until a customer service representative is available. This is synchronous programming and very wasteful of your time. Some newer customer service lines enable you to enter your telephone number so the customer service representative can call you back when a customer representative becomes available. This approach is much less wasteful of your time because you can do useful things while waiting for the callback. There are several patterns that you can use to write code which executes asynchronously. The most popular pattern in JavaScript is the callback pattern. When you call a function which might take a long time to return a result, you pass a callback function to the function. For example, the following code (which uses jQuery) includes a function named getFlickrPhotos which returns photos from the Flickr website which match a set of tags (such as “dog” and “funny”): function getFlickrPhotos(tags, callback) { $.getJSON( "http://api.flickr.com/services/feeds/photos_public.gne?jsoncallback=?", { tags: tags, tagmode: "all", format: "json" }, function (data) { if (callback) { callback(data.items); } } ); } getFlickrPhotos("funny, dogs", function(data) { $.each(data, function(index, item) { console.log(item); }); }); The getFlickr() function includes a callback parameter. When you call the getFlickr() function, you pass a function to the callback parameter which gets executed when the getFlicker() function finishes retrieving the list of photos from the Flickr web service. In the code above, the callback function simply iterates through the results and writes each result to the console. Using callbacks is a natural way to perform asynchronous programming with JavaScript. Instead of waiting for an operation to complete, sitting there and listening to really bad music, you can get a callback when the operation is complete. Using Promises The CommonJS website defines a promise like this (http://wiki.commonjs.org/wiki/Promises): “Promises provide a well-defined interface for interacting with an object that represents the result of an action that is performed asynchronously, and may or may not be finished at any given point in time. By utilizing a standard interface, different components can return promises for asynchronous actions and consumers can utilize the promises in a predictable manner.” A promise provides a standard pattern for specifying callbacks. In the WinJS library, when you create a promise, you can specify three callbacks: a complete callback, a failure callback, and a progress callback. Promises are used extensively in the WinJS library. The methods in the animation library, the control library, and the binding library all use promises. For example, the xhr() method included in the WinJS base library returns a promise. The xhr() method wraps calls to the standard XmlHttpRequest object in a promise. The following code illustrates how you can use the xhr() method to perform an Ajax request which retrieves a file named Photos.txt: var options = { url: "/data/photos.txt" }; WinJS.xhr(options).then( function (xmlHttpRequest) { console.log("success"); var data = JSON.parse(xmlHttpRequest.responseText); console.log(data); }, function(xmlHttpRequest) { console.log("fail"); }, function(xmlHttpRequest) { console.log("progress"); } ) The WinJS.xhr() method returns a promise. The Promise class includes a then() method which accepts three callback functions: a complete callback, an error callback, and a progress callback: Promise.then(completeCallback, errorCallback, progressCallback) In the code above, three anonymous functions are passed to the then() method. The three callbacks simply write a message to the JavaScript Console. The complete callback also dumps all of the data retrieved from the photos.txt file. Creating Promises You can create your own promises by creating a new instance of the Promise class. The constructor for the Promise class requires a function which accepts three parameters: a complete, error, and progress function parameter. For example, the code below illustrates how you can create a method named wait10Seconds() which returns a promise. The progress function is called every second and the complete function is not called until 10 seconds have passed: (function () { "use strict"; var app = WinJS.Application; function wait10Seconds() { return new WinJS.Promise(function (complete, error, progress) { var seconds = 0; var intervalId = window.setInterval(function () { seconds++; progress(seconds); if (seconds > 9) { window.clearInterval(intervalId); complete(); } }, 1000); }); } app.onactivated = function (eventObject) { if (eventObject.detail.kind === Windows.ApplicationModel.Activation.ActivationKind.launch) { wait10Seconds().then( function () { console.log("complete") }, function () { console.log("error") }, function (seconds) { console.log("progress:" + seconds) } ); } } app.start(); })(); All of the work happens in the constructor function for the promise. The window.setInterval() method is used to execute code every second. Every second, the progress() callback method is called. If more than 10 seconds have passed then the complete() callback method is called and the clearInterval() method is called. When you execute the code above, you can see the output in the Visual Studio JavaScript Console. Creating a Timeout Promise In the previous section, we created a custom Promise which uses the window.setInterval() method to complete the promise after 10 seconds. We really did not need to create a custom promise because the Promise class already includes a static method for returning promises which complete after a certain interval. The code below illustrates how you can use the timeout() method. The timeout() method returns a promise which completes after a certain number of milliseconds. WinJS.Promise.timeout(3000).then( function(){console.log("complete")}, function(){console.log("error")}, function(){console.log("progress")} ); In the code above, the Promise completes after 3 seconds (3000 milliseconds). The Promise returned by the timeout() method does not support progress events. Therefore, the only message written to the console is the message “complete” after 10 seconds. Canceling Promises Some promises, but not all, support cancellation. When you cancel a promise, the promise’s error callback is executed. For example, the following code uses the WinJS.xhr() method to perform an Ajax request. However, immediately after the Ajax request is made, the request is cancelled. // Specify Ajax request options var options = { url: "/data/photos.txt" }; // Make the Ajax request var request = WinJS.xhr(options).then( function (xmlHttpRequest) { console.log("success"); }, function (xmlHttpRequest) { console.log("fail"); }, function (xmlHttpRequest) { console.log("progress"); } ); // Cancel the Ajax request request.cancel(); When you run the code above, the message “fail” is written to the Visual Studio JavaScript Console. Composing Promises You can build promises out of other promises. In other words, you can compose promises. There are two static methods of the Promise class which you can use to compose promises: the join() method and the any() method. When you join promises, a promise is complete when all of the joined promises are complete. When you use the any() method, a promise is complete when any of the promises complete. The following code illustrates how to use the join() method. A new promise is created out of two timeout promises. The new promise does not complete until both of the timeout promises complete: WinJS.Promise.join([WinJS.Promise.timeout(1000), WinJS.Promise.timeout(5000)]) .then(function () { console.log("complete"); }); The message “complete” will not be written to the JavaScript Console until both promises passed to the join() method completes. The message won’t be written for 5 seconds (5,000 milliseconds). The any() method completes when any promise passed to the any() method completes: WinJS.Promise.any([WinJS.Promise.timeout(1000), WinJS.Promise.timeout(5000)]) .then(function () { console.log("complete"); }); The code above writes the message “complete” to the JavaScript Console after 1 second (1,000 milliseconds). The message is written to the JavaScript console immediately after the first promise completes and before the second promise completes. Summary The goal of this blog entry was to describe WinJS promises. First, we discussed how promises enable you to easily write code which performs asynchronous actions. You learned how to use a promise when performing an Ajax request. Next, we discussed how you can create your own promises. You learned how to create a new promise by creating a constructor function with complete, error, and progress parameters. Finally, you learned about several advanced methods of promises. You learned how to use the timeout() method to create promises which complete after an interval of time. You also learned how to cancel promises and compose promises from other promises.

    Read the article

  • Software Engineering Practices &ndash; Different Projects should have different maturity levels

    - by Dylan Smith
    I’ve had a lot of discussions at the office lately about the drastically different sets of software engineering practices used on our various projects, if what we are doing is appropriate, and what factors should you be considering when determining what practices are most appropriate in a given context. I wanted to write up my thoughts in a little more detail on this subject, so here we go: If you compare any two software projects (specifically comparing their codebases) you’ll often see very different levels of maturity in the software engineering practices employed. By software engineering practices, I’m specifically referring to the quality of the code and the amount of technical debt present in the project. Things such as Test Driven Development, Domain Driven Design, Behavior Driven Development, proper adherence to the SOLID principles, etc. are all practices that you would expect at the mature end of the spectrum. At the other end of the spectrum would be the quick-and-dirty solutions that are done using something like an Access Database, Excel Spreadsheet, or maybe some quick “drag-and-drop coding”. For this blog post I’m going to refer to this as the Software Engineering Maturity Spectrum (SEMS). I believe there is a time and a place for projects at every part of that SEMS. The risks and costs associated with under-engineering solutions have been written about a million times over so I won’t bother going into them again here, but there are also (unnecessary) costs with over-engineering a solution. Sometimes putting multiple layers, and IoC containers, and abstracting out the persistence, etc is complete overkill if a one-time use Access database could solve the problem perfectly well. A lot of software developers I talk to seem to automatically jump to the very right-hand side of this SEMS in everything they do. A common rationalization I hear is that it may seem like a small trivial application today, but these things always grow and stick around for many years, then you’re stuck maintaining a big ball of mud. I think this is a cop-out. Sure you can’t always anticipate how an application will be used or grow over its lifetime (can you ever??), but that doesn’t mean you can’t manage it and evolve the underlying software architecture as necessary (even if that means having to toss the code out and re-write it at some point…maybe even multiple times). My thoughts are that we should be making a conscious decision around the start of each project approximately where on the SEMS we want the project to exist. I believe this decision should be based on 3 factors: 1. Importance - How important to the business is this application? What is the impact if the application were to suddenly stop working? 2. Complexity - How complex is the application functionality? 3. Life-Expectancy - How long is this application expected to be in use? Is this a one-time use application, does it fill a short-term need, or is it more strategic and is expected to be in-use for many years to come? Of course this isn’t an exact science. You can’t say that Project X should be at the 73% mark on the SEMS and expect that to be helpful. My point is not that you need to precisely figure out what point on the SEMS the project should be at then translate that into some prescriptive set of practices and techniques you should be using. Rather my point is that we need to be aware that there is a spectrum, and that not everything is going to be (or should be) at the edges of that spectrum, indeed a large number of projects should probably fall somewhere within the middle; and different projects should adopt a different level of software engineering practices and maturity levels based on the needs of that project. To give an example of this way of thinking from my day job: Every couple of years my company plans and hosts a large event where ~400 of our customers all fly in to one location for a multi-day event with various activities. We have some staff whose job it is to organize the logistics of this event, which includes tracking which flights everybody is booked on, arranging for transportation to/from airports, arranging for hotel rooms, name tags, etc The last time we arranged this event all these various pieces of data were tracked in separate spreadsheets and reconciliation and cross-referencing of all the data was literally done by hand using printed copies of the spreadsheets and several people sitting around a table going down each list row by row. Obviously there is some room for improvement in how we are using software to manage the event’s logistics. The next time this event occurs we plan to provide the event planning staff with a more intelligent tool (either an Excel spreadsheet or probably an Access database) that can track all the information in one location and make sure that the various pieces of data are properly linked together (so for example if a person cancels you only need to delete them from one place, and not a dozen separate lists). This solution would fall at or near the very left end of the SEMS meaning that we will just quickly create something with very little attention paid to using mature software engineering practices. If we examine this project against the 3 criteria I listed above for determining it’s place within the SEMS we can see why: Importance – If this application were to stop working the business doesn’t grind to a halt, revenue doesn’t stop, and in fact our customers wouldn’t even notice since it isn’t a customer facing application. The impact would simply be more work for our event planning staff as they revert back to the previous way of doing things (assuming we don’t have any data loss). Complexity – The use cases for this project are pretty straightforward. It simply needs to manage several lists of data, and link them together appropriately. Precisely the task that access (and/or Excel) can do with minimal custom development required. Life-Expectancy – For this specific project we’re only planning to create something to be used for the one event (we only hold these events every 2 years). If it works well this may change (see below). Let’s assume we hack something out quickly and it works great when we plan the next event. We may decide that we want to make some tweaks to the tool and adopt it for planning all future events of this nature. In that case we should examine where the current application is on the SEMS, and make a conscious decision whether something needs to be done to move it further to the right based on the new objectives and goals for this application. This may mean scrapping the access database and re-writing it as an actual web or windows application. In this case, the life-expectancy changed, but let’s assume the importance and complexity didn’t change all that much. We can still probably get away with not adopting a lot of the so-called “best practices”. For example, we can probably still use some of the RAD tooling available and might have an Autonomous View style design that connects directly to the database and binds to typed datasets (we might even choose to simply leave it as an access database and continue using it; this is a decision that needs to be made on a case-by-case basis). At Anvil Digital we have aspirations to become a primarily product-based company. So let’s say we use this tool to plan a handful of events internally, and everybody loves it. Maybe a couple years down the road we decide we want to package the tool up and sell it as a product to some of our customers. In this case the project objectives/goals change quite drastically. Now the tool becomes a source of revenue, and the impact of it suddenly stopping working is significantly less acceptable. Also as we hold focus groups, and gather feedback from customers and potential customers there’s a pretty good chance the feature-set and complexity will have to grow considerably from when we were using it only internally for planning a small handful of events for one company. In this fictional scenario I would expect the target on the SEMS to jump to the far right. Depending on how we implemented the previous release we may be able to refactor and evolve the existing codebase to introduce a more layered architecture, a robust set of automated tests, introduce a proper ORM and IoC container, etc. More likely in this example the jump along the SEMS would be so large we’d probably end up scrapping the current code and re-writing. Although, if it was a slow phased roll-out to only a handful of customers, where we collected feedback, made some tweaks, and then rolled out to a couple more customers, we may be able to slowly refactor and evolve the code over time rather than tossing it out and starting from scratch. The key point I’m trying to get across is not that you should be throwing out your code and starting from scratch all the time. But rather that you should be aware of when and how the context and objectives around a project changes and periodically re-assess where the project currently falls on the SEMS and whether that needs to be adjusted based on changing needs. Note: There is also the idea of “spectrum decay”. Since our industry is rapidly evolving, what we currently accept as mature software engineering practices (the right end of the SEMS) probably won’t be the same 3 years from now. If you have a project that you were to assess at somewhere around the 80% mark on the SEMS today, but don’t touch the code for 3 years and come back and re-assess its position, it will almost certainly have changed since the right end of the SEMS will have moved farther out (maybe the project is now only around 60% due to decay). Developer Skills Another important aspect to this whole discussion is around the skill sets of your architects and lead developers. When talking about the progression of a developers skills from junior->intermediate->senior->… they generally start by only being able to write code that belongs on the left side of the SEMS and as they gain more knowledge and skill they become capable of working at a higher and higher level along the SEMS. We all realize that the learning never stops, but eventually you’ll get to the point where you can comfortably develop at the right-end of the SEMS (the exact practices and techniques that translates to is constantly changing, but that’s not the point here). A critical skill that I’d love to see more evidence of in our industry is the most senior guys not only being able to work at the right-end of the SEMS, but more importantly be able to consciously work at any point along the SEMS as project needs dictate. An even more valuable skill would be if you could make the conscious decision to move a projects code further right on the SEMS (based on changing needs) and do so in an incremental manner without having to start from scratch. An exercise that I’m planning to go through with all of our projects here at Anvil in the near future is to map out where I believe each project currently falls within this SEMS, where I believe the project *should* be on the SEMS based on the business needs, and for those that don’t match up (i.e. most of them) come up with a plan to improve the situation.

    Read the article

  • Operator of the week - Assert

    - by Fabiano Amorim
    Well my friends, I was wondering how to help you in a practical way to understand execution plans. So I think I'll talk about the Showplan Operators. Showplan Operators are used by the Query Optimizer (QO) to build the query plan in order to perform a specified operation. A query plan will consist of many physical operators. The Query Optimizer uses a simple language that represents each physical operation by an operator, and each operator is represented in the graphical execution plan by an icon. I'll try to talk about one operator every week, but so as to avoid having to continue to write about these operators for years, I'll mention only of those that are more common: The first being the Assert. The Assert is used to verify a certain condition, it validates a Constraint on every row to ensure that the condition was met. If, for example, our DDL includes a check constraint which specifies only two valid values for a column, the Assert will, for every row, validate the value passed to the column to ensure that input is consistent with the check constraint. Assert  and Check Constraints: Let's see where the SQL Server uses that information in practice. Take the following T-SQL: IF OBJECT_ID('Tab1') IS NOT NULL   DROP TABLE Tab1 GO CREATE TABLE Tab1(ID Integer, Gender CHAR(1))  GO  ALTER TABLE TAB1 ADD CONSTRAINT ck_Gender_M_F CHECK(Gender IN('M','F'))  GO INSERT INTO Tab1(ID, Gender) VALUES(1,'X') GO To the command above the SQL Server has generated the following execution plan: As we can see, the execution plan uses the Assert operator to check that the inserted value doesn't violate the Check Constraint. In this specific case, the Assert applies the rule, 'if the value is different to "F" and different to "M" than return 0 otherwise returns NULL'. The Assert operator is programmed to show an error if the returned value is not NULL; in other words, the returned value is not a "M" or "F". Assert checking Foreign Keys Now let's take a look at an example where the Assert is used to validate a foreign key constraint. Suppose we have this  query: ALTER TABLE Tab1 ADD ID_Genders INT GO  IF OBJECT_ID('Tab2') IS NOT NULL   DROP TABLE Tab2 GO CREATE TABLE Tab2(ID Integer PRIMARY KEY, Gender CHAR(1))  GO  INSERT INTO Tab2(ID, Gender) VALUES(1, 'F') INSERT INTO Tab2(ID, Gender) VALUES(2, 'M') INSERT INTO Tab2(ID, Gender) VALUES(3, 'N') GO  ALTER TABLE Tab1 ADD CONSTRAINT fk_Tab2 FOREIGN KEY (ID_Genders) REFERENCES Tab2(ID) GO  INSERT INTO Tab1(ID, ID_Genders, Gender) VALUES(1, 4, 'X') Let's look at the text execution plan to see what these Assert operators were doing. To see the text execution plan just execute SET SHOWPLAN_TEXT ON before run the insert command. |--Assert(WHERE:(CASE WHEN NOT [Pass1008] AND [Expr1007] IS NULL THEN (0) ELSE NULL END))      |--Nested Loops(Left Semi Join, PASSTHRU:([Tab1].[ID_Genders] IS NULL), OUTER REFERENCES:([Tab1].[ID_Genders]), DEFINE:([Expr1007] = [PROBE VALUE]))           |--Assert(WHERE:(CASE WHEN [Tab1].[Gender]<>'F' AND [Tab1].[Gender]<>'M' THEN (0) ELSE NULL END))           |    |--Clustered Index Insert(OBJECT:([Tab1].[PK]), SET:([Tab1].[ID] = RaiseIfNullInsert([@1]),[Tab1].[ID_Genders] = [@2],[Tab1].[Gender] = [Expr1003]), DEFINE:([Expr1003]=CONVERT_IMPLICIT(char(1),[@3],0)))           |--Clustered Index Seek(OBJECT:([Tab2].[PK]), SEEK:([Tab2].[ID]=[Tab1].[ID_Genders]) ORDERED FORWARD) Here we can see the Assert operator twice, first (looking down to up in the text plan and the right to left in the graphical plan) validating the Check Constraint. The same concept showed above is used, if the exit value is "0" than keep running the query, but if NULL is returned shows an exception. The second Assert is validating the result of the Tab1 and Tab2 join. It is interesting to see the "[Expr1007] IS NULL". To understand that you need to know what this Expr1007 is, look at the Probe Value (green text) in the text plan and you will see that it is the result of the join. If the value passed to the INSERT at the column ID_Gender exists in the table Tab2, then that probe will return the join value; otherwise it will return NULL. So the Assert is checking the value of the search at the Tab2; if the value that is passed to the INSERT is not found  then Assert will show one exception. If the value passed to the column ID_Genders is NULL than the SQL can't show a exception, in that case it returns "0" and keeps running the query. If you run the INSERT above, the SQL will show an exception because of the "X" value, but if you change the "X" to "F" and run again, it will show an exception because of the value "4". If you change the value "4" to NULL, 1, 2 or 3 the insert will be executed without any error. Assert checking a SubQuery: The Assert operator is also used to check one subquery. As we know, one scalar subquery can't validly return more than one value: Sometimes, however, a  mistake happens, and a subquery attempts to return more than one value . Here the Assert comes into play by validating the condition that a scalar subquery returns just one value. Take the following query: INSERT INTO Tab1(ID_TipoSexo, Sexo) VALUES((SELECT ID_TipoSexo FROM Tab1), 'F')    INSERT INTO Tab1(ID_TipoSexo, Sexo) VALUES((SELECT ID_TipoSexo FROM Tab1), 'F')    |--Assert(WHERE:(CASE WHEN NOT [Pass1016] AND [Expr1015] IS NULL THEN (0) ELSE NULL END))        |--Nested Loops(Left Semi Join, PASSTHRU:([tempdb].[dbo].[Tab1].[ID_TipoSexo] IS NULL), OUTER REFERENCES:([tempdb].[dbo].[Tab1].[ID_TipoSexo]), DEFINE:([Expr1015] = [PROBE VALUE]))              |--Assert(WHERE:([Expr1017]))             |    |--Compute Scalar(DEFINE:([Expr1017]=CASE WHEN [tempdb].[dbo].[Tab1].[Sexo]<>'F' AND [tempdb].[dbo].[Tab1].[Sexo]<>'M' THEN (0) ELSE NULL END))              |         |--Clustered Index Insert(OBJECT:([tempdb].[dbo].[Tab1].[PK__Tab1__3214EC277097A3C8]), SET:([tempdb].[dbo].[Tab1].[ID_TipoSexo] = [Expr1008],[tempdb].[dbo].[Tab1].[Sexo] = [Expr1009],[tempdb].[dbo].[Tab1].[ID] = [Expr1003]))              |              |--Top(TOP EXPRESSION:((1)))              |                   |--Compute Scalar(DEFINE:([Expr1008]=[Expr1014], [Expr1009]='F'))              |                        |--Nested Loops(Left Outer Join)              |                             |--Compute Scalar(DEFINE:([Expr1003]=getidentity((1856985942),(2),NULL)))              |                             |    |--Constant Scan              |                             |--Assert(WHERE:(CASE WHEN [Expr1013]>(1) THEN (0) ELSE NULL END))              |                                  |--Stream Aggregate(DEFINE:([Expr1013]=Count(*), [Expr1014]=ANY([tempdb].[dbo].[Tab1].[ID_TipoSexo])))             |                                       |--Clustered Index Scan(OBJECT:([tempdb].[dbo].[Tab1].[PK__Tab1__3214EC277097A3C8]))              |--Clustered Index Seek(OBJECT:([tempdb].[dbo].[Tab2].[PK__Tab2__3214EC27755C58E5]), SEEK:([tempdb].[dbo].[Tab2].[ID]=[tempdb].[dbo].[Tab1].[ID_TipoSexo]) ORDERED FORWARD)  You can see from this text showplan that SQL Server as generated a Stream Aggregate to count how many rows the SubQuery will return, This value is then passed to the Assert which then does its job by checking its validity. Is very interesting to see that  the Query Optimizer is smart enough be able to avoid using assert operators when they are not necessary. For instance: INSERT INTO Tab1(ID_TipoSexo, Sexo) VALUES((SELECT ID_TipoSexo FROM Tab1 WHERE ID = 1), 'F') INSERT INTO Tab1(ID_TipoSexo, Sexo) VALUES((SELECT TOP 1 ID_TipoSexo FROM Tab1), 'F')  For both these INSERTs, the Query Optimiser is smart enough to know that only one row will ever be returned, so there is no need to use the Assert. Well, that's all folks, I see you next week with more "Operators". Cheers, Fabiano

    Read the article

  • Going Paperless

    - by Jesse
    One year ago I came to work for a company where the entire development team is 100% “remote”; we’re spread over 3 time zones and each of us works from home. This seems to be an increasingly popular way for people to work and there are many articles and blog posts out there enumerating the advantages and disadvantages of working this way. I had read a lot about telecommuting before accepting this job and felt as if I had a pretty decent idea of what I was getting into, but I’ve encountered a few things over the past year that I did not expect. Among the most surprising by-products of working from home for me has been a dramatic reduction in the amount of paper that I use on a weekly basis. Hoarding In The Workplace Prior to my current telecommute job I worked in what most would consider pretty traditional office environments. I sat in cubicles furnished with an enormous plastic(ish) modular desks, had a mediocre (at best) PC workstation, and had ready access to a seemingly endless supply of legal pads, pens, staplers and paper clips. The ready access to paper, countless conference room meetings, and abundance of available surface area on my desk and in drawers created a perfect storm for wasting paper. I brought a pad of paper with me to every meeting I ever attended, scrawled some brief notes, and then tore that sheet off to keep next to my keyboard to follow up on any needed action items. Once my immediate need for the notes was fulfilled, that sheet would get shuffled off into a corner of my desk or filed away in a drawer “just in case”. I would guess that for all of the notes that I ever filed away, I might have actually had to dig up and refer to 2% of them (and that’s probably being very generous). That said, on those rare occasions that I did have to dig something up from old notes, it was usually pretty important and I ended up being very glad that I saved them. It was only when I would leave a job or move desks that I would finally gather all those notes together and take them to shredding bin to be disposed of. When I left my last job the amount of paper I had accumulated over my three years there was absurd, and I knew coworkers who had substance-abuse caliber paper wasting addictions that made my bad habit look like nail-biting in comparison. A Product Of My Environment I always hated using all of this paper, but simply couldn’t bring myself to stop. It would look bad if I showed up to an important conference room meeting without a pad of paper. What if someone said something profound! Plus, everyone else always brought paper with them. If you saw someone walking down the hallway with a pad of paper in hand you knew they must be on their way to a conference room meeting. Some people even had fancy looking portfolio notebook sheaths that gave their legal pads all the prestige of a briefcase. No one ever worried about running out of fresh paper because there was an endless supply, and there certainly was no shortage of places to store and file used paper. In short, the traditional office was setup for using tons and tons of paper; it’s baked into the culture there. For that reason, it didn’t take long for me to kick the paper habit once I started working from home. In my home office, desk and drawer space are at a premium. I don’t have the budget (or the tolerance) for huge modular office furniture in my spare bedroom. I also no longer have access to a bottomless pit of office supplies stock piled in cabinets and closets. If I want to use some paper, I have to go out and buy it. Finally (and most importantly), all of the meetings that I have to attend these days are “virtual”. We use instant messaging, VOIP, video conferencing, and e-mail to communicate with each other. All I need to take notes during a meeting is my computer, which I happen to be sitting right in front of all day. I don’t have any hard numbers for this, but my gut feeling is that I actually take a lot more notes now than I ever did when I worked in an office. The big difference is I don’t have to use any paper to do so. This makes it far easier to keep important information safe and organized. The Right Tool For The Job When I first started working from home I tried to find a single application that would fill the gap left by the pen and paper that I always had at my desk when I worked in an office. Well, there are no silver bullets and I’ve evolved my approach over time to try and find the best tool for the job at hand. Here’s a quick summary of how I take notes and keep everything organized. Notepad++ – This is the first application I turn to when I feel like there’s some bit of information that I need to write down and save. I use Launchy, so opening Notepad++ and creating a new file only takes a few keystrokes. If I find that the information I’m trying to get down requires a more sophisticated application I escalate as needed. The Desktop – By default, I save every file or other bit of information to the desktop. Anyone who has ever had to fix their parents computer before knows that this is a dangerous game (any file my mother has ever worked on is saved directly to the desktop and rarely moves anywhere else). I agree that storing things on the desktop isn’t a great long term approach to keeping organized, which is why I treat my desktop a bit like my e-mail inbox. I strive to keep both empty (or as close to empty as I possibly can). If something is on my desktop, it means that it’s something relevant to a task or project that I’m currently working on. About once a week I take things that I’m not longer working on and put them into my ‘Notes’ folder. The ‘Notes’ Folder – As I work on a task, I tend to accumulate multiple files associated with that task. For example, I might have a bit of SQL that I’m working on to gather data for a new report, a quick C# method that I came up with but am not yet ready to commit to source control, a bulleted list of to-do items in a .txt file, etc. If the desktop starts to get too cluttered, I create a new sub-folder in my ‘Notes’ folder. Each sub-folder’s name is the current date followed by a brief description of the task or project. Then all files related to that task or project go into that sub folder. By using the date as the first part of the folder name, these folders are automatically sorted in reverse chronological order. This means that things I worked on recently will generally be near the top of the list. Using the built-in Windows search functionality I now have a pretty quick and easy way to try and find something that I worked on a week ago or six months ago. Dropbox – Dropbox is a free service that lets you store up to 2GB of files “in the cloud” and have those files synced to all of the different computers that you use. My ‘Notes’ folder lives in Dropbox, meaning that it’s contents are constantly backed up and are always available to me regardless of which computer I’m using. They also have a pretty decent iPhone application that lets you browse and view all of the files that you have stored there. The free 2GB edition is probably enough for just storing notes, but I also pay $99/year for the 50GB storage upgrade and keep all of my music, e-books, pictures, and documents in Dropbox. It’s a fantastic service and I highly recommend it. Evernote – I use Evernote mostly to organize information that I access on a fairly regular basis. For example, my Evernote account has a running grocery shopping list, recipes that my wife and I use a lot, and contact information for people I contact infrequently enough that I don’t want to keep them in my phone. I know some people that keep nearly everything in Evernote, but there’s something about it that I find a bit clunky, so I tend to use it sparingly. Google Tasks – One of my biggest paper wasting habits was keeping a running task-list next to my computer at work. Every morning I would sit down, look at my task list, cross off what was done and add new tasks that I thought of during my morning commute. This usually resulted in having to re-copy the task list onto a fresh sheet of paper when I was done. I still keep a running task list at my desk, but I’ve started using Google Tasks instead. This is a dead-simple web-based application for quickly adding, deleting, and organizing tasks in a simple checklist style. You can quickly move tasks up and down on the list (which I use for prioritizing), and even create sub-tasks for breaking down larger tasks into smaller pieces. Balsamiq Mockups – This is a simple and lightweight tool for creating drawings of user interfaces. It’s great for sketching out a new feature, brainstorm the layout of a interface, or even draw up a quick sequence diagram. I’m terrible at drawing, so Balsamiq Mockups not only lets me create sketches that other people can actually understand, but it’s also handy because you can upload a sketch to a common location for other team members to access. I can honestly say that using these tools (and having limited resources at home) have lead me to cut my paper usage down to virtually none. If I ever were to return to a traditional office workplace (hopefully never!) I’d try to employ as many of these applications and techniques as I could to keep paper usage low. I feel far less cluttered and far better organized now.

    Read the article

  • 10 Windows Tweaking Myths Debunked

    - by Chris Hoffman
    Windows is big, complicated, and misunderstood. You’ll still stumble across bad advice from time to time when browsing the web. These Windows tweaking, performance, and system maintenance tips are mostly just useless, but some are actively harmful. Luckily, most of these myths have been stomped out on mainstream sites and forums. However, if you start searching the web, you’ll still find websites that recommend you do these things. Erase Cache Files Regularly to Speed Things Up You can free up disk space by running an application like CCleaner, another temporary-file-cleaning utility, or even the Windows Disk Cleanup tool. In some cases, you may even see an old computer speed up when you erase a large amount of useless files. However, running CCleaner or similar utilities every day to erase your browser’s cache won’t actually speed things up. It will slow down your web browsing as your web browser is forced to redownload the files all over again, and reconstruct the cache you regularly delete. If you’ve installed CCleaner or a similar program and run it every day with the default settings, you’re actually slowing down your web browsing. Consider at least preventing the program from wiping out your web browser cache. Enable ReadyBoost to Speed Up Modern PCs Windows still prompts you to enable ReadyBoost when you insert a USB stick or memory card. On modern computers, this is completely pointless — ReadyBoost won’t actually speed up your computer if you have at least 1 GB of RAM. If you have a very old computer with a tiny amount of RAM — think 512 MB — ReadyBoost may help a bit. Otherwise, don’t bother. Open the Disk Defragmenter and Manually Defragment On Windows 98, users had to manually open the defragmentation tool and run it, ensuring no other applications were using the hard drive while it did its work. Modern versions of Windows are capable of defragmenting your file system while other programs are using it, and they automatically defragment your disks for you. If you’re still opening the Disk Defragmenter every week and clicking the Defragment button, you don’t need to do this — Windows is doing it for you unless you’ve told it not to run on a schedule. Modern computers with solid-state drives don’t have to be defragmented at all. Disable Your Pagefile to Increase Performance When Windows runs out of empty space in RAM, it swaps out data from memory to a pagefile on your hard disk. If a computer doesn’t have much memory and it’s running slow, it’s probably moving data to the pagefile or reading data from it. Some Windows geeks seem to think that the pagefile is bad for system performance and disable it completely. The argument seems to be that Windows can’t be trusted to manage a pagefile and won’t use it intelligently, so the pagefile needs to be removed. As long as you have enough RAM, it’s true that you can get by without a pagefile. However, if you do have enough RAM, Windows will only use the pagefile rarely anyway. Tests have found that disabling the pagefile offers no performance benefit. Enable CPU Cores in MSConfig Some websites claim that Windows may not be using all of your CPU cores or that you can speed up your boot time by increasing the amount of cores used during boot. They direct you to the MSConfig application, where you can indeed select an option that appears to increase the amount of cores used. In reality, Windows always uses the maximum amount of processor cores your CPU has. (Technically, only one core is used at the beginning of the boot process, but the additional cores are quickly activated.) Leave this option unchecked. It’s just a debugging option that allows you to set a maximum number of cores, so it would be useful if you wanted to force Windows to only use a single core on a multi-core system — but all it can do is restrict the amount of cores used. Clean Your Prefetch To Increase Startup Speed Windows watches the programs you run and creates .pf files in its Prefetch folder for them. The Prefetch feature works as a sort of cache — when you open an application, Windows checks the Prefetch folder, looks at the application’s .pf file (if it exists), and uses that as a guide to start preloading data that the application will use. This helps your applications start faster. Some Windows geeks have misunderstood this feature. They believe that Windows loads these files at boot, so your boot time will slow down due to Windows preloading the data specified in the .pf files. They also argue you’ll build up useless files as you uninstall programs and .pf files will be left over. In reality, Windows only loads the data in these .pf files when you launch the associated application and only stores .pf files for the 128 most recently launched programs. If you were to regularly clean out the Prefetch folder, not only would programs take longer to open because they won’t be preloaded, Windows will have to waste time recreating all the .pf files. You could also modify the PrefetchParameters setting to disable Prefetch, but there’s no reason to do that. Let Windows manage Prefetch on its own. Disable QoS To Increase Network Bandwidth Quality of Service (QoS) is a feature that allows your computer to prioritize its traffic. For example, a time-critical application like Skype could choose to use QoS and prioritize its traffic over a file-downloading program so your voice conversation would work smoothly, even while you were downloading files. Some people incorrectly believe that QoS always reserves a certain amount of bandwidth and this bandwidth is unused until you disable it. This is untrue. In reality, 100% of bandwidth is normally available to all applications unless a program chooses to use QoS. Even if a program does choose to use QoS, the reserved space will be available to other programs unless the program is actively using it. No bandwidth is ever set aside and left empty. Set DisablePagingExecutive to Make Windows Faster The DisablePagingExecutive registry setting is set to 0 by default, which allows drivers and system code to be paged to the disk. When set to 1, drivers and system code will be forced to stay resident in memory. Once again, some people believe that Windows isn’t smart enough to manage the pagefile on its own and believe that changing this option will force Windows to keep important files in memory rather than stupidly paging them out. If you have more than enough memory, changing this won’t really do anything. If you have little memory, changing this setting may force Windows to push programs you’re using to the page file rather than push unused system files there — this would slow things down. This is an option that may be helpful for debugging in some situations, not a setting to change for more performance. Process Idle Tasks to Free Memory Windows does things, such as creating scheduled system restore points, when you step away from your computer. It waits until your computer is “idle” so it won’t slow your computer and waste your time while you’re using it. Running the “Rundll32.exe advapi32.dll,ProcessIdleTasks” command forces Windows to perform all of these tasks while you’re using the computer. This is completely pointless and won’t help free memory or anything like that — all you’re doing is forcing Windows to slow your computer down while you’re using it. This command only exists so benchmarking programs can force idle tasks to run before performing benchmarks, ensuring idle tasks don’t start running and interfere with the benchmark. Delay or Disable Windows Services There’s no real reason to disable Windows services anymore. There was a time when Windows was particularly heavy and computers had little memory — think Windows Vista and those “Vista Capable” PCs Microsoft was sued over. Modern versions of Windows like Windows 7 and 8 are lighter than Windows Vista and computers have more than enough memory, so you won’t see any improvements from disabling system services included with Windows. Some people argue for not disabling services, however — they recommend setting services from “Automatic” to “Automatic (Delayed Start)”. By default, the Delayed Start option just starts services two minutes after the last “Automatic” service starts. Setting services to Delayed Start won’t really speed up your boot time, as the services will still need to start — in fact, it may lengthen the time it takes to get a usable desktop as services will still be loading two minutes after booting. Most services can load in parallel, and loading the services as early as possible will result in a better experience. The “Delayed Start” feature is primarily useful for system administrators who need to ensure a specific service starts later than another service. If you ever find a guide that recommends you set a little-known registry setting to improve performance, take a closer look — the change is probably useless. Want to actually speed up your PC? Try disabling useless startup programs that run on boot, increasing your boot time and consuming memory in the background. This is a much better tip than doing any of the above, especially considering most Windows PCs come packed to the brim with bloatware.     

    Read the article

  • Introduction to LinqPad Driver for StreamInsight 2.1

    - by Roman Schindlauer
    We are announcing the availability of the LinqPad driver for StreamInsight 2.1. The purpose of this blog post is to offer a quick introduction into the new features that we added to the StreamInsight LinqPad driver. We’ll show you how to connect to a remote server, how to inspect the entities present of that server, how to compose on top of them and how to manage their lifetime. Installing the driver Info on how to install the driver can be found in an earlier blog post here. Establishing connections As you click on the “Add Connection” link in the left pane you will notice that now it’s possible to build the data context automatically. The new driver appears as an option in the upper list, and if you pick it you will open a connection dialog that lets you connect to a remote StreamInsight server. The connection dialog lets you specify the address of the remote server. You will notice that it’s possible to pick up the binding information from the configuration file of the LinqPad application (which is normally in the same folder as LinqPad.exe and is called LinqPad.exe.config). In order for the context to be generated you need to pick an application from the server. The control is editable hence you can create a new application if you don’t want to make changes to an existing application. If you choose a new application name you will be prompted for confirmation before this gets created. Once you click OK the connection is created and you can start issuing queries against the remote server. If there’s any connectivity error the connection is marked with a red X and you can see the error message informing you what went wrong (i.e., the remote server could not be reached etc.). The context for remote servers Let’s take a look at what happens after we are connected successfully. Every LinqPad query runs inside a context – think of it as a class that wraps all the code that you’re writing. If you’re connecting to a live server the context will contain the following: The application object itself. All entities present in this application (sources, sinks, subjects and processes). The picture below shows a snapshot of the left pane of LinqPad after a successful connection. Every entity on the server has a different icon which will allow users to figure out its purpose. You will also notice that some entities have a string in parentheses following the name. It should be interpreted as such: the first name is the name of the property of the context class and the second name is the name of the entity as it exists on the server. Not all valid entity names are valid identifier names so in cases where we had to make a transformation you see both. Note also that as you hover over the entities you get IntelliSense with their types – more on that later. Remoting is not supported As you play with the entities exposed by the context you will notice that you can’t read and write directly to/from them. If for instance you’re trying to dump the content of an entity you will get an error message telling you that in the current version remoting is not supported. This is because the entity lives on the remote server and dumping its content means reading the events produced by this entity into the local process. ObservableSource.Dump(); Will yield the following error: Reading from a remote 'System.Reactive.Linq.IQbservable`1[System.Int32]' is not supported. Use the 'Microsoft.ComplexEventProcessing.Linq.RemoteProvider.Bind' method to read from the source using a remote observer. This basically tells you that you can call the Bind() method to direct the output of this source to a sink that has to be defined on the remote machine as well. You can’t bring the results to the LinqPad window unless you write code specifically for that. Compose queries You may ask – what's the purpose of all that? After all the same information is present in the EventFlowDebugger, why bother with showing it in LinqPad? First of all, What gets exposed in LinqPad is not what you see in the debugger. In LinqPad we have a property on the context class for every entity that lives on the server. Because LinqPad offers IntelliSense we in fact have much more information about the entity, and more importantly we can compose with that entity very easily. For example, let’s say that this code creates an entity: using (var server = Server.Connect(...)) {     var a = server.CreateApplication("WhiteFish");     var src = a         .DefineObservable<int>(() => Observable.Range(0, 3))         .Deploy("ObservableSource"); If later we want to compose with the source we have to fetch it and then we can bind something to     a.GetObservable<int>("ObservableSource)").Bind(... This means that we had to know a bunch of things about this: that it’s a source, that it’s an observable, it produces a result with payload Int32 and it’s named “ObservableSource”. Only the second and last bits of information are present in the debugger, by the way. As you type in the query window you see that all the entities are present, you get IntelliSense support for them and it’s much easier to make sense of what’s available. Let’s look at a scenario where composition is plausible. With the new programming model it’s possible to create “cold” sources that are parameterized. There was a way to accomplish that even in the previous version by passing parameters to the adapters, but this time it’s much more elegant because the expression declares what parameters are required. Say that we hover the mouse over the ThrottledSource source – we will see that its type is Func<int, int, IQbservable<int>> - this in effect means that we need to pass two int parameters before we can get a source that produces events, and the type for those events is int – in the particular case of my example I had the source produce a range of integers and the two parameters were the start and end of the range. So we see how a developer can create a source that is not running yet. Then someone else (e.g. an administrator) can pass whatever parameters appropriate and run the process. Proxy Types Here’s an interesting scenario – what if someone created a source on a server but they forgot to tell you what type they used. Worse yet, they might have used an anonymous type and even though they can refer to it by name you can’t figure out how to use that type. Let’s walk through an example that shows how you can compose against types you don’t need to have the definition of. This is how we can create a source that returns an anonymous type: Application.DefineObservable(() => Observable.Range(1, 10).Select(i => new { I = i })).Deploy("O1"); Now if we refresh the connection we can see the new source named O1 appear in the list. But what’s more important is that we now have a type to work with. So we can compose a query that refers to the anonymous type. var threshold = new StreamInsightDynamicDriver.TypeProxies.AnonymousType1_0<int>(5); var filter = from i in O1              where i > threshold              select i; filter.Deploy("O2"); You will notice that the anonymous type defined with this statement: new { I = i } can now be manipulated by a client that does not have access to it because the LinqPad driver has generated another type in its stead, named StreamInsightDynamicDriver.TypeProxies.AnonymousType1_0. This type has all the properties and fields of the type defined on the server, except in this case we can instantiate values and use it to compose more queries. It is worth noting that the same thing works for types that are not anonymous – the test is if the LinqPad driver can resolve the type or not. If it’s not possible then a new type will be generated that approximates the type that exists on the server. Control metadata In addition to composing processes on top of the existing entities we can do other useful things. We can delete them – nothing new here as we simply access the entities through the Entities collection of the application class. Here is where having their real name in parentheses comes handy. There’s another way to find out what’s behind a property – dump its expression. The first line in the output tells us what’s the name of the entity used to build this property in the context. Runtime information So let’s create a process to see what happens. We can bind a source to a sink and run the resulting process. If you right click on the connection you can refresh it and see the process present in the list of entities. Then you can drag the process to the query window and see that you can have access to process object in the Processes collection of the application. You can then manipulate the process (delete it, read its diagnostic view etc.). Regards, The StreamInsight Team

    Read the article

  • Code Reuse is (Damn) Hard

    - by James Michael Hare
    Being a development team lead, the task of interviewing new candidates was part of my job.  Like any typical interview, we started with some easy questions to get them warmed up and help calm their nerves before hitting the hard stuff. One of those easier questions was almost always: “Name some benefits of object-oriented development.”  Nearly every time, the candidate would chime in with a plethora of canned answers which typically included: “it helps ease code reuse.”  Of course, this is a gross oversimplification.  Tools only ease reuse, its developers that ultimately can cause code to be reusable or not, regardless of the language or methodology. But it did get me thinking…  we always used to say that as part of our mantra as to why Object-Oriented Programming was so great.  With polymorphism, inheritance, encapsulation, etc. we in essence set up the concepts to help facilitate reuse as much as possible.  And yes, as a developer now of many years, I unquestionably held that belief for ages before it really struck me how my views on reuse have jaded over the years.  In fact, in many ways Agile rightly eschews reuse as taking a backseat to developing what's needed for the here and now.  It used to be I was in complete opposition to that view, but more and more I've come to see the logic in it.  Too many times I've seen developers (myself included) get lost in design paralysis trying to come up with the perfect abstraction that would stand all time.  Nearly without fail, all of these pieces of code become obsolete in a matter of months or years. It’s not that I don’t like reuse – it’s just that reuse is hard.  In fact, reuse is DAMN hard.  Many times it is just a distraction that eats up architect and developer time, and worse yet can be counter-productive and force wrong decisions.  Now don’t get me wrong, I love the idea of reusable code when it makes sense.  These are in the few cases where you are designing something that is inherently reusable.  The problem is, most business-class code is inherently unfit for reuse! Furthermore, the code that is reusable will often fail to be reused if you don’t have the proper framework in place for effective reuse that includes standardized versioning, building, releasing, and documenting the components.  That should always be standard across the board when promoting reusable code.  All of this is hard, and it should only be done when you have code that is truly reusable or you will be exerting a large amount of development effort for very little bang for your buck. But my goal here is not to get into how to reuse (that is a topic unto itself) but what should be reused.  First, let’s look at an extension method.  There’s many times where I want to kick off a thread to handle a task, then when I want to reign that thread in of course I want to do a Join on it.  But what if I only want to wait a limited amount of time and then Abort?  Well, I could of course write that logic out by hand each time, but it seemed like a great extension method: 1: public static class ThreadExtensions 2: { 3: public static bool JoinOrAbort(this Thread thread, TimeSpan timeToWait) 4: { 5: bool isJoined = false; 6:  7: if (thread != null) 8: { 9: isJoined = thread.Join(timeToWait); 10:  11: if (!isJoined) 12: { 13: thread.Abort(); 14: } 15: } 16: return isJoined; 17: } 18: } 19:  When I look at this code, I can immediately see things that jump out at me as reasons why this code is very reusable.  Some of them are standard OO principles, and some are kind-of home grown litmus tests: Single Responsibility Principle (SRP) – The only reason this extension method need change is if the Thread class itself changes (one responsibility). Stable Dependencies Principle (SDP) – This method only depends on classes that are more stable than it is (System.Threading.Thread), and in itself is very stable, hence other classes may safely depend on it. It is also not dependent on any business domain, and thus isn't subject to changes as the business itself changes. Open-Closed Principle (OCP) – This class is inherently closed to change. Small and Stable Problem Domain – This method only cares about System.Threading.Thread. All-or-None Usage – A user of a reusable class should want the functionality of that class, not parts of that functionality.  That’s not to say they most use every method, but they shouldn’t be using a method just to get half of its result. Cost of Reuse vs. Cost to Recreate – since this class is highly stable and minimally complex, we can offer it up for reuse very cheaply by promoting it as “ready-to-go” and already unit tested (important!) and available through a standard release cycle (very important!). Okay, all seems good there, now lets look at an entity and DAO.  I don’t know about you all, but there have been times I’ve been in organizations that get the grand idea that all DAOs and entities should be standardized and shared.  While this may work for small or static organizations, it’s near ludicrous for anything large or volatile. 1: namespace Shared.Entities 2: { 3: public class Account 4: { 5: public int Id { get; set; } 6:  7: public string Name { get; set; } 8:  9: public Address HomeAddress { get; set; } 10:  11: public int Age { get; set;} 12:  13: public DateTime LastUsed { get; set; } 14:  15: // etc, etc, etc... 16: } 17: } 18:  19: ... 20:  21: namespace Shared.DataAccess 22: { 23: public class AccountDao 24: { 25: public Account FindAccount(int id) 26: { 27: // dao logic to query and return account 28: } 29:  30: ... 31:  32: } 33: } Now to be fair, I’m not saying there doesn’t exist an organization where some entites may be extremely static and unchanging.  But at best such entities and DAOs will be problematic cases of reuse.  Let’s examine those same tests: Single Responsibility Principle (SRP) – The reasons to change for these classes will be strongly dependent on what the definition of the account is which can change over time and may have multiple influences depending on the number of systems an account can cover. Stable Dependencies Principle (SDP) – This method depends on the data model beneath itself which also is largely dependent on the business definition of an account which can be very inherently unstable. Open-Closed Principle (OCP) – This class is not really closed for modification.  Every time the account definition may change, you’d need to modify this class. Small and Stable Problem Domain – The definition of an account is inherently unstable and in fact may be very large.  What if you are designing a system that aggregates account information from several sources? All-or-None Usage – What if your view of the account encompasses data from 3 different sources but you only care about one of those sources or one piece of data?  Should you have to take the hit of looking up all the other data?  On the other hand, should you have ten different methods returning portions of data in chunks people tend to ask for?  Neither is really a great solution. Cost of Reuse vs. Cost to Recreate – DAOs are really trivial to rewrite, and unless your definition of an account is EXTREMELY stable, the cost to promote, support, and release a reusable account entity and DAO are usually far higher than the cost to recreate as needed. It’s no accident that my case for reuse was a utility class and my case for non-reuse was an entity/DAO.  In general, the smaller and more stable an abstraction is, the higher its level of reuse.  When I became the lead of the Shared Components Committee at my workplace, one of the original goals we looked at satisfying was to find (or create), version, release, and promote a shared library of common utility classes, frameworks, and data access objects.  Now, of course, many of you will point to nHibernate and Entity for the latter, but we were looking at larger, macro collections of data that span multiple data sources of varying types (databases, web services, etc). As we got deeper and deeper in the details of how to manage and release these items, it quickly became apparent that while the case for reuse was typically a slam dunk for utilities and frameworks, the data access objects just didn’t “smell” right.  We ended up having session after session of design meetings to try and find the right way to share these data access components. When someone asked me why it was taking so long to iron out the shared entities, my response was quite simple, “Reuse is hard...”  And that’s when I realized, that while reuse is an awesome goal and we should strive to make code maintainable, often times you end up creating far more work for yourself than necessary by trying to force code to be reusable that inherently isn’t. Think about classes the times you’ve worked in a company where in the design session people fight over the best way to implement a class to make it maximally reusable, extensible, and any other buzzwordable.  Then think about how quickly that design became obsolete.  Many times I set out to do a project and think, “yes, this is the best design, I can extend it easily!” only to find out the business requirements change COMPLETELY in such a way that the design is rendered invalid.  Code, in general, tends to rust and age over time.  As such, writing reusable code can often be difficult and many times ends up being a futile exercise and worse yet, sometimes makes the code harder to maintain because it obfuscates the design in the name of extensibility or reusability. So what do I think are reusable components? Generic Utility classes – these tend to be small classes that assist in a task and have no business context whatsoever. Implementation Abstraction Frameworks – home-grown frameworks that try to isolate changes to third party products you may be depending on (like writing a messaging abstraction layer for publishing/subscribing that is independent of whether you use JMS, MSMQ, etc). Simplification and Uniformity Frameworks – To some extent this is similar to an abstraction framework, but there may be one chosen provider but a development shop mandate to perform certain complex items in a certain way.  Or, perhaps to simplify and dumb-down a complex task for the average developer (such as implementing a particular development-shop’s method of encryption). And what are less reusable? Application and Business Layers – tend to fluctuate a lot as requirements change and new features are added, so tend to be an unstable dependency.  May be reused across applications but also very volatile. Entities and Data Access Layers – these tend to be tuned to the scope of the application, so reusing them can be hard unless the abstract is very stable. So what’s the big lesson?  Reuse is hard.  In fact it’s damn hard.  And much of the time I’m not convinced we should focus too hard on it. If you’re designing a utility or framework, then by all means design it for reuse.  But you most also really set down a good versioning, release, and documentation process to maximize your chances.  For anything else, design it to be maintainable and extendable, but don’t waste the effort on reusability for something that most likely will be obsolete in a year or two anyway.

    Read the article

  • Introduction to LinqPad Driver for StreamInsight 2.1

    - by Roman Schindlauer
    We are announcing the availability of the LinqPad driver for StreamInsight 2.1. The purpose of this blog post is to offer a quick introduction into the new features that we added to the StreamInsight LinqPad driver. We’ll show you how to connect to a remote server, how to inspect the entities present of that server, how to compose on top of them and how to manage their lifetime. Installing the driver Info on how to install the driver can be found in an earlier blog post here. Establishing connections As you click on the “Add Connection” link in the left pane you will notice that now it’s possible to build the data context automatically. The new driver appears as an option in the upper list, and if you pick it you will open a connection dialog that lets you connect to a remote StreamInsight server. The connection dialog lets you specify the address of the remote server. You will notice that it’s possible to pick up the binding information from the configuration file of the LinqPad application (which is normally in the same folder as LinqPad.exe and is called LinqPad.exe.config). In order for the context to be generated you need to pick an application from the server. The control is editable hence you can create a new application if you don’t want to make changes to an existing application. If you choose a new application name you will be prompted for confirmation before this gets created. Once you click OK the connection is created and you can start issuing queries against the remote server. If there’s any connectivity error the connection is marked with a red X and you can see the error message informing you what went wrong (i.e., the remote server could not be reached etc.). The context for remote servers Let’s take a look at what happens after we are connected successfully. Every LinqPad query runs inside a context – think of it as a class that wraps all the code that you’re writing. If you’re connecting to a live server the context will contain the following: The application object itself. All entities present in this application (sources, sinks, subjects and processes). The picture below shows a snapshot of the left pane of LinqPad after a successful connection. Every entity on the server has a different icon which will allow users to figure out its purpose. You will also notice that some entities have a string in parentheses following the name. It should be interpreted as such: the first name is the name of the property of the context class and the second name is the name of the entity as it exists on the server. Not all valid entity names are valid identifier names so in cases where we had to make a transformation you see both. Note also that as you hover over the entities you get IntelliSense with their types – more on that later. Remoting is not supported As you play with the entities exposed by the context you will notice that you can’t read and write directly to/from them. If for instance you’re trying to dump the content of an entity you will get an error message telling you that in the current version remoting is not supported. This is because the entity lives on the remote server and dumping its content means reading the events produced by this entity into the local process. ObservableSource.Dump(); Will yield the following error: Reading from a remote 'System.Reactive.Linq.IQbservable`1[System.Int32]' is not supported. Use the 'Microsoft.ComplexEventProcessing.Linq.RemoteProvider.Bind' method to read from the source using a remote observer. This basically tells you that you can call the Bind() method to direct the output of this source to a sink that has to be defined on the remote machine as well. You can’t bring the results to the LinqPad window unless you write code specifically for that. Compose queries You may ask – what's the purpose of all that? After all the same information is present in the EventFlowDebugger, why bother with showing it in LinqPad? First of all, What gets exposed in LinqPad is not what you see in the debugger. In LinqPad we have a property on the context class for every entity that lives on the server. Because LinqPad offers IntelliSense we in fact have much more information about the entity, and more importantly we can compose with that entity very easily. For example, let’s say that this code creates an entity: using (var server = Server.Connect(...)) {     var a = server.CreateApplication("WhiteFish");     var src = a         .DefineObservable<int>(() => Observable.Range(0, 3))         .Deploy("ObservableSource"); If later we want to compose with the source we have to fetch it and then we can bind something to     a.GetObservable<int>("ObservableSource)").Bind(... This means that we had to know a bunch of things about this: that it’s a source, that it’s an observable, it produces a result with payload Int32 and it’s named “ObservableSource”. Only the second and last bits of information are present in the debugger, by the way. As you type in the query window you see that all the entities are present, you get IntelliSense support for them and it’s much easier to make sense of what’s available. Let’s look at a scenario where composition is plausible. With the new programming model it’s possible to create “cold” sources that are parameterized. There was a way to accomplish that even in the previous version by passing parameters to the adapters, but this time it’s much more elegant because the expression declares what parameters are required. Say that we hover the mouse over the ThrottledSource source – we will see that its type is Func<int, int, IQbservable<int>> - this in effect means that we need to pass two int parameters before we can get a source that produces events, and the type for those events is int – in the particular case of my example I had the source produce a range of integers and the two parameters were the start and end of the range. So we see how a developer can create a source that is not running yet. Then someone else (e.g. an administrator) can pass whatever parameters appropriate and run the process. Proxy Types Here’s an interesting scenario – what if someone created a source on a server but they forgot to tell you what type they used. Worse yet, they might have used an anonymous type and even though they can refer to it by name you can’t figure out how to use that type. Let’s walk through an example that shows how you can compose against types you don’t need to have the definition of. This is how we can create a source that returns an anonymous type: Application.DefineObservable(() => Observable.Range(1, 10).Select(i => new { I = i })).Deploy("O1"); Now if we refresh the connection we can see the new source named O1 appear in the list. But what’s more important is that we now have a type to work with. So we can compose a query that refers to the anonymous type. var threshold = new StreamInsightDynamicDriver.TypeProxies.AnonymousType1_0<int>(5); var filter = from i in O1              where i > threshold              select i; filter.Deploy("O2"); You will notice that the anonymous type defined with this statement: new { I = i } can now be manipulated by a client that does not have access to it because the LinqPad driver has generated another type in its stead, named StreamInsightDynamicDriver.TypeProxies.AnonymousType1_0. This type has all the properties and fields of the type defined on the server, except in this case we can instantiate values and use it to compose more queries. It is worth noting that the same thing works for types that are not anonymous – the test is if the LinqPad driver can resolve the type or not. If it’s not possible then a new type will be generated that approximates the type that exists on the server. Control metadata In addition to composing processes on top of the existing entities we can do other useful things. We can delete them – nothing new here as we simply access the entities through the Entities collection of the application class. Here is where having their real name in parentheses comes handy. There’s another way to find out what’s behind a property – dump its expression. The first line in the output tells us what’s the name of the entity used to build this property in the context. Runtime information So let’s create a process to see what happens. We can bind a source to a sink and run the resulting process. If you right click on the connection you can refresh it and see the process present in the list of entities. Then you can drag the process to the query window and see that you can have access to process object in the Processes collection of the application. You can then manipulate the process (delete it, read its diagnostic view etc.). Regards, The StreamInsight Team

    Read the article

  • Constant game speed independent of variable FPS in OpenGL with GLUT?

    - by Nazgulled
    I've been reading Koen Witters detailed article about different game loop solutions but I'm having some problems implementing the last one with GLUT, which is the recommended one. After reading a couple of articles, tutorials and code from other people on how to achieve a constant game speed, I think that what I currently have implemented (I'll post the code below) is what Koen Witters called Game Speed dependent on Variable FPS, the second on his article. First, through my searching experience, there's a couple of people that probably have the knowledge to help out on this but don't know what GLUT is and I'm going to try and explain (feel free to correct me) the relevant functions for my problem of this OpenGL toolkit. Skip this section if you know what GLUT is and how to play with it. GLUT Toolkit: GLUT is an OpenGL toolkit and helps with common tasks in OpenGL. The glutDisplayFunc(renderScene) takes a pointer to a renderScene() function callback, which will be responsible for rendering everything. The renderScene() function will only be called once after the callback registration. The glutTimerFunc(TIMER_MILLISECONDS, processAnimationTimer, 0) takes the number of milliseconds to pass before calling the callback processAnimationTimer(). The last argument is just a value to pass to the timer callback. The processAnimationTimer() will not be called each TIMER_MILLISECONDS but just once. The glutPostRedisplay() function requests GLUT to render a new frame so we need call this every time we change something in the scene. The glutIdleFunc(renderScene) could be used to register a callback to renderScene() (this does not make glutDisplayFunc() irrelevant) but this function should be avoided because the idle callback is continuously called when events are not being received, increasing the CPU load. The glutGet(GLUT_ELAPSED_TIME) function returns the number of milliseconds since glutInit was called (or first call to glutGet(GLUT_ELAPSED_TIME)). That's the timer we have with GLUT. I know there are better alternatives for high resolution timers, but let's keep with this one for now. I think this is enough information on how GLUT renders frames so people that didn't know about it could also pitch in this question to try and help if they fell like it. Current Implementation: Now, I'm not sure I have correctly implemented the second solution proposed by Koen, Game Speed dependent on Variable FPS. The relevant code for that goes like this: #define TICKS_PER_SECOND 30 #define MOVEMENT_SPEED 2.0f const int TIMER_MILLISECONDS = 1000 / TICKS_PER_SECOND; int previousTime; int currentTime; int elapsedTime; void renderScene(void) { (...) // Setup the camera position and looking point SceneCamera.LookAt(); // Do all drawing below... (...) } void processAnimationTimer(int value) { // setups the timer to be called again glutTimerFunc(TIMER_MILLISECONDS, processAnimationTimer, 0); // Get the time when the previous frame was rendered previousTime = currentTime; // Get the current time (in milliseconds) and calculate the elapsed time currentTime = glutGet(GLUT_ELAPSED_TIME); elapsedTime = currentTime - previousTime; /* Multiply the camera direction vector by constant speed then by the elapsed time (in seconds) and then move the camera */ SceneCamera.Move(cameraDirection * MOVEMENT_SPEED * (elapsedTime / 1000.0f)); // Requests to render a new frame (this will call my renderScene() once) glutPostRedisplay(); } void main(int argc, char **argv) { glutInit(&argc, argv); (...) glutDisplayFunc(renderScene); (...) // Setup the timer to be called one first time glutTimerFunc(TIMER_MILLISECONDS, processAnimationTimer, 0); // Read the current time since glutInit was called currentTime = glutGet(GLUT_ELAPSED_TIME); glutMainLoop(); } This implementation doesn't fell right. It works in the sense that helps the game speed to be constant dependent on the FPS. So that moving from point A to point B takes the same time no matter the high/low framerate. However, I believe I'm limiting the game framerate with this approach. Each frame will only be rendered when the time callback is called, that means the framerate will be roughly around TICKS_PER_SECOND frames per second. This doesn't feel right, you shouldn't limit your powerful hardware, it's wrong. It's my understanding though, that I still need to calculate the elapsedTime. Just because I'm telling GLUT to call the timer callback every TIMER_MILLISECONDS, it doesn't mean it will always do that on time. I'm not sure how can I fix this and to be completely honest, I have no idea what is the game loop in GLUT, you know, the while( game_is_running ) loop in Koen's article. But it's my understanding that GLUT is event-driven and that game loop starts when I call glutMainLoop() (which never returns), yes? I thought I could register an idle callback with glutIdleFunc() and use that as replacement of glutTimerFunc(), only rendering when necessary (instead of all the time as usual) but when I tested this with an empty callback (like void gameLoop() {}) and it was basically doing nothing, only a black screen, the CPU spiked to 25% and remained there until I killed the game and it went back to normal. So I don't think that's the path to follow. Using glutTimerFunc() is definitely not a good approach to perform all movements/animations based on that, as I'm limiting my game to a constant FPS, not cool. Or maybe I'm using it wrong and my implementation is not right? How exactly can I have a constant game speed with variable FPS? More exactly, how do I correctly implement Koen's Constant Game Speed with Maximum FPS solution (the fourth one on his article) with GLUT? Maybe this is not possible at all with GLUT? If not, what are my alternatives? What is the best approach to this problem (constant game speed) with GLUT? I originally posted this question on Stack Overflow before being pointed out about this site. The following is a different approach I tried after creating the question in SO, so I'm posting it here too. Another Approach: I've been experimenting and here's what I was able to achieve now. Instead of calculating the elapsed time on a timed function (which limits my game's framerate) I'm now doing it in renderScene(). Whenever changes to the scene happen I call glutPostRedisplay() (ie: camera moving, some object animation, etc...) which will make a call to renderScene(). I can use the elapsed time in this function to move my camera for instance. My code has now turned into this: int previousTime; int currentTime; int elapsedTime; void renderScene(void) { (...) // Setup the camera position and looking point SceneCamera.LookAt(); // Do all drawing below... (...) } void renderScene(void) { (...) // Get the time when the previous frame was rendered previousTime = currentTime; // Get the current time (in milliseconds) and calculate the elapsed time currentTime = glutGet(GLUT_ELAPSED_TIME); elapsedTime = currentTime - previousTime; /* Multiply the camera direction vector by constant speed then by the elapsed time (in seconds) and then move the camera */ SceneCamera.Move(cameraDirection * MOVEMENT_SPEED * (elapsedTime / 1000.0f)); // Setup the camera position and looking point SceneCamera.LookAt(); // All drawing code goes inside this function drawCompleteScene(); glutSwapBuffers(); /* Redraw the frame ONLY if the user is moving the camera (similar code will be needed to redraw the frame for other events) */ if(!IsTupleEmpty(cameraDirection)) { glutPostRedisplay(); } } void main(int argc, char **argv) { glutInit(&argc, argv); (...) glutDisplayFunc(renderScene); (...) currentTime = glutGet(GLUT_ELAPSED_TIME); glutMainLoop(); } Conclusion, it's working, or so it seems. If I don't move the camera, the CPU usage is low, nothing is being rendered (for testing purposes I only have a grid extending for 4000.0f, while zFar is set to 1000.0f). When I start moving the camera the scene starts redrawing itself. If I keep pressing the move keys, the CPU usage will increase; this is normal behavior. It drops back when I stop moving. Unless I'm missing something, it seems like a good approach for now. I did find this interesting article on iDevGames and this implementation is probably affected by the problem described on that article. What's your thoughts on that? Please note that I'm just doing this for fun, I have no intentions of creating some game to distribute or something like that, not in the near future at least. If I did, I would probably go with something else besides GLUT. But since I'm using GLUT, and other than the problem described on iDevGames, do you think this latest implementation is sufficient for GLUT? The only real issue I can think of right now is that I'll need to keep calling glutPostRedisplay() every time the scene changes something and keep calling it until there's nothing new to redraw. A little complexity added to the code for a better cause, I think. What do you think?

    Read the article

  • Clusterware 11gR2 &ndash; Setting up an Active/Passive failover configuration

    - by Gilles Haro
    Oracle provides many interesting ways to ensure High Availability. Dataguard configurations, RAC configurations or even both (as recommended for a Maximum Available Architecture - MAA) are the most frequently found. But when it comes to protecting a system with an Active/Passive architecture with failover capabilities, one often thinks to expensive third party cluster systems. Oracle Clusterware technology, which comes free with Oracle Database, is – in the knowing of most people - often linked to Oracle RAC and therefore, is rarely used to implement failover solutions. 11gR2 Clusterware – which is part of Oracle Grid Infrastructure - provides a comprehensive framework to setup automatic failover configurations. It is actually possible to make “failover-able'” and, therefore to protect, almost every kind of application (from xclock to the more complex Application Server) In the next couple of lines, I will try to present the different steps to achieve this goal : Have a fully operational 11gR2 database protected by automatic failover capabilities. I assume you are fluent in installing Oracle Database 11gR2, Oracle Grid Infrastructure 11gR2 on a Linux system and that ASM is not a problem for you (as I am using it as a shared storage). If not, please have a look at Oracle Documentation. As often, I made my tests using an Oracle VirtualBox environment. The scripts are tested and functional. Unfortunately, there can always be a typo or a mistake. This blog entry is not a course around the Clusterware Framework. I just hope it will let you see how powerful it is and that it will give you the whilst to go further with it…   Prerequisite 2 Linux boxes (OELCluster01 and OELCluster02) at the same OS level. I used OEL 5 Update 5 with Enterprise Kernel. Shared Storage (SAN). On my VirtualBox system, I used Openfiler to simulate the SAN Oracle 11gR2 Database (11.2.0.1) Oracle 11gR2 Grid Infrastructure (11.2.0.1)   Step 1 – Install the software Using asmlib, create 3 ASM disks (ASM_CRS, ASM_DTA and ASM_FRA) Install Grid Infrastructure for a cluster (OELCluster01 and OELCluster02 are the 2 nodes of the cluster) Use ASM_CRS to store Voting Disk and OCR. Use SCAN. Install Oracle Database Standalone binaries on both nodes. Use asmca to check/mount the disk groups on 2 nodes Use dbca to create and configure a database on the primary node Let’s name it DB11G. Copy the pfile, password file to the second node. Create adump directoty on the second node.   Step 2 - Setup the resource to be protected After its creation with dbca, the database is automatically protected by the Oracle Restart technology available with Grid Infrastructure. Consequently, it restarts automatically (if possible) after a crash (ex: kill –9 smon). A database resource has been created for that in the Cluster Registry. We can observe this with the command : crsctl status resource that shows and ora.dba11g.db entry. Let’s save the definition of this resource, for future use : mkdir –p /crs/11.2.0/HA_scripts chown oracle:oinstall /crs/11.2.0/HA_scripts crsctl status resource ora.db11g.db -p > /crs/11.2.0/HA_scripts/myResource.txt Although very interesting, Oracle Restart is not cluster aware and cannot restart the database on any other node of the cluster. So, let’s remove it from the OCR definitions, we don’t need it ! srvctl stop database -d DB11G srvctl remove database -d DB11G Instead of it, we need to create a new resource of a more general type : cluster_resource. Here are the steps to achieve this : Create an action script :  /crs/11.2.0/HA_scripts/my_ActivePassive_Cluster.sh #!/bin/bash export ORACLE_HOME=/oracle/product/11.2.0/dbhome_1 export ORACLE_SID=DB11G case $1 in 'start')   $ORACLE_HOME/bin/sqlplus /nolog <<EOF   connect / as sysdba   startup EOF   RET=0   ;; 'stop')   $ORACLE_HOME/bin/sqlplus /nolog <<EOF   connect / as sysdba   shutdown immediate EOF   RET=0   ;; 'check')    ok=`ps -ef | grep smon | grep $ORACLE_SID | wc -l`    if [ $ok = 0 ]; then      RET=1    else      RET=0    fi    ;; '*')      RET=0   ;; esac if [ $RET -eq 0 ]; then    exit 0 else    exit 1 fi   This script must provide, at least, methods to start, stop and check the database. It is self-explaining and contains nothing special. Just be aware that it is run as Oracle user (because of the ACL property – see later) and needs to know about the environment. It also needs to be present on every node of the cluster. chmod +x /crs/11.2.0/HA_scripts/my_ActivePassive_Cluster.sh scp  /crs/11.2.0/HA_scripts/my_ActivePassive_Cluster.sh   oracle@OELCluster02:/crs/11.2.0/HA_scripts Create a new resource file, based on the information we got from previous  myResource.txt . Name it myNewResource.txt. myResource.txt  is shown below. As we can see, it defines an ora.database.type resource, named ora.db11g.db. A lot of properties are related to this type of resource and do not need to be used for a cluster_resource. NAME=ora.db11g.db TYPE=ora.database.type ACL=owner:oracle:rwx,pgrp:oinstall:rwx,other::r-- ACTION_FAILURE_TEMPLATE= ACTION_SCRIPT= ACTIVE_PLACEMENT=1 AGENT_FILENAME=%CRS_HOME%/bin/oraagent%CRS_EXE_SUFFIX% AUTO_START=restore CARDINALITY=1 CHECK_INTERVAL=1 CHECK_TIMEOUT=600 CLUSTER_DATABASE=false DB_UNIQUE_NAME=DB11G DEFAULT_TEMPLATE=PROPERTY(RESOURCE_CLASS=database) PROPERTY(DB_UNIQUE_NAME= CONCAT(PARSE(%NAME%, ., 2), %USR_ORA_DOMAIN%, .)) ELEMENT(INSTANCE_NAME= %GEN_USR_ORA_INST_NAME%) DEGREE=1 DESCRIPTION=Oracle Database resource ENABLED=1 FAILOVER_DELAY=0 FAILURE_INTERVAL=60 FAILURE_THRESHOLD=1 GEN_AUDIT_FILE_DEST=/oracle/admin/DB11G/adump GEN_USR_ORA_INST_NAME= GEN_USR_ORA_INST_NAME@SERVERNAME(oelcluster01)=DB11G HOSTING_MEMBERS= INSTANCE_FAILOVER=0 LOAD=1 LOGGING_LEVEL=1 MANAGEMENT_POLICY=AUTOMATIC NLS_LANG= NOT_RESTARTING_TEMPLATE= OFFLINE_CHECK_INTERVAL=0 ORACLE_HOME=/oracle/product/11.2.0/dbhome_1 PLACEMENT=restricted PROFILE_CHANGE_TEMPLATE= RESTART_ATTEMPTS=2 ROLE=PRIMARY SCRIPT_TIMEOUT=60 SERVER_POOLS=ora.DB11G SPFILE=+DTA/DB11G/spfileDB11G.ora START_DEPENDENCIES=hard(ora.DTA.dg,ora.FRA.dg) weak(type:ora.listener.type,uniform:ora.ons,uniform:ora.eons) pullup(ora.DTA.dg,ora.FRA.dg) START_TIMEOUT=600 STATE_CHANGE_TEMPLATE= STOP_DEPENDENCIES=hard(intermediate:ora.asm,shutdown:ora.DTA.dg,shutdown:ora.FRA.dg) STOP_TIMEOUT=600 UPTIME_THRESHOLD=1h USR_ORA_DB_NAME=DB11G USR_ORA_DOMAIN=haroland USR_ORA_ENV= USR_ORA_FLAGS= USR_ORA_INST_NAME=DB11G USR_ORA_OPEN_MODE=open USR_ORA_OPI=false USR_ORA_STOP_MODE=immediate VERSION=11.2.0.1.0 I removed database type related entries from myResource.txt and modified some other to produce the following myNewResource.txt. Notice the NAME property that should not have the ora. prefix Notice the TYPE property that is not ora.database.type but cluster_resource. Notice the definition of ACTION_SCRIPT. Notice the HOSTING_MEMBERS that enumerates the members of the cluster (as returned by the olsnodes command). NAME=DB11G.db TYPE=cluster_resource DESCRIPTION=Oracle Database resource ACL=owner:oracle:rwx,pgrp:oinstall:rwx,other::r-- ACTION_SCRIPT=/crs/11.2.0/HA_scripts/my_ActivePassive_Cluster.sh PLACEMENT=restricted ACTIVE_PLACEMENT=0 AUTO_START=restore CARDINALITY=1 CHECK_INTERVAL=10 DEGREE=1 ENABLED=1 HOSTING_MEMBERS=oelcluster01 oelcluster02 LOGGING_LEVEL=1 RESTART_ATTEMPTS=1 START_DEPENDENCIES=hard(ora.DTA.dg,ora.FRA.dg) weak(type:ora.listener.type,uniform:ora.ons,uniform:ora.eons) pullup(ora.DTA.dg,ora.FRA.dg) START_TIMEOUT=600 STOP_DEPENDENCIES=hard(intermediate:ora.asm,shutdown:ora.DTA.dg,shutdown:ora.FRA.dg) STOP_TIMEOUT=600 UPTIME_THRESHOLD=1h Register the resource. Take care of the resource type. It needs to be a cluster_resource and not a ora.database.type resource (Oracle recommendation) .   crsctl add resource DB11G.db  -type cluster_resource -file /crs/11.2.0/HA_scripts/myNewResource.txt Step 3 - Start the resource crsctl start resource DB11G.db This command launches the ACTION_SCRIPT with a start and a check parameter on the primary node of the cluster. Step 4 - Test this We will test the setup using 2 methods. crsctl relocate resource DB11G.db This command calls the ACTION_SCRIPT  (on the two nodes)  to stop the database on the active node and start it on the other node. Once done, we can revert back to the original node, but, this time we can use a more “MS$ like” method :Turn off the server on which the database is running. After short delay, you should observe that the database is relocated on node 1. Conclusion Once the software installed and the standalone database created (which is a rather common and usual task), the steps to reach the objective are quite easy : Create an executable action script on every node of the cluster. Create a resource file. Create/Register the resource with OCR using the resource file. Start the resource. This solution is a very interesting alternative to licensable third party solutions.   References Clusterware 11gR2 documentation Oracle Clusterware Resource Reference   Gilles Haro Technical Expert - Core Technology, Oracle Consulting   

    Read the article

  • How Visual Studio 2010 and Team Foundation Server enable Compliance

    - by Martin Hinshelwood
    One of the things that makes Team Foundation Server (TFS) the most powerful Application Lifecycle Management (ALM) platform is the traceability it provides to those that use it. This traceability is crucial to enable many companies to adhere to many of the Compliance regulations to which they are bound (e.g. CFR 21 Part 11 or Sarbanes–Oxley.)   From something as simple as relating Tasks to Check-in’s or being able to see the top 10 files in your codebase that are causing the most Bugs, to identifying which Bugs and Requirements are in which Release. All that information is available and more in TFS. Although all of this tradability is available within TFS you do need to understand that it is not for free. Well… I say that, but if you are using TFS properly you will have this information with no additional work except for firing up the reporting. Using Visual Studio ALM and Team Foundation Server you can relate every line of code changes all the way up to requirements and back down through Test Cases to the Test Results. Figure: The only thing missing is Build In order to build the relationship model below we need to examine how each of the relationships get there. Each member of your team from programmer to tester and Business Analyst to Business have their roll to play to knit this together. Figure: The relationships required to make this work can get a little confusing If Build is added to this to relate Work Items to Builds and with knowledge of which builds are in which environments you can easily identify what is contained within a Release. Figure: How are things progressing Along with the ability to produce the progress and trend reports the tractability that is built into TFS can be used to fulfil most audit requirements out of the box, and augmented to fulfil the rest. In order to understand the relationships, lets look at each of the important Artifacts and how they are associated with each other… Requirements – The root of all knowledge Requirements are the thing that the business cares about delivering. These could be derived as User Stories or Business Requirements Documents (BRD’s) but they should be what the Business asks for. Requirements can be related to many of the Artifacts in TFS, so lets look at the model: Figure: If the centre of the world was a requirement We can track which releases Requirements were scheduled in, but this can change over time as more details come to light. Figure: Who edited the Requirement and when There is also the ability to query Work Items based on the History of changed that were made to it. This is particularly important with Requirements. It might not be enough to say what Requirements were completed in a given but also to know which Requirements were ever assigned to a particular release. Figure: Some magic required, but result still achieved As an augmentation to this it is also possible to run a query that shows results from the past, just as if we had a time machine. You can take any Query in the system and add a “Asof” clause at the end to query historical data in the operational store for TFS. select <fields> from WorkItems [where <condition>] [order by <fields>] [asof <date>] Figure: Work Item Query Language (WIQL) format In order to achieve this you do need to save the query as a *.wiql file to your local computer and edit it in notepad, but one imported into TFS you run it any time you want. Figure: Saving Queries locally can be useful All of these Audit features are available throughout the Work Item Tracking (WIT) system within TFS. Tasks – Where the real work gets done Tasks are the work horse of the development team, but they only as useful as Excel if you do not relate them properly to other Artifacts. Figure: The Task Work Item Type has its own relationships Requirements should be broken down into Tasks that the development team work from to build what is required by the business. This may be done by a small dedicated group or by everyone that will be working on the software team but however it happens all of the Tasks create should be a Child of a Requirement Work Item Type. Figure: Tasks are related to the Requirement Tasks should be used to track the day-to-day activities of the team working to complete the software and as such they should be kept simple and short lest developers think they are more trouble than they are worth. Figure: Task Work Item Type has a narrower purpose Although the Task Work Item Type describes the work that will be done the actual development work involves making changes to files that are under Source Control. These changes are bundled together in a single atomic unit called a Changeset which is committed to TFS in a single operation. During this operation developers can associate Work Item with the Changeset. Figure: Tasks are associated with Changesets   Changesets – Who wrote this crap Changesets themselves are just an inventory of the changes that were made to a number of files to complete a Task. Figure: Changesets are linked by Tasks and Builds   Figure: Changesets tell us what happened to the files in Version Control Although comments can be changed after the fact, the inventory and Work Item associations are permanent which allows us to Audit all the way down to the individual change level. Figure: On Check-in you can resolve a Task which automatically associates it Because of this we can view the history on any file within the system and see how many changes have been made and what Changesets they belong to. Figure: Changes are tracked at the File level What would be even more powerful would be if we could view these changes super imposed over the top of the lines of code. Some people call this a blame tool because it is commonly used to find out which of the developers introduced a bug, but it can also be used as another method of Auditing changes to the system. Figure: Annotate shows the lines the Annotate functionality allows us to visualise the relationship between the individual lines of code and the Changesets. In addition to this you can create a Label and apply it to a version of your version control. The problem with Label’s is that they can be changed after they have been created with no tractability. This makes them practically useless for any sort of compliance audit. So what do you use? Branches – And why we need them Branches are a really powerful tool for development and release management, but they are most important for audits. Figure: One way to Audit releases The R1.0 branch can be created from the Label that the Build creates on the R1 line when a Release build was created. It can be created as soon as the Build has been signed of for release. However it is still possible that someone changed the Label between this time and its creation. Another better method can be to explicitly link the Build output to the Build. Builds – Lets tie some more of this together Builds are the glue that helps us enable the next level of tractability by tying everything together. Figure: The dashed pieces are not out of the box but can be enabled When the Build is called and starts it looks at what it has been asked to build and determines what code it is going to get and build. Figure: The folder identifies what changes are included in the build The Build sets a Label on the Source with the same name as the Build, but the Build itself also includes the latest Changeset ID that it will be building. At the end of the Build the Build Agent identifies the new Changesets it is building by looking at the Check-ins that have occurred since the last Build. Figure: What changes have been made since the last successful Build It will then use that information to identify the Work Items that are associated with all of the Changesets Changesets are associated with Build and change the “Integrated In” field of those Work Items . Figure: Find all of the Work Items to associate with The “Integrated In” field of all of the Work Items identified by the Build Agent as being integrated into the completed Build are updated to reflect the Build number that successfully integrated that change. Figure: Now we know which Work Items were completed in a build Now that we can link a single line of code changed all the way back through the Task that initiated the action to the Requirement that started the whole thing and back down to the Build that contains the finished Requirement. But how do we know wither that Requirement has been fully tested or even meets the original Requirements? Test Cases – How we know we are done The only way we can know wither a Requirement has been completed to the required specification is to Test that Requirement. In TFS there is a Work Item type called a Test Case Test Cases enable two scenarios. The first scenario is the ability to track and validate Acceptance Criteria in the form of a Test Case. If you agree with the Business a set of goals that must be met for a Requirement to be accepted by them it makes it both difficult for them to reject a Requirement when it passes all of the tests, but also provides a level of tractability and validation for audit that a feature has been built and tested to order. Figure: You can have many Acceptance Criteria for a single Requirement It is crucial for this to work that someone from the Business has to sign-off on the Test Case moving from the  “Design” to “Ready” states. The Second is the ability to associate an MS Test test with the Test Case thereby tracking the automated test. This is useful in the circumstance when you want to Track a test and the test results of a Unit Test designed to test the existence of and then re-existence of a a Bug. Figure: Associating a Test Case with an automated Test Although it is possible it may not make sense to track the execution of every Unit Test in your system, there are many Integration and Regression tests that may be automated that it would make sense to track in this way. Bug – Lets not have regressions In order to know wither a Bug in the application has been fixed and to make sure that it does not reoccur it needs to be tracked. Figure: Bugs are the centre of their own world If the fix to a Bug is big enough to require that it is broken down into Tasks then it is probably a Requirement. You can associate a check-in with a Bug and have it tracked against a Build. You would also have one or more Test Cases to prove the fix for the Bug. Figure: Bugs have many associations This allows you to track Bugs / Defects in your system effectively and report on them. Change Request – I am not a feature In the CMMI Process template Change Requests can also be easily tracked through the system. In some cases it can be very important to track Change Requests separately as an Auditor may want to know what was changed and who authorised it. Again and similar to Bugs, if the Change Request is big enough that it would require to be broken down into Tasks it is in reality a new feature and should be tracked as a Requirement. Figure: Make sure your Change Requests only Affect Requirements and not rewrite them Conclusion Visual Studio 2010 and Team Foundation Server together provide an exceptional Application Lifecycle Management platform that can help your team comply with even the harshest of Compliance requirements while still enabling them to be Agile. Most Audits are heavy on required documentation but most of that information is captured for you as long a you do it right. You don’t even need every team member to understand it all as each of the Artifacts are relevant to a different type of team member. Business Analysts manage Requirements and Change Requests Programmers manage Tasks and check-in against Change Requests and Bugs Testers manage Bugs and Test Cases Build Masters manage Builds Although there is some crossover there are still rolls or “hats” that are worn. Do you thing this is all achievable? Have I missed anything that you think should be there?

    Read the article

  • evaluating a code of a graph [migrated]

    - by mazen.r.f
    This is relatively a long code,if you have the tolerance and the will to find out how to make this code work then take a look please, i will appreciate your feed back. i have spent two days trying to come up with a code to represent a graph , then calculate the shortest path using dijkastra algorithm , but i am not able to get the right result , even the code runs without errors , but the result is not correct , always i am getting 0. briefly,i have three classes , Vertex, Edge, Graph , the Vertex class represents the nodes in the graph and it has id and carried ( which carry the weight of the links connected to it while using dijkastra algorithm ) and a vector of the ids belong to other nodes the path will go through before arriving to the node itself , this vector is named previous_nodes. the Edge class represents the edges in the graph it has two vertices ( one in each side ) and a wight ( the distance between the two vertices ). the Graph class represents the graph , it has two vectors one is the vertices included in this graph , and the other is the edges included in the graph. inside the class Graph there is a method its name shortest takes the sources node id and the destination and calculates the shortest path using dijkastra algorithm, and i think that it is the most important part of the code. my theory about the code is that i will create two vectors one for the vertices in the graph i will name it vertices and another vector its name is ver_out it will include the vertices out of calculation in the graph, also i will have two vectors of type Edge , one its name edges for all the edges in the graph and the other its name is track to contain temporarily the edges linked to the temporarily source node in every round , after the calculation of every round the vector track will be cleared. in main() i created five vertices and 10 edges to simulate a graph , the result of the shortest path supposedly to be 4 , but i am always getting 0 , that means i am having something wrong in my code , so if you are interesting in helping me find my mistake and how to make the code work , please take a look. the way shortest work is as follow at the beginning all the edges will be included in the vector edges , we select the edges related to the source and put them in the vector track , then we iterate through track and add the wight of every edge to the vertex (node ) related to it ( not the source vertex ) , then after we clear track and remove the source vertex from the vector vertices and select a new source , and start over again select the edges related to the new source , put them in track , iterate over edges in tack , adding the weights to the corresponding vertices then remove this vertex from the vector vertices, and clear track , and select a new source , and so on . here is the code. #include<iostream> #include<vector> #include <stdlib.h> // for rand() using namespace std; class Vertex { private: unsigned int id; // the name of the vertex unsigned int carried; // the weight a vertex may carry when calculating shortest path vector<unsigned int> previous_nodes; public: unsigned int get_id(){return id;}; unsigned int get_carried(){return carried;}; void set_id(unsigned int value) {id = value;}; void set_carried(unsigned int value) {carried = value;}; void previous_nodes_update(unsigned int val){previous_nodes.push_back(val);}; void previous_nodes_erase(unsigned int val){previous_nodes.erase(previous_nodes.begin() + val);}; Vertex(unsigned int init_val = 0, unsigned int init_carried = 0) :id (init_val), carried(init_carried) // constructor { } ~Vertex() {}; // destructor }; class Edge { private: Vertex first_vertex; // a vertex on one side of the edge Vertex second_vertex; // a vertex on the other side of the edge unsigned int weight; // the value of the edge ( or its weight ) public: unsigned int get_weight() {return weight;}; void set_weight(unsigned int value) {weight = value;}; Vertex get_ver_1(){return first_vertex;}; Vertex get_ver_2(){return second_vertex;}; void set_first_vertex(Vertex v1) {first_vertex = v1;}; void set_second_vertex(Vertex v2) {second_vertex = v2;}; Edge(const Vertex& vertex_1 = 0, const Vertex& vertex_2 = 0, unsigned int init_weight = 0) : first_vertex(vertex_1), second_vertex(vertex_2), weight(init_weight) { } ~Edge() {} ; // destructor }; class Graph { private: std::vector<Vertex> vertices; std::vector<Edge> edges; public: Graph(vector<Vertex> ver_vector, vector<Edge> edg_vector) : vertices(ver_vector), edges(edg_vector) { } ~Graph() {}; vector<Vertex> get_vertices(){return vertices;}; vector<Edge> get_edges(){return edges;}; void set_vertices(vector<Vertex> vector_value) {vertices = vector_value;}; void set_edges(vector<Edge> vector_ed_value) {edges = vector_ed_value;}; unsigned int shortest(unsigned int src, unsigned int dis) { vector<Vertex> ver_out; vector<Edge> track; for(unsigned int i = 0; i < edges.size(); ++i) { if((edges[i].get_ver_1().get_id() == vertices[src].get_id()) || (edges[i].get_ver_2().get_id() == vertices[src].get_id())) { track.push_back (edges[i]); edges.erase(edges.begin()+i); } }; for(unsigned int i = 0; i < track.size(); ++i) { if(track[i].get_ver_1().get_id() != vertices[src].get_id()) { track[i].get_ver_1().set_carried((track[i].get_weight()) + track[i].get_ver_2().get_carried()); track[i].get_ver_1().previous_nodes_update(vertices[src].get_id()); } else { track[i].get_ver_2().set_carried((track[i].get_weight()) + track[i].get_ver_1().get_carried()); track[i].get_ver_2().previous_nodes_update(vertices[src].get_id()); } } for(unsigned int i = 0; i < vertices.size(); ++i) if(vertices[i].get_id() == src) vertices.erase(vertices.begin() + i); // removing the sources vertex from the vertices vector ver_out.push_back (vertices[src]); track.clear(); if(vertices[0].get_id() != dis) {src = vertices[0].get_id();} else {src = vertices[1].get_id();} for(unsigned int i = 0; i < vertices.size(); ++i) if((vertices[i].get_carried() < vertices[src].get_carried()) && (vertices[i].get_id() != dis)) src = vertices[i].get_id(); //while(!edges.empty()) for(unsigned int round = 0; round < vertices.size(); ++round) { for(unsigned int k = 0; k < edges.size(); ++k) { if((edges[k].get_ver_1().get_id() == vertices[src].get_id()) || (edges[k].get_ver_2().get_id() == vertices[src].get_id())) { track.push_back (edges[k]); edges.erase(edges.begin()+k); } }; for(unsigned int n = 0; n < track.size(); ++n) if((track[n].get_ver_1().get_id() != vertices[src].get_id()) && (track[n].get_ver_1().get_carried() > (track[n].get_ver_2().get_carried() + track[n].get_weight()))) { track[n].get_ver_1().set_carried((track[n].get_weight()) + track[n].get_ver_2().get_carried()); track[n].get_ver_1().previous_nodes_update(vertices[src].get_id()); } else if(track[n].get_ver_2().get_carried() > (track[n].get_ver_1().get_carried() + track[n].get_weight())) { track[n].get_ver_2().set_carried((track[n].get_weight()) + track[n].get_ver_1().get_carried()); track[n].get_ver_2().previous_nodes_update(vertices[src].get_id()); } for(unsigned int t = 0; t < vertices.size(); ++t) if(vertices[t].get_id() == src) vertices.erase(vertices.begin() + t); track.clear(); if(vertices[0].get_id() != dis) {src = vertices[0].get_id();} else {src = vertices[1].get_id();} for(unsigned int tt = 0; tt < edges.size(); ++tt) { if(vertices[tt].get_carried() < vertices[src].get_carried()) { src = vertices[tt].get_id(); } } } return vertices[dis].get_carried(); } }; int main() { cout<< "Hello, This is a graph"<< endl; vector<Vertex> vers(5); vers[0].set_id(0); vers[1].set_id(1); vers[2].set_id(2); vers[3].set_id(3); vers[4].set_id(4); vector<Edge> eds(10); eds[0].set_first_vertex(vers[0]); eds[0].set_second_vertex(vers[1]); eds[0].set_weight(5); eds[1].set_first_vertex(vers[0]); eds[1].set_second_vertex(vers[2]); eds[1].set_weight(9); eds[2].set_first_vertex(vers[0]); eds[2].set_second_vertex(vers[3]); eds[2].set_weight(4); eds[3].set_first_vertex(vers[0]); eds[3].set_second_vertex(vers[4]); eds[3].set_weight(6); eds[4].set_first_vertex(vers[1]); eds[4].set_second_vertex(vers[2]); eds[4].set_weight(2); eds[5].set_first_vertex(vers[1]); eds[5].set_second_vertex(vers[3]); eds[5].set_weight(5); eds[6].set_first_vertex(vers[1]); eds[6].set_second_vertex(vers[4]); eds[6].set_weight(7); eds[7].set_first_vertex(vers[2]); eds[7].set_second_vertex(vers[3]); eds[7].set_weight(1); eds[8].set_first_vertex(vers[2]); eds[8].set_second_vertex(vers[4]); eds[8].set_weight(8); eds[9].set_first_vertex(vers[3]); eds[9].set_second_vertex(vers[4]); eds[9].set_weight(3); unsigned int path; Graph graf(vers, eds); path = graf.shortest(2, 4); cout<< path << endl; return 0; }

    Read the article

  • Why Are Minimized Programs Often Slow to Open Again?

    - by Jason Fitzpatrick
    It seems particularly counterintuitive: you minimize an application because you plan on returning to it later and wish to skip shutting the application down and restarting it later, but sometimes maximizing it takes even longer than launching it fresh. What gives? Today’s Question & Answer session comes to us courtesy of SuperUser—a subdivision of Stack Exchange, a community-driven grouping of Q&A web sites. The Question SuperUser reader Bart wants to know why he’s not saving any time with application minimization: I’m working in Photoshop CS6 and multiple browsers a lot. I’m not using them all at once, so sometimes some applications are minimized to taskbar for hours or days. The problem is, when I try to maximize them from the taskbar – it sometimes takes longer than starting them! Especially Photoshop feels really weird for many seconds after finally showing up, it’s slow, unresponsive and even sometimes totally freezes for minute or two. It’s not a hardware problem as it’s been like that since always on all on my PCs. Would I also notice it after upgrading my HDD to SDD and adding RAM (my main PC holds 4 GB currently)? Could guys with powerful pcs / macs tell me – does it also happen to you? I guess OSes somehow “focus” on active software and move all the resources away from the ones that run, but are not used. Is it possible to somehow set RAM / CPU / HDD priorities or something, for let’s say, Photoshop, so it won’t slow down after long period of inactivity? So what is the deal? Why does he find himself waiting to maximize a minimized app? The Answer SuperUser contributor Allquixotic explains why: Summary The immediate problem is that the programs that you have minimized are being paged out to the “page file” on your hard disk. This symptom can be improved by installing a Solid State Disk (SSD), adding more RAM to your system, reducing the number of programs you have open, or upgrading to a newer system architecture (for instance, Ivy Bridge or Haswell). Out of these options, adding more RAM is generally the most effective solution. Explanation The default behavior of Windows is to give active applications priority over inactive applications for having a spot in RAM. When there’s significant memory pressure (meaning the system doesn’t have a lot of free RAM if it were to let every program have all the RAM it wants), it starts putting minimized programs into the page file, which means it writes out their contents from RAM to disk, and then makes that area of RAM free. That free RAM helps programs you’re actively using — say, your web browser — run faster, because if they need to claim a new segment of RAM (like when you open a new tab), they can do so. This “free” RAM is also used as page cache, which means that when active programs attempt to read data on your hard disk, that data might be cached in RAM, which prevents your hard disk from being accessed to get that data. By using the majority of your RAM for page cache, and swapping out unused programs to disk, Windows is trying to improve responsiveness of the program(s) you are actively using, by making RAM available to them, and caching the files they access in RAM instead of the hard disk. The downside of this behavior is that minimized programs can take a while to have their contents copied from the page file, on disk, back into RAM. The time increases the larger the program’s footprint in memory. This is why you experience that delay when maximizing Photoshop. RAM is many times faster than a hard disk (depending on the specific hardware, it can be up to several orders of magnitude). An SSD is considerably faster than a hard disk, but it is still slower than RAM by orders of magnitude. Having your page file on an SSD will help, but it will also wear out the SSD more quickly than usual if your page file is heavily utilized due to RAM pressure. Remedies Here is an explanation of the available remedies, and their general effectiveness: Installing more RAM: This is the recommended path. If your system does not support more RAM than you already have installed, you will need to upgrade more of your system: possibly your motherboard, CPU, chassis, power supply, etc. depending on how old it is. If it’s a laptop, chances are you’ll have to buy an entire new laptop that supports more installed RAM. When you install more RAM, you reduce memory pressure, which reduces use of the page file, which is a good thing all around. You also make available more RAM for page cache, which will make all programs that access the hard disk run faster. As of Q4 2013, my personal recommendation is that you have at least 8 GB of RAM for a desktop or laptop whose purpose is anything more complex than web browsing and email. That means photo editing, video editing/viewing, playing computer games, audio editing or recording, programming / development, etc. all should have at least 8 GB of RAM, if not more. Run fewer programs at a time: This will only work if the programs you are running do not use a lot of memory on their own. Unfortunately, Adobe Creative Suite products such as Photoshop CS6 are known for using an enormous amount of memory. This also limits your multitasking ability. It’s a temporary, free remedy, but it can be an inconvenience to close down your web browser or Word every time you start Photoshop, for instance. This also wouldn’t stop Photoshop from being swapped when minimizing it, so it really isn’t a very effective solution. It only helps in some specific situations. Install an SSD: If your page file is on an SSD, the SSD’s improved speed compared to a hard disk will result in generally improved performance when the page file has to be read from or written to. Be aware that SSDs are not designed to withstand a very frequent and constant random stream of writes; they can only be written over a limited number of times before they start to break down. Heavy use of a page file is not a particularly good workload for an SSD. You should install an SSD in combination with a large amount of RAM if you want maximum performance while preserving the longevity of the SSD. Use a newer system architecture: Depending on the age of your system, you may be using an out of date system architecture. The “system architecture” is generally defined as the “generation” (think generations like children, parents, grandparents, etc.) of the motherboard and CPU. Newer generations generally support faster I/O (input/output), better memory bandwidth, lower latency, and less contention over shared resources, instead providing dedicated links between components. For example, starting with the “Nehalem” generation (around 2009), the Front-Side Bus (FSB) was eliminated, which removed a common bottleneck, because almost all system components had to share the same FSB for transmitting data. This was replaced with a “point to point” architecture, meaning that each component gets its own dedicated “lane” to the CPU, which continues to be improved every few years with new generations. You will generally see a more significant improvement in overall system performance depending on the “gap” between your computer’s architecture and the latest one available. For example, a Pentium 4 architecture from 2004 is going to see a much more significant improvement upgrading to “Haswell” (the latest as of Q4 2013) than a “Sandy Bridge” architecture from ~2010. Links Related questions: How to reduce disk thrashing (paging)? Windows Swap (Page File): Enable or Disable? Also, just in case you’re considering it, you really shouldn’t disable the page file, as this will only make matters worse; see here. And, in case you needed extra convincing to leave the Windows Page File alone, see here and here. Have something to add to the explanation? Sound off in the the comments. Want to read more answers from other tech-savvy Stack Exchange users? Check out the full discussion thread here.     

    Read the article

  • Evaluating code for a graph [migrated]

    - by mazen.r.f
    This is relatively long code. Please take a look at this code if you are still willing to do so. I will appreciate your feedback. I have spent two days trying to come up with code to represent a graph, calculating the shortest path using Dijkstra's algorithm. But I am not able to get the right result, even though the code runs without errors. The result is not correct and I am always getting 0. I have three classes: Vertex, Edge, and Graph. The Vertex class represents the nodes in the graph and it has id and carried (which carry the weight of the links connected to it while using Dijkstra's algorithm) and a vector of the ids belong to other nodes the path will go through before arriving to the node itself. This vector is named previous_nodes. The Edge class represents the edges in the graph and has two vertices (one in each side) and a width (the distance between the two vertices). The Graph class represents the graph. It has two vectors, where one is the vertices included in this graph, and the other is the edges included in the graph. Inside the class Graph, there is a method named shortest() that takes the sources node id and the destination and calculates the shortest path using Dijkstra's algorithm. I think that it is the most important part of the code. My theory about the code is that I will create two vectors, one for the vertices in the graph named vertices, and another vector named ver_out (it will include the vertices out of calculation in the graph). I will also have two vectors of type Edge, where one is named edges (for all the edges in the graph), and the other is named track (to temporarily contain the edges linked to the temporary source node in every round). After the calculation of every round, the vector track will be cleared. In main(), I've created five vertices and 10 edges to simulate a graph. The result of the shortest path supposedly is 4, but I am always getting 0. That means I have something wrong in my code. If you are interesting in helping me find my mistake and making the code work, please take a look. The way shortest work is as follow: at the beginning, all the edges will be included in the vector edges. We select the edges related to the source and put them in the vector track, then we iterate through track and add the width of every edge to the vertex (node) related to it (not the source vertex). After that, we clear track and remove the source vertex from the vector vertices and select a new source. Then we start over again and select the edges related to the new source, put them in track, iterate over edges in track, adding the weights to the corresponding vertices, then remove this vertex from the vector vertices. Then clear track, and select a new source, and so on. #include<iostream> #include<vector> #include <stdlib.h> // for rand() using namespace std; class Vertex { private: unsigned int id; // the name of the vertex unsigned int carried; // the weight a vertex may carry when calculating shortest path vector<unsigned int> previous_nodes; public: unsigned int get_id(){return id;}; unsigned int get_carried(){return carried;}; void set_id(unsigned int value) {id = value;}; void set_carried(unsigned int value) {carried = value;}; void previous_nodes_update(unsigned int val){previous_nodes.push_back(val);}; void previous_nodes_erase(unsigned int val){previous_nodes.erase(previous_nodes.begin() + val);}; Vertex(unsigned int init_val = 0, unsigned int init_carried = 0) :id (init_val), carried(init_carried) // constructor { } ~Vertex() {}; // destructor }; class Edge { private: Vertex first_vertex; // a vertex on one side of the edge Vertex second_vertex; // a vertex on the other side of the edge unsigned int weight; // the value of the edge ( or its weight ) public: unsigned int get_weight() {return weight;}; void set_weight(unsigned int value) {weight = value;}; Vertex get_ver_1(){return first_vertex;}; Vertex get_ver_2(){return second_vertex;}; void set_first_vertex(Vertex v1) {first_vertex = v1;}; void set_second_vertex(Vertex v2) {second_vertex = v2;}; Edge(const Vertex& vertex_1 = 0, const Vertex& vertex_2 = 0, unsigned int init_weight = 0) : first_vertex(vertex_1), second_vertex(vertex_2), weight(init_weight) { } ~Edge() {} ; // destructor }; class Graph { private: std::vector<Vertex> vertices; std::vector<Edge> edges; public: Graph(vector<Vertex> ver_vector, vector<Edge> edg_vector) : vertices(ver_vector), edges(edg_vector) { } ~Graph() {}; vector<Vertex> get_vertices(){return vertices;}; vector<Edge> get_edges(){return edges;}; void set_vertices(vector<Vertex> vector_value) {vertices = vector_value;}; void set_edges(vector<Edge> vector_ed_value) {edges = vector_ed_value;}; unsigned int shortest(unsigned int src, unsigned int dis) { vector<Vertex> ver_out; vector<Edge> track; for(unsigned int i = 0; i < edges.size(); ++i) { if((edges[i].get_ver_1().get_id() == vertices[src].get_id()) || (edges[i].get_ver_2().get_id() == vertices[src].get_id())) { track.push_back (edges[i]); edges.erase(edges.begin()+i); } }; for(unsigned int i = 0; i < track.size(); ++i) { if(track[i].get_ver_1().get_id() != vertices[src].get_id()) { track[i].get_ver_1().set_carried((track[i].get_weight()) + track[i].get_ver_2().get_carried()); track[i].get_ver_1().previous_nodes_update(vertices[src].get_id()); } else { track[i].get_ver_2().set_carried((track[i].get_weight()) + track[i].get_ver_1().get_carried()); track[i].get_ver_2().previous_nodes_update(vertices[src].get_id()); } } for(unsigned int i = 0; i < vertices.size(); ++i) if(vertices[i].get_id() == src) vertices.erase(vertices.begin() + i); // removing the sources vertex from the vertices vector ver_out.push_back (vertices[src]); track.clear(); if(vertices[0].get_id() != dis) {src = vertices[0].get_id();} else {src = vertices[1].get_id();} for(unsigned int i = 0; i < vertices.size(); ++i) if((vertices[i].get_carried() < vertices[src].get_carried()) && (vertices[i].get_id() != dis)) src = vertices[i].get_id(); //while(!edges.empty()) for(unsigned int round = 0; round < vertices.size(); ++round) { for(unsigned int k = 0; k < edges.size(); ++k) { if((edges[k].get_ver_1().get_id() == vertices[src].get_id()) || (edges[k].get_ver_2().get_id() == vertices[src].get_id())) { track.push_back (edges[k]); edges.erase(edges.begin()+k); } }; for(unsigned int n = 0; n < track.size(); ++n) if((track[n].get_ver_1().get_id() != vertices[src].get_id()) && (track[n].get_ver_1().get_carried() > (track[n].get_ver_2().get_carried() + track[n].get_weight()))) { track[n].get_ver_1().set_carried((track[n].get_weight()) + track[n].get_ver_2().get_carried()); track[n].get_ver_1().previous_nodes_update(vertices[src].get_id()); } else if(track[n].get_ver_2().get_carried() > (track[n].get_ver_1().get_carried() + track[n].get_weight())) { track[n].get_ver_2().set_carried((track[n].get_weight()) + track[n].get_ver_1().get_carried()); track[n].get_ver_2().previous_nodes_update(vertices[src].get_id()); } for(unsigned int t = 0; t < vertices.size(); ++t) if(vertices[t].get_id() == src) vertices.erase(vertices.begin() + t); track.clear(); if(vertices[0].get_id() != dis) {src = vertices[0].get_id();} else {src = vertices[1].get_id();} for(unsigned int tt = 0; tt < edges.size(); ++tt) { if(vertices[tt].get_carried() < vertices[src].get_carried()) { src = vertices[tt].get_id(); } } } return vertices[dis].get_carried(); } }; int main() { cout<< "Hello, This is a graph"<< endl; vector<Vertex> vers(5); vers[0].set_id(0); vers[1].set_id(1); vers[2].set_id(2); vers[3].set_id(3); vers[4].set_id(4); vector<Edge> eds(10); eds[0].set_first_vertex(vers[0]); eds[0].set_second_vertex(vers[1]); eds[0].set_weight(5); eds[1].set_first_vertex(vers[0]); eds[1].set_second_vertex(vers[2]); eds[1].set_weight(9); eds[2].set_first_vertex(vers[0]); eds[2].set_second_vertex(vers[3]); eds[2].set_weight(4); eds[3].set_first_vertex(vers[0]); eds[3].set_second_vertex(vers[4]); eds[3].set_weight(6); eds[4].set_first_vertex(vers[1]); eds[4].set_second_vertex(vers[2]); eds[4].set_weight(2); eds[5].set_first_vertex(vers[1]); eds[5].set_second_vertex(vers[3]); eds[5].set_weight(5); eds[6].set_first_vertex(vers[1]); eds[6].set_second_vertex(vers[4]); eds[6].set_weight(7); eds[7].set_first_vertex(vers[2]); eds[7].set_second_vertex(vers[3]); eds[7].set_weight(1); eds[8].set_first_vertex(vers[2]); eds[8].set_second_vertex(vers[4]); eds[8].set_weight(8); eds[9].set_first_vertex(vers[3]); eds[9].set_second_vertex(vers[4]); eds[9].set_weight(3); unsigned int path; Graph graf(vers, eds); path = graf.shortest(2, 4); cout<< path << endl; return 0; }

    Read the article

  • Software development is (mostly) a trade, and what to do about it

    - by Jeff
    (This is another cross-post from my personal blog. I don’t even remember when I first started to write it, but I feel like my opinion is well enough baked to share.) I've been sitting on this for a long time, particularly as my opinion has changed dramatically over the last few years. That I've encountered more crappy code than maintainable, quality code in my career as a software developer only reinforces what I'm about to say. Software development is just a trade for most, and not a huge academic endeavor. For those of you with computer science degrees readying your pitchforks and collecting your algorithm interview questions, let me explain. This is not an assault on your way of life, and if you've been around, you know I'm right about the quality problem. You also know the HR problem is very real, or we wouldn't be paying top dollar for mediocre developers and importing people from all over the world to fill the jobs we can't fill. I'm going to try and outline what I see as some of the problems, and hopefully offer my views on how to address them. The recruiting problem I think a lot of companies are doing it wrong. Over the years, I've had two kinds of interview experiences. The first, and right, kind of experience involves talking about real life achievements, followed by some variation on white boarding in pseudo-code, drafting some basic system architecture, or even sitting down at a comprooder and pecking out some basic code to tackle a real problem. I can honestly say that I've had a job offer for every interview like this, save for one, because the task was to debug something and they didn't like me asking where to look ("everyone else in the company died in a plane crash"). The other interview experience, the wrong one, involves the classic torture test designed to make the candidate feel stupid and do things they never have, and never will do in their job. First they will question you about obscure academic material you've never seen, or don't care to remember. Then they'll ask you to white board some ridiculous algorithm involving prime numbers or some kind of string manipulation no one would ever do. In fact, if you had to do something like this, you'd Google for a solution instead of waste time on a solved problem. Some will tell you that the academic gauntlet interview is useful to see how people respond to pressure, how they engage in complex logic, etc. That might be true, unless of course you have someone who brushed up on the solutions to the silly puzzles, and they're playing you. But here's the real reason why the second experience is wrong: You're evaluating for things that aren't the job. These might have been useful tactics when you had to hire people to write machine language or C++, but in a world dominated by managed code in C#, or Java, people aren't managing memory or trying to be smarter than the compilers. They're using well known design patterns and techniques to deliver software. More to the point, these puzzle gauntlets don't evaluate things that really matter. They don't get into code design, issues of loose coupling and testability, knowledge of the basics around HTTP, or anything else that relates to building supportable and maintainable software. The first situation, involving real life problems, gives you an immediate idea of how the candidate will work out. One of my favorite experiences as an interviewee was with a guy who literally brought his work from that day and asked me how to deal with his problem. I had to demonstrate how I would design a class, make sure the unit testing coverage was solid, etc. I worked at that company for two years. So stop looking for algorithm puzzle crunchers, because a guy who can crush a Fibonacci sequence might also be a guy who writes a class with 5,000 lines of untestable code. Fashion your interview process on ways to reveal a developer who can write supportable and maintainable code. I would even go so far as to let them use the Google. If they want to cut-and-paste code, pass on them, but if they're looking for context or straight class references, hire them, because they're going to be life-long learners. The contractor problem I doubt anyone has ever worked in a place where contractors weren't used. The use of contractors seems like an obvious way to control costs. You can hire someone for just as long as you need them and then let them go. You can even give them the work that no one else wants to do. In practice, most places I've worked have retained and budgeted for the contractor year-round, meaning that the $90+ per hour they're paying (of which half goes to the person) would have been better spent on a full-time person with a $100k salary and benefits. But it's not even the cost that is an issue. It's the quality of work delivered. The accountability of a contractor is totally transient. They only need to deliver for as long as you keep them around, and chances are they'll never again touch the code. There's no incentive for them to get things right, there's little incentive to understand your system or learn anything. At the risk of making an unfair generalization, craftsmanship doesn't matter to most contractors. The education problem I don't know what they teach in college CS courses. I've believed for most of my adult life that a college degree was an essential part of being successful. Of course I would hold that bias, since I did it, and have the paper to show for it in a box somewhere in the basement. My first clue that maybe this wasn't a fully qualified opinion comes from the fact that I double-majored in journalism and radio/TV, not computer science. Eventually I worked with people who skipped college entirely, many of them at Microsoft. Then I worked with people who had a masters degree who sucked at writing code, next to the high school diploma types that rock it every day. I still think there's a lot to be said for the social development of someone who has the on-campus experience, but for software developers, college might not matter. As I mentioned before, most of us are not writing compilers, and we never will. It's actually surprising to find how many people are self-taught in the art of software development, and that should reveal some interesting truths about how we learn. The first truth is that we learn largely out of necessity. There's something that we want to achieve, so we do what I call just-in-time learning to meet those goals. We acquire knowledge when we need it. So what about the gaps in our knowledge? That's where the most valuable education occurs, via our mentors. They're the people we work next to and the people who write blogs. They are critical to our professional development. They don't need to be an encyclopedia of jargon, but they understand the craft. Even at this stage of my career, I probably can't tell you what SOLID stands for, but you can bet that I practice the principles behind that acronym every day. That comes from experience, augmented by my peers. I'm hell bent on passing that experience to others. Process issues If you're a manager type and don't do much in the way of writing code these days (shame on you for not messing around at least), then your job is to isolate your tradespeople from nonsense, while bringing your business into the realm of modern software development. That doesn't mean you slap up a white board with sticky notes and start calling yourself agile, it means getting all of your stakeholders to understand that frequent delivery of quality software is the best way to deal with change and evolving expectations. It also means that you have to play technical overlord to make sure the education and quality issues are dealt with. That's why I make the crack about sticky notes, because without the right technique being practiced among your code monkeys, you're just a guy with sticky notes. You're asking your business to accept frequent and iterative delivery, now make sure that the folks writing the code can handle the same thing. This means unit testing, the right instrumentation, integration tests, automated builds and deployments... all of the stuff that makes it easy to see when change breaks stuff. The prognosis I strongly believe that education is the most important part of what we do. I'm encouraged by things like The Starter League, and it's the kind of thing I'd love to see more of. I would go as far as to say I'd love to start something like this internally at an existing company. Most of all though, I can't emphasize enough how important it is that we mentor each other and share our knowledge. If you have people on your staff who don't want to learn, fire them. Seriously, get rid of them. A few months working with someone really good, who understands the craftsmanship required to build supportable and maintainable code, will change that person forever and increase their value immeasurably.

    Read the article

  • iPhone Serialization problem

    - by Jenicek
    Hi, I need to save my own created class to file, I found on the internet, that good approach is to use NSKeyedArchiver and NSKeyedUnarchiver My class definition looks like this: @interface Game : NSObject <NSCoding> { NSMutableString *strCompleteWord; NSMutableString *strWordToGuess; NSMutableArray *arGuessedLetters; //This array stores characters NSMutableArray *arGuessedLettersPos; //This array stores CGRects NSInteger iScore; NSInteger iLives; NSInteger iRocksFallen; BOOL bGameCompleted; BOOL bGameOver; } I've implemented methods initWithCoder: and encodeWithCoder: this way: - (id)initWithCoder:(NSCoder *)coder { if([coder allowsKeyedCoding]) { strCompleteWord = [[coder decodeObjectForKey:@"CompletedWord"] copy]; strWordToGuess = [[coder decodeObjectForKey:@"WordToGuess"] copy]; arGuessedLetters = [[coder decodeObjectForKey:@"GuessedLetters"] retain]; // arGuessedLettersPos = [[coder decodeObjectForKey:@"GuessedLettersPos"] retain]; iScore = [coder decodeIntegerForKey:@"Score"]; iLives = [coder decodeIntegerForKey:@"Lives"]; iRocksFallen = [coder decodeIntegerForKey:@"RocksFallen"]; bGameCompleted = [coder decodeBoolForKey:@"GameCompleted"]; bGameOver = [coder decodeBoolForKey:@"GameOver"]; } else { strCompleteWord = [[coder decodeObject] retain]; strWordToGuess = [[coder decodeObject] retain]; arGuessedLetters = [[coder decodeObject] retain]; // arGuessedLettersPos = [[coder decodeObject] retain]; [coder decodeValueOfObjCType:@encode(NSInteger) at:&iScore]; [coder decodeValueOfObjCType:@encode(NSInteger) at:&iLives]; [coder decodeValueOfObjCType:@encode(NSInteger) at:&iRocksFallen]; [coder decodeValueOfObjCType:@encode(BOOL) at:&bGameCompleted]; [coder decodeValueOfObjCType:@encode(BOOL) at:&bGameOver]; } return self; } - (void)encodeWithCoder:(NSCoder *)coder { if([coder allowsKeyedCoding]) { [coder encodeObject:strCompleteWord forKey:@"CompleteWord"]; [coder encodeObject:strWordToGuess forKey:@"WordToGuess"]; [coder encodeObject:arGuessedLetters forKey:@"GuessedLetters"]; //[coder encodeObject:arGuessedLettersPos forKey:@"GuessedLettersPos"]; [coder encodeInteger:iScore forKey:@"Score"]; [coder encodeInteger:iLives forKey:@"Lives"]; [coder encodeInteger:iRocksFallen forKey:@"RocksFallen"]; [coder encodeBool:bGameCompleted forKey:@"GameCompleted"]; [coder encodeBool:bGameOver forKey:@"GameOver"]; } else { [coder encodeObject:strCompleteWord]; [coder encodeObject:strWordToGuess]; [coder encodeObject:arGuessedLetters]; //[coder encodeObject:arGuessedLettersPos]; [coder encodeValueOfObjCType:@encode(NSInteger) at:&iScore]; [coder encodeValueOfObjCType:@encode(NSInteger) at:&iLives]; [coder encodeValueOfObjCType:@encode(NSInteger) at:&iRocksFallen]; [coder encodeValueOfObjCType:@encode(BOOL) at:&bGameCompleted]; [coder encodeValueOfObjCType:@encode(BOOL) at:&bGameOver]; } } And I use these methods to archive and unarchive data: [NSKeyedArchiver archiveRootObject:currentGame toFile:strPath]; Game *currentGame = [NSKeyedUnarchiver unarchiveObjectWithFile:strPath]; I have two problems. 1) As you can see, lines with arGuessedLettersPos is commented, it's because every time I try to encode this array, error comes up(this archiver cannot encode structs), and this array is used for storing CGRect structs. I've seen solution on the internet. The thing is, that every CGRect in the array is converted to an NSString (using NSStringFromCGRect()) and then saved. Is it a good approach? 2)This is bigger problem for me. Even if I comment this line and then run the code successfully, then save(archive) the data and then try to load (unarchive) them, no data is loaded. There aren't any error but currentGame object does not have data that should be loaded. Could you please give me some advice? This is first time I'm using archivers and unarchivers. Thanks a lot for every reply.

    Read the article

  • How to populate a core data store programmatically?

    - by jdmuys
    I have ran out of hairs to pull with a crash in this routine that populates a core data store from a 9000+ line plist file. The crash happened at the very end of the routine inside the call to [managedObjectContext save:&error]. While if I save after every object insertion, the crash doesn't happen. Of course, saving after every object insertion totally kills the performance (from less than a second to many minutes). I modified my code so that it saves every K insertions, and the crash happens as soon as K = 2. The crash is an out-of-bound exception for an NSArray: Serious application error. Exception was caught during Core Data change processing: *** -[NSCFArray objectAtIndex:]: index (1) beyond bounds (1) with userInfo (null) Also maybe relevant, when the exception happen, my fetch result controller controllerDidChangeContent: delegate routine is in the call stack. It simply calls my table view endUpdate routine. I am now running out of ideas. How am I supposed to populate a core data store with a table view? Here is the call stack: #0 0x901ca4e6 in objc_exception_throw #1 0x01d86c3b in +[NSException raise:format:arguments:] #2 0x01d86b9a in +[NSException raise:format:] #3 0x00072cb9 in _NSArrayRaiseBoundException #4 0x00010217 in -[NSCFArray objectAtIndex:] #5 0x002eaaa7 in -[UITableView(_UITableViewPrivate) _endCellAnimationsWithContext:] #6 0x002def02 in -[UITableView endUpdates] #7 0x00004863 in -[AirportViewController controllerDidChangeContent:] at AirportViewController.m:463 #8 0x01c43be1 in -[NSFetchedResultsController(PrivateMethods) _managedObjectContextDidChange:] #9 0x0001462a in _nsnote_callback #10 0x01d31005 in _CFXNotificationPostNotification #11 0x00011ee0 in -[NSNotificationCenter postNotificationName:object:userInfo:] #12 0x01ba417d in -[NSManagedObjectContext(_NSInternalNotificationHandling) _postObjectsDidChangeNotificationWithUserInfo:] #13 0x01c03763 in -[NSManagedObjectContext(_NSInternalChangeProcessing) _createAndPostChangeNotification:withDeletions:withUpdates:withRefreshes:] #14 0x01b885ea in -[NSManagedObjectContext(_NSInternalChangeProcessing) _processRecentChanges:] #15 0x01bbe728 in -[NSManagedObjectContext save:] #16 0x000039ea in -[AirportViewController populateAirports] at AirportViewController.m:112 Here is the code to the routine. I apologize because a number of lines are probably irrelevant, but I'd rather err on that side. The crash happens the very first time it calls [managedObjectContext save:&error]: - (void) populateAirports { NSBundle *meBundle = [NSBundle mainBundle]; NSString *dbPath = [meBundle pathForResource:@"DuckAirportsBin" ofType:@"plist"]; NSArray *initialAirports = [[NSArray alloc] initWithContentsOfFile:dbPath]; //********************************************************************************* // get existing countries NSMutableDictionary *countries = [[NSMutableDictionary alloc] initWithCapacity:200]; NSFetchRequest *fetchRequest = [[NSFetchRequest alloc] init]; NSEntityDescription *entity = [NSEntityDescription entityForName:@"Country" inManagedObjectContext:managedObjectContext]; [fetchRequest setEntity:entity]; NSError *error = nil; NSArray *values = [managedObjectContext executeFetchRequest:fetchRequest error:&error]; if (!values) { NSLog(@"Unresolved error %@, %@", error, [error userInfo]); abort(); } int numCountries = [values count]; NSLog(@"We have %d countries in store", numCountries); for (Country *aCountry in values) { [countries setObject:aCountry forKey:aCountry.code]; } [fetchRequest release]; //********************************************************************************* // read airports int numAirports = 0; int numUnsavedAirports = 0; #define MAX_UNSAVED_AIRPORTS_BEFORE_SAVE 2 numCountries = 0; for (NSDictionary *anAirport in initialAirports) { NSAutoreleasePool * pool = [[NSAutoreleasePool alloc] init]; NSString *countryCode = [anAirport objectForKey:@"country"]; Country *thatCountry = [countries objectForKey:countryCode]; if (!thatCountry) { thatCountry = [NSEntityDescription insertNewObjectForEntityForName:@"Country" inManagedObjectContext:managedObjectContext]; thatCountry.code = countryCode; thatCountry.name = [anAirport objectForKey:@"country_name"]; thatCountry.population = 0; [countries setObject:thatCountry forKey:countryCode]; numCountries++; NSLog(@"Found %dth country %@=%@", numCountries, countryCode, thatCountry.name); } // now that we have the country, we create the airport Airport *newAirport = [NSEntityDescription insertNewObjectForEntityForName:@"Airport" inManagedObjectContext:managedObjectContext]; newAirport.city = [anAirport objectForKey:@"city"]; newAirport.code = [anAirport objectForKey:@"code"]; newAirport.name = [anAirport objectForKey:@"name"]; newAirport.country_name = [anAirport objectForKey:@"country_name"]; newAirport.latitude = [NSNumber numberWithDouble:[[anAirport objectForKey:@"latitude"] doubleValue]]; newAirport.longitude = [NSNumber numberWithDouble:[[anAirport objectForKey:@"longitude"] doubleValue]]; newAirport.altitude = [NSNumber numberWithDouble:[[anAirport objectForKey:@"altitude"] doubleValue]]; newAirport.country = thatCountry; // [thatCountry addAirportsObject:newAirport]; numAirports++; numUnsavedAirports++; if (numUnsavedAirports >= MAX_UNSAVED_AIRPORTS_BEFORE_SAVE) { if (![managedObjectContext save:&error]) { NSLog(@"Unresolved error %@, %@", error, [error userInfo]); abort(); } numUnsavedAirports = 0; } [pool release]; }

    Read the article

  • Delphi: Using Enumerators to filter TList<T: class> by class type?

    - by afrazier
    Okay, this might be confusing. What I'm trying to do is use an enumerator to only return certain items in a generic list based on class type. Given the following hierarchy: type TShapeClass = class of TShape; TShape = class(TObject) private FId: Integer; public function ToString: string; override; property Id: Integer read FId write FId; end; TCircle = class(TShape) private FDiameter: Integer; public property Diameter: Integer read FDiameter write FDiameter; end; TSquare = class(TShape) private FSideLength: Integer; public property SideLength: Integer read FSideLength write FSideLength; end; TShapeList = class(TObjectList<TShape>) end; How can I extend TShapeList such that I can do something similar to the following: procedure Foo; var ShapeList: TShapeList; Shape: TShape; Circle: TCircle; Square: TSquare; begin // Create ShapeList and fill with TCircles and TSquares for Circle in ShapeList<TCircle> do begin // do something with each TCircle in ShapeList end; for Square in ShapeList<TSquare> do begin // do something with each TSquare in ShapeList end; for Shape in ShapeList<TShape> do begin // do something with every object in TShapeList end; end; I've tried extending TShapeList using an adapted version of Primoz Gabrijelcic's bit on Parameterized Enumerators using a factory record as follows: type TShapeList = class(TObjectList<TShape>) public type TShapeFilterEnumerator<T: TShape> = record private FShapeList: TShapeList; FClass: TShapeClass; FIndex: Integer; function GetCurrent: T; public constructor Create(ShapeList: TShapeList); function MoveNext: Boolean; property Current: T read GetCurrent; end; TShapeFilterFactory<T: TShape> = record private FShapeList: TShapeList; public constructor Create(ShapeList: TShapeList); function GetEnumerator: TShapeFilterEnumerator<T>; end; function FilteredEnumerator<T: TShape>: TShapeFilterFactory<T>; end; Then I modified Foo to be: procedure Foo; var ShapeList: TShapeList; Shape: TShape; Circle: TCircle; Square: TSquare; begin // Create ShapeList and fill with TCircles and TSquares for Circle in ShapeList.FilteredEnumerator<TCircle> do begin // do something with each TCircle in ShapeList end; for Square in ShapeList.FilteredEnumerator<TSquare> do begin // do something with each TSquare in ShapeList end; for Shape in ShapeList.FilteredEnumerator<TShape> do begin // do something with every object in TShapeList end; end; However, Delphi 2010 is throwing an error when I try to compile Foo about Incompatible types: TCircle and TShape. If I comment out the TCircle loop, then I get a similar error about TSquare. If I comment the TSquare loop out as well, the code compiles and works. Well, it works in the sense that it enumerates every object since they all descend from TShape. The strange thing is that the line number that the compiler indicates is 2 lines beyond the end of my file. In my demo project, it indicated line 177, but there's only 175 lines. Is there any way to make this work? I'd like to be able to assign to Circle directly without going through any typecasts or checking in my for loop itself.

    Read the article

  • When should I use indexed arrays of OpenGL vertices?

    - by Tartley
    I'm trying to get a clear idea of when I should be using indexed arrays of OpenGL vertices, drawn with gl[Multi]DrawElements and the like, versus when I should simply use contiguous arrays of vertices, drawn with gl[Multi]DrawArrays. (Update: The consensus in the replies I got is that one should always be using indexed vertices.) I have gone back and forth on this issue several times, so I'm going to outline my current understanding, in the hopes someone can either tell me I'm now finally more or less correct, or else point out where my remaining misunderstandings are. Specifically, I have three conclusions, in bold. Please correct them if they are wrong. One simple case is if my geometry consists of meshes to form curved surfaces. In this case, the vertices in the middle of the mesh will have identical attributes (position, normal, color, texture coord, etc) for every triangle which uses the vertex. This leads me to conclude that: 1. For geometry with few seams, indexed arrays are a big win. Follow rule 1 always, except: For geometry that is very 'blocky', in which every edge represents a seam, the benefit of indexed arrays is less obvious. To take a simple cube as an example, although each vertex is used in three different faces, we can't share vertices between them, because for a single vertex, the surface normals (and possible other things, like color and texture co-ord) will differ on each face. Hence we need to explicitly introduce redundant vertex positions into our array, so that the same position can be used several times with different normals, etc. This means that indexed arrays are of less use. e.g. When rendering a single face of a cube: 0 1 o---o |\ | | \ | | \| o---o 3 2 (this can be considered in isolation, because the seams between this face and all adjacent faces mean than none of these vertices can be shared between faces) if rendering using GL_TRIANGLE_FAN (or _STRIP), then each face of the cube can be rendered thus: verts = [v0, v1, v2, v3] colors = [c0, c0, c0, c0] normal = [n0, n0, n0, n0] Adding indices does not allow us to simplify this. From this I conclude that: 2. When rendering geometry which is all seams or mostly seams, when using GL_TRIANGLE_STRIP or _FAN, then I should never use indexed arrays, and should instead always use gl[Multi]DrawArrays. (Update: Replies indicate that this conclusion is wrong. Even though indices don't allow us to reduce the size of the arrays here, they should still be used because of other performance benefits, as discussed in the comments) The only exception to rule 2 is: When using GL_TRIANGLES (instead of strips or fans), then half of the vertices can still be re-used twice, with identical normals and colors, etc, because each cube face is rendered as two separate triangles. Again, for the same single cube face: 0 1 o---o |\ | | \ | | \| o---o 3 2 Without indices, using GL_TRIANGLES, the arrays would be something like: verts = [v0, v1, v2, v2, v3, v0] normals = [n0, n0, n0, n0, n0, n0] colors = [c0, c0, c0, c0, c0, c0] Since a vertex and a normal are often 3 floats each, and a color is often 3 bytes, that gives, for each cube face, about: verts = 6 * 3 floats = 18 floats normals = 6 * 3 floats = 18 floats colors = 6 * 3 bytes = 18 bytes = 36 floats and 18 bytes per cube face. (I understand the number of bytes might change if different types are used, the exact figures are just for illustration.) With indices, we can simplify this a little, giving: verts = [v0, v1, v2, v3] (4 * 3 = 12 floats) normals = [n0, n0, n0, n0] (4 * 3 = 12 floats) colors = [c0, c0, c0, c0] (4 * 3 = 12 bytes) indices = [0, 1, 2, 2, 3, 0] (6 shorts) = 24 floats + 12 bytes, and maybe 6 shorts, per cube face. See how in the latter case, vertices 0 and 2 are used twice, but only represented once in each of the verts, normals and colors arrays. This sounds like a small win for using indices, even in the extreme case of every single geometry edge being a seam. This leads me to conclude that: 3. When using GL_TRIANGLES, one should always use indexed arrays, even for geometry which is all seams. Please correct my conclusions in bold if they are wrong.

    Read the article

  • What can I do to improve a project if there is a no-listening situation. Developers vs Management

    - by NazGul
    Hi all, I hope that I'm not the only one and I can get a answer from someone with more experience than me, so I can think cleaner and I don't get depressed with this developer's life. I'm working as developer for a small company three years now. In that three years I'm working in the same project and sincerely, I think this project could be used as a CASE STUDY because it has all the situations that cannot happen in a project and that makes a project fails. To begin with, and I believe you've already noticed, the project has 3 years already (develoment only) and is still unfinished, because in every meeting there is a "new priority" ,or a "new problem" to be solve or a "new feature" to be add. So, first problem is no target set. How can you know when something is finished if you don't know what you want? I understand Management, because they see an oportunity and try to get that, but I don't understand how can they not see (or hear us) that they'll lose all they already have and what they'll eventually get. Second, there is no team group. My team consists of three people, a Senior Developer, a DBA and, finally, I for all the work (support, testing, new features, bug fixing, meeting, projet management of clients, etc) aka Junior Developer. The first (senior developer), does not perform any tests on his changes, so, most of the time, his changes give us problems (us = me, since I'm the one who will fix it). The second (DBA) is an uncompromising person and you can not talk to him, believe me, I tried! In his view, everything he does is fantastic... even if it is the most complicated to make it... And he does everything he wants, even if we need that only for 5 months later and would help some extra-hand to do the things we have to do for now. As you can see, there is very hard to work with no help... Third, there is no testings. Every... I repeat, Every release of the project, the customers wants to kill us, because there is a lot of bugs. Management? They say that they want tests before the release. Us? We say the same. Time? No time. Management? There is always some time to open the application and click in some buttons. Us? Try to explain that it is not so simple. Management doesn't care... end of story. Actually, must of the bugs could be avoid with a rigorous work... Some people just want to do the show to the Management. "Did you ask for this? Cool, it's done. Bugs? The Do-all-the-work guy will solve." Unfortunally for me, sometimes the Do-all-the-work also has to finish it. And to makes this all better, I'm the person who will listen the complaints from the customers. Cool, huh? I know, everyone makes mistakes. But there is mistakes and mistakes... To complete, in the Management view, "the problem is the lack of an individual project management", because we cannot do all the stuff they ask, even if there is no PM for the project itself. And ask us to work overtime without any reward... I do say all this stuff to the management and others members, but by telling this, the I'm the bad guy, the guy who is complain when everything is going well... but we need to work overtime... sigh What can I do to make it works? Anyone has a situation like this, what did you do? I hope you could understand my problem, my English is a little rusty. Thanks.

    Read the article

< Previous Page | 163 164 165 166 167 168 169 170 171 172 173 174  | Next Page >