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  • SQL SERVER – Color Coding SQL Server Management Studio Status Bar – SQL in Sixty Seconds #023 – Video

    - by pinaldave
    I often see developers executing the unplanned code on production server when they actually want to execute on the development server. Developers and DBAs get confused because when they use SQL Server Management Studio (SSMS) they forget to pay attention to the server they are connecting. It is very easy to fix this problem. You can select different color for a different server. Once you have different color for different server in the status bar, it will be easier for developer easily notice the server against which they are about to execute the script. Personally when I work on SQL Server development, here is the color code, which I follow. I keep Green for my development server, blue for my staging server and red for my production server. Honestly color coding does not signify much but different color for different server is the key here. More Tips on SSMS in SQL in Sixty Seconds: Generate Script for Schema and Data in SQL Server – SQL in Sixty Seconds #021  Remove Debug Button in SQL Server Management Studio – SQL in Sixty Seconds #020  Three Tricks to Comment T-SQL in SQL Server Management Studio – SQL in Sixty Seconds #019  Importing CSV into SQL Server – SQL in Sixty Seconds #018   Tricks to Replace SELECT * with Column Names – SQL in Sixty Seconds #017 I encourage you to submit your ideas for SQL in Sixty Seconds. We will try to accommodate as many as we can. If we like your idea we promise to share with you educational material. Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: Database, Pinal Dave, PostADay, SQL, SQL Authority, SQL in Sixty Seconds, SQL Query, SQL Scripts, SQL Server, SQL Server Management Studio, SQL Tips and Tricks, T SQL, Technology, Video

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  • NSClient++: external script with optional arguments

    - by syneticon-dj
    I am trying to define an external script which would take optional arguments in NSClient++ 0.4.1 on Windows. Following the nsclient-full.ini example code I have defined mycheck=cmd /C echo C:\mydir\myscript.ps1 %ARGS% | powershell.exe -command - which simply yields the string %ARGS% passed as the only argument to myscript.ps1, no matter what I specify in my call through NRPE (using Nagios' check_nrpe if that matters). I then tried to rewrite the definition to mycheck=cmd /C echo C:\mydir\myscript.ps1 $ARG1$ $ARG2$ | powershell.exe -command - (myscript.ps1 would take up to two arguments), which does help a bit. At least, if two arguments are provided, I can fetch them via the args[] array. The trouble starts when the call has less than two arguments - in this case the literal strings $ARG2 and $ARG1$ are passed through as arguments. Handling this case in the code of myscript.ps1 makes the whole argument processing routine ugly at best. Is there a sane way of defining optional parameters to an external script which would not pass NSClient's variable names if no parameter has been specified?

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  • Parallelism in .NET – Part 9, Configuration in PLINQ and TPL

    - by Reed
    Parallel LINQ and the Task Parallel Library contain many options for configuration.  Although the default configuration options are often ideal, there are times when customizing the behavior is desirable.  Both frameworks provide full configuration support. When working with Data Parallelism, there is one primary configuration option we often need to control – the number of threads we want the system to use when parallelizing our routine.  By default, PLINQ and the TPL both use the ThreadPool to schedule tasks.  Given the major improvements in the ThreadPool in CLR 4, this default behavior is often ideal.  However, there are times that the default behavior is not appropriate.  For example, if you are working on multiple threads simultaneously, and want to schedule parallel operations from within both threads, you might want to consider restricting each parallel operation to using a subset of the processing cores of the system.  Not doing this might over-parallelize your routine, which leads to inefficiencies from having too many context switches. In the Task Parallel Library, configuration is handled via the ParallelOptions class.  All of the methods of the Parallel class have an overload which accepts a ParallelOptions argument. We configure the Parallel class by setting the ParallelOptions.MaxDegreeOfParallelism property.  For example, let’s revisit one of the simple data parallel examples from Part 2: Parallel.For(0, pixelData.GetUpperBound(0), row => { for (int col=0; col < pixelData.GetUpperBound(1); ++col) { pixelData[row, col] = AdjustContrast(pixelData[row, col], minPixel, maxPixel); } }); .csharpcode, .csharpcode pre { font-size: small; color: black; font-family: consolas, "Courier New", courier, monospace; background-color: #ffffff; /*white-space: pre;*/ } .csharpcode pre { margin: 0em; } .csharpcode .rem { color: #008000; } .csharpcode .kwrd { color: #0000ff; } .csharpcode .str { color: #006080; } .csharpcode .op { color: #0000c0; } .csharpcode .preproc { color: #cc6633; } .csharpcode .asp { background-color: #ffff00; } .csharpcode .html { color: #800000; } .csharpcode .attr { color: #ff0000; } .csharpcode .alt { background-color: #f4f4f4; width: 100%; margin: 0em; } .csharpcode .lnum { color: #606060; } Here, we’re looping through an image, and calling a method on each pixel in the image.  If this was being done on a separate thread, and we knew another thread within our system was going to be doing a similar operation, we likely would want to restrict this to using half of the cores on the system.  This could be accomplished easily by doing: var options = new ParallelOptions(); options.MaxDegreeOfParallelism = Math.Max(Environment.ProcessorCount / 2, 1); Parallel.For(0, pixelData.GetUpperBound(0), options, row => { for (int col=0; col < pixelData.GetUpperBound(1); ++col) { pixelData[row, col] = AdjustContrast(pixelData[row, col], minPixel, maxPixel); } }); Now, we’re restricting this routine to using no more than half the cores in our system.  Note that I included a check to prevent a single core system from supplying zero; without this check, we’d potentially cause an exception.  I also did not hard code a specific value for the MaxDegreeOfParallelism property.  One of our goals when parallelizing a routine is allowing it to scale on better hardware.  Specifying a hard-coded value would contradict that goal. Parallel LINQ also supports configuration, and in fact, has quite a few more options for configuring the system.  The main configuration option we most often need is the same as our TPL option: we need to supply the maximum number of processing threads.  In PLINQ, this is done via a new extension method on ParallelQuery<T>: ParallelEnumerable.WithDegreeOfParallelism. Let’s revisit our declarative data parallelism sample from Part 6: double min = collection.AsParallel().Min(item => item.PerformComputation()); Here, we’re performing a computation on each element in the collection, and saving the minimum value of this operation.  If we wanted to restrict this to a limited number of threads, we would add our new extension method: int maxThreads = Math.Max(Environment.ProcessorCount / 2, 1); double min = collection .AsParallel() .WithDegreeOfParallelism(maxThreads) .Min(item => item.PerformComputation()); This automatically restricts the PLINQ query to half of the threads on the system. PLINQ provides some additional configuration options.  By default, PLINQ will occasionally revert to processing a query in parallel.  This occurs because many queries, if parallelized, typically actually cause an overall slowdown compared to a serial processing equivalent.  By analyzing the “shape” of the query, PLINQ often decides to run a query serially instead of in parallel.  This can occur for (taken from MSDN): Queries that contain a Select, indexed Where, indexed SelectMany, or ElementAt clause after an ordering or filtering operator that has removed or rearranged original indices. Queries that contain a Take, TakeWhile, Skip, SkipWhile operator and where indices in the source sequence are not in the original order. Queries that contain Zip or SequenceEquals, unless one of the data sources has an originally ordered index and the other data source is indexable (i.e. an array or IList(T)). Queries that contain Concat, unless it is applied to indexable data sources. Queries that contain Reverse, unless applied to an indexable data source. If the specific query follows these rules, PLINQ will run the query on a single thread.  However, none of these rules look at the specific work being done in the delegates, only at the “shape” of the query.  There are cases where running in parallel may still be beneficial, even if the shape is one where it typically parallelizes poorly.  In these cases, you can override the default behavior by using the WithExecutionMode extension method.  This would be done like so: var reversed = collection .AsParallel() .WithExecutionMode(ParallelExecutionMode.ForceParallelism) .Select(i => i.PerformComputation()) .Reverse(); Here, the default behavior would be to not parallelize the query unless collection implemented IList<T>.  We can force this to run in parallel by adding the WithExecutionMode extension method in the method chain. Finally, PLINQ has the ability to configure how results are returned.  When a query is filtering or selecting an input collection, the results will need to be streamed back into a single IEnumerable<T> result.  For example, the method above returns a new, reversed collection.  In this case, the processing of the collection will be done in parallel, but the results need to be streamed back to the caller serially, so they can be enumerated on a single thread. This streaming introduces overhead.  IEnumerable<T> isn’t designed with thread safety in mind, so the system needs to handle merging the parallel processes back into a single stream, which introduces synchronization issues.  There are two extremes of how this could be accomplished, but both extremes have disadvantages. The system could watch each thread, and whenever a thread produces a result, take that result and send it back to the caller.  This would mean that the calling thread would have access to the data as soon as data is available, which is the benefit of this approach.  However, it also means that every item is introducing synchronization overhead, since each item needs to be merged individually. On the other extreme, the system could wait until all of the results from all of the threads were ready, then push all of the results back to the calling thread in one shot.  The advantage here is that the least amount of synchronization is added to the system, which means the query will, on a whole, run the fastest.  However, the calling thread will have to wait for all elements to be processed, so this could introduce a long delay between when a parallel query begins and when results are returned. The default behavior in PLINQ is actually between these two extremes.  By default, PLINQ maintains an internal buffer, and chooses an optimal buffer size to maintain.  Query results are accumulated into the buffer, then returned in the IEnumerable<T> result in chunks.  This provides reasonably fast access to the results, as well as good overall throughput, in most scenarios. However, if we know the nature of our algorithm, we may decide we would prefer one of the other extremes.  This can be done by using the WithMergeOptions extension method.  For example, if we know that our PerformComputation() routine is very slow, but also variable in runtime, we may want to retrieve results as they are available, with no bufferring.  This can be done by changing our above routine to: var reversed = collection .AsParallel() .WithExecutionMode(ParallelExecutionMode.ForceParallelism) .WithMergeOptions(ParallelMergeOptions.NotBuffered) .Select(i => i.PerformComputation()) .Reverse(); On the other hand, if are already on a background thread, and we want to allow the system to maximize its speed, we might want to allow the system to fully buffer the results: var reversed = collection .AsParallel() .WithExecutionMode(ParallelExecutionMode.ForceParallelism) .WithMergeOptions(ParallelMergeOptions.FullyBuffered) .Select(i => i.PerformComputation()) .Reverse(); Notice, also, that you can specify multiple configuration options in a parallel query.  By chaining these extension methods together, we generate a query that will always run in parallel, and will always complete before making the results available in our IEnumerable<T>.

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  • Image Preview in ASP.NET MVC

    - by imran_ku07
      Introduction :         Previewing an image is a great way to improve the UI of your site. Also it is always best to check the file type, size and see a preview before submitting the whole form. There are some ways to do this using simple JavaScript but not work in all browsers (like FF3).In this Article I will show you how do this using ASP.NET MVC application. You also see how this will work in case of nested form.   Description :          Create a new ASP.NET MVC project and then add a file upload and image control into your View. <form id="form1" method="post" action="NerdDinner/ImagePreview/AjaxSubmit">            <table>                <tr>                    <td>                        <input type="file" name="imageLoad1" id="imageLoad1"  onchange="ChangeImage(this,'#imgThumbnail')" />                    </td>                </tr>                <tr>                    <td align="center">                        <img src="images/TempImage.gif" id="imgThumbnail" height="200px" width="200px">                     </td>                </tr>            </table>        </form>           Note that here NerdDinner is refers to the virtual directory name, ImagePreview is the Controller and ImageLoad is the action name which you will see shortly          I will use the most popular jQuery form plug-in, that turns a form into an AJAX form with very little code. Therefore you must get these from Jquery site and then add these files into your page.          <script src="NerdDinner/Scripts/jquery-1.3.2.js" type="text/javascript"></script>        <script src="NerdDinner/Scripts/jquery.form.js" type="text/javascript"></script>            Then add the javascript function. <script type="text/javascript">function ChangeImage(fileId,imageId){ $("#form1").ajaxSubmit({success: function(responseText){ var d=new Date(); $(imageId)[0].src="NerdDinner/ImagePreview/ImageLoad?a="+d.getTime(); } });}</script>             This function simply submit the form named form1 asynchronously to ImagePreviewController's method AjaxSubmit and after successfully receiving the response, it will set the image src property to the action method ImageLoad. Here I am also adding querystring, preventing the browser to serve the cached image.           Now I will create a new Controller named ImagePreviewController. public class ImagePreviewController : Controller { [AcceptVerbs(HttpVerbs.Post)] public ActionResult AjaxSubmit(int? id) { Session["ContentLength"] = Request.Files[0].ContentLength; Session["ContentType"] = Request.Files[0].ContentType; byte[] b = new byte[Request.Files[0].ContentLength]; Request.Files[0].InputStream.Read(b, 0, Request.Files[0].ContentLength); Session["ContentStream"] = b; return Content( Request.Files[0].ContentType+";"+ Request.Files[0].ContentLength ); } public ActionResult ImageLoad(int? id) { byte[] b = (byte[])Session["ContentStream"]; int length = (int)Session["ContentLength"]; string type = (string)Session["ContentType"]; Response.Buffer = true; Response.Charset = ""; Response.Cache.SetCacheability(HttpCacheability.NoCache); Response.ContentType = type; Response.BinaryWrite(b); Response.Flush(); Session["ContentLength"] = null; Session["ContentType"] = null; Session["ContentStream"] = null; Response.End(); return Content(""); } }             The AjaxSubmit action method will save the image in Session and return content type and content length in response. ImageLoad action method will return the contents of image in response.Then clear these Sessions.           Just run your application and see the effect.   Checking Size and Content Type of File:          You may notice that AjaxSubmit action method is returning both content type and content length. You can check both properties before submitting your complete form.     $(myform).ajaxSubmit({success: function(responseText)            {                                var contentType=responseText.substring(0,responseText.indexOf(';'));                var contentLength=responseText.substring(responseText.indexOf(';')+1);                // Here you can do your validation                var d=new Date();                $(imageId)[0].src="http://weblogs.asp.net/MoneypingAPP/ImagePreview/ImageLoad?a="+d.getTime();            }        });  Handling Nested Form Case:          The above code will work if you have only one form. But this is not the case always.You may have a form control which wraps all the controls and you do not want to submit the whole form, just for getting a preview effect.           In this case you need to create a dynamic form control using JavaScript, and then add file upload control to this form and submit the form asynchronously  function ChangeImage(fileId,imageId)         {            var myform=document.createElement("form");                    myform.action="NerdDinner/ImagePreview/AjaxSubmit";            myform.enctype="multipart/form-data";            myform.method="post";            var imageLoad=document.getElementById(fileId).cloneNode(true);            myform.appendChild(imageLoad);            document.body.appendChild(myform);            $(myform).ajaxSubmit({success: function(responseText)                {                                    var contentType=responseText.substring(0,responseText.indexOf(';'));                    var contentLength=responseText.substring(responseText.indexOf(';')+1);                    var d=new Date();                    $(imageId)[0].src="http://weblogs.asp.net/MoneypingAPP/ImagePreview/ImageLoad?a="+d.getTime();                    document.body.removeChild(myform);                }            });        }            You also need append the child in order to send request and remove them after receiving response.

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  • SQL SERVER – A Puzzle – Illusion – Confusion – April Fools’ Day

    - by pinaldave
    Today is April 1st and just like every other year, I like to bring something interesting and light for the day. Atleast there should be days in every one’s life when they should feel easy. Here is a quick puzzle for you and I believe it will make you feel extremely smart if you can figure out the result behind the same. Run following in SQL Server Management Studio and observe the output: SELECT 30.0/(-2.0)/5.0; SELECT 30.0/-2.0/5.0; Here are few questions for you: 1) What will be the result of above two queries? 2) Why? If you think you can figure out the result without executing them – I encourage you to execute BOTH of them in SSMS and see if they give you same result or different result. Well, now I am waiting for your answer here – why? I often post similar things on my facebook page http://facebook.com/SQLAuth – you are welcome to play with me there. Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: PostADay, SQL, SQL Authority, SQL Puzzle, SQL Query, SQL Scripts, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • The HTG Guide to Using a Bluetooth Keyboard with Your Android Device

    - by Matt Klein
    Android devices aren’t usually associated with physical keyboards. But, since Google is now bundling their QuickOffice app with the newly-released Kit-Kat, it appears inevitable that at least some Android tablets (particularly 10-inch models) will take on more productivity roles. In recent years, physical keyboards have been rendered obsolete by swipe style input methods such as Swype and Google Keyboard. Physical keyboards tend to make phones thick and plump, and that won’t fly today when thin (and even flexible and curved) is in vogue. So, you’ll be hard-pressed to find smartphone manufacturers launching new models with physical keyboards, thus rendering sliders to a past chapter in mobile phone evolution. It makes sense to ditch the clunky keyboard phone in favor of a lighter, thinner model. You’re going to carry around in your pocket or purse all day, why have that extra bulk and weight? That said, there is sound logic behind pairing tablets with keyboards. Microsoft continues to plod forward with its Surface models, and while critics continue to lavish praise on the iPad, its functionality is obviously enhanced and extended when you add a physical keyboard. Apple even has an entire page devoted specifically to iPad-compatible keyboards. But an Android tablet and a keyboard? Does such a thing even exist? They do actually. There are docking keyboards and keyboard/case combinations, there’s the Asus Transformer family, Logitech markets a Windows 8 keyboard that speaks “Android”, and these are just to name a few. So we know that keyboard products that are designed to work with Android exist, but what about an everyday Bluetooth keyboard you might use with Windows or OS X? How-To Geek wanted look at how viable it is to use such a keyboard with Android. We conducted some research and examined some lists of Android keyboard shortcuts. Most of what we found was long outdated. Many of the shortcuts don’t even apply anymore, while others just didn’t work. Regardless, after a little experimentation and a dash of customization, it turns out using a keyboard with Android is kind of fun, and who knows, maybe it will catch on. Setting things up Setting up a Bluetooth keyboard with Android is very easy. First, you’ll need a Bluetooth keyboard and of course an Android device, preferably running version 4.1 (Jelly Bean) or higher. For our test, we paired a second-generation Google Nexus 7 running Android 4.3 with a Samsung Series 7 keyboard. In Android, enable Bluetooth if it isn’t already on. We’d like to note that if you don’t normally use Bluetooth accessories and peripherals with your Android device (or any device really), it’s best practice to leave Bluetooth off because, like GPS, it drains the device’s battery more quickly. To enable Bluetooth, simply go to “Settings” -> “Bluetooth” and tap the slider button to “On”. To set up the keyboard, make sure it is on and then tap “Bluetooth” in the Android settings. On the resulting screen, your Android device should automatically search for and hopefully find your keyboard. If you don’t get it right the first time, simply turn the keyboard on again and then tap “Search for Devices” to try again. If it still doesn’t work, make sure you have fresh batteries and the keyboard isn’t paired to another device. If it is, you will need to unpair it before it will work with your Android device (consult your keyboard manufacturer’s documentation or Google if you don’t know how to do this). When Android finds your keyboard, select it under “Available Devices” … … and you should be prompted to type in a code: If successful, you will see that device is now “Connected” and you’re ready to go. If you want to test things out, try pressing the “Windows” key (“Apple” or “Command”) + ESC, and you will be whisked to your Home screen. So, what can you do? Traditional Mac and Windows users know there’s usually a keyboard shortcut for just about everything (and if there isn’t, there’s all kinds of ways to remap keys to do a variety of commands, tasks, and functions). So where does Android fall in terms of baked-in keyboard commands? There answer to that is kind of enough, but not too much. There are definitely established combos you can use to get around, but they aren’t clear and there doesn’t appear to be any one authority on what they are. Still, there is enough keyboard functionality in Android to make it a viable option, if only for those times when you need to get something done (long e-mail or important document) and an on-screen keyboard simply won’t do. It’s important to remember that Android is, and likely always will be a touch-first interface. That said, it does make some concessions to physical keyboards. In other words, you can get around Android fairly well without having to lift your hands off the keys, but you will still have to tap the screen regularly, unless you add a mouse. For example, you can wake your device by tapping a key rather than pressing its power button. However, if your device is slide or pattern-locked, then you’ll have to use the touchscreen to unlock it – a password or PIN however, works seamlessly with a keyboard – other things like widgets and app controls and features, have to be tapped. You get the idea. Keyboard shortcuts and navigation As we said, baked-in keyboard shortcut combos aren’t necessarily abundant nor apparent. The one thing you can always do is search. Any time you want to Google something, start typing from the Home screen and the search screen will automatically open and begin displaying results. Other than that, here is what we were able to figure out: ESC = go back CTRL + ESC = menu CTRL + ALT + DEL = restart (no questions asked) ALT + SPACE = search page (say “OK Google” to voice search) ALT + TAB (ALT + SHIFT + TAB) = switch tasks Also, if you have designated volume function keys, those will probably work too. There’s also some dedicated app shortcuts like calculator, Gmail, and a few others: CMD + A = calculator CMD + C = contacts CMD + E = e-mail CMD + G = Gmail CMD + L = Calendar CMD + P = Play Music CMD + Y = YouTube Overall, it’s not a long comprehensive list and there’s no dedicated keyboard combos for the full array of Google’s products. Granted, it’s hard to imagine getting a lot of mileage out of a keyboard with Maps but with something like Keep, you could type out long, detailed lists on your tablet, and then view them on your smartphone when you go out shopping. You can also use the arrow keys to navigate your Home screen over shortcuts and open the app drawer. When something on the screen is selected, it will be highlighted in blue. Press “Enter” to open your selection. Additionally, if an app has its own set of shortcuts, e.g. Gmail has quite a few unique shortcuts to it, as does Chrome, some – though not many – will work in Android (not for YouTube though). Also, many “universal” shortcuts such as Copy (CTRL + C), Cut (CTRL + X), Paste (CTRL + V), and Select All (CTRL + A) work where needed – such as in instant messaging, e-mail, social media apps, etc. Creating custom application shortcuts What about custom shortcuts? When we were researching this article, we were under the impression that it was possible to assign keyboard combinations to specific apps, such as you could do on older Android versions such as Gingerbread. This no long seems to be the case and nowhere in “Settings” could we find a way to assign hotkey combos to any of our favorite, oft-used apps or functions. If you do want custom keyboard shortcuts, what can you do? Luckily, there’s an app on Google Play that allows you to, among other things, create custom app shortcuts. It is called External Keyboard Helper (EKH) and while there is a free demo version, the pay version is only a few bucks. We decided to give EKH a whirl and through a little experimentation and finally reading the developer’s how-to, we found we could map custom keyboard combos to just about anything. To do this, first open the application and you’ll see the main app screen. Don’t worry about choosing a custom layout or anything like that, you want to go straight to the “Advanced settings”: In the “Advanced settings” select “Application shortcuts” to continue: You can have up to 16 custom application shortcuts. We are going to create a custom shortcut to the Facebook app. We choose “A0”, and from the resulting list, Facebook. You can do this for any number of apps, services, and settings. As you can now see, the Facebook app has now been linked to application-zero (A0): Go back to the “Advanced settings” and choose “Customize keyboard mappings”: You will be prompted to create a custom keyboard layout so we choose “Custom 1”: When you choose to create a custom layout, you can do a great many more things with your keyboard. For example, many keyboards have predefined function (Fn) keys, which you can map to your tablet’s brightness controls, toggle WiFi on/off, and much more. A word of advice, the application automatically remaps certain keys when you create a custom layout. This might mess up some existing keyboard combos. If you simply want to add some functionality to your keyboard, you can go ahead and delete EKH’s default changes and start your custom layout from scratch. To create a new combo, select “Add new key mapping”: For our new shortcut, we are going to assign the Facebook app to open when we key in “ALT + F”. To do this, we press the “F” key while in the “Scancode” field and we see it returns a value of “33”. If we wanted to use a different key, we can press “Change” and scan another key’s numerical value. We now want to assign the “ALT” key to application “A0”, previously designated as the Facebook app. In the “AltGr” field, we enter “A0” and then “Save” our custom combo. And now we see our new application shortcut. Now, as long as we’re using our custom layout, every time we press “ALT + F”, the Facebook app will launch: External Keyboard Helper extends far beyond simple application shortcuts and if you are looking for deeper keyboard customization options, you should definitely check it out. Among other things, EKH also supports dozens of languages, allows you to quickly switch between layouts using a key or combo, add up to 16 custom text shortcuts, and much more! It can be had on Google Play for $2.53 for the full version, but you can try the demo version for free. More extensive documentation on how to use the app is also available. Android? Keyboard? Sure, why not? Unlike traditional desktop operating systems, you don’t need a physical keyboard and mouse to use a mobile operating system. You can buy an iPad or Nexus 10 or Galaxy Note, and never need another accessory or peripheral – they work as intended right out of the box. It’s even possible you can write the next great American novel on one these devices, though that might require a lot of practice and patience. That said, using a keyboard with Android is kind of fun. It’s not revelatory but it does elevate the experience. You don’t even need to add customizations (though they are nice) because there are enough existing keyboard shortcuts in Android to make it usable. Plus, when it comes to inputting text such as in an editor or terminal application, we fully advocate big, physical keyboards. Bottom line, if you’re looking for a way to enhance your Android tablet, give a keyboard a chance. Do you use your Android device for productivity? Is a physical keyboard an important part of your setup? Do you have any shortcuts that we missed? Sound off in the comments and let us know what you think.     

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  • Integrate Bing Search API into ASP.Net application

    - by sreejukg
    Couple of months back, I wrote an article about how to integrate Bing Search engine (API 2.0) with ASP.Net website. You can refer the article here http://weblogs.asp.net/sreejukg/archive/2012/04/07/integrate-bing-api-for-search-inside-asp-net-web-application.aspx Things are changing rapidly in the tech world and Bing has also changed! The Bing Search API 2.0 will work until August 1, 2012, after that it will not return results. Shocked? Don’t worry the API has moved to Windows Azure market place and available for you to sign up and continue using it and there is a free version available based on your usage. In this article, I am going to explain how you can integrate the new Bing API that is available in the Windows Azure market place with your website. You can access the Windows Azure market place from the below link https://datamarket.azure.com/ There is lot of applications available for you to subscribe and use. Bing is one of them. You can find the new Bing Search API from the below link https://datamarket.azure.com/dataset/5BA839F1-12CE-4CCE-BF57-A49D98D29A44 To get access to Bing Search API, first you need to register an account with Windows Azure market place. Sign in to the Windows Azure market place site using your windows live account. Once you sign in with your windows live account, you need to register to Windows Azure Market place account. From the Windows Azure market place, you will see the sign in button it the top right of the page. Clicking on the sign in button will take you to the Windows live ID authentication page. You can enter a windows live ID here to login. Once logged in you will see the Registration page for the Windows Azure market place as follows. You can agree or disagree for the email address usage by Microsoft. I believe selecting the check box means you will get email about what is happening in Windows Azure market place. Click on continue button once you are done. In the next page, you should accept the terms of use, it is not optional, you must agree to terms and conditions. Scroll down to the page and select the I agree checkbox and click on Register Button. Now you are a registered member of Windows Azure market place. You can subscribe to data applications. In order to use BING API in your application, you must obtain your account Key, in the previous version of Bing you were required an API key, the current version uses Account Key instead. Once you logged in to the Windows Azure market place, you can see “My Account” in the top menu, from the Top menu; go to “My Account” Section. From the My Account section, you can manage your subscriptions and Account Keys. Account Keys will be used by your applications to access the subscriptions from the market place. Click on My Account link, you can see Account Keys in the left menu and then Add an account key or you can use the default Account key available. Creating account key is very simple process. Also you can remove the account keys you create if necessary. The next step is to subscribe to BING Search API. At this moment, Bing Offers 2 APIs for search. The available options are as follows. 1. Bing Search API - https://datamarket.azure.com/dataset/5ba839f1-12ce-4cce-bf57-a49d98d29a44 2. Bing Search API – Web Results only - https://datamarket.azure.com/dataset/8818f55e-2fe5-4ce3-a617-0b8ba8419f65 The difference is that the later will give you only web results where the other you can specify the source type such as image, video, web, news etc. Carefully choose the API based on your application requirements. In this article, I am going to use Web Results Only API, but the steps will be similar to both. Go to the API page https://datamarket.azure.com/dataset/8818f55e-2fe5-4ce3-a617-0b8ba8419f65, you can see the subscription options in the right side. And in the bottom of the page you can see the free option Since I am going to use the free options, just Click the Sign Up link for that. Just select I agree check box and click on the Sign Up button. You will get a recipt pagethat detail your subscription. Now you are ready Bing Search API – Web results. The next step is to integrate the API into your ASP.Net application. Now if you go to the Search API page (as well as in the Receipt page), you can see a .Net C# Class Library link, click on the link, you will get a code file named “BingSearchContainer.cs”. In the following sections I am going to demonstrate the use of Bing Search API from an ASP.Net application. Create an empty ASP.Net web application. In the solution explorer, the application will looks as follows. Now add the downloaded code file (“BingSearchContainer.cs”) to the project. Right click your project in solution explorer, Add -> existing item, then browse to the downloaded location, select the “BingSearchContainer.cs” file and add it to the project. To build the code file you need to add reference to the following library. System.Data.Services.Client You can find the library in the .Net tab, when you select Add -> Reference Try to build your project now; it should build without any errors. Add an ASP.Net page to the project. I have included a text box and a button, then a Grid View to the page. The idea is to Search the text entered and display the results in the gridview. The page will look in the Visual Studio Designer as follows. The markup of the page is as follows. In the button click event handler for the search button, I have used the following code. Now run your project and enter some text in the text box and click the Search button, you will see the results coming from Bing, cool. I entered the text “Microsoft” in the textbox and clicked on the button and I got the following results. Searching Specific Websites If you want to search a particular website, you pass the site url with site:<site url name> and if you have more sites, use pipe (|). e.g. The following search query site:microsoft.com | site:adobe.com design will search the word design and return the results from Microsoft.com and Adobe.com See the sample code that search only Microsoft.com for the text entered for the above sample. var webResults = bingContainer.Web("site:www.Microsoft.com " + txtSearch.Text, null, null, null, null, null, null); Paging the results returned by the API By default the BING API will return 100 results based on your query. The default code file that you downloaded from BING doesn’t include any option for this. You can modify the downloaded code to perform this paging. The BING API supports two parameters $top (for number of results to return) and $skip (for number of records to skip). So if you want 3rd page of results with page size = 10, you need to pass $top = 10 and $skip=20. Open the BingSearchContainer.cs in the editor. You can see the Web method in it as follows. public DataServiceQuery<WebResult> Web(String Query, String Market, String Adult, Double? Latitude, Double? Longitude, String WebFileType, String Options) {  In the method signature, I have added two more parameters public DataServiceQuery<WebResult> Web(String Query, String Market, String Adult, Double? Latitude, Double? Longitude, String WebFileType, String Options, int resultCount, int pageNo) { and in the method, you need to pass the parameters to the query variable. query = query.AddQueryOption("$top", resultCount); query = query.AddQueryOption("$skip", (pageNo -1)*resultCount); return query; Note that I didn’t perform any validation, but you need to check conditions such as resultCount and pageCount should be greater than or equal to 1. If the parameters are not valid, the Bing Search API will throw the error. The modified method is as follows. The changes are highlighted. Now see the following code in the SearchPage.aspx.cs file protected void btnSearch_Click(object sender, EventArgs e) {     var bingContainer = new Bing.BingSearchContainer(new Uri(https://api.datamarket.azure.com/Bing/SearchWeb/));     // replace this value with your account key     var accountKey = "your key";     // the next line configures the bingContainer to use your credentials.     bingContainer.Credentials = new NetworkCredential(accountKey, accountKey);     var webResults = bingContainer.Web("site:microsoft.com" +txtSearch.Text , null, null, null, null, null, null,3,2);     lstResults.DataSource = webResults;     lstResults.DataBind(); } The following code will return 3 results starting from second page (by skipping first 3 results). See the result page as follows. Bing provides complete integration to its offerings. When you develop search based applications, you can use the power of Bing to perform the search. Integrating Bing Search API to ASP.Net application is a simple process and without investing much time, you can develop a good search based application. Make sure you read the terms of use before designing the application and decide which API usage is suitable for you. Further readings BING API Migration Guide http://go.microsoft.com/fwlink/?LinkID=248077 Bing API FAQ http://go.microsoft.com/fwlink/?LinkID=252146 Bing API Schema Guide http://go.microsoft.com/fwlink/?LinkID=252151

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  • saslauthd + PostFix producing password verification and authentication errors

    - by Aram Papazian
    So I'm trying to setup PostFix while using SASL (Cyrus variety preferred, I was using dovecot earlier but I'm switching from dovecot to courier so I want to use cyrus instead of dovecot) but I seem to be having issues. Here are the errors I'm receiving: ==> mail.log <== Aug 10 05:11:49 crazyinsanoman postfix/smtpd[779]: warning: SASL authentication failure: Password verification failed Aug 10 05:11:49 crazyinsanoman postfix/smtpd[779]: warning: ipname[xx.xx.xx.xx]: SASL PLAIN authentication failed: authentication failure ==> mail.info <== Aug 10 05:11:49 crazyinsanoman postfix/smtpd[779]: warning: SASL authentication failure: Password verification failed Aug 10 05:11:49 crazyinsanoman postfix/smtpd[779]: warning: ipname[xx.xx.xx.xx]: SASL PLAIN authentication failed: authentication failure ==> mail.warn <== Aug 10 05:11:49 crazyinsanoman postfix/smtpd[779]: warning: SASL authentication failure: Password verification failed Aug 10 05:11:49 crazyinsanoman postfix/smtpd[779]: warning: ipname[xx.xx.xx.xx]: SASL PLAIN authentication failed: authentication failure I tried $testsaslauthd -u xxxx -p xxxx 0: OK "Success." So I know that the password/user I'm using is correct. I'm thinking that most likely I have a setting wrong somewhere, but can't seem to find where. Here is my files. Here is my main.cf for postfix: # See /usr/share/postfix/main.cf.dist for a commented, more complete version # Debian specific: Specifying a file name will cause the first # line of that file to be used as the name. The Debian default # is /etc/mailname. myorigin = /etc/mailname # This is already done in /etc/mailname #myhostname = crazyinsanoman.xxxxx.com smtpd_banner = $myhostname ESMTP $mail_name #biff = no # appending .domain is the MUA's job. #append_dot_mydomain = no readme_directory = /usr/share/doc/postfix # TLS parameters smtpd_tls_cert_file = /etc/postfix/smtpd.cert smtpd_tls_key_file = /etc/postfix/smtpd.key smtpd_use_tls = yes smtpd_tls_session_cache_database = btree:${data_directory}/smtpd_scache smtp_tls_session_cache_database = btree:${data_directory}/smtp_scache # Relay smtp through another server or leave blank to do it yourself #relayhost = smtp.yourisp.com # Network details; Accept connections from anywhere, and only trust this machine mynetworks = 127.0.0.0/8 inet_interfaces = all #mynetworks_style = host #As we will be using virtual domains, these need to be empty local_recipient_maps = mydestination = # how long if undelivered before sending "delayed mail" warning update to sender delay_warning_time = 4h # will it be a permanent error or temporary unknown_local_recipient_reject_code = 450 # how long to keep message on queue before return as failed. # some have 3 days, I have 16 days as I am backup server for some people # whom go on holiday with their server switched off. maximal_queue_lifetime = 7d # max and min time in seconds between retries if connection failed minimal_backoff_time = 1000s maximal_backoff_time = 8000s # how long to wait when servers connect before receiving rest of data smtp_helo_timeout = 60s # how many address can be used in one message. # effective stopper to mass spammers, accidental copy in whole address list # but may restrict intentional mail shots. smtpd_recipient_limit = 16 # how many error before back off. smtpd_soft_error_limit = 3 # how many max errors before blocking it. smtpd_hard_error_limit = 12 # Requirements for the HELO statement smtpd_helo_restrictions = permit_mynetworks, warn_if_reject reject_non_fqdn_hostname, reject_invalid_hostname, permit # Requirements for the sender details smtpd_sender_restrictions = permit_mynetworks, warn_if_reject reject_non_fqdn_sender, reject_unknown_sender_domain, reject_unauth_pipelining, permit # Requirements for the connecting server smtpd_client_restrictions = reject_rbl_client sbl.spamhaus.org, reject_rbl_client blackholes.easynet.nl, reject_rbl_client dnsbl.njabl.org # Requirement for the recipient address smtpd_recipient_restrictions = reject_unauth_pipelining, permit_mynetworks, reject_non_fqdn_recipient, reject_unknown_recipient_domain, reject_unauth_destination, permit smtpd_data_restrictions = reject_unauth_pipelining # require proper helo at connections smtpd_helo_required = yes # waste spammers time before rejecting them smtpd_delay_reject = yes disable_vrfy_command = yes # not sure of the difference of the next two # but they are needed for local aliasing alias_maps = hash:/etc/postfix/aliases alias_database = hash:/etc/postfix/aliases # this specifies where the virtual mailbox folders will be located virtual_mailbox_base = /var/spool/mail/vmail # this is for the mailbox location for each user virtual_mailbox_maps = mysql:/etc/postfix/mysql_mailbox.cf # and this is for aliases virtual_alias_maps = mysql:/etc/postfix/mysql_alias.cf # and this is for domain lookups virtual_mailbox_domains = mysql:/etc/postfix/mysql_domains.cf # this is how to connect to the domains (all virtual, but the option is there) # not used yet # transport_maps = mysql:/etc/postfix/mysql_transport.cf # Setup the uid/gid of the owner of the mail files - static:5000 allows virtual ones virtual_uid_maps = static:5000 virtual_gid_maps = static:5000 inet_protocols=all # Cyrus SASL Support smtpd_sasl_path = smtpd smtpd_sasl_local_domain = xxxxx.com ####################### ## OLD CONFIGURATION ## ####################### #myorigin = /etc/mailname #mydestination = crazyinsanoman.xxxxx.com, localhost, localhost.localdomain #mailbox_size_limit = 0 #recipient_delimiter = + #html_directory = /usr/share/doc/postfix/html message_size_limit = 30720000 #virtual_alias_domains = ##virtual_alias_maps = hash:/etc/postfix/virtual #virtual_mailbox_base = /home/vmail ##luser_relay = webmaster #smtpd_sasl_type = dovecot #smtpd_sasl_path = private/auth smtpd_sasl_auth_enable = yes smtpd_sasl_security_options = noanonymous broken_sasl_auth_clients = yes #smtpd_sasl_authenticated_header = yes smtpd_recipient_restrictions = permit_mynetworks, permit_sasl_authenticated, reject_unauth_destination #virtual_create_maildirsize = yes #virtual_maildir_extended = yes #proxy_read_maps = $local_recipient_maps $mydestination $virtual_alias_maps $virtual_alias_domains $virtual_mailbox_maps $virtual_mailbox_domains $relay_recipient_maps $relay_domains $canonical_maps $sender_canonical_maps $recipient_canonical_maps $relocated_maps $transport_maps $mynetworks $virtual_mailbox_limit_maps #virtual_transport = dovecot #dovecot_destination_recipient_limit = 1 Here is my master.cf: # # Postfix master process configuration file. For details on the format # of the file, see the master(5) manual page (command: "man 5 master"). # # Do not forget to execute "postfix reload" after editing this file. # # ========================================================================== # service type private unpriv chroot wakeup maxproc command + args # (yes) (yes) (yes) (never) (100) # ========================================================================== smtp inet n - - - - smtpd submission inet n - - - - smtpd -o smtpd_tls_security_level=encrypt -o smtpd_sasl_auth_enable=yes -o smtpd_client_restrictions=permit_sasl_authenticated,reject # -o milter_macro_daemon_name=ORIGINATING #smtps inet n - - - - smtpd # -o smtpd_tls_wrappermode=yes # -o smtpd_sasl_auth_enable=yes # -o smtpd_client_restrictions=permit_sasl_authenticated,reject # -o milter_macro_daemon_name=ORIGINATING #628 inet n - - - - qmqpd pickup fifo n - - 60 1 pickup cleanup unix n - - - 0 cleanup qmgr fifo n - n 300 1 qmgr #qmgr fifo n - - 300 1 oqmgr tlsmgr unix - - - 1000? 1 tlsmgr rewrite unix - - - - - trivial-rewrite bounce unix - - - - 0 bounce defer unix - - - - 0 bounce trace unix - - - - 0 bounce verify unix - - - - 1 verify flush unix n - - 1000? 0 flush proxymap unix - - n - - proxymap proxywrite unix - - n - 1 proxymap smtp unix - - - - - smtp # When relaying mail as backup MX, disable fallback_relay to avoid MX loops relay unix - - - - - smtp -o smtp_fallback_relay= # -o smtp_helo_timeout=5 -o smtp_connect_timeout=5 showq unix n - - - - showq error unix - - - - - error retry unix - - - - - error discard unix - - - - - discard local unix - n n - - local virtual unix - n n - - virtual lmtp unix - - - - - lmtp anvil unix - - - - 1 anvil scache unix - - - - 1 scache # # ==================================================================== # Interfaces to non-Postfix software. Be sure to examine the manual # pages of the non-Postfix software to find out what options it wants. # # Many of the following services use the Postfix pipe(8) delivery # agent. See the pipe(8) man page for information about ${recipient} # and other message envelope options. # ==================================================================== # # maildrop. See the Postfix MAILDROP_README file for details. # Also specify in main.cf: maildrop_destination_recipient_limit=1 # maildrop unix - n n - - pipe flags=DRhu user=vmail argv=/usr/bin/maildrop -d ${recipient} # # ==================================================================== # # Recent Cyrus versions can use the existing "lmtp" master.cf entry. # # Specify in cyrus.conf: # lmtp cmd="lmtpd -a" listen="localhost:lmtp" proto=tcp4 # # Specify in main.cf one or more of the following: # mailbox_transport = lmtp:inet:localhost # virtual_transport = lmtp:inet:localhost # # ==================================================================== # # Cyrus 2.1.5 (Amos Gouaux) # Also specify in main.cf: cyrus_destination_recipient_limit=1 # cyrus unix - n n - - pipe user=cyrus argv=/cyrus/bin/deliver -e -r ${sender} -m ${extension} ${user} # # ==================================================================== # Old example of delivery via Cyrus. # #old-cyrus unix - n n - - pipe # flags=R user=cyrus argv=/cyrus/bin/deliver -e -m ${extension} ${user} # # ==================================================================== # # See the Postfix UUCP_README file for configuration details. # uucp unix - n n - - pipe flags=Fqhu user=uucp argv=uux -r -n -z -a$sender - $nexthop!rmail ($recipient) # # Other external delivery methods. # ifmail unix - n n - - pipe flags=F user=ftn argv=/usr/lib/ifmail/ifmail -r $nexthop ($recipient) bsmtp unix - n n - - pipe flags=Fq. user=bsmtp argv=/usr/lib/bsmtp/bsmtp -t$nexthop -f$sender $recipient scalemail-backend unix - n n - 2 pipe flags=R user=scalemail argv=/usr/lib/scalemail/bin/scalemail-store ${nexthop} ${user} ${extension} mailman unix - n n - - pipe flags=FR user=list argv=/usr/lib/mailman/bin/postfix-to-mailman.py ${nexthop} ${user} #dovecot unix - n n - - pipe # flags=DRhu user=vmail:vmail argv=/usr/lib/dovecot/deliver -d ${recipient} Here is what I'm using for /etc/postfix/sasl/smtpd.conf log_level: 7 pwcheck_method: saslauthd pwcheck_method: auxprop mech_list: PLAIN LOGIN CRAM-MD5 DIGEST-MD5 allow_plaintext: true auxprop_plugin: mysql sql_hostnames: 127.0.0.1 sql_user: xxxxx sql_passwd: xxxxx sql_database: maildb sql_select: select crypt from users where id = '%u' As you can see I'm trying to use mysql as my authentication method. The password in 'users' is set through the 'ENCRYPT()' function. I also followed the methods found in http://www.jimmy.co.at/weblog/?p=52 in order to redo /var/spool/postfix/var/run/saslauthd as that seems to be a lot of people's problems, but that didn't help at all. Also, here is my /etc/default/saslauthd START=yes DESC="SASL Authentication Daemon" NAME="saslauthd" # Which authentication mechanisms should saslauthd use? (default: pam) # # Available options in this Debian package: # getpwent -- use the getpwent() library function # kerberos5 -- use Kerberos 5 # pam -- use PAM # rimap -- use a remote IMAP server # shadow -- use the local shadow password file # sasldb -- use the local sasldb database file # ldap -- use LDAP (configuration is in /etc/saslauthd.conf) # # Only one option may be used at a time. See the saslauthd man page # for more information. # # Example: MECHANISMS="pam" MECHANISMS="pam" MECH_OPTIONS="" THREADS=5 OPTIONS="-c -m /var/spool/postfix/var/run/saslauthd -r" I had heard that potentially changing MECHANISM to MECHANISMS="mysql" but obviously that didn't help as is shown by the options listed above and also by trying it out anyway in case the documentation was outdated. So, I'm now at a loss... I have no idea where to go from here or what steps I need to do to get this working =/ Anyone have any ideas? EDIT: Here is the error that is coming from auth.log ... I don't know if this will help at all, but here you go: Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql auxprop plugin using mysql engine Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin Parse the username [email protected] Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin try and connect to a host Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin trying to open db 'maildb' on host '127.0.0.1' Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin Parse the username [email protected] Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin try and connect to a host Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin trying to open db 'maildb' on host '127.0.0.1' Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: begin transaction Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin create statement from userPassword user xxxxxx.com Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin doing query select crypt from users where id = '[email protected]'; Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin create statement from cmusaslsecretPLAIN user xxxxxx.com Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin doing query select crypt from users where id = '[email protected]'; Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: commit transaction Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin Parse the username [email protected] Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin try and connect to a host Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin trying to open db 'maildb' on host '127.0.0.1' Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin Parse the username [email protected] Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin try and connect to a host Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin trying to open db 'maildb' on host '127.0.0.1' Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin Parse the username [email protected] Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin try and connect to a host Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin trying to open db 'maildb' on host '127.0.0.1' Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin Parse the username [email protected] Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin try and connect to a host Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin trying to open db 'maildb' on host '127.0.0.1' Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: begin transaction Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin create statement from userPassword user xxxxxx.com Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin doing query select crypt from users where id = '[email protected]'; Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin create statement from cmusaslsecretPLAIN user xxxxxx.com Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin doing query select crypt from users where id = '[email protected]'; Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: commit transaction Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin Parse the username [email protected] Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin try and connect to a host Aug 11 17:19:56 crazyinsanoman postfix/smtpd[9503]: sql plugin trying to open db 'maildb' on host '127.0.0.1'

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  • SQL SERVER – Difference Between GRANT and WITH GRANT

    - by pinaldave
    This was very interesting question recently asked me to during my session at TechMela Nepal. The question is what is the difference between GRANT and WITH GRANT when giving permissions to user. Let us first see syntax for the same. GRANT: USE master; GRANT VIEW ANY DATABASE TO username; GO WITH GRANT: USE master; GRANT VIEW ANY DATABASE TO username WITH GRANT OPTION; GO The difference between both of this option is very simple. In case of only GRANT – username can not grant the same permission to other users. In case, of the option of WITH GRANT – username will be able to give the permission it has received to other users. This is very basic definition of the subject. I would like to request my readers to come up with working script to prove this scenario. If can submit your script to me by email (pinal ‘at’ sqlauthority.com) or in comment field. Reference : Pinal Dave (http://blog.SQLAuthority.com) Filed under: Pinal Dave, SQL, SQL Authority, SQL Query, SQL Scripts, SQL Security, SQL Server, SQL Tips and Tricks, T SQL, Technology Tagged: SQL Permissions

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  • SQL SERVER – Solution – Challenge – Puzzle – Usage of FAST Hint

    - by pinaldave
    Earlier I had posted quick puzzle and I had received wonderful response to the same from Brad Schulz. Today we will go over the solution. The puzzle was posted here: SQL SERVER – Challenge – Puzzle – Usage of FAST Hint The question was in what condition the hint FAST will be useful. In the response to this puzzle blog post here is what SQL Server Expert Brad Schulz has pointed me to his blog post where he explain how FAST hint can be useful. I strongly recommend to read his blog post over here. With the permission of the Brad, I am reproducing following queries here. He has come up with example where FAST hint improves the performance. USE AdventureWorks GO DECLARE @DesiredDateAtMidnight DATETIME = '20010709' DECLARE @NextDateAtMidnight DATETIME = DATEADD(DAY,1,@DesiredDateAtMidnight) -- Query without FAST SELECT OrderID=h.SalesOrderID ,h.OrderDate ,h.TerritoryID ,TerritoryName=t.Name ,c.CardType ,c.CardNumber ,CardExpire=RIGHT(STR(100+ExpMonth),2)+'/'+STR(ExpYear,4) ,h.TotalDue FROM Sales.SalesOrderHeader h LEFT JOIN Sales.SalesTerritory t ON h.TerritoryID=t.TerritoryID LEFT JOIN Sales.CreditCard c ON h.CreditCardID=c.CreditCardID WHERE OrderDate>=@DesiredDateAtMidnight AND OrderDate<@NextDateAtMidnight ORDER BY h.SalesOrderID; -- Query with FAST(10) SELECT OrderID=h.SalesOrderID ,h.OrderDate ,h.TerritoryID ,TerritoryName=t.Name ,c.CardType ,c.CardNumber ,CardExpire=RIGHT(STR(100+ExpMonth),2)+'/'+STR(ExpYear,4) ,h.TotalDue FROM Sales.SalesOrderHeader h LEFT JOIN Sales.SalesTerritory t ON h.TerritoryID=t.TerritoryID LEFT JOIN Sales.CreditCard c ON h.CreditCardID=c.CreditCardID WHERE OrderDate>=@DesiredDateAtMidnight AND OrderDate<@NextDateAtMidnight ORDER BY h.SalesOrderID OPTION(FAST 10) Now when you check the execution plan for the same, you will find following visible difference. You will find query with FAST returns results with much lower cost. Thank you Brad for excellent post and teaching us something. I request all of you to read original blog post written by Brad for much more information. Reference: Pinal Dave (http://blog.sqlauthority.com) Filed under: Pinal Dave, Readers Contribution, Readers Question, SQL, SQL Authority, SQL Puzzle, SQL Query, SQL Scripts, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • Entity Framework version 1- Brief Synopsis and Tips &ndash; Part 1

    - by Rohit Gupta
    To Do Eager loading use Projections (for e.g. from c in context.Contacts select c, c.Addresses)  or use Include Query Builder Methods (Include(“Addresses”)) If there is multi-level hierarchical Data then to eager load all the relationships use Include Query Builder methods like customers.Include("Order.OrderDetail") to include Order and OrderDetail collections or use customers.Include("Order.OrderDetail.Location") to include all Order, OrderDetail and location collections with a single include statement =========================================================================== If the query uses Joins then Include() Query Builder method will be ignored, use Nested Queries instead If the query does projections then Include() Query Builder method will be ignored Use Address.ContactReference.Load() OR Contact.Addresses.Load() if you need to Deferred Load Specific Entity – This will result in extra round trips to the database ObjectQuery<> cannot return anonymous types... it will return a ObjectQuery<DBDataRecord> Only Include method can be added to Linq Query Methods Any Linq Query method can be added to Query Builder methods. If you need to append a Query Builder Method (other than Include) after a LINQ method  then cast the IQueryable<Contact> to ObjectQuery<Contact> and then append the Query Builder method to it =========================================================================== Query Builder methods are Select, Where, Include Methods which use Entity SQL as parameters e.g. "it.StartDate, it.EndDate" When Query Builder methods do projection then they return ObjectQuery<DBDataRecord>, thus to iterate over this collection use contact.Item[“Name”].ToString() When Linq To Entities methods do projection, they return collection of anonymous types --- thus the collection is strongly typed and supports Intellisense EF Object Context can track changes only on Entities, not on Anonymous types. If you use a Defining Query for a EntitySet then the EntitySet becomes readonly since a Defining Query is the same as a View (which is treated as a ReadOnly by default). However if you want to use this EntitySet for insert/update/deletes then we need to map stored procs (as created in the DB) to the insert/update/delete functions of the Entity in the Designer You can use either Execute method or ToList() method to bind data to datasources/bindingsources If you use the Execute Method then remember that you can traverse through the ObjectResult<> collection (returned by Execute) only ONCE. In WPF use ObservableCollection to bind to data sources , for keeping track of changes and letting EF send updates to the DB automatically. Use Extension Methods to add logic to Entities. For e.g. create extension methods for the EntityObject class. Create a method in ObjectContext Partial class and pass the entity as a parameter, then call this method as desired from within each entity. ================================================================ DefiningQueries and Stored Procedures: For Custom Entities, one can use DefiningQuery or Stored Procedures. Thus the Custom Entity Collection will be populated using the DefiningQuery (of the EntitySet) or the Sproc. If you use Sproc to populate the EntityCollection then the query execution is immediate and this execution happens on the Server side and any filters applied will be applied in the Client App. If we use a DefiningQuery then these queries are composable, meaning the filters (if applied to the entityset) will all be sent together as a single query to the DB, returning only filtered results. If the sproc returns results that cannot be mapped to existing entity, then we first create the Entity/EntitySet in the CSDL using Designer, then create a dummy Entity/EntitySet using XML in the SSDL. When creating a EntitySet in the SSDL for this dummy entity, use a TSQL that does not return any results, but does return the relevant columns e.g. select ContactID, FirstName, LastName from dbo.Contact where 1=2 Also insure that the Entity created in the SSDL uses the SQL DataTypes and not .NET DataTypes. If you are unable to open the EDMX file in the designer then note the Errors ... they will give precise info on what is wrong. The Thrid option is to simply create a Native Query in the SSDL using <Function Name="PaymentsforContact" IsComposable="false">   <CommandText>SELECT ActivityId, Activity AS ActivityName, ImagePath, Category FROM dbo.Activities </CommandText></FuncTion> Then map this Function to a existing Entity. This is a quick way to get a custom Entity which is regular Entity with renamed columns or additional columns (which are computed columns). The disadvantage to using this is that It will return all the rows from the Defining query and any filter (if defined) will be applied only at the Client side (after getting all the rows from DB). If you you DefiningQuery instead then we can use that as a Composable Query. The Fourth option (for mapping a READ stored proc results to a non-existent Entity) is to create a View in the Database which returns all the fields that the sproc also returns, then update the Model so that the model contains this View as a Entity. Then map the Read Sproc to this View Entity. The other option would be to simply create the View and remove the sproc altogether. ================================================================ To Execute a SProc that does not return a entity, use a EntityCommand to execute that proc. You cannot call a sproc FunctionImport that does not return Entities From Code, the only way is to use SSDL function calls using EntityCommand.  This changes with EntityFramework Version 4 where you can return Scalar Types, Complex Types, Entities or NonQuery ================================================================ UDF when created as a Function in SSDL, we need to set the Name & IsComposable properties for the Function element. IsComposable is always false for Sprocs, for UDF's set this to true. You cannot call UDF "Function" from within code since you cannot import a UDF Function into the CSDL Model (with Version 1 of EF). only stored procedures can be imported and then mapped to a entity ================================================================ Entity Framework requires properties that are involved in association mappings to be mapped in all of the function mappings for the entity (Insert, Update and Delete). Because Payment has an association to Reservation... hence we need to pass both the paymentId and reservationId to the Delete sproc even though just the paymentId is the PK on the Payment Table. ================================================================ When mapping insert, update and delete procs to a Entity, insure that all the three or none are mapped. Further if you have a base class and derived class in the CSDL, then you must map (ins, upd, del) sprocs to all parent and child entities in the inheritance relationship. Note that this limitation that base and derived entity methods must all must be mapped does not apply when you are mapping Read Stored Procedures.... ================================================================ You can write stored procedures SQL directly into the SSDL by creating a Function element in the SSDL and then once created, you can map this Function to a CSDL Entity directly in the designer during Function Import ================================================================ You can do Entity Splitting such that One Entity maps to multiple tables in the DB. For e.g. the Customer Entity currently derives from Contact Entity...in addition it also references the ContactPersonalInfo Entity. One can copy all properties from the ContactPersonalInfo Entity into the Customer Entity and then Delete the CustomerPersonalInfo entity, finall one needs to map the copied properties to the ContactPersonalInfo Table in Table Mapping (by adding another table (ContactPersonalInfo) to the Table Mapping... this is called Entity Splitting. Thus now when you insert a Customer record, it will automatically create SQL to insert records into the Contact, Customers and ContactPersonalInfo tables even though you have a Single Entity called Customer in the CSDL =================================================================== There is Table by Type Inheritance where another EDM Entity can derive from another EDM entity and absorb the inherted entities properties, for example in the Break Away Geek Adventures EDM, the Customer entity derives (inherits) from the Contact Entity and absorbs all the properties of Contact entity. Thus when you create a Customer Entity in Code and then call context.SaveChanges the Object Context will first create the TSQL to insert into the Contact Table followed by a TSQL to insert into the Customer table =================================================================== Then there is the Table per Hierarchy Inheritance..... where different types are created based on a condition (similar applying a condition to filter a Entity to contain filtered records)... the diference being that the filter condition populates a new Entity Type (derived from the base Entity). In the BreakAway sample the example is Lodging Entity which is a Abstract Entity and Then Resort and NonResort Entities which derive from Lodging Entity and records are filtered based on the value of the Resort Boolean field =================================================================== Then there is Table per Concrete Type Hierarchy where we create a concrete Entity for each table in the database. In the BreakAway sample there is a entity for the Reservation table and another Entity for the OldReservation table even though both the table contain the same number of fields. The OldReservation Entity can then inherit from the Reservation Entity and configure the OldReservation Entity to remove all Scalar Properties from the Entity (since it inherits the properties from Reservation and filters based on ReservationDate field) =================================================================== Complex Types (Complex Properties) Entities in EF can also contain Complex Properties (in addition to Scalar Properties) and these Complex Properties reference a ComplexType (not a EntityType) DropdownList, ListBox, RadioButtonList, CheckboxList, Bulletedlist are examples of List server controls (not data bound controls) these controls cannot use Complex properties during databinding, they need Scalar Properties. So if a Entity contains Complex properties and you need to bind those to list server controls then use projections to return Scalar properties and bind them to the control (the disadvantage is that projected collections are not tracked by the Object Context and hence cannot persist changes to the projected collections bound to controls) ObjectDataSource and EntityDataSource do account for Complex properties and one can bind entities with Complex Properties to Data Source controls and they will be tracked for changes... with no additional plumbing needed to persist changes to these collections bound to controls So DataBound controls like GridView, FormView need to use EntityDataSource or ObjectDataSource as a datasource for entities that contain Complex properties so that changes to the datasource done using the GridView can be persisted to the DB (enabling the controls for updates)....if you cannot use the EntityDataSource you need to flatten the ComplexType Properties using projections With EF Version 4 ComplexTypes are supported by the Designer and can add/remove/compose Complex Types directly using the Designer =================================================================== Conditional Mapping ... is like Table per Hierarchy Inheritance where Entities inherit from a base class and then used conditions to populate the EntitySet (called conditional Mapping). Conditional Mapping has limitations since you can only use =, is null and IS NOT NULL Conditions to do conditional mapping. If you need more operators for filtering/mapping conditionally then use QueryView(or possibly Defining Query) to create a readonly entity. QueryView are readonly by default... the EntitySet created by the QueryView is enabled for change tracking by the ObjectContext, however the ObjectContext cannot create insert/update/delete TSQL statements for these Entities when SaveChanges is called since it is QueryView. One way to get around this limitation is to map stored procedures for the insert/update/delete operations in the Designer. =================================================================== Difference between QueryView and Defining Query : QueryView is defined in the (MSL) Mapping File/section of the EDM XML, whereas the DefiningQuery is defined in the store schema (SSDL). QueryView is written using Entity SQL and is this database agnostic and can be used against any database/Data Layer. DefiningQuery is written using Database Lanaguage i.e. TSQL or PSQL thus you have more control =================================================================== Performance: Lazy loading is deferred loading done automatically. lazy loading is supported with EF version4 and is on by default. If you need to turn it off then use context.ContextOptions.lazyLoadingEnabled = false To improve Performance consider PreCompiling the ObjectQuery using the CompiledQuery.Compile method

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  • Agile PLM on Developing Agile PLM: Software Lifecycle Management

    - by Kerrie Foy
    Change is constant.  That saying couldn’t be truer when applied to software development.   And with all that change comes extensive product complexity.  How do you manage it all?  As software developers ourselves, we can certainly empathize with the challenge. On April 3, 2012 Stephen Van Lare, VP of PLM Product Development, hosted a webcast to share how Oracle uses Agile to develop Agile – a PLM solution for managing a PLM solution!   Stephen passionately shared his unique insight based on 10 years of using Agile PLM to manage the development process, as well as customer use cases.  He shared our time-proven view of the software’s relationship to the product record, while pointing out that PLM is not source control.  He began with the challenges of software development, which boiled down to the deduction that “despite many great tools in the software development industry, it takes a lot more than good source control, more than good bug tracking, to get to an on-time, on-budget and quality release in your marketplace.   It requires defining the right things you want to do, managing the scope, managing your schedule, and, most importantly, managing the change to all those things over the lifecycle of the process. And this is the definition of PLM.”   Stephen then defined the relationship of PLM to the software development process by detailing the two main use cases –  Product Lifecycle and Mechatronics – which can be used simultaneously and in fact are already used in most industries today.  The Product Lifecycle use case is used to manage artifacts and change throughout product development, while the Mechatronics use case involves the software, hardware and electrical design in the BOM.  In essence, PLM is just as relevant to software as the rest of the BOM when trying to maximize profits during any phase of the lifecycle. Please take the opportunity to watch Stephen Van Lare as he details how and why based on his own experience developing Agile with Agile, as well as a lively Q&A session, in the Software PLM Webcast Replay.

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  • SQL SERVER – GUID vs INT – Your Opinion

    - by pinaldave
    I think the title is clear what I am going to write in your post. This is age old problem and I want to compile the list stating advantages and disadvantages of using GUID and INT as a Primary Key or Clustered Index or Both (the usual case). Let me start a list by suggesting one advantage and one disadvantage in each case. INT Advantage: Numeric values (and specifically integers) are better for performance when used in joins, indexes and conditions. Numeric values are easier to understand for application users if they are displayed. Disadvantage: If your table is large, it is quite possible it will run out of it and after some numeric value there will be no additional identity to use. GUID Advantage: Unique across the server. Disadvantage: String values are not as optimal as integer values for performance when used in joins, indexes and conditions. More storage space is required than INT. Please note that I am looking to create list of all the generic comparisons. There can be special cases where the stated information is incorrect, feel free to comment on the same. Please leave your opinion and advice in comment section. I will combine a final list and update this blog after a week. By listing your name in post, I will also give due credit. Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: Pinal Dave, SQL, SQL Authority, SQL Constraint and Keys, SQL Data Storage, SQL Performance, SQL Query, SQL Server, SQL Tips and Tricks, SQLServer, T SQL, Technology

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  • OBIEE 11.1.1 - Disable Wrap Data Types in WebLogic Server 10.3.x

    - by Ahmed Awan
    By default, JDBC data type’s objects are wrapped with a WebLogic wrapper. This allows for features like debugging output and track connection usage to be done by the server. The wrapping can be turned off by setting this value to false. This improves performance, in some cases significantly, and allows for the application to use the native driver objects directly. Tip: How to Disable Wrapping in WLS Administration Console You can use the Administration Console to disable data type wrapping for following JDBC data sources in bifoundation_domain domain: Data Source Name bip_datasource mds-owsm EPMSystemRegistry   To disable wrapping for each JDBC data source (as stated in above table): 1.     If you have not already done so, in the Change Center of the Administration Console, click Lock & Edit. 2.     In the Domain Structure tree, expand Services, then select Data Sources. 3.     On the Summary of Data Sources page, click the data source name for example “mds-owsm”. 4.     Select the Configuration: Connection Pool tab. 5.     Scroll down and click Advanced to show the advanced connection pool options. 6.     In Wrap Data Types, deselect the checkbox to disable wrapping. 7.     Click Save. 8.     To activate these changes, in the Change Center of the Administration Console, click Activate Changes. Important Note: This change does not take effect immediately—it requires the server be restarted.

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  • Add Free Google Apps to Your Website or Blog

    - by Matthew Guay
    Would you like to have an email address from your own domain, but prefer Gmail’s interface and integration with Google Docs?  Here’s how you can add the free Google Apps Standard to your site and get the best of both worlds. Note: To signup for Google Apps and get it setup on your domain, you will need to be able to add info to your WordPress blog or change Domain settings manually. Getting Started Head to the Google Apps signup page (link below), and click the Get Started button on the right.  Note that we are signing up for the free Google Apps which allows a max of 50 users; if you need more than 50 email addresses for your domain, you can choose Premiere Edition instead for $50/year. Select that you are the Administrator of the domain, and enter the domain or subdomain you want to use with Google Apps.  Here we’re adding Google Apps to the techinch.com site, but we could instead add Apps to mail.techinch.com if needed…click Get Started. Enter your name, phone number, an existing email address, and other Administrator information.  The Apps signup page also includes some survey questions about your organization, but you only have to fill in the required fields. On the next page, enter a username and password for the administrator account.  Note that the user name will also be the administrative email address as [email protected]. Now you’re ready to authenticate your Google Apps account with your domain.  The steps are slightly different depending on whether your site is on WordPress.com or on your own hosting service or server, so we’ll show how to do it both ways.   Authenticate and Integrate Google Apps with WordPress.com To add Google Apps to a domain you have linked to your WordPress.com blog, select Change yourdomain.com CNAME record and click Continue. Copy the code under #2, which should be something like googleabcdefg123456.  Do not click the button at the bottom; wait until we’ve completed the next step.   Now, in a separate browser window or tab, open your WordPress Dashboard.  Click the arrow beside Upgrades, and select Domains from the menu. Click the Edit DNS link beside the domain name you’re adding to Google Apps. Scroll down to the Google Apps section, and paste your code from Google Apps into the verification code field.  Click Generate DNS records when you’re done. This will add the needed DNS settings to your records in the box above the Google Apps section.  Click Save DNS records. Now, go back to the Google Apps signup page, and click I’ve completed the steps above. Authenticate Google Apps on Your Own Server If your website is hosted on your own server or hosting account, you’ll need to take a few more steps to add Google Apps to your domain.  You can add a CNAME record to your domain host using the same information that you would use with a WordPress account, or you can upload an HTML file to your site’s main directory.  In this test we’re going to upload an HTML file to our site for verification. Copy the code under #1, which should be something like googleabcdefg123456.  Do not click the button at the bottom; wait until we’ve completed the next step first. Create a new HTML file and paste the code in it.  You can do this easily in Notepad: create a new document, paste the code, and then save as googlehostedservice.html.  Make sure to select the type as All Files or otherwise the file will have a .txt extension. Upload this file to your web server via FTP or a web dashboard for your site.  Make sure it is in the top level of your site’s directory structure, and try visiting it at yoursite.com/googlehostedservice.html. Now, go back to the Google Apps signup page, and click I’ve completed the steps above. Setup Your Email on Google Apps When this is done, your Google Apps account should be activated and ready to finish setting up.  Google Apps will offer to launch a guide to step you through the rest of the process; you can click Launch guide if you want, or click Skip this guide to continue on your own and go directly to the Apps dashboard.   If you choose to open the guide, you’ll be able to easily learn the ropes of Google Apps administration.  Once you’ve completed the tutorial, you’ll be taken to the Google Apps dashboard. Most of the Google Apps will be available for immediate use, but Email may take a bit more setup.  Click Activate email to get your Gmail-powered email running on your domain.    Add Google MX Records to Your Server You will need to add Google MX records to your domain registrar in order to have your mail routed to Google.  If your domain is hosted on WordPress.com, you’ve already made these changes so simply click I have completed these steps.  Otherwise, you’ll need to manually add these records before clicking that button.   Adding MX Entries is fairly easy, but the steps may depend on your hosting company or registrar.  With some hosts, you may have to contact support to have them add the MX records for you.  Our site’s host uses the popular cPanel for website administration, so here’s how we added the MX Entries through cPanel. Add MX Entries through cPanel Login to your site’s cPanel, and click the MX Entry link under Mail. Delete any existing MX Records for your domain or subdomain first to avoid any complications or interactions with Google Apps.  If you think you may want to revert to your old email service in the future, save a copy of the records so you can switch back if you need. Now, enter the MX Records that Google listed.  Here’s our account after we added all of the entries to our account. Finally, return to your Google Apps Dashboard and click the I have completed these steps button at the bottom of the page. Activating Service You’re now officially finished activating and setting up your Google Apps account.  Google will first have to check the MX records for your domain; this only took around an hour in our test, but Google warns it can take up to 48 hours in some cases. You may then see that Google is updating its servers with your account information.  Once again, this took much less time than Google’s estimate. When everything’s finished, you can click the link to access the inbox of your new Administrator email account in Google Apps. Welcome to Gmail … at your own domain!  All of the Google Apps work just the same in this version as they do in the public @gmail.com version, so you should feel right at home. You can return to the Google Apps dashboard from the Administrative email account by clicking the Manage this domain at the top right. In the Dashboard, you can easily add new users and email accounts, as well as change settings in your Google Apps account and add your site’s branding to your Apps. Your Google Apps will work just like their standard @gmail.com counterparts.  Here’s an example of an inbox customized with the techinch logo and a Gmail theme. Links to Remember Here are the common links to your Google Apps online.  Substitute your domain or subdomain for yourdomain.com. Dashboard https://www.google.com/a/cpanel/yourdomain.com Email https://mail.google.com/a/yourdomain.com Calendar https://www.google.com/calendar/hosted/yourdomain.com Docs https://docs.google.com/a/yourdomain.com Sites https://sites.google.com/a/yourdomain.com Conclusion Google Apps offers you great webapps and webmail for your domain, and let’s you take advantage of Google’s services while still maintaining the professional look of your own domain.  Setting up your account can be slightly complicated, but once it’s finished, it will run seamlessly and you’ll never have to worry about email or collaboration with your team again. Signup for the free Google Apps Standard Similar Articles Productive Geek Tips Mysticgeek Blog: Create Your Own Simple iGoogle GadgetAccess Your Favorite Google Services in Chrome the Easy WayRevo Uninstaller Pro [REVIEW]Mysticgeek Blog: A Look at Internet Explorer 8 Beta 1 on Windows XPFind Similar Websites in Google Chrome TouchFreeze Alternative in AutoHotkey The Icy Undertow Desktop Windows Home Server – Backup to LAN The Clear & Clean Desktop Use This Bookmarklet to Easily Get Albums Use AutoHotkey to Assign a Hotkey to a Specific Window Latest Software Reviews Tinyhacker Random Tips Xobni Plus for Outlook All My Movies 5.9 CloudBerry Online Backup 1.5 for Windows Home Server Snagit 10 Video preview of new Windows Live Essentials 21 Cursor Packs for XP, Vista & 7 Map the Stars with Stellarium Use ILovePDF To Split and Merge PDF Files TimeToMeet is a Simple Online Meeting Planning Tool Easily Create More Bookmark Toolbars in Firefox

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  • Connecting Linux to WatchGuard Firebox SSL (OpenVPN client)

    Recently, I got a new project assignment that requires to connect permanently to the customer's network through VPN. They are using a so-called SSL VPN. As I am using OpenVPN since more than 5 years within my company's network I was quite curious about their solution and how it would actually be different from OpenVPN. Well, short version: It is a disguised version of OpenVPN. Unfortunately, the company only offers a client for Windows and Mac OS which shouldn't bother any Linux user after all. OpenVPN is part of every recent distribution and can be activated in a couple of minutes - both client as well as server (if necessary). WatchGuard Firebox SSL - About dialog Borrowing some files from a Windows client installation Initially, I didn't know about the product, so therefore I went through the installation on Windows 8. No obstacles (and no restart despite installation of TAP device drivers!) here and the secured VPN channel was up and running in less than 2 minutes or so. Much appreciated from both parties - customer and me. Of course, this whole client package and my long year approved and stable installation ignited my interest to have a closer look at the WatchGuard client. Compared to the original OpenVPN client (okay, I have to admit this is years ago) this commercial product is smarter in terms of file locations during installation. You'll be able to access the configuration and key files below your roaming application data folder. To get there, simply enter '%AppData%\WatchGuard\Mobile VPN' in your Windows/File Explorer and confirm with Enter/Return. This will display the following files: Application folder below user profile with configuration and certificate files From there we are going to borrow four files, namely: ca.crt client.crt client.ovpn client.pem and transfer them to the Linux system. You might also be able to isolate those four files from a Mac OS client. Frankly, I'm just too lazy to run the WatchGuard client installation on a Mac mini only to find the folder location, and I'm going to describe why a little bit further down this article. I know that you can do that! Feedback in the comment section is appreciated. Configuration of OpenVPN (console) Depending on your distribution the following steps might be a little different but in general you should be able to get the important information from it. I'm going to describe the steps in Ubuntu 13.04 (Raring Ringtail). As usual, there are two possibilities to achieve your goal: console and UI. Let's what it is necessary to be done. First of all, you should ensure that you have OpenVPN installed on your system. Open your favourite terminal application and run the following statement: $ sudo apt-get install openvpn network-manager-openvpn network-manager-openvpn-gnome Just to be on the safe side. The four above mentioned files from your Windows machine could be copied anywhere but either you place them below your own user directory or you put them (as root) below the default directory: /etc/openvpn At this stage you would be able to do a test run already. Just in case, run the following command and check the output (it's the similar information you would get from the 'View Logs...' context menu entry in Windows: $ sudo openvpn --config client.ovpn Pay attention to the correct path to your configuration and certificate files. OpenVPN will ask you to enter your Auth Username and Auth Password in order to establish the VPN connection, same as the Windows client. Remote server and user authentication to establish the VPN Please complete the test run and see whether all went well. You can disconnect pressing Ctrl+C. Simplifying your life - authentication file In my case, I actually set up the OpenVPN client on my gateway/router. This establishes a VPN channel between my network and my client's network and allows me to switch machines easily without having the necessity to install the WatchGuard client on each and every machine. That's also very handy for my various virtualised Windows machines. Anyway, as the client configuration, key and certificate files are located on a headless system somewhere under the roof, it is mandatory to have an automatic connection to the remote site. For that you should first change the file extension '.ovpn' to '.conf' which is the default extension on Linux systems for OpenVPN, and then open the client configuration file in order to extend an existing line. $ sudo mv client.ovpn client.conf $ sudo nano client.conf You should have a similar content to this one here: dev tunclientproto tcp-clientca ca.crtcert client.crtkey client.pemtls-remote "/O=WatchGuard_Technologies/OU=Fireware/CN=Fireware_SSLVPN_Server"remote-cert-eku "TLS Web Server Authentication"remote 1.2.3.4 443persist-keypersist-tunverb 3mute 20keepalive 10 60cipher AES-256-CBCauth SHA1float 1reneg-sec 3660nobindmute-replay-warningsauth-user-pass auth.txt Note: I changed the IP address of the remote directive above (which should be obvious, right?). Anyway, the required change is marked in red and we have to create a new authentication file 'auth.txt'. You can give the directive 'auth-user-pass' any file name you'd like to. Due to my existing OpenVPN infrastructure my setup differs completely from the above written content but for sake of simplicity I just keep it 'as-is'. Okay, let's create this file 'auth.txt' $ sudo nano auth.txt and just put two lines of information in it - username on the first, and password on the second line, like so: myvpnusernameverysecretpassword Store the file, change permissions, and call openvpn with your configuration file again: $ sudo chmod 0600 auth.txt $ sudo openvpn --config client.conf This should now work without being prompted to enter username and password. In case that you placed your files below the system-wide location /etc/openvpn you can operate your VPNs also via service command like so: $ sudo service openvpn start client $ sudo service openvpn stop client Using Network Manager For newer Linux users or the ones with 'console-phobia' I'm going to describe now how to use Network Manager to setup the OpenVPN client. For this move your mouse to the systray area and click on Network Connections => VPN Connections => Configure VPNs... which opens your Network Connections dialog. Alternatively, use the HUD and enter 'Network Connections'. Network connections overview in Ubuntu Click on 'Add' button. On the next dialog select 'Import a saved VPN configuration...' from the dropdown list and click on 'Create...' Choose connection type to import VPN configuration Now you navigate to your folder where you put the client files from the Windows system and you open the 'client.ovpn' file. Next, on the tab 'VPN' proceed with the following steps (directives from the configuration file are referred): General Check the IP address of Gateway ('remote' - we used 1.2.3.4 in this setup) Authentication Change Type to 'Password with Certificates (TLS)' ('auth-pass-user') Enter User name to access your client keys (Auth Name: myvpnusername) Enter Password (Auth Password: verysecretpassword) and choose your password handling Browse for your User Certificate ('cert' - should be pre-selected with client.crt) Browse for your CA Certificate ('ca' - should be filled as ca.crt) Specify your Private Key ('key' - here: client.pem) Then click on the 'Advanced...' button and check the following values: Use custom gateway port: 443 (second value of 'remote' directive) Check the selected value of Cipher ('cipher') Check HMAC Authentication ('auth') Enter the Subject Match: /O=WatchGuard_Technologies/OU=Fireware/CN=Fireware_SSLVPN_Server ('tls-remote') Finally, you have to confirm and close all dialogs. You should be able to establish your OpenVPN-WatchGuard connection via Network Manager. For that, click on the 'VPN Connections => client' entry on your Network Manager in the systray. It is advised that you keep an eye on the syslog to see whether there are any problematic issues that would require some additional attention. Advanced topic: routing As stated above, I'm running the 'WatchGuard client for Linux' on my head-less server, and since then I'm actually establishing a secure communication channel between two networks. In order to enable your network clients to get access to machines on the remote side there are two possibilities to enable that: Proper routing on both sides of the connection which enables both-direction access, or Network masquerading on the 'client side' of the connection Following, I'm going to describe the second option a little bit more in detail. The Linux system that I'm using is already configured as a gateway to the internet. I won't explain the necessary steps to do that, and will only focus on the additional tweaks I had to do. You can find tons of very good instructions and tutorials on 'How to setup a Linux gateway/router' - just use Google. OK, back to the actual modifications. First, we need to have some information about the network topology and IP address range used on the 'other' side. We can get this very easily from /var/log/syslog after we established the OpenVPN channel, like so: $ sudo tail -n20 /var/log/syslog Or if your system is quite busy with logging, like so: $ sudo less /var/log/syslog | grep ovpn The output should contain PUSH received message similar to the following one: Jul 23 23:13:28 ios1 ovpn-client[789]: PUSH: Received control message: 'PUSH_REPLY,topology subnet,route 192.168.1.0 255.255.255.0,dhcp-option DOMAIN ,route-gateway 192.168.6.1,topology subnet,ping 10,ping-restart 60,ifconfig 192.168.6.2 255.255.255.0' The interesting part for us is the route command which I highlighted already in the sample PUSH_REPLY. Depending on your remote server there might be multiple networks defined (172.16.x.x and/or 10.x.x.x). Important: The IP address range on both sides of the connection has to be different, otherwise you will have to shuffle IPs or increase your the netmask. {loadposition content_adsense} After the VPN connection is established, we have to extend the rules for iptables in order to route and masquerade IP packets properly. I created a shell script to take care of those steps: #!/bin/sh -eIPTABLES=/sbin/iptablesDEV_LAN=eth0DEV_VPNS=tun+VPN=192.168.1.0/24 $IPTABLES -A FORWARD -i $DEV_LAN -o $DEV_VPNS -d $VPN -j ACCEPT$IPTABLES -A FORWARD -i $DEV_VPNS -o $DEV_LAN -s $VPN -j ACCEPT$IPTABLES -t nat -A POSTROUTING -o $DEV_VPNS -d $VPN -j MASQUERADE I'm using the wildcard interface 'tun+' because I have multiple client configurations for OpenVPN on my server. In your case, it might be sufficient to specify device 'tun0' only. Simplifying your life - automatic connect on boot Now, that the client connection works flawless, configuration of routing and iptables is okay, we might consider to add another 'laziness' factor into our setup. Due to kernel updates or other circumstances it might be necessary to reboot your system. Wouldn't it be nice that the VPN connections are established during the boot procedure? Yes, of course it would be. To achieve this, we have to configure OpenVPN to automatically start our VPNs via init script. Let's have a look at the responsible 'default' file and adjust the settings accordingly. $ sudo nano /etc/default/openvpn Which should have a similar content to this: # This is the configuration file for /etc/init.d/openvpn## Start only these VPNs automatically via init script.# Allowed values are "all", "none" or space separated list of# names of the VPNs. If empty, "all" is assumed.# The VPN name refers to the VPN configutation file name.# i.e. "home" would be /etc/openvpn/home.conf#AUTOSTART="all"#AUTOSTART="none"#AUTOSTART="home office"## ... more information which remains unmodified ... With the OpenVPN client configuration as described above you would either set AUTOSTART to "all" or to "client" to enable automatic start of your VPN(s) during boot. You should also take care that your iptables commands are executed after the link has been established, too. You can easily test this configuration without reboot, like so: $ sudo service openvpn restart Enjoy stable VPN connections between your Linux system(s) and a WatchGuard Firebox SSL remote server. Cheers, JoKi

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  • How can I use GPRename's regex feature to reinsert the matched-group into the 'replace'?

    - by David Thomas
    I've been using GPRename to batch-rename files; this is rather more efficient than individually correcting each file, but still seems to be less efficient than it might be, primarily because either I don't understand the regex syntax used, or because the regex implementation is incomplete1 Given a list of files of the following syntax: (01) - title of file1.avi (02) - title of file2.avi (03) - title of file3.avi I attempted to use the 'replace' (with the regex option selected, the case-sensitive option deselected): (\(\d{2}\)) The preview then shows (given that I've specified no 'replace with' option as yet): title of file1.avi title of file2.avi title of file3.avi Which is great, clearly the regex is identifying the correct group (the (01)). Now, what I was hoping to do (using the JavaScript syntax) in the 'replace with' option is use: $1 (I also tried using '$1', \1 and '\1') This was just to check that I could access the matched group, and it seems I can't, the matched group is, as I suppose might be expected, replaced with the literal replacement string. So, my question: is it possible to match a particular group of characters, in this case the numbers within the brackets, and then insert those into the replacement string? Therefore: (01) title of file1.avi (02) title of file2.avi (03) title of file3.avi Becomes: 01 title of file1.avi 02 title of file2.avi 03 title of file3.avi I absolutely suspect the former, personally.

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  • Google tweets – Now search twitter archives using Google

    - by samsudeen
    Google has launched a Twitter archive service which allows you to  search tweets in real time as well as on its huge public archive (remember Twitter crossed 10 billionth tweet last month). The search results are displayed as tweets with twitter logo. To explore the twitter search go to Google.com homepage  and select   “Show options” on the search results page, then select “Updates.”.  The search is similar to the Google search with options to dig through the tweets by timeframe. You can explore results by zooming through a particular time range  or date. In addition to the time chart, it also displays the relative volume of an activity on Twitter about the topic. as you can see there is a spike about GSLV launch after 3 PM today.There is also a short cut link “Now” on the left corner which displays the latest results on the topics searched.The tweets also gets refreshed automatically.   Considering the huge volume of activity (50 million messages per day) on twitter, the archive is going to more and bigger. By providing such feature Google has once again proved it is way ahead of others in search Related Posts:None FoundJoin us on Facebook to read all our stories right inside your Facebook news feed.

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  • Google tweets – Now search twitter archives using Google

    - by samsudeen
    Google has launched a Twitter archive service which allows you to  search tweets in real time as well as on its huge public archive (remember Twitter crossed 10 billionth tweet last month). The search results are displayed as tweets with twitter logo. To explore the twitter search go to Google.com homepage  and select   “Show options” on the search results page, then select “Updates.”.  The search is similar to the Google search with options to dig through the tweets by timeframe. You can explore results by zooming through a particular time range  or date. In addition to the time chart, it also displays the relative volume of an activity on Twitter about the topic. as you can see there is a spike about GSLV launch after 3 PM today.There is also a short cut link “Now” on the left corner which displays the latest results on the topics searched.The tweets also gets refreshed automatically.   Considering the huge volume of activity (50 million messages per day) on twitter, the archive is going to more and bigger. By providing such feature Google has once again proved it is way ahead of others in search Related Posts:None FoundJoin us on Facebook to read all our stories right inside your Facebook news feed.

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  • If-Modified-Since vs If-None-Match

    - by Roger
    This question is based on this article response header HTTP/1.1 200 OK Last-Modified: Tue, 12 Dec 2006 03:03:59 GMT ETag: "10c24bc-4ab-457e1c1f" Content-Length: 12195 request header GET /i/yahoo.gif HTTP/1.1 Host: us.yimg.com If-Modified-Since: Tue, 12 Dec 2006 03:03:59 GMT If-None-Match: "10c24bc-4ab-457e1c1f" HTTP/1.1 304 Not Modified In this case browser is sending both If-None-Match and If-Modified-Since. My question is on the server side do I need to match BOTH etag and If-Modified-Since before I send 304. Or Should I just look at etag and send 304 if etag is a match. In this case I am ignoring If-Modified-Since .

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  • Revisiting ANTS Performance Profiler 7.4

    - by James Michael Hare
    Last year, I did a small review on the ANTS Performance Profiler 6.3, now that it’s a year later and a major version number higher, I thought I’d revisit the review and revise my last post. This post will take the same examples as the original post and update them to show what’s new in version 7.4 of the profiler. Background A performance profiler’s main job is to keep track of how much time is typically spent in each unit of code. This helps when we have a program that is not running at the performance we expect, and we want to know where the program is experiencing issues. There are many profilers out there of varying capabilities. Red Gate’s typically seem to be the very easy to “jump in” and get started with very little training required. So let’s dig into the Performance Profiler. I’ve constructed a very crude program with some obvious inefficiencies. It’s a simple program that generates random order numbers (or really could be any unique identifier), adds it to a list, sorts the list, then finds the max and min number in the list. Ignore the fact it’s very contrived and obviously inefficient, we just want to use it as an example to show off the tool: 1: // our test program 2: public static class Program 3: { 4: // the number of iterations to perform 5: private static int _iterations = 1000000; 6: 7: // The main method that controls it all 8: public static void Main() 9: { 10: var list = new List<string>(); 11: 12: for (int i = 0; i < _iterations; i++) 13: { 14: var x = GetNextId(); 15: 16: AddToList(list, x); 17: 18: var highLow = GetHighLow(list); 19: 20: if ((i % 1000) == 0) 21: { 22: Console.WriteLine("{0} - High: {1}, Low: {2}", i, highLow.Item1, highLow.Item2); 23: Console.Out.Flush(); 24: } 25: } 26: } 27: 28: // gets the next order id to process (random for us) 29: public static string GetNextId() 30: { 31: var random = new Random(); 32: var num = random.Next(1000000, 9999999); 33: return num.ToString(); 34: } 35: 36: // add it to our list - very inefficiently! 37: public static void AddToList(List<string> list, string item) 38: { 39: list.Add(item); 40: list.Sort(); 41: } 42: 43: // get high and low of order id range - very inefficiently! 44: public static Tuple<int,int> GetHighLow(List<string> list) 45: { 46: return Tuple.Create(list.Max(s => Convert.ToInt32(s)), list.Min(s => Convert.ToInt32(s))); 47: } 48: } So let’s run it through the profiler and see what happens! Visual Studio Integration First, let’s look at how the ANTS profilers integrate with Visual Studio’s menu system. Once you install the ANTS profilers, you will get an ANTS menu item with several options: Notice that you can either Profile Performance or Launch ANTS Performance Profiler. These sound similar but achieve two slightly different actions: Profile Performance: this immediately launches the profiler with all defaults selected to profile the active project in Visual Studio. Launch ANTS Performance Profiler: this launches the profiler much the same way as starting it from the Start Menu. The profiler will pre-populate the application and path information, but allow you to change the settings before beginning the profile run. So really, the main difference is that Profile Performance immediately begins profiling with the default selections, where Launch ANTS Performance Profiler allows you to change the defaults and attach to an already-running application. Let’s Fire it Up! So when you fire up ANTS either via Start Menu or Launch ANTS Performance Profiler menu in Visual Studio, you are presented with a very simple dialog to get you started: Notice you can choose from many different options for application type. You can profile executables, services, web applications, or just attach to a running process. In fact, in version 7.4 we see two new options added: ASP.NET Web Application (IIS Express) SharePoint web application (IIS) So this gives us an additional way to profile ASP.NET applications and the ability to profile SharePoint applications as well. You can also choose your level of detail in the Profiling Mode drop down. If you choose Line-Level and method-level timings detail, you will get a lot more detail on the method durations, but this will also slow down profiling somewhat. If you really need the profiler to be as unintrusive as possible, you can change it to Sample method-level timings. This is performing very light profiling, where basically the profiler collects timings of a method by examining the call-stack at given intervals. Which method you choose depends a lot on how much detail you need to find the issue and how sensitive your program issues are to timing. So for our example, let’s just go with the line and method timing detail. So, we check that all the options are correct (if you launch from VS2010, the executable and path are filled in already), and fire it up by clicking the [Start Profiling] button. Profiling the Application Once you start profiling the application, you will see a real-time graph of CPU usage that will indicate how much your application is using the CPU(s) on your system. During this time, you can select segments of the graph and bookmark them, giving them mnemonic names. This can be useful if you want to compare performance in one part of the run to another part of the run. Notice that once you select a block, it will give you the call tree breakdown for that selection only, and the relative performance of those calls. Once you feel you have collected enough information, you can click [Stop Profiling] to stop the application run and information collection and begin a more thorough analysis. Analyzing Method Timings So now that we’ve halted the run, we can look around the GUI and see what we can see. By default, the times are shown in terms of percentage of time of the total run of the application, though you can change it in the View menu item to milliseconds, ticks, or seconds as well. This won’t affect the percentages of methods, it only affects what units the times are shown. Notice also that the major hotspot seems to be in a method without source, ANTS Profiler will filter these out by default, but you can right-click on the line and remove the filter to see more detail. This proves especially handy when a bottleneck is due to a method in the BCL. So now that we’ve removed the filter, we see a bit more detail: In addition, ANTS Performance Profiler gives you the ability to decompile the methods without source so that you can dive even deeper, though typically this isn’t necessary for our purposes. When looking at timings, there are generally two types of timings for each method call: Time: This is the time spent ONLY in this method, not including calls this method makes to other methods. Time With Children: This is the total of time spent in both this method AND including calls this method makes to other methods. In other words, the Time tells you how much work is being done exclusively in this method, and the Time With Children tells you how much work is being done inclusively in this method and everything it calls. You can also choose to display the methods in a tree or in a grid. The tree view is the default and it shows the method calls arranged in terms of the tree representing all method calls and the parent method that called them, etc. This is useful for when you find a hot-spot method, you can see who is calling it to determine if the problem is the method itself, or if it is being called too many times. The grid method represents each method only once with its totals and is useful for quickly seeing what method is the trouble spot. In addition, you can choose to display Methods with source which are generally the methods you wrote (as opposed to native or BCL code), or Any Method which shows not only your methods, but also native calls, JIT overhead, synchronization waits, etc. So these are just two ways of viewing the same data, and you’re free to choose the organization that best suits what information you are after. Analyzing Method Source If we look at the timings above, we see that our AddToList() method (and in particular, it’s call to the List<T>.Sort() method in the BCL) is the hot-spot in this analysis. If ANTS sees a method that is consuming the most time, it will flag it as a hot-spot to help call out potential areas of concern. This doesn’t mean the other statistics aren’t meaningful, but that the hot-spot is most likely going to be your biggest bang-for-the-buck to concentrate on. So let’s select the AddToList() method, and see what it shows in the source window below: Notice the source breakout in the bottom pane when you select a method (from either tree or grid view). This shows you the timings in this method per line of code. This gives you a major indicator of where the trouble-spot in this method is. So in this case, we see that performing a Sort() on the List<T> after every Add() is killing our performance! Of course, this was a very contrived, duh moment, but you’d be surprised how many performance issues become duh moments. Note that this one line is taking up 86% of the execution time of this application! If we eliminate this bottleneck, we should see drastic improvement in the performance. So to fix this, if we still wanted to maintain the List<T> we’d have many options, including: delay Sort() until after all Add() methods, using a SortedSet, SortedList, or SortedDictionary depending on which is most appropriate, or forgoing the sorting all together and using a Dictionary. Rinse, Repeat! So let’s just change all instances of List<string> to SortedSet<string> and run this again through the profiler: Now we see the AddToList() method is no longer our hot-spot, but now the Max() and Min() calls are! This is good because we’ve eliminated one hot-spot and now we can try to correct this one as well. As before, we can then optimize this part of the code (possibly by taking advantage of the fact the list is now sorted and returning the first and last elements). We can then rinse and repeat this process until we have eliminated as many bottlenecks as possible. Calls by Web Request Another feature that was added recently is the ability to view .NET methods grouped by the HTTP requests that caused them to run. This can be helpful in determining which pages, web services, etc. are causing hot spots in your web applications. Summary If you like the other ANTS tools, you’ll like the ANTS Performance Profiler as well. It is extremely easy to use with very little product knowledge required to get up and running. There are profilers built into the higher product lines of Visual Studio, of course, which are also powerful and easy to use. But for quickly jumping in and finding hot spots rapidly, Red Gate’s Performance Profiler 7.4 is an excellent choice. Technorati Tags: Influencers,ANTS,Performance Profiler,Profiler

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  • SQLAuthority News – Best SQLAuthority Posts of May

    - by pinaldave
    Month of May is always interesting and full of enthusiasm. Lots of good articles shared and lots of enthusiast communication on technology. This month we had 140 Character Cartoon Challenge Winner. We also had interesting conversation on what kind of lock WITH NOLOCK takes on objects as well. A quick tutorial on how to import CSV files into Database using SSIS started few other related questions. I also had fun time with community activities. I attended MVP Open Day. Vijay Raj also took awesome photos of my daughter – Shaivi. I have gain my faith back in Social Media and have created my Facebook Page, if you like SQLAuthority.com I request you to Like Facebook page as well. I am very active on twitter (@pinaldave) and answer lots of technical question if I am online during that time. During this month couple of old thing, I did learn by accident 1) Restart and Shutdown Remote Computer 2) SSMS has web browser. If you have made it till here – I suggest you to take participation in very interesting conversation here – Why SELECT * throws an error but SELECT COUNT(*) does not? Reference: Pinal Dave (http://blog.SQLAuthority.com) Filed under: About Me, Pinal Dave, PostADay, SQL, SQL Authority, SQL Query, SQL Server, SQL Tips and Tricks, T SQL, Technology

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  • Getting Started with Employee Info Starter Kit (v4.0.0)

    - by joycsharp
    The new release of Employee Info Starter Kit contains lots of exciting features available in Visual Studio 2010 and .NET 4.0. To get started with the new version, you will need less than 5 minutes. Minimum System Requirements Before getting started, please make sure you have installed Visual Studio 2010 RC (or higher) and Sql Server 2005 Express edition (or higher installed on your machine. Running the Starter Kit for First Time 1. Download the starter kit 4.0.0 version form here and extract it. 2. Go to <extraction folder>\Source\Eisk.Solution and click the solution file 3. From the solution explorer, right click the “Eisk.Web” web site project node and select “Set as Startup Project” and hit Ctrl + F5   4. You will be prompted to install database, just follow the instruction. That’s it! You are ready to use this starter kit. Running the Tests Employee Info Starter Kit contains a infrastructure for Integration and Unit Testing, by utilizing cool test tools in Visual Studio 2010. Once you complete the steps, mentioned above, take a minute to run the test cases on the fly. 1. From the solution explorer, to go “Solution Items\e-i-s-k-2010.vsmdi” and click it. You will see the available Tests in the Visual Studio Test Lists. Select all, except the “Load Tests” node (since Load Tests takes a bit time) 2. Click “Run Checked Tests” control from the upper left corner. You will see the tests running and finally the status of the tests, which indicates the current health of you application from different scenarios. Technorati Tags: asp.net,architecture,starter kit,employee info starter kit,visual studio 2010,.net 4.0,entity framework

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  • Getting Started with Employee Info Starter Kit (v4.0.0)

    - by Mohammad Ashraful Alam
    The new release of Employee Info Starter Kit contains lots of exciting features available in Visual Studio 2010 and .NET 4.0. To get started with the new version, you will need less than 5 minutes. Minimum System Requirements Before getting started, please make sure you have installed Visual Studio 2010 RC (or higher) and Sql Server 2005 Express edition (or higher installed on your machine. Running the Starter Kit for First Time 1. Download the starter kit 4.0.0 version form here and extract it. 2. Go to <extraction folder>\Source\Eisk.Solution and click the solution file 3. From the solution explorer, right click the “Eisk.Web” web site project node and select “Set as Startup Project” and hit Ctrl + F5   4. You will be prompted to install database, just follow the instruction. That’s it! You are ready to use this starter kit. Running the Tests Employee Info Starter Kit contains a infrastructure for Integration and Unit Testing, by utilizing cool test tools in Visual Studio 2010. Once you complete the steps, mentioned above, take a minute to run the test cases on the fly. 1. From the solution explorer, to go “Solution Items\e-i-s-k-2010.vsmdi” and click it. You will see the available Tests in the Visual Studio Test Lists. Select all, except the “Load Tests” node (since Load Tests takes a bit time) 2. Click “Run Checked Tests” control from the upper left corner. You will see the tests running and finally the status of the tests, which indicates the current health of you application from different scenarios. Technorati Tags: asp.net,architecture,starter kit,employee info starter kit,visual studio 2010,.net 4.0,entity framework

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  • Using Oracle Enterprise Manager Ops Center to Update Solaris via Live Upgrade

    - by LeonShaner
    Introduction: This Oracle Enterprise Manager Ops Center blog entry provides tips for using Ops Center to update Solaris using Live Upgrade on Solaris 10 and Boot Environments on Solaris 11. Why use Live Upgrade? Live Upgrade (LU) can significantly reduce downtime associated with patching Live Upgrade avoids dropping to single-user mode for long periods of time during patching Live Upgrade relies on an Alternate Boot Environment (ABE)/(BE), which is patched while in multi-user mode; thereby allowing normal system operations to continue with the active BE, while the alternate BE is being patched Activating an newly patched (A)BE is essentially a reboot; therefore the downtime is ~= reboot Admins can easily revert to the prior Boot Environment (BE) as a safeguard / fallback. Why use Ops Center to patch via Live Upgrade, Alternate Boot Environments, and Solaris 11 equivalents? All the benefits of Ops Center's extensive patch and package knowledge base can be leveraged on top of Live Upgrade Ops Center can orchestrate patching based on Live Upgrade and Solaris 11 features, which all works together to minimize downtime Ops Centers advanced inventory and reporting features assurance that each OS is updated to a verifiable, consistent standard, rather than relying on ad-hoc (error prone) procedures and scripts Ops Center gives admins control over the boot environment specifications or they can let Ops Center decide when a BE is necessary, thereby reducing complexity and lowering the opportunity for user error Preparing to use Live Upgrade-like features in Solaris 11 Requirements and information you should know: Global Zone Root file-systems must be separate from Solaris Container / Zone filesystems Solaris 11 has features which are similar in concept to Live Upgrade on Solaris 10, but differ greatly in implementationImportant distinctions: Solaris 11 assumes ZFS root Solaris 11 adds Boot Environments (BE's) as an integrated feature (see beadm) Solaris 11 BE's avoid single-user patching (vs. Solaris 10 w/ ZFS snapshot=ABE). Solaris 11 Image Packaging System (IPS) has hooks for BE creation, as needed Solaris 11 allows pkgs to be installed + upgraded in alternate BE (e.g. instead of the live system) but it is controlled on a per-pkg basis Boot Environments are activated across a reboot; instead of spending long periods installing + upgrading packages in single user mode. Fallback to a prior BE is a function of the BE infrastructure (a la beadm). (Generally) Reboot + BE activation can be much much faster on Solaris 11 Preparing to use Live Upgrade on Solaris 10 Requirements and information you should know: Global Zone Root file-systems must be separate from Solaris Container / Zone filesystems Live Upgrade Pre-requisite patches must be applied before the first Live Upgrade Alternate Boot Environments are created (see "Pre-requisite Patches" section, below...) Solaris 10 Update 6 or newer on ZFS root is the practical starting point for Live Upgrade Live Upgrade with ZFS root is far more straight-forward than any scheme based on Alternative Boot Environments in slices or temporarily breaking mirrors Use Solaris best practices to upgrade the OS to at least Solaris 10 Update 4 (outside of Ops Center) UFS root can (technically) be used, but it is significantly more involved (e.g. discouraged) -- there are many reasons to move to ZFS while going through the process to update to Solaris 10 Update 6 or newer (out side of Ops Center) Recommendation: Start with Solaris 10 Update 6 or newer on ZFS root Recommendation: Start with Ops Center 12c or newer Ops Center 12c can automatically create your ABE's for you, without the need for custom scripts Ops Center 12c Update 2 avoids kernel panic on unpatched Solaris 10 update 9 (and older) -- unrelated to Live Upgrade, but more on the issue, below. NOTE: There is no magic!  If you have systems running Solaris 10 Update 5 or older on UFS root, and you don't know how to get them updated to Solaris 10 on ZFS root, then there are services available from Oracle Advanced Customer Support (ACS), which specialize in this area. Live Upgrade Pre-requisite Patches (Solaris 10) Certain Live Upgrade related patches must be present before the first Live Upgrade ABE's are created on Solaris 10.Use the following MOS Search String to find the “living document” that outlines the required patch minimums, which are necessary before using any Live Upgrade features: Solaris Live Upgrade Software Patch Requirements(Click above – the link is valid as of this writing, but search in MOS for the same "Solaris Live Upgrade Software Patch Requirements" string if necessary) It is a very good idea to check the document periodically and adapt to its contents, accordingly.IMPORTANT:  In case it wasn't clear in the above document, some direct patching of the active OS, including a reboot, may be required before Live Upgrade can be successfully used the first time.HINT: You can use Ops Center to determine what to expect for a given system, and to schedule the “pre-patching” during a maintenance window if necessary. Preparing to use Ops Center Discover + Manage (Install + Configure the Ops Center agent in) each Global Zone Recommendation:  Begin by using OCDoctor --agent-prereq to determine whether OS meets OC prerequisites (resolve any issues) See prior requirements and recommendations w.r.t. starting with Solaris 10 Update 6 or newer on ZFS (or at least Solaris 10 Update 4 on UFS, with caveats) WARNING: Systems running unpatched Solaris 10 update 9 (or older) should run the Ops Center 12c Update 2 agent to avoid a potential kernel panic The 12c Update 2 agent will check patch minimums and disable certain process accounting features if the kernel is not sufficiently patched to avoid the panic SPARC: 142900-05 Obsoleted by: 142900-06 SunOS 5.10: kernel patch 10 Oracle Solaris on SPARC (32-bit) X64: 142901-05 Obsoleted by: 142901-06 SunOS 5.10_x86: kernel patch 10 Oracle Solaris on x86 (32-bit) OR SPARC: 142909-17 SunOS 5.10: kernel patch 10 Oracle Solaris on SPARC (32-bit) X64: 142910-17 SunOS 5.10_x86: kernel patch 10 Oracle Solaris on x86 (32-bit) Ops Center 12c (initial release) and 12c Update 1 agent can also be safely used with a workaround (to be performed BEFORE installing the agent): # mkdir -p /etc/opt/sun/oc # echo "zstat_exacct_allowed=false" > /etc/opt/sun/oc/zstat.conf # chmod 755 /etc/opt/sun /etc/opt/sun/oc # chmod 644 /etc/opt/sun/oc/zstat.conf # chown -Rh root:sys /etc/opt/sun/oc NOTE: Remove the above after patching the OS sufficiently, or after upgrading to the 12c Update 2 agent Using Ops Center to apply Live Upgrade-related Pre-Patches (Solaris 10)Overview: Create an OS Update Profile containing the minimum LU-related pre-patches, based on the Solaris Live Upgrade Software Patch Requirements, previously mentioned. SIMULATE the deployment of the LU-related pre-patches Observe whether any of the LU-related pre-patches will require a reboot The job details for each Global Zone will advise whether a reboot step will be required ACTUALLY deploy the LU-related pre-patches, according to your change control process (e.g. if no reboot, maybe okay to do now; vs. must do later because of the reboot). You can schedule the job to occur later, during a maintenance window Check the job status for each node, resolving any issues found Once the LU-related pre-patches are applied, you can Ops Center to patch using Live Upgrade on Solaris 10 Using Ops Center to patch Solaris 10 with LU/ABE's -- the GOODS!(this is the heart of the tip): Create an OS Update Profile containing the patches that make up your standard build Use Solaris Baselines when possible Add other individual patches as needed ACTUALLY deploy the OS Update Profile Specify the appropriate Live Upgrade options, e.g. Synchronize the active BE to the alternate BE before patching Do not activate the BE after patching Check the job status for each node, resolving any issues found Activate the newly patched BE according to your change control process Activate = Reboot to the ABE, making the ABE the new active BE Ops Center does not separate LU activate from reboot, so expect a reboot! Check the job status for each node, resolving any issues found Examples (w/Screenshots) Solaris 10 and Live Upgrade: Auto-Create the Alternate Boot Environment (ZFS root only) ABE to be created on ZFS with name S10_12_07REC (Example) Uses built in feature to call “lucreate -n S10_12_07REC” behind scenes if not already present NOTE: Leave “lucreate” params blank (if you do specify options, the will be appended after -n $ABEName) Solaris 10 and Live Upgrade: Alternate Boot Environment Creation via Operational Profile (script) The Alternate Boot Environment is to be created via custom, user-supplied script, which does whatever is needed for the system where Live Upgrade will be used. Operational Profile, which provides the script to create an ABE: Very similar to the automatic case, but with a Script (Operational Profile), which is used to create the ABE Relies on user-supplied script in the form of an Operational Profile Could be used to prepare an ABE based on a UFS root in a slice, or on a separate device (e.g. by breaking a mirror first) – it is up to the script author to do the right thing! EXAMPLE: Same result as the ZFS case, but illustrating the Operational Profile (e.g. script) approach to call: # lucreate -n S10_1207REC NOTE: OC special variable is $ABEName Boot Environment Profile, which references the Operational Profile Script = Operational Profile on this screen Refers to Operational Profile shown in the previous section The user-supplied S10_Create_BE Operational Profile will be run The Operational Profile must send a non-zero exit code if there is a problem (so that the OS Update job will not proceed) Solaris 10 OS Update Profile (to provide the actual patch specifications) Solaris 10 Baseline “Recommended” chosen for “Install” Solaris 10 OS Update Plan (two-steps in this case) “Create a Boot Environment” + “Update OS” are chosen. Using Ops Center to patch Solaris 11 with Boot Environments (as needed) Create a Solaris 11 OS Update Profile containing the packages that make up your standard build ACTUALLY deploy the Solaris 11 OS Update Profile BE will be created if needed (or you can stipulate no BE) BE name will be auto-generated (if needed), or you may specify a BE name Check the job status for each node, resolving any issues found Check if a BE was created; if so, activate the new BE Activate = Reboot to the BE, making the new BE the active BE Ops Center does not separate BE activate from reboot NOTE: Not every Solaris 11 OS Update will require a new BE, so a reboot may not be necessary. Solaris 11: Auto BE Create (as Needed -- let Ops Center decide) BE to be created as needed BE to be named automatically Reboot (if necessary) deferred to separate step Solaris 11: OS Profile Solaris 11 “entire” chosen for a particular SRU Solaris 11: OS Update Plan (w/BE)  “Create a Boot Environment” + “Update OS” are chosen. Summary: Solaris 10 Live Upgrade, Alternate Boot Environments, and their equivalents on Solaris 11 can be very powerful tools to help minimize the downtime associated with updating your servers.  For very old Solaris, there are some important prerequisites to adhere to, but once the initial preparation is complete, Live Upgrade can be used going forward.  For Solaris 11, the built-in Boot Environment handling is leveraged directly by the Image Packaging System, and the result is a much more straight forward way to patch, and far fewer prerequisites to satisfy in getting there.  Ops Center simplifies using either approach, and helps you improve consistency from system to system, which ultimately helps you improve the overall up-time across all the Solaris systems in your environment. Please let us know what you think?  Until next time...\Leon-- Leon Shaner | Senior IT/Product ArchitectSystems Management | Ops Center Engineering @ Oracle The views expressed on this [blog; Web site] are my own and do not necessarily reflect the views of Oracle. For more information, please go to Oracle Enterprise Manager  web page or  follow us at :  Twitter | Facebook | YouTube | Linkedin | Newsletter

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